ni 43-101 pre-feasibility study – kitsault molybdenum

7/24/2019 NI 43-101 Pre-feasibility Study – Kitsault Molybdenum

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Report to:

NI 43-101 Pre-feasibility Study –Avanti Mining Inc.Kitsault Molybdenum PropertyBritish Columbia, Canada

Document No. 0954780100-REP-R0007-02



0954780100-REP-R0007-01

Report to:

NI 43-101 PRE-FEASIBILITY STUDY –

AVANTI MINING INC. KITSAULTMOLYBDENUM PROPERTYBRITISH COLUMBIA, CANADA

DECEMBER 15, 2009

Prepared by Frank Grills, P.Eng., M.Sc.(Eng), PMP, Wardrop Engineering Inc.

Marinus Andre de Ruijter, P.Eng., Wardrop Engineering Inc.

Miloje Vicentijevic, P.Eng., M.Eng., Wardrop Engineering Inc.

Jeffrey Volk, CPG, FAusIMM, M.Sc., SRK Consulting (U.S.) Inc.

Michael Levy, P.E., P.G., SRK Consulting (U.S.) Inc.

Peter Healey, P.Eng., SRK Consulting (Canada) Inc.

Stephen Day, P.Geo., SRK Consulting (Canada) Inc.

Michael Royle, M.App.Sci., P.Geo.(BC), SRK Consulting (Canada) Inc.

Ken J. Brouwer, P.Eng., Knight Piésold Ltd.

Harold Rolf Schmitt, M.Sc., P.Geo., Rescan Environmental Services Ltd.

Deepak Malhotra, MMSA, Ph.D., M.Sc., Resource Development Inc.

FG/alm

Suite 800, 555 West Hastings Street, Vancouver, British Columbia V6B 1M1Phone: 604-408-3788 Fax: 604-408-3722 E-mail: [email protected]



0954780100-REP-R0007-02

N O T I C E

This report was prepared for Avanti Mining Inc. (Avanti) by Wardrop Engineering Inc. (Wardrop),Rescan Environmental Services Ltd. (Rescan), Knight Piésold Ltd. (KP), Resource

Development Inc. (RDi), SRK Consulting (Canada) Inc. (SRK Canada) and SRK Consulting(U.S.) Inc. (SRK US), (collectively the Project Consultants). This document is meant to be read

as a whole. This document contains the expression of the professional opinion of Wardrop,

Rescan, KP, RDi, SRK Canada and SRK (U.S.) based on (i) information available at the time of preparation, (ii) data supplied by outside sources, (iii) conclusions of other technical specialistsnamed in this report, and (iv) the assumptions, conditions, and qualifications in this report. The

quality of the information, conclusions, and estimates contained herein are based on industrystandards for engineering and evaluation of a mineral project, and are consistent with the

intended level of accuracy for a Pre-feasibility Study.



i 0954780100-REP-R0007-02

T A B L E O F C O N T E N T S

1.0 SUMMARY ............................................................................................................................ 1-1

1.1 INTRODUCTION ........................................................................................................................1-1

1.2 PROPERTY DESCRIPTION AND LOCATION................................................................................1-2

1.3 OWNERSHIP.............................................................................................................................1-3

1.4 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY .......1-6

1.5 HISTORY..................................................................................................................................1-8

1.6 GEOLOGICAL SETTING...........................................................................................................1-10

1.7 DEPOSIT TYPES.....................................................................................................................1-11

1.8 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATE......................................................1-111.8.1 MINERAL RESOURCE ESTIMATE ..........................................................................1-111.8.2 MINERAL RESERVE ESTIMATE.............................................................................1-13

1.9 OVERALL GENERAL ARRANGEMENT ......................................................................................1-14

1.10 MINERAL PROCESSING AND METALLURGICAL TESTING.........................................................1-161.10.1 METALLURGICAL TESTWORK TIME PERIODS .......................................................1-161.10.2 MINERALOGICAL EXAMINATION............................................................................1-161.10.3 HISTORICAL METALLURGICAL REVIEW.................................................................1-161.10.4 RDI TEST PROGRAM ...........................................................................................1-18

1.11 MILL.......................................................................................................................................1-191.11.1 MAJOR DESIGN CRITERIA....................................................................................1-19

1.11.2 PROCESS

PLANT

DESCRIPTION

...........................................................................1-201.12 MINING ..................................................................................................................................1-22

1.13 PIT HYDROGEOLOGY .............................................................................................................1-24

1.14 GEOTECHNICAL.....................................................................................................................1-25

1.15 TAILING, WASTE ROCK, AND WATER MANAGEMENT..............................................................1-25

1.16 LIME CREEK HYDROELECTRIC POWER FACILITY ...................................................................1-27

1.17 INFRASTRUCTURE..................................................................................................................1-271.17.1 POWER SUPPLY AND DISTRIBUTION....................................................................1-28

1.18 ACCESS ROADS.....................................................................................................................1-28

1.19 LOGISTICS AND PORT FACILITY .............................................................................................1-29

1.20 ENVIRONMENTAL...................................................................................................................1-291.20.1 OVERVIEW...........................................................................................................1-291.20.2 ACID ROCK DRAINAGE POTENTIAL AND METAL LEACHING ..................................1-311.20.3 TAILING................................................................................................................1-32

1.21 RECLAMATION AND CLOSURE................................................................................................1-321.21.1 TAILING AND WATER MANAGEMENT FACILITIES...................................................1-32

1.22 PROJECT EXECUTION PLAN...................................................................................................1-34



ii 0954780100-REP-R0007-02

1.23 CAPITAL AND OPERATING COST ESTIMATES .........................................................................1-361.23.1 CAPITAL COST ESTIMATE ....................................................................................1-361.23.2 OPERATING COST ESTIMATE...............................................................................1-37

1.24 FINANCIAL ANALYSIS .............................................................................................................1-401.24.1 POST-TAX ANALYSIS............................................................................................1-41

1.24.2 SENSITIVITY ANALYSIS ........................................................................................1-411.25 RECOMMENDATIONS..............................................................................................................1-43

1.25.1 GEOLOGY ............................................................................................................1-431.25.2 MINING ................................................................................................................1-431.25.3 TAILING AND WATER MANAGEMENT ....................................................................1-441.25.4 PROCESS.............................................................................................................1-451.25.5 INFRASTRUCTURE ................................................................................................1-46

2.0 INTRODUCTION................................................................................................................... 2-1

3.0 RELIANCE ON OTHER EXPERTS .....................................................................................3-1

4.0 PROPERTY DESCRIPTION AND LOCATION................................................................... 4-1

4.1 PROPERTY LOCATION..............................................................................................................4-1

4.2 UNDERLYING AGREEMENT.......................................................................................................4-4

4.3 MINERAL TITLES AND SURFACE RIGHTS..................................................................................4-44.3.1 SURFACE LANDS ...................................................................................................4-4

4.4 LOCATION OF MINERALIZATION................................................................................................4-5

4.5 ROYALTIES, AGREEMENTS, AND ENCUMBRANCES...................................................................4-54.5.1 ROYALTIES ............................................................................................................4-54.5.2 ENCUMBRANCES FROM OTHER AGREEMENTS, LIENS, AND DOCUMENTS..............4-6

4.6 ENVIRONMENTAL LIABILITIES...................................................................................................4-6

4.7 PERMITTING.............................................................................................................................4-74.7.1 PERMITS OBTAINED FROM THE VENDOR OF THE MINE ..........................................4-74.7.2 PERMITS TO BE OBTAINED TO COMMENCE WORK AT THE MINE ............................4-8

4.8 CONSULTATION WITH NISGA’A AND FIRST NATIONS ................................................................4-9

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND

PHYSIOGRAPHY.................................................................................................................. 5-1

5.1 CLIMATE AND LENGTH OF OPERATING SEASON ......................................................................5-1

5.2 PHYSIOGRAPHY.......................................................................................................................5-1

5.3 ACCESS...................................................................................................................................5-3

5.4 LOCAL RESOURCE AND INFRASTRUCTURE ..............................................................................5-35.4.1 ACCESS ROAD AND TRANSPORTATION ..................................................................5-3

5.4.2 POWER SUPPLY.....................................................................................................5-45.4.3 PORT .....................................................................................................................5-45.4.4 BUILDINGS AND ANCILLARY FACILITIES..................................................................5-55.4.5 PERMANENT ACCOMMODATION FACILITY ..............................................................5-55.4.6 TAILINGS MANAGEMENT FACILITY .........................................................................5-55.4.7 WASTE DISPOSAL AREA........................................................................................5-55.4.8 HUMAN RESOURCES .............................................................................................5-5



iii 0954780100-REP-R0007-02

6.0 HISTORY...............................................................................................................................6-1

6.1 HISTORICAL RESOURCE ESTIMATES........................................................................................6-2

6.2 HISTORIC FEASIBILITY AND ECONOMIC REVIEWS ....................................................................6-4

7.0 GEOLOGICAL SETTING ..................................................................................................... 7-1

7.1 REGIONAL GEOLOGY ...............................................................................................................7-1

7.2 GENERAL GEOLOGY ................................................................................................................7-4

7.3 KITSAULT DEPOSIT GEOLOGY – GENERAL ..............................................................................7-4

7.4 KITSAULT DEPOSIT GEOLOGY – DETAILED..............................................................................7-6

7.5 REVISED GEOLOGIC INTERPRETATION ....................................................................................7-8

7.6 BELL MOLY PROSPECT GEOLOGY .........................................................................................7-17

7.7 ROUNDY CREEK GEOLOGY....................................................................................................7-19

7.8 REVIEW OF GEOLOGY............................................................................................................7-20

8.0 DEPOSIT TYPES.................................................................................................................. 8-1

9.0 MINERALIZATION................................................................................................................9-1

9.1 KITSAULT MINERALIZATION ......................................................................................................9-1

9.2 MOLYBDENUM DISTRIBUTION – PLAN AND SECTION................................................................9-3

9.3 KITSAULT GEOCHEMISTRY.......................................................................................................9-3

9.4 KITSAULT HYDROTHERMAL AND THERMAL ALTERATION..........................................................9-6

9.5 BELL MOLY MINERALIZATION ...................................................................................................9-7

9.6 ROUNDY CREEK MINERALIZATION...........................................................................................9-8

10.0 EXPLORATION................................................................................................................... 10-1

10.1 KITSAULT...............................................................................................................................10-1

10.2 BELL MOLY............................................................................................................................10-1

10.3 ROUNDY CREEK ....................................................................................................................10-2

10.4 FUTURE EXPLORATION WORK...............................................................................................10-210.4.1 KITSAULT.............................................................................................................10-210.4.2 BELL MOLY..........................................................................................................10-310.4.3 ROUNDY CREEK ..................................................................................................10-4

11.0 DRILLING............................................................................................................................11-1

11.1 KITSAULT HISTORIC...............................................................................................................11-1

11.2 KITSAULT – 2008 PROGRAM..................................................................................................11-2

11.3 BELL MOLY............................................................................................................................11-2

11.4 ROUNDY CREEK ....................................................................................................................11-2

11.5 2008 KITSAULT DRILL PROGRAM OBJECTIVES ......................................................................11-2

11.6 2008 KITSAULT DRILL PROGRAM ..........................................................................................11-3

12.0 SAMPLING METHOD AND APPROACH .........................................................................12-1

12.1 FACTORS IMPACTING ACCURACY OF RESULTS......................................................................12-1



iv 0954780100-REP-R0007-02

12.2 SAMPLE QUALITY...................................................................................................................12-1

13.0 SAMPLE PREPARATION, ANALYSES, AND SECURITY.............................................. 13-1

13.1 SAMPLE PREPARATION AND ASSAYING METHODS.................................................................13-113.1.1 HISTORIC DRILLING PROGRAMS ..........................................................................13-1

13.1.2 2008 AVANTI DRILL PROGRAM ............................................................................13-213.2 QUALITY CONTROL AND QUALITY ASSURANCE......................................................................13-3

13.2.1 HISTORIC DRILL PROGRAMS................................................................................13-313.2.2 2008 AVANTI DRILL PROGRAM ............................................................................13-4

13.3 INTERPRETATION...................................................................................................................13-913.3.1 HISTORIC ASSAY DATA........................................................................................13-913.3.2 2008 KITSAULT ASSAY DATA............................................................................ 13-10

14.0 DATA VERIFICATION........................................................................................................14-1

14.1 HISTORIC DRILL PROGRAMS..................................................................................................14-1

14.2 2008 AVANTI DRILL PROGRAMS ............................................................................................14-3

14.3 LIMITATIONS ..........................................................................................................................14-315.0 ADJACENT PROPERTIES ................................................................................................15-1

16.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATE.....................................16-1

16.1 MINERAL RESOURCES ...........................................................................................................16-116.1.1 DRILLHOLE DATABASE.........................................................................................16-116.1.2 EXPLORATORY DATA ANALYSIS...........................................................................16-416.1.3 TOPOGRAPHY ................................................................................................... 16-1116.1.4 GEOLOGY ......................................................................................................... 16-1216.1.5 COMPOSITING................................................................................................... 16-1416.1.6 EVALUATION OF OUTLIER GRADES ................................................................... 16-14

16.1.7 SPECIFIC GRAVITY............................................................................................ 16-1516.1.8 VARIOGRAM ANALYSIS AND MODELLING........................................................... 16-1516.1.9 PROBABILITY SHELL ......................................................................................... 16-1716.1.10 BLOCK MODEL LIMITS....................................................................................... 16-2016.1.11 GRADE ESTIMATION ......................................................................................... 16-2116.1.12 MODEL VALIDATION.......................................................................................... 16-2216.1.13 SWATH PLOTS (DRIFT ANALYSIS)..................................................................... 16-2716.1.14 RESOURCE CLASSIFICATION............................................................................. 16-3116.1.15 MINERAL RESOURCE STATEMENT .................................................................... 16-3116.1.16 RESULTS OF SRK AUDIT .................................................................................. 16-3316.1.17 MINERAL RESOURCE SENSITIVITY.................................................................... 16-3316.1.18 DISCUSSION AND CONCLUSIONS ...................................................................... 16-34

16.2 MINERAL RESERVES........................................................................................................... 16-3417.0 METALLURGICAL TESTING ............................................................................................17-1

17.1 INTRODUCTION ......................................................................................................................17-1

17.2 METALLURGICAL TESTWORK TIME PERIODS .........................................................................17-1

17.3 MINERALOGICAL EXAMINATION..............................................................................................17-3

17.4 HISTORICAL METALLURGICAL REVIEW...................................................................................17-4



v 0954780100-REP-R0007-02

17.4.1 BC MOLYBDENUM PERIOD – 1963 THROUGH TO 1976.......................................17-417.4.2 KITSAULT CONCENTRATOR PRODUCTION DATA 1967 TO 1972...........................17-717.4.3 SECOND PRE-PRODUCTION PERIOD FROM 1976 TO 1980 ..................................17-717.4.4 KITSAULT CONCENTRATOR PRODUCTION DATA – 1981 TO 1982..................... 17-1717.4.5 SGS LAKEFIELD PILOT PLANT STUDY – 1983.................................................. 17-19

17.4.6 SRK PEA – 2008............................................................................................. 17-2017.4.7 LEAD LEACHING TESTWORK REVIEW................................................................ 17-2017.4.8 MOLYBDENUM ANALYSES................................................................................. 17-2717.4.9 ROCK TYPE DISTRIBUTION ............................................................................... 17-28

17.5 RDI REVIEW ....................................................................................................................... 17-2817.5.1 HISTORICAL REVIEW......................................................................................... 17-2817.5.2 METALLURGICAL TESTING ................................................................................ 17-29

17.6 CONCLUSIONS .................................................................................................................... 17-44

17.7 RECOMMENDATIONS........................................................................................................... 17-44

18.0 MINERAL PROCESSING...................................................................................................18-1

18.1 INTRODUCTION ......................................................................................................................18-118.2 SUMMARY..............................................................................................................................18-1

18.3 MAJOR DESIGN CRITERIA......................................................................................................18-4

18.4 PLANT DESIGN.......................................................................................................................18-518.4.1 OPERATING SCHEDULE AND AVAILABILITY...........................................................18-5

18.5 PROCESS PLANT DESCRIPTION.............................................................................................18-518.5.1 PRIMARY CRUSHING............................................................................................18-518.5.2 CRUSHED ORE STOCKPILE AND RECLAIM ...........................................................18-618.5.3 GRINDING AND CLASSIFICATION..........................................................................18-618.5.4 FLOTATION CIRCUIT.............................................................................................18-818.5.5 MOLYBDENUM CONCENTRATE THICKENING ..................................................... 18-1118.5.6 LEAD LEACH CIRCUIT ....................................................................................... 18-1218.5.7 MOLYBDENUM CONCENTRATE DRYING ............................................................ 18-1318.5.8 TAILINGS HANDLING AND DISPOSAL ................................................................. 18-1418.5.9 REAGENT HANDLING AND STORAGE................................................................. 18-1418.5.10 ASSAY AND METALLURGICAL LABORATORY...................................................... 18-1618.5.11 WATER SUPPLY................................................................................................ 18-1618.5.12 AIR SUPPLY ...................................................................................................... 18-17

19.0 OTHER RELEVANT DATA AND INFORMATION............................................................19-1

19.1 MINING ..................................................................................................................................19-119.1.1 PIT OPTIMIZATION................................................................................................19-319.1.2 OPEN PIT DESIGN............................................................................................. 19-1519.1.3 MINE PRODUCTION SCHEDULE......................................................................... 19-2219.1.4 EQUIPMENT ANALYSIS AND SELECTION ............................................................ 19-3319.1.5 MINING EQUIPMENT FLEET PRODUCTIVITIES .................................................... 19-3719.1.6 DRILLING .......................................................................................................... 19-3819.1.7 BLASTING ......................................................................................................... 19-3919.1.8 LOADING........................................................................................................... 19-4019.1.9 TRUCK HAULAGE .............................................................................................. 19-4119.1.10 PIT DRAINAGE AND DEWATERING..................................................................... 19-44



vi 0954780100-REP-R0007-02

19.2 GEOTECHNICAL ENGINEERING – PIT SLOPES..................................................................... 19-4519.2.1 GEOTECHNICAL DATA COLLECTION.................................................................. 19-4519.2.2 GEOTECHNICAL MODEL .................................................................................... 19-4819.2.3 DATA ANALYSIS ................................................................................................ 19-4819.2.4 INTERRAMP/OVERALL STABILITY ANALYSES..................................................... 19-49

19.2.5 BENCH STABILITY CONSIDERATIONS ................................................................ 19-5019.2.6 SLOPE DESIGN RECOMMENDATIONS ................................................................ 19-51

19.3 PIT HYDROGEOLOGY .......................................................................................................... 19-52

19.4 SITE WATER MANAGEMENT................................................................................................ 19-5419.4.1 GENERAL.......................................................................................................... 19-5419.4.2 HYDROLOGY AND FLOOD FLOWS ..................................................................... 19-5619.4.3 DIVERSION STRUCTURES AND LIME CREEK HYDROELECTRIC PROJECT .......... 19-5719.4.4 MINE SITE WATER MANAGEMENT..................................................................... 19-6119.4.5 WATER STORAGE AND RELEASE ...................................................................... 19-6219.4.6 WATER BALANCE.............................................................................................. 19-6219.4.7 CLOSURE.......................................................................................................... 19-66

19.5 TAILING

MANAGEMENT

....................................................................................................... 19-6619.5.1 INTRODUCTION ................................................................................................. 19-6619.5.2 TAILING MANAGEMENT FACILITY DESIGN ......................................................... 19-6719.5.3 EMBANKMENT CONSTRUCTION......................................................................... 19-7519.5.4 TAILING, RECLAIM AND FRESH WATER SYSTEMS ............................................. 19-82

19.6 WASTE ROCK AND TAILING GEOCHEMICAL CHARACTERIZATION........................................ 19-9919.6.1 WASTE ROCK METAL LEACHING AND ACID ROCK DRAINAGE POTENTIAL......... 19-9919.6.2 TAILING........................................................................................................... 19-102

19.7 RECLAMATION AND CLOSURE...........................................................................................19-10219.7.1 GENERAL........................................................................................................ 19-10219.7.2 RECLAMATION UNITS...................................................................................... 19-10419.7.3 RECLAMATION OBJECTIVES ............................................................................19-106

19.7.4 BASELINE STUDIES ......................................................................................... 19-10619.7.5 RECLAMATION AND CLOSURE ACTIVITIES.......................................................19-10819.7.6 CLOSURE COST ESTIMATES ...........................................................................19-111

19.8 ON-SITE INFRASTRUCTURE...............................................................................................19-11619.8.1 SITE LAYOUT .................................................................................................. 19-11619.8.2 LIME CREEK DIVERSION TUNNEL.................................................................... 19-11719.8.3 PLANT CONVEYORS........................................................................................ 19-11819.8.4 SITE ROADS ................................................................................................... 19-11819.8.5 DIVERSION TUNNEL ACCESS ROADS.............................................................. 19-11819.8.6 PROCESS PLANT ............................................................................................19-11819.8.7 ANCILLARY BUILDINGS.................................................................................... 19-11919.8.8 ASSAY OFFICE................................................................................................19-119

19.8.9 CONCENTRATE STORAGE............................................................................... 19-11919.8.10 WAREHOUSE/TRUCK SHOP/MINE DRY ........................................................... 19-11919.8.11 TRUCK WASH/TIRE CHANGE BUILDING ........................................................... 19-12019.8.12 SEWAGE ......................................................................................................... 19-12019.8.13 COMMUNICATIONS SYSTEM ............................................................................19-12019.8.14 POTABLE WATER SUPPLY .............................................................................. 19-12019.8.15 EXPLOSIVES STORAGE AND HANDLING ..........................................................19-12119.8.16 GEOTECHNICAL CONDITIONS..........................................................................19-121



vii 0954780100-REP-R0007-02

19.9 OFF-SITE INFRASTRUCTURE .............................................................................................19-121

19.10 POWER SUPPLY AND DISTRIBUTION................................................................................. 19-12319.10.1 MAIN SITE SUBSTATION.................................................................................. 19-12319.10.2 POWER DISTRIBUTION.................................................................................... 19-123

19.11 MAIN ACCESS ROADS....................................................................................................... 19-12619.11.1 PLANT ROAD DESIGN REQUIREMENTS ........................................................... 19-12619.11.2 PORT SITE ROAD AND EARTHWORKS............................................................. 19-12719.11.3 EXPLOSIVES STORAGE FACILITY EARTHWORKS/ACCESS ROAD .................... 19-12719.11.4 OTHER CIVIL WORKS...................................................................................... 19-12719.11.5 EARTHWORK SIDE SLOPES.............................................................................19-128

19.12 LOGISTICS ........................................................................................................................ 19-128

19.13 CAPITAL COST ESTIMATE ................................................................................................. 19-13019.13.1 EXCLUSIONS................................................................................................... 19-13019.13.2 MINE CAPITAL COSTS..................................................................................... 19-13119.13.3 PROCESS CAPITAL COST................................................................................ 19-13519.13.4 PERMANENT ACCOMMODATION AND CONSTRUCTION CAMPS ........................ 19-136

19.13.5 LABOUR RATES ..............................................................................................19-13719.13.6 TAXES............................................................................................................. 19-13719.13.7 LOGISTICS ...................................................................................................... 19-13719.13.8 OWNERS’ COSTS (INCLUDING OWNERS COMMISSIONING ALLOWANCE)......... 19-13719.13.9 EXCLUSIONS................................................................................................... 19-13719.13.10 ASSUMPTIONS ................................................................................................19-13819.13.11 CONTINGENCY................................................................................................19-13819.13.12 TMF AND RELATED WATER MANAGEMENT SYSTEMS ....................................19-13819.13.13 PORT FACILITIES ............................................................................................19-140

19.14 OPERATING COST ESTIMATE............................................................................................19-14019.14.1 MINING OPERATING COSTS ............................................................................19-14119.14.2 PROCESS OPERATING COSTS ........................................................................19-145

19.14.3 GENERAL AND ADMINISTRATIVE ..................................................................... 19-15119.14.4 TMF OPERATING AND WATER MANAGEMENT COSTS ....................................19-15319.14.5 UNIT CASH OPERATION COSTS ...................................................................... 19-154

19.15 FINANCIAL ANALYSIS ........................................................................................................ 19-15419.15.1 INTRODUCTION ...............................................................................................19-15419.15.2 PRE-TAX MODEL ............................................................................................19-15519.15.3 POST-TAX MODEL...........................................................................................19-15819.15.4 SENSITIVITY ANALYSIS ................................................................................... 19-15919.15.5 ROYALTIES ..................................................................................................... 19-16019.15.6 SMELTER TERMS ............................................................................................19-16119.15.7 CONCENTRATE TRANSPORT LOGISTICS .........................................................19-161

20.0 PROJECT EXECUTION PLAN.......................................................................................... 20-121.0 ENVIRONMENTAL CONSIDERATIONS..........................................................................21-1

21.1 REGULATORY APPROVAL PROCESS ......................................................................................21-121.1.1 OVERVIEW...........................................................................................................21-121.1.2 BRITISH COLUMBIA AUTHORIZATIONS, LICENSES, AND PERMITS.........................21-221.1.3 FEDERAL AUTHORIZATIONS, LICENSES AND PERMITS .........................................21-421.1.4 RECLAMATION SECURITY BOND ..........................................................................21-5



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21.2 ENVIRONMENTAL STUDIES ....................................................................................................21-521.2.1 SOILS AND TERRAIN MAPPING.............................................................................21-621.2.2 WASTE ROCK METAL LEACHING AND ACID ROCK DRAINAGE POTENTIAL............21-721.2.3 METEOROLOGY AND ATMOSPHERE .....................................................................21-721.2.4 SURFACE HYDROLOGY ........................................................................................21-9

21.2.5 HYDROGEOLOGY.............................................................................................. 21-1021.2.6 OCEANOGRAPHY .............................................................................................. 21-1221.2.7 FRESHWATER AQUATICS .................................................................................. 21-1321.2.8 FRESHWATER AND MARINE FISHERIES............................................................. 21-1421.2.9 WETLANDS ....................................................................................................... 21-1721.2.10 ECOSYSTEMS AND VEGETATION MAPPING ....................................................... 21-1821.2.11 TERRESTRIAL AND MARINE WILDLIFE............................................................... 21-2021.2.12 VISUAL QUALITY ............................................................................................... 21-2421.2.13 ARCHAEOLOGY ................................................................................................. 21-2621.2.14 HUMAN HEALTH................................................................................................ 21-27

21.3 SOCIOECONOMIC STUDIES ................................................................................................. 21-2821.3.1 SOCIOECONOMIC.............................................................................................. 21-28

21.3.2 NON-TRADITIONAL LAND AND RESOURCE USE................................................. 21-3021.3.3 NISGA’A USE AND TRADITIONAL KNOWLEDGE .................................................. 21-3121.3.4 COMMUNITY ENGAGEMENT AND CONSULTATION REQUIREMENTS.................... 21-3421.3.5 CONSULTATION WITH NISGA'A NATION AND FIRST NATIONS............................. 21-34

22.0 MARKETING AND CONTRACTS .....................................................................................22-1

22.1 INTRODUCTION ......................................................................................................................22-1

22.2 MOLYBDENUM MARKET OVERVIEW .......................................................................................22-3

22.3 MINE PRODUCTION ................................................................................................................22-5

22.4 OPERATING COSTS ...............................................................................................................22-6

22.5 SECONDARY SUPPLIES..........................................................................................................22-6

22.6 CPM’S SUPPLY OUTLOOK .....................................................................................................22-722.7 MOLYBDENUM DEMAND.........................................................................................................22-7

22.7.1 MOLYBDENUM CONSUMPTION BY THE STEEL INDUSTRY .................................. 22-1022.7.2 OTHER APPLICATIONS ...................................................................................... 22-14

22.8 SUBSTITUTES ..................................................................................................................... 22-16

22.9 CPM’S SUPPLY AND DEMAND BALANCE............................................................................. 22-17

22.10 PRICE OUTLOOK – CPM’S BASE CASE SCENARIO ............................................................. 22-18

22.11 PRICE OUTLOOK – CPM’S ALTERNATIVE SCENARIOS ........................................................ 22-1922.11.1 CPM’S HIGH CASE SCENARIO.......................................................................... 22-1922.11.2 CPM’S LOW CASE SCENARIO .......................................................................... 22-19

23.0 RISK AND MITIGATION.....................................................................................................23-1

24.0 OPPORTUNITIES AND RECOMMENDATIONS..............................................................24-1

24.1 GEOLOGY ..............................................................................................................................24-1

24.2 MINING ..................................................................................................................................24-1

24.3 TAILING AND WATER MANAGEMENT ......................................................................................24-2

24.4 PROCESS...............................................................................................................................24-3



ix 0954780100-REP-R0007-02

24.5 INFRASTRUCTURE..................................................................................................................24-4

24.6 ENVIROMENTAL .....................................................................................................................24-4

25.0 INTERPRETATION AND CONCLUSIONS.......................................................................25-1

26.0 REFERENCES ....................................................................................................................26-1

26.1 ELECTRONIC DOCUMENTATION .............................................................................................26-5

L I S T O F A P P E N D I C E S

APPENDIX A CERTIFICATES OF QUALIFIED PERSONS

APPENDIX B “PRINCE RUPERT ROASTER FACILITY PORT FACILITIES SCOPING STUDY” BY

WORLEYPARSONS WESTMAR

APPENDIX C “AN INVESTIGATION INTO THE METALLURGICAL TESTING OF MOLYBDENUM MATERIAL FROM

THE KITSAULT DEPOSIT” BY SGS CANADA INC.

APPENDIX D PROCESS FLOWSHEET DIAGRAMS, PROCESS DESIGN CRITERIA, GENERAL ARRANGEMENT

DRAWINGS, PLANT MOBILE EQUIPMENT LIST, AND EQUIPMENT LOAD LIST

APPENDIX E ECONOMICS OF THE BASE AND FINAL CASE OPTIMIZATIONS

APPENDIX F MINING PHASE DEVELOPMENTS – PHASES 1 THROUGH 4

APPENDIX G MINING EQUIPMENT SELECTION CALCULATIONS

APPENDIX H HAUL CENTROIDS, PROFILES, AND CYCLE TIMES

APPENDIX I CALCULATED CYCLE TIME BY PHASE, PERIOD, AND DESTINATION

APPENDIX J “KITSAULT PRE-FEASIBILITY STUDY – PIT HYDROGEOLOGY” BY SRK CONSULTING (CANADA)

APPENDIX K “AVANTI KITSAULT MINE LTD. LIME CREEK HYDROELECTRIC PROJECT PREFEASIBILITY

STUDY” BY KNIGHT PIÉSOLD CONSULTING

APPENDIX L ENVIRONMENTAL, DECOMMISSIONING, AND POST-CLOSURE COSTS

APPENDIX M DRAWINGS

APPENDIX N “KITSAULT MOLYBDENUM PRE-FEASIBILITY STUDY PRELIMINARY LOGISTICS TRADE-OFF

STUDY” BY WARDROP ENGINEERING INC.

APPENDIX O CAPITAL COST ESTIMATEAPPENDIX P PROJECT SITE CONDITIONS SPECIFICATION

APPENDIX Q OPERATING COST ESTIMATE

APPENDIX R FINANCIAL ANALYSIS

NOTE: APPENDICES B THROUGH R ARE AVAILABLE ON REQUEST AT

THE AVANTI MINING INC. OFFICES IN VANCOUVER



x 0954780100-REP-R0007-02

L I S T O F T A B L E S

Table 1.1 Mineral Resource Statement* for the Kitsault Molybdenum Deposit –

March 31, 2009 ................................................................................................1-12Table 1.2 Mo Cut-off Grade Sensitivity Analysis within Resource Pit – Measured and

Indicated Resources.........................................................................................1-13Table 1.3 Mineral Reserves by Class ..............................................................................1-14Table 1.4 Major Process Design Criteria .........................................................................1-19Table 1.5 Summarized Production Schedule...................................................................1-23Table 1.6 Capital Cost Summary .....................................................................................1-37Table 1.7 Unit Cash Operation Costs (Life of Mine Average)..........................................1-37Table 1.8 Process Operating Cost Summary ..................................................................1-38Table 1.9 Mining Operating Costs....................................................................................1-39Table 1.10 TMF Annual Operating Costs per Tonne Ore ..................................................1-39Table 1.11 Metal Production from Kitsault Project.............................................................1-40Table 1.12 Molybdenum Pricing Cases (US$)...................................................................1-40

Table 1.13 Summary of the Post-Tax Financial Analysis ..................................................1-41Table 1.14 Base Case Sensitivity to Pre-tax NPV (US$000s) at 8% Discount Rate.........1-42Table 1.15 Sensitivity to Post-Tax NPV (US$000s)...........................................................1-42Table 2.1 Summary of Qualified Persons ..........................................................................2-1Table 6.1 CMC Historic Resource & Reserve Estimate for the Kitsault Deposit* .............6-3Table 6.2 CMC Historic Resource Estimate for Bell Moly Clary Creek Deposit* ..............6-3Table 6.3 Woodcock and Carter Historic Resource Estimate for Roundy Creek

Deposit, 1976*....................................................................................................6-3Table 6.4 J.W Mustard Roundy Creek Resource Estimate; 1983*....................................6-4Table 6.5 CMC Resource Model vs. Blast Hole Model Comparison*................................6-4Table 14.1 CMC Resource Model vs. Blast Hole Model Comparison*..............................14-2Table 16.1 Grid to UTM Conversion ..................................................................................16-2

Table 16.2 Simplified Rock Types....................................................................................16-13Table 16.3 Summary of Outlier Limitations......................................................................16-15Table 16.4 Variogram Parameters – Molybdenum, Inside Shell .....................................16-16Table 16.5 Variogram Parameters – Molybdenum, Outside Shell ..................................16-16Table 16.6 Variogram Parameters – WO3 .......................................................................16-16Table 16.7 Variogram Parameters – Copper...................................................................16-17Table 16.8 Variogram Parameters – Lead.......................................................................16-17Table 16.9 Variogram Parameters – Iron.........................................................................16-17Table 16.10 Variogram Parameters – Silver......................................................................16-17Table 16.11 0.05% Molybdenum Indicator Variogram Parameters ...................................16-20Table 16.12 Indicator Interpolation Parameters.................................................................16-20Table 16.13 Block Model Limits .........................................................................................16-20Table 16.14 Interpolation Parameters................................................................................16-21

Table 16.15 Mineral Resource Statement* for the Kitsault Molybdenum Deposit –March 31, 2009 ..............................................................................................16-32

Table 16.16 Mo Cut-off Grade Sensitivity Analysis within Resource Pit – Measure andIndicated Resources.......................................................................................16-33

Table 16.17 Mo Cut-off Grade Sensitivity Analysis within Resource Pit – InferredResources ......................................................................................................16-34

Table 16.18 Mineral Reserves by Class ............................................................................16-35Table 17.1 Production Summary Data – Kitsault Concentrator 1967 to 1972...................17-7



xi 0954780100-REP-R0007-02

Table 17.2 Ore Head Analysis – AMAX 1976....................................................................17-9Table 17.3 Summary of Work and Abrasion Indices – Allis-Chalmers 1977.....................17-9Table 17.4 Summary of Grindability Studies – AMAX 1976 ............................................17-10Table 17.5 Grinding Design Criteria – Climax Molybdenum Feasibility Study 1977 .......17-10Table 17.6 Metallurgical Summary of Laboratory Batch Flotation Tests – Olin 1977 ....17-12

Table 17.7 Metallurgical Summary of Pilot Plant Rougher Flotation – Olin 1977............17-13Table 17.8 Summary of Regrind-Cleaner Flotation Upgrading of Pilot PlantConcentrates – Olin 1977 ..............................................................................17-14

Table 17.9 Combined Rougher and Cleaner Flotation Results .......................................17-16Table 17.10 Production Summary — Kitsault Concentrator 1981-1982............................17-17Table 17.11 Monthly Production Data – Kitsault Concentrator 1982.................................17-18Table 17.12 Reagent Consumptions – Production Data 1981 to 1982 .............................17-19Table 17.13 Lead Leaching of BC Molybdenum Kitsault Mill Concentrate........................17-22Table 17.14 Lead Leaching from AMAX Laboratory Concentrate Samples......................17-23Table 17.15 Lead Leach Test Results — SGS Lakefield 1983 .........................................17-25Table 17.16 Metal Concentration of Leach Plant Liquor Before and After Precipitation...17-26Table 17.17 Summary of Leach Liquor Heavy Metal Concentration .................................17-26Table 17.18 Comparison of Molybdenum Analyses ..........................................................17-27

Table 17.19 Rock Type Distribution ...................................................................................17-28Table 17.20 Composite Sample Details.............................................................................17-30Table 17.21 Summary of Hazen Drop Weight Breakage Evaluation.................................17-30Table 17.22 Summary of Hazen SMC Breakage Evaluation.............................................17-31Table 17.23 Summary of Hazen BWi, RWi, CWi, and Ai Evaluation.................................17-31Table 17.24 Rougher Flotation Test Results......................................................................17-34Table 17.25 10 kg Rougher/Scavenger Flotation Tests.....................................................17-35Table 17.26 First Cleaner Flotation – Three Minutes.........................................................17-35Table 17.27 Combined Reground Cleaner Tails and Scavenger Concentrate .................17-36Table 17.28 Open Circuit 4-Stage Cleaner Tests with 1st Cleaner Concentrate

Regrind, and With and Without 3rd Cleaner Concentrate Regrind ...............17-37Table 17.29 Open Circuit Flotation on the Master Composite...........................................17-37Table 17.30 Locked Cycle Flotation Test (LCT1) on the Master Composite.....................17-38Table 17.31 Locked Cycle Flotation Test (LCT2) on the Master Composite.....................17-38Table 17.32 Pyrite and Sulfide Flotation Test Data ...........................................................17-42Table 17.33 By-product Recovery by Gravity Concentration.............................................17-43Table 17.34 By-product Recovery using AP3477 for Flotation..........................................17-43Table 17.35 By-product Recovery using SIPX for Flotation ..............................................17-43Table 18.1 Major Process Design Criteria .........................................................................18-4Table 19.1 Block Model Limits ...........................................................................................19-3Table 19.2 Whittle Base Optimization Parameters............................................................19-4Table 19.3 Nested Pits – Base Optimization......................................................................19-6Table 19.4 Nested Pits – Optimized Cut-off Scenarios......................................................19-9Table 19.5 Comparison of Base Case versus Final Case...............................................19-11Table 19.6 Design Parameters for Pit Optimization.........................................................19-13

Table 19.7 Discounted Open Pit Cash Flow for Different Sizes ......................................19-14Table 19.8 Mineable Ore Reserves..................................................................................19-19Table 19.9 Summarized Production Schedule.................................................................19-23Table 19.10 Ranking of Equipment Fleet...........................................................................19-37Table 19.11 Mine Equipment Operating Schedule ............................................................19-38Table 19.12 Blasthole Drill Productivity..............................................................................19-39Table 19.13 Blasting Parameters for 269 mm Production Blastholes ...............................19-40Table 19.14 Average Productivity of Shovels in Ore and Waste Rock..............................19-40



xii 0954780100-REP-R0007-02

Table 19.15 Facility Locations............................................................................................19-42Table 19.16 Cat 789C Haul Truck Productivity ..................................................................19-44Table 19.17 Summary of Drillholes Orientated & Logged for Geotechnical Data .............19-46Table 19.18 Summary of BEJ and Geomean Groundwater Inflow Rates .........................19-54Table 19.19 Water Balance Input Values...........................................................................19-63

Table 19.20 Water Balance Catchment Areas...................................................................19-64Table 19.21 Site-wide Annual Water Balance Flows – Year 1 to Closure.........................19-65Table 19.22 Overall Design Basis for the TMF – Page 1 of 3 ...........................................19-68Table 19.23 Material Requirements for the Various Stages of the TMF ...........................19-72Table 19.24 Annual Fill Placing Requirements ..................................................................19-73Table 19.25 Tailing System Design Periods ......................................................................19-83Table 19.26 Design Data....................................................................................................19-84Table 19.27 Process Battery Limits....................................................................................19-92Table 19.28 Capital Cost Summary .................................................................................19-131Table 19.29 Mine Capital Costs .......................................................................................19-132Table 19.30 Mine Mobile and Support Equipment Capital ..............................................19-133Table 19.31 Power Supply Construction Cost Summary.................................................19-134Table 19.32 Process Capital Cost Basis of Estimate.......................................................19-135

Table 19.33 Mine Hourly Labour Requirements ..............................................................19-142Table 19.34 Mine Salaried Labour Requirements ...........................................................19-143Table 19.35 Mine Operating and Maintenance Hourly Labour Rates .............................19-143Table 19.36 Mine Staff Salaries .......................................................................................19-144Table 19.37 Mining Costs per Tonne ...............................................................................19-145Table 19.38 Process Operating Costs and G&A Costs Summary...................................19-146Table 19.39 Mill Labour Requirements ............................................................................19-147Table 19.40 Mill Maintenance Labour Requirements ......................................................19-148Table 19.41 Plant Operating Supplies .............................................................................19-149Table 19.42 Maintenance Supplies Allowances ..............................................................19-151Table 19.43 Total Mill Power Supply................................................................................19-151Table 19.44 General & Administrative Expenses ............................................................19-152Table 19.45 TMF Operating and Water Management Costs...........................................19-154Table 19.46 Unit Cash Operation Costs (LOM Average).................................................19-154Table 19.47 Summary of the Post-tax Financial Analysis ...............................................19-155Table 19.48 Metal Production from Kitsault Project.........................................................19-156Table 19.49 Summary of Metal Price and Exchange Rate Scenarios.............................19-158Table 19.50 Pre- and Post-tax NPV, IRR, and Payback by Metal Price Scenario ..........19-158Table 19.51 Sensitivity Analysis to Post Tax NPV (US$ M) ............................................19-160Table 21.1 BC Authorizations, Licenses, and Permits Required for the Kitsault

Project ..............................................................................................................21-3Table 21.2 Federal Authorizations, Licenses, and Permits Required for the Project ........21-4Table 22.1 World Molybdenum Reserves and Reserve Base, 2008 (Billion Pounds) ......22-4Table 22.2 Growth in Molybdenum Demand by End Use – Compound Annual Growth

Rates for Selected Years .................................................................................22-8

Table 22.3 Molybdenum End-use Profiles .........................................................................22-9Table 22.4 Real Annual Molybdenum Prices per Pound.................................................22-20Table 23.1 Risks and Mitigations .......................................................................................23-1



xiii 0954780100-REP-R0007-02

L I S T O F F I G U R E S

Figure 1.1 Kitsault Location and Regional Geology............................................................1-4

Figure 1.2 Mineral Tenure Map and Location of Molybdenum Deposits on AvantiKitsault Mine Ltd. Tenures .................................................................................1-5

Figure 1.3 General View of the Kitsault Mine Area .............................................................1-7Figure 1.4 Overall General Arrangement ..........................................................................1-15Figure 1.5 Final Pit.............................................................................................................1-23Figure 1.6 Project Schedule ..............................................................................................1-35Figure 4.1 Kitsault Location and Regional Geology............................................................4-2Figure 4.2 Claim Map and Location of Molybdenum Deposits on Avanti Claims ...............4-3Figure 5.1 General View of the Kitsault Mine Area .............................................................5-2Figure 7.1 Geology of the Lime Creek (Kitsault Area) ........................................................7-2Figure 7.2 Geology of the Lime Creek Property..................................................................7-3Figure 7.3 Generalized Geology and Mineralization Projected to Surface from DDHs....7-10Figure 7.4 Generalized Geology – Mineralization Level Plan 450 m................................7-11

Figure 7.5 Molybdenum Assay/Geology Section 6141850N – Looking North..................7-12Figure 7.6 Molybdenum Assay/Geology Section 6141900N – Looking North..................7-13Figure 7.7 Molybdenum Assay/Geology Section 473350E – Looking West.....................7-14Figure 7.8 Bell Molybdenum Area Geology.......................................................................7-18Figure 7.9 Geology of the Roundy Creek Area .................................................................7-19Figure 9.1 2008 DDH Location Map....................................................................................9-5Figure 13.1 Scatterplot of the Bell Moly Duplicate Assay Data...........................................13-4Figure 13.2 CDN Standard AV1 Control Chart ...................................................................13-5Figure 13.3 CDN Standard AV2 Control Chart ...................................................................13-6Figure 13.4 CDN Standard AV3 Control Chart ...................................................................13-6Figure 13.5 Performance of Kitsault Blank Material............................................................13-7Figure 13.6 Correspondence of Molybdenum in Kitsault Coarse Reject Duplicates..........13-8

Figure 13.7 Correspondence of Silver in Kitsault Coarse Reject Duplicates......................13-8Figure 13.8 Performance of Check Assays from Pulp Duplicates ......................................13-9Figure 14.1 QQ Plot.............................................................................................................14-2Figure 16.1 Drillhole Plan ....................................................................................................16-3Figure 16.2 Kitsault Rock Types..........................................................................................16-4Figure 16.3 Contact Profile for Kitsault Molybdenum Major 4-NE Porphyry Contact .........16-5Figure 16.4 Contact Profile for Kitsault Molybdenum Diorite-Granodiorite Contact ...........16-6Figure 16.5 Contact Profile for Kitsault Molybdenum Hornfels/Diorite/ Granodiorite-

Alaskite Contacts..............................................................................................16-7Figure 16.6 Contact Profile for Kitsault Molybdenum Hornfels-Granodiorite Contact ........16-8Figure 16.7 Contact Profile for Kitsault Molybdenum Hornfels-Diorite Contact ..................16-9Figure 16.8 Contact Profile for Kitsault Molybdenum Major 3-Alaskite Contact ...............16-10Figure 16.9 Histogram for Molybdenum Grade.................................................................16-11

Figure 16.10 Drillholes – Topographical Surface with 5 m Contour Lines..........................16-12Figure 16.11 Distribution of Intrusive Phases......................................................................16-13Figure 16.12 Cumulative Probability Plot ............................................................................16-18Figure 16.13 Probability Shell..............................................................................................16-19Figure 16.14 Plan 450 m Elevation .....................................................................................16-23Figure 16.15 Section 6141850N, East-West Vertical Cross Sections ................................16-24Figure 16.16 Recovered Molybdenum 10 x 10 x 10 m SMU – 2009 Model Inside

Shell................................................................................................................16-26



xiv 0954780100-REP-R0007-02

Figure 16.17 Grade/Tonnage Graph ...................................................................................16-27Figure 16.18 Swath Plot East-West 50 m Swaths ..............................................................16-28Figure 16.19 Swath Plot North-South 50 m Swaths............................................................16-29Figure 16.20 Swath Plot Vertical 20 m Swaths...................................................................16-30Figure 17.1 Molybdenum and Lead Recovery during Rougher Flotation as a Function

of Primary Grind – Climax 1977.....................................................................17-11Figure 17.2 Mo and Pb Recovery during Rougher Flotation as a Function of FlotationTime at a Grind of 45 to 50% plus 150 µm – Olin 1977.................................17-12

Figure 17.3 Lead Removal from Molybdenite Concentrate during Each Stage of Beneficiation – Olin 1977 ...............................................................................17-15

Figure 17.4 Flowsheet Evaluated in the 2009 SGS Test Program ...................................17-33Figure 18.1 Simplified Process Flowsheet ..........................................................................18-3Figure 19.1 General Arrangement.......................................................................................19-2Figure 19.2 Potential Final Pit Shells in the Base Optimization..........................................19-8Figure 19.3 Nested Pits – Optimized Cut-off Scenario .....................................................19-10Figure 19.4 Base Case and Optimized Cut-off – NPV Analysis .......................................19-12Figure 19.5 Decision Analysis ...........................................................................................19-13Figure 19.6 Nested Pits – Discounted Cash Flow for Different Pit Sizes .........................19-15

Figure 19.7 Final Pit...........................................................................................................19-21Figure 19.8 Mining Status as of Year 0 .............................................................................19-25Figure 19.9 Mining Status as of Year 1 .............................................................................19-26Figure 19.10 Mining Status as of Year 2.............................................................................19-27Figure 19.11 Mining Status as of Year 3.............................................................................19-28Figure 19.12 Mining Status as of Year 4.............................................................................19-29Figure 19.13 Mining Status as of Year 5.............................................................................19-30Figure 19.14 Mining Status as of Year 9.............................................................................19-31Figure 19.15 Mining Status as of Year 14 ...........................................................................19-32Figure 19.16 Sectors of Potential Bench Scale Instability ..................................................19-51Figure 19.17 General Layout of the Kitsault Mine Site........................................................19-55Figure 19.18 Water Diversion Tunnel – Plan and Profile ....................................................19-58Figure 19.19 Lime Creek Diversion Dam – Plan and Cross-section...................................19-59Figure 19.20 Construction Access Road to Upstream End of Tunnel and Lime Creek

Diversion Dam................................................................................................19-60Figure 19.21 Conceptual Water Balance Model .................................................................19-62Figure 19.22 Depth-Capacity Relationship for the TMF......................................................19-71Figure 19.23 Filling Curve for the TMF................................................................................19-72Figure 19.24 Construction Access Roads to Borrow Areas................................................19-77Figure 19.25 Construction Access Roads to TMF Embankment and Downstream Portal

of Diversion Tunnel ........................................................................................19-78Figure 19.26 Main Embankment Construction Stream Diversion.......................................19-79Figure 19.27 Haul Road from Open Pit to TMF Embankment – Stage 1 ...........................19-80Figure 19.28 Block Diagrams for the Various Process Streams.........................................19-85Figure 19.29 Tailing, Reclaim, and Water Management Pipelines – Stage 1 (Year -1).....19-86

Figure 19.30 Tailing, Reclaim, and Water Management Pipelines – Stage 2 (Year 3) ......19-87Figure 19.31 Tailing, Reclaim, and Water Management Pipelines – Stage 3 (Year 8) ......19-88Figure 19.32 Tailing, Reclaim, and Water Management Pipelines – Stage 3 (Year 15) ....19-89Figure 19.33 Camp Potable Water Supply..........................................................................19-90Figure 19.34 Typical Pipework Details for TMF ..................................................................19-91Figure 19.35 Distribution of NP/AP Indicated by Block Modelling ....................................19-101Figure 19.36 Closure Activities..........................................................................................19-103Figure 19.37 Port Facility Layout.......................................................................................19-122



xv 0954780100-REP-R0007-02

Figure 19.38 Undiscounted Annual and Cumulative Cash Flows.....................................19-157Figure 19.39 Pre-tax NPV Sensitivity Analysis..................................................................19-159Figure 19.40 IRR Sensitivity Analysis................................................................................19-160Figure 20.1 Project Schedule ..............................................................................................20-3Figure 22.1 Molybdenum Prices – Monthly, through November 2009................................22-2

Figure 22.2 World Molybdenum Reserves, 2008................................................................22-3Figure 22.3 World Molybdenum Reserve Base, 2008 ........................................................22-4Figure 22.4 Annual Mine Production of Molybdenum – Projected through 2018p .............22-6Figure 22.5 Molybdenum End Uses, 2008..........................................................................22-8Figure 22.6 Share of Alloy Steel as a Percentage of Crude Steel Production, 2008 .......22-10Figure 22.7 World Energy Consumption ...........................................................................22-13Figure 22.8 Average Annual Growth in Crude Oil Production – 1996 to 2007, by Crude

Quality ............................................................................................................22-14Figure 22.9 Base Case: Real Molybdenum Prices and World Supply and Demand

Balance – Annual, Projected through 2018p .................................................22-18

G L O S S A R Y

UN I T S O F ME A S U R E

Above mean sea level............. ............... ........... .............. ............ .............. ............... ........... ....... amsl

Acre ............. .............. ............ .............. ............ .............. .............. ............ .............. ............ ....... ac

Ampere ........ .............. ............ .............. .............. ............ .............. ............ .............. .............. ..... A

Annum (year)............. .............. ............ .............. ............ .............. .............. ............ .............. ...... a

Billion ........................................................................................................................................ B

Billion tonnes............................................................................................................................. BtBillion years ago ........................................................................................................................ Ga

British thermal unit ..................................................................................................................... BTU

Centimetre................................................................................................................................. cm

Cubic centimetre........................................................................................................................ cm3

Cubic feet per minute................................................................................................................. cfm

Cubic feet per second ................................................................................................................ ft3/s

Cubic foot .................................................................................................................................. ft

Cubic inch ................................................................................................................................. in3

Cubic metre............................................................................................................................... m3

Cubic yard ................................................................................................................................. yd

Coefficients of Variation ............................................................................................................. CVs

Day ........................................................................................................................................... d

Days per week ........................................................................................................................... d/wk

Days per year (annum)............................................................................................................... d/a

Dead weight tonnes ................................................................................................................... DWT

Decibel adjusted ........................................................................................................................ dBa

Decibel...................................................................................................................................... dB

Degree ...................................................................................................................................... °



xvi 0954780100-REP-R0007-02

Degrees Celsius ........................................................................................................................ °C

Diameter.................................................................................................................................... ø

Dollar (American)....................................................................................................................... US$

Dollar (Canadian)....................................................................................................................... Cdn$

Dry metric ton............................................................................................................................ dmt

Foot........................................................................................................................................... ft

Gallon........................................................................................................................................ gal

Gallons per minute (US)............................................................................................................. gpm

Gigajoule................................................................................................................................... GJ

Gigapascal ................................................................................................................................ GPa

Gigawatt.................................................................................................................................... GW

Gram......................................................................................................................................... g

Grams per litre........................................................................................................................... g/L

Grams per tonne........................................................................................................................ g/t

Greater than .............................................................................................................................. >

Hectare (10,000 m ) ................................................................................................................... ha

Hertz ......................................................................................................................................... HzHorsepower ............................................................................................................................... hp

Hour.......................................................................................................................................... h

Hours per day ............................................................................................................................ h/d

Hours per week ......................................................................................................................... h/wk

Hours per year........................................................................................................................... h/a

Inch........................................................................................................................................... "

Kilo (thousand) .......................................................................................................................... k

Kilogram.................................................................................................................................... kg

Kilograms per cubic metre.......................................................................................................... kg/m

Kilograms per hour..................................................................................................................... kg/h

Kilograms per square metre ....................................................................................................... kg/m2

Kilometre................................................................................................................................... km

Kilometres per hour.................................................................................................................... km/h

Kilopascal.................................................................................................................................. kPa

Kilotonne................................................................................................................................... kt

Kilovolt ...................................................................................................................................... kV

Kilovolt-ampere.......................................................................................................................... kVA

Kilovolts..................................................................................................................................... kV

Kilowatt ..................................................................................................................................... kW

Kilowatt hour.............................................................................................................................. kWh

Kilowatt hours per tonne (metric ton) .......................................................................................... kWh/t

Kilowatt hours per year............................................................................................................... kWh/a

Less than................................................................................................................................... <Litre........................................................................................................................................... L

Litres per minute........................................................................................................................ L/m

Loose cubic metre...................................................................................................................... lcm

Megabytes per second............................................................................................................... Mb/s

Megapascal ............................................................................................................................... MPa

Megavolt-ampere....................................................................................................................... MVA



xvii 0954780100-REP-R0007-02

Megawatt................................................................................................................................... MW

Metre......................................................................................................................................... m

Metres above sea level ............................................................................................................. masl

Metres Baltic sea level ............................................................................................................... mbsl

Metres per minute...................................................................................................................... m/min

Metres per second..................................................................................................................... m/s

Metric ton (tonne)....................................................................................................................... t

Microns ..................................................................................................................................... µm

Milligram.................................................................................................................................... mg

Milligrams per litre...................................................................................................................... mg/L

Millilitre...................................................................................................................................... mL

Millimetre................................................................................................................................... mm

Million........................................................................................................................................ M

Million bank cubic metres ........................................................................................................... Mbm3

Million bank cubic metres per annum.......................................................................................... Mbm3/a

Million tonnes............................................................................................................................. Mt

Minute (plane angle) .................................................................................................................. 'Minute (time) ............................................................................................................................. min

Month........................................................................................................................................ mo

Net operating hour ..................................................................................................................... noh

Ounce ....................................................................................................................................... oz

Pascal ....................................................................................................................................... Pa

Centipoise................................................................................................................................. mPa∙s

Parts per million......................................................................................................................... ppm

Parts per billion.......................................................................................................................... ppb

Percent...................................................................................................................................... %

Pound(s).................................................................................................................................... lb

Pounds per square inch ............................................................................................................. psi

Revolutions per minute............................................................................................................... rpm

Second (plane angle)................................................................................................................. "

Second (time) ............................................................................................................................ s

Specific gravity .......................................................................................................................... SG

Square centimetre...................................................................................................................... cm2

Square foot................................................................................................................................ ft2

Square inch ............................................................................................................................... in2

Square kilometre........................................................................................................................ km2

Square metre............................................................................................................................. m2

Thousand tonnes ....................................................................................................................... kt

Three Dimensional..................................................................................................................... 3D

Three Dimensional Model........................................................................................................... 3DMTonne (1,000 kg) ....................................................................................................................... t

Tonnes per day.......................................................................................................................... t/d

Tonnes per hour ........................................................................................................................ t/h

Tonnes per year......................................................................................................................... t/a

Tonnes seconds per hour metre cubed....................................................................................... ts/hm

Volt ........................................................................................................................................... V



xviii 0954780100-REP-R0007-02

Week......................................................................................................................................... wk

Weight/weight............................................................................................................................ w/w

Wet metric ton ........................................................................................................................... wmt

Year (annum)............................................................................................................................. a

AB B R E V I A T I O N S A N D A C R ON Y MS

acid potential.............................................................................................................................. AP

acid rock drainage ...................................................................................................................... ARD

Alcoa Inc. .............. .............. ............ .............. ............ .............. .............. ............ .............. ........... Alcoa

Aluminerie Lauralco, Inc................. .............. ............ .............. ............ .............. .............. ............. ALI

aluminum conductor steel reinforced ........................................................................................... ACSR

Amax of Canada Ltd. .............. ............... ........... ............... ........... ............... .............. ........... ........ Amax

Amax Zinc (Newfoundland) Limited ............. .............. ............ .............. ............ .............. .............. AZN

Archaeological Impact Assessment ............. .............. ............ .............. ............ .............. .............. AIA

atomic absorption spectrophotometer.......................................................................................... AAS

Avanti Kitsault Mines Ltd. ............. .............. ............ .............. ............ .............. .............. ............ ... AKM Avanti Mining Inc. ............. .............. ............ .............. ............ .............. .............. ............ .............. Avanti

BC Conservation Data Centre ..................................................................................................... BC CDC

BC Ministry of Forests and Range............................................................................................... MFR

Bell Molybdenum Mine Limited.................................................................................................... Bell

best judgement........................................................................................................................... BEJ

Bond Abrasion Index................................................................................................................... Ai

Bond Ball Mill Work Index ........................................................................................................... BWi

Bond Crushing Index .................................................................................................................. CWi

Bond Rod Mill Work Index........................................................................................................... RWi

British Columbia Environmental Assessment Office ..................................................................... BC EAO

British Columbia Ministry of Energy, Mines, and Petroleum Resources .............. .............. ............ . BC MEMPR

British Columbia Molybdenum Ltd. .............................................................................................. BC Molybdenum

British Columbia ......................................................................................................................... BC

Canadian Dam Association......................................................................................................... CDA

Canadian Environmental Assessment Act ................................................................................... CEAA

Canadian Institute of Mining, Metallurgy, and Petroleum .............................................................. CIM

Capital Cost Allowance ............................................................................................................... CCA

Climax Molybdenum Company of British Columbia ...................................................................... CMC

Committee on the Status of Endangered Wildlife in Canada......................................................... COSEWIC

compound annual growth rate ..................................................................................................... CAGR

Contract Support Services Inc..................................................................................................... CSS

Conveyor Equipment Manufacturer’s Association .............. .............. ............ .............. ............ ...... CEMA

CPM Group ................................................................................................................................ CPM

Dowfroth 250.............................................................................................................................. DF250

Earthquake Design Ground Motion.............................................................................................. EDGM

Energy & Metals Consensus Forecasts ....................................................................................... EMCF

engineering, procurement, and construction management.................... ............ .............. .............. EPCM

Environmental Assessment Act................................................................................................... EAA

Environmental Assessment......................................................................................................... EA



xix 0954780100-REP-R0007-02

exploratory date analysis ............................................................................................................ EDA

explosives storage facility............................................................................................................ ESF

Fleet Production and Cost Analysis ............................................................................................. FPC

Forest Service Road................................................................................................................... FSR

Freeport McMorran Inc................................................................................................................ Freeport

Gemcom Surpac™ ..................................................................................................................... Surpac

Gemcom Whittle™ v4.1.3 ........................................................................................................... Whittle

general and administrative.......................................................................................................... G&A

Geologic Strength Index.............................................................................................................. GSI

gravity .............. ............... ........... ............... ........... ............... .............. ........... ............... ........... .... g

Harmful Alteration, Disruption, or Destruction of Fish Habitat..................... ........... ............... ......... HADD

Hazen Research Inc. .................................................................................................................. Hazen

heating, ventilation, and air conditioning ...................................................................................... HVAC

Heritage Conservation Act ............. .............. ............ .............. ............ .............. .............. ............ . HCA

High Pressure Grinding Rolls ...................................................................................................... HPGR

high-density polyethylene............................................................................................................ HDPE

high-strength low-alloy................................................................................................................ HSLAhydroelectric power plant ............................................................................................................ HEP

inductively-coupled plasma ......................................................................................................... ICP

Inflow Design Flood .................................................................................................................... IDF

internal rate of return................................................................................................................... IRR

inter-ramp slope angle ................................................................................................................ ISA

Invasive Plant Council of BC ....................................................................................................... IPC BC

inverse distance weighting .......................................................................................................... ID

Kennco Explorations (Western) Ltd. ............................................................................................ Kennco

Kennecott Exploration Company ................................................................................................. KEL

Kilborn Engineering Ltd............................................................................................................... Kilborn

Kitimat-Stikine Regional District...................... ............ .............. .............. ............ .............. ........... KSRD

Kitsault Molybdenum project ....................................................................................................... KM project

Kitsault Resort Ltd. ..................................................................................................................... KRL

Kitsault Resort Ltd. ..................................................................................................................... KRL

Knight Piésold Ltd....................................................................................................................... KP

Land and Resource Management Plan........................................................................................ LRMP

leach-solvent extraction-electrowinning .............. .............. ............ .............. ............ .............. ....... SX/EW

life of mine.................................................................................................................................. LOM

London Metals Exchange............................................................................................................ LME

low-grade ore stockpile............................................................................................................... LGO stockpile

material coefficient...................................................................................................................... mi

Material Safety Data Sheet ......................................................................................................... MSDS

Maximum Design Earthquake ..................................................................................................... MDEmetal leaching and acid rock drainage......................................................................................... ML/ARD

Metal Mining Effluent Regulations ............................................................................................... MMER

Meteorological Services of Canada ............................................................................................. MSC

methyl isobutyl carbonyl.............................................................................................................. MIBC

Mineral Titles Online BC ............................................................................................................. MTO

Ministry of Energy and Mines ...................................................................................................... MEM



xx 0954780100-REP-R0007-02

Ministry of Environment............................................................................................................... MoE

Ministry of Environment, Lands, and Parks .................................................................................. MELP

Mintec, Inc.................................................................................................................................. Mintec

motor control centres .................................................................................................................. MCCs

National Instrument 43-101......................................................................................................... NI 43-101

nearest neighbour....................................................................................................................... NN

net present value........................................................................................................................ NPV

net smelter royalty ...................................................................................................................... NSR

neutralization potential................................................................................................................ NP

Nisga’a Final Agreement............................................................................................................. NFA

Nisga’a Lisims Government......................................................................................................... NLG

Nisga'a Final Agreement ............................................................................................................. NFA

non-potentially acid generating.................................................................................................... NPAG

Operating Basis Earthquake........................................................................................................ OBE

ordinary kriging........................................................................................................................... OK

Peak Ground Acceleration .......................................................................................................... PGA

potassium amyl xanthate ............................................................................................................ PAXpotentially acid generating........................................................................................................... PAG

Pre-feasibility Study .................................................................................................................... PFS

Preliminary Economic Assessment.............................................................................................. PEA

PricewaterhouseCoopers LLP..................................................................................................... PWC

probability of failure..................................................................................................................... POF

Probable Maximum Flood ........................................................................................................... PMF

Probable Maximum Precipitation................................................................................................. PMP

Provincial Sales Tax ................................................................................................................... PST

Qualified Person......................................................................................................................... QP

quality assurance/quality control.................................................................................................. QA/QC

quantile-quantile ......................................................................................................................... QQ

Quartz Monzonite Porphyry 1...................................................................................................... QMP1

Quartz Monzonite Porphyry ll ...................................................................................................... QMPll

Rainfall Frequency Atlas of Canada............................................................................................. RFAC

Remote Access to Archaeological Data....................................................................................... RAAD

Rescan Environmental Services Ltd. ........................................................................................... Rescan

Resource Development Inc......................................................................................................... RDi

Resource Inventory Committee ................................................................................................... RIC

Resource Inventory Standards Committee....... .............. ............ .............. .............. ............ .......... RISC

responsible authorities................................................................................................................ RA

return on investment ................................................................................................................... ROI

rock quality designation............................................................................................................... RQD

rotating biological contactor......................................................................................................... RBCrun-of-mine................................................................................................................................. ROM

SAG mill/ball mill circuit with pebble crushing............................................................................... SABC

Securities and Exchange Commission......................................................................................... SEC

selective mining unit ................................................................................................................... SMU

semi-autogenous grinding........................................................................................................... SAG

semi-autogenous mill comminution.............................................................................................. SMC



xxi 0954780100-REP-R0007-02

SGS Canada Inc......................................................................................................................... SGS

specific gravity............................................................................................................................ SG

SRK Consulting (Canada) Inc...................................................................................................... SRK Canada

SRK Consulting (U.S.) Inc........................................................................................................... SRK US

Statutory Right of Way No. BX201679......................................................................................... the ROW

Sustainable Resource Management Plan.................................................................................... SRMP

Tailing Management Facility........................................................................................................ TMF

Terrestrial Ecosystem Mapping ................................................................................................... TEM

traditional knowledge.................................................................................................................. TK

triaxial compressive strength....................................................................................................... TCS

unconfined compressive strength................................................................................................ UCS

United States Geological Survey ................................................................................................. USGS

vacation, sickness, and absenteeism........................................................................................... VS&A

Valued Ecosystem Components.................................................................................................. VEC

valued socioeconomic components ............................................................................................. VSC

variable frequency drive.............................................................................................................. VFD

variable radius point counts......................................................................................................... VRPCWardrop Engineering Inc............................................................................................................. Wardrop

Water Survey of Canada............................................................................................................. WSC

Work Breakdown Structure ......................................................................................................... WBS

Workplace Hazardous Materials Information Systems............... .............. ............ .............. ........... WHMIS

WorleyParsons Canada Ltd., W estmar Division ........................................................................... WPW

x-ray fluorescence spectrometer ................................................................................................. XRF

Xstrata plc.................................................................................................................................. Xstrata



Avanti Mining I nc. 1-1 0954780100-REP-R0007-02 NI 43-101 Pre-feasibility Study – Avanti Mining Inc.

Kitsault Molybdenum Property, British Columbia, Canada

1 . 0 S U M M A R Y

1 . 1 I N T R O D U C T I O N

This National Instrument 43-101 (NI 43-101) compliant Pre-feasibility Study (PFS) on

the Kitsault Molybdenum (KM) project has been prepared by Wardrop EngineeringInc. (Wardrop) for Avanti Mining Inc. (Avanti) based on work by the following

independent consultants:

SRK Consulting (Canada) Inc. (SRK Canada)

SRK Consulting (U.S.) Inc. (SRK US)

Knight Piésold Ltd. (KP)

Resource Development Inc. (RDi)

Rescan Environmental Services Ltd. (Rescan)

WorleyParsons Canada Ltd., Westmar Division (WPW)

CPM Group (CPM).

Mr. Jeffrey Volk (P.Geo., FAusIMM) of SRK (US) visited the property on August 9,2008, and is the Qualified Person (QP) for all matters relating to the mineral resource

estimate.

Mr. Michael Levy (P.E., P.G.) of SRK (US) visited the property on September 8 to 11,2008 and is the QP for matters relating to the pit slopes.

Mr. Peter Healey (P.Eng) of SRK (Canada) has visited the property frequently, mostrecently on June 29, 2009 and is the QP for matters and costs relating to mineclosure and reclamation.

Mr. Stephen Day (P.Geo.) of SRK (Canada) has visited the property on September 21 to 22, 2009, and is the QP for matters relating to metal leaching and acid rock

drainage.

Mr. Miloje Vicentijevic (P.Eng.) of Wardrop visited the property on June 28 to 29,2009 and is the QP for matters relating to mineral reserve statements mining, mining

capital, and mine operating costs.

Mr. Andre de Ruijter (P.Eng.) of Wardrop visited the property on July 20 to 21, 2009and is the QP for matters relating to the metallurgical testing review, mineral

processing, and process operating costs.





Mr. Deepak Malhotra (P.Eng.) of RDi is the QP for matters relating to metallurgicaltesting in 2009.

Mr. Frank Grills (P.Eng.) of Wardrop visited the property on July 20 to 21, 2009 and

is the QP for matters relating to the process capital cost estimate and infrastructure.

Mr. Ken Brouwer (P.Eng.) of KP visited the property on July 20 to 21, and is the QP

for matters and costs relating to plant site geotechnical conditions, surface water

diversions (including the diversion tunnel), tailing embankments, tailing access roads,tailing/reclaim pipeworks, and the hydroelectric power plant.

Mr. Rolf Schmitt (P.Geo.) of Rescan is the QP for matters relating to environmental

assessment, which involved qualified professionals conducting investigations on theproperty from October 2008 through November of 2009.

This PFS capital cost estimate has been prepared to a ±30% level of accuracy and

the operating costs, an accuracy of ±25%. All dollar figures are expressed in US

dollars, unless otherwise specified. An exchange rate of US$0.92 to Cdn$1.00 hasbeen utilized where applicable.

The authors are relying on disclosures by Avanti on the recent change in ownershipand financial agreements, and have not reviewed any of the pertinent legaldocuments.

On September 16, 2008, Avanti announced that it had been denied access to theghost town of Kitsault by its owner, Kitsault Resort Ltd. (KRL). At the time, Avantiheld a license and permit from the owner (Aluminerie Lauralco, Inc. (ALI)) of a

statutory right of way to use the roads within the town of Kitsault for mine related

activities. As a result, Avanti, in its capacity as a licensee and permittee,commenced an action in British Columbia Supreme Court seeking:

1. a declaration that it is entitled to use the roads in the town, and

2. a permanent injunction preventing KRL from denying Avanti use of the rightof way.

Avanti is now the registered owner of the right of way and has amended its status inthe above proceedings.

1 . 2 P R O P E R T Y D E S C R I P T I O N A N D L O C A T I O N

The Kitsault property is located about 140 km north of Prince Rupert, BC, and southof the head of Alice Arm, an inlet of the Pacific Ocean (Figure 1.1). The property

includes three known molybdenum deposits: Kitsault, Bell Moly, and Roundy Creek.

The Kitsault mine site is located within NTS maps 103P044 and 043, atapproximately latitude 55°25'19″N and longitude 129°25'10″W, at about 600 m





elevation. The 8,286 ha of mineral leases and mineral tenures are fully described inthe following sections. The principal mining feature on the property is the Kitsault

open pit mine, which is not currently operating; it is shown in Figure 1.2, in relation tothe mineral tenures in the area.

The molybdenum deposit at Bell Moly is located at approximately latitude 55°28'Nand longitude 129°20'W, at about 750 m elevation. The Roundy Creek molybdenumdeposit is located at about latitude 55°24'49″N and longitude 129°29'32″W at

approximately 320 m elevation (Figure 1.2). The Bell Moly and Roundy Creek

deposits do not form part of this study.

1 . 3 O W N E R S H I P

Avanti had previously signed a Definitive Purchase and Sales Agreement with ALI, to

acquire an undivided, 100% direct interest in the Kitsault molybdenum mine and

surrounding mineral tenures, located in northern BC, subject to a 120-dayconfirmatory due diligence period and regulatory approval (June 20, 2008 Press

Release). The agreement structure requires Avanti to pay US$20 M to ALI for a

100% interest in the Kitsault property. ALI has a 1% Net Smelter Royalty on futureproduction subject to the right, within 90 days from the presentation of a Bankable

Feasibility Study, to elect to surrender the Net Smelter Royalty in exchange for either an additional US$10 M payment (payable at Commercial Production or in Avanti

shares at their election). On October 20, 2008, Avanti announced that they had

completed the purchase (on October 17, 2008) of the Kitsault property.

A finder’s fee for 2,000,000 fully-paid and not-assessable Avanti common shares

was payable to a third party on the closing of the Kitsault acquisition. A success fee

of 500,000 Avanti common shares was paid to a Financial Advisor. Both of thesefees were paid to the respective parties at the closing of the Kitsault acquisition onOctober 17, 2008.





Figure 1.1 Kitsault Location and Regional Geology



Avanti Mining Inc. 1-5 0954780100-REP-R0007-02 NI 43-101 Pre-feasibility Study – Avanti Mining Inc.


Figure 1.2 Mineral Tenure Map and Location of Molybdenum Deposits on Avanti Kitsault Mine Ltd. Tenures





1 . 4 A C C E S S I B I L I T Y , C L I M A T E , L O C A L R E S O U R C E S ,

I N F R A S T R U C T U R E , A N D P H Y S I O G R A P H Y

The Kitsault property is in upland, hilly plateau country characterized by thick stands

of timber interspersed with small lakes, meadows, and swamps. The dominanttopographic features are a series of eroded basaltic lava flows that commonly formcliffs up to 100 m high.

Generally, topography rises quickly from tidewater at Alice Arm to an elevation of 600 to 800 m at a plateau. The Kitsault and Bell Moly sites are on the plateau(Figure 1.3) and the Roundy Creek site is midway up the elevation change from

tidewater to the plateau.

Bedrock is generally blanketed by a few metres of glacial till, which in turn iscommonly overlain by a layer of peat bog up to 1 m thick. The rock outcrop in this

area, except for the basalt cliffs, commonly averages less than 1%.

The climate in the Kitsault area is temperate coastal, verging on a rain forestenvironment. Annual average precipitation at Alice Arm and Anyox is about 2 m. At

the Kitsault mine site, a majority of the precipitation is snow. The mine can operate

365 d/a, with no days lost to snow stoppages.

The KM property can be accessed via water, float plane, or by a 100-km northbound

paved road from Terrace to Nass Camp, and then a further 95 km via an upgraded

gravel road to site.

Alternatively, the site can be accessed by an alternative road route from Terrace via

Highway 37 to Cranberry Junction (175 km) then via a gravel connector road fromCranberry Junction to the KM site (approximately 106 km).

For the purposes of this PFS, the port facility is the preferred access proposed to

ship equipment and supplies via barge to the Kitsault site. The equipment and

supplies will be offloaded at the port site and transported to the mine site via theupgraded access road. The new port is expected to be located on a leased 60-ha

Crown- land site south of Kitsault, BC. The port’s berth facilities and theconstruction/permanent accommodation complex facility will be located in this area.





Figure 1.3 General View of the Kitsault Mine Area





1 . 5 H I S T O R Y

Historical mineral resource and reserve estimates presented in this report are based

on information collected and compiled from the 1950s to the early 1980s. Historic

resource estimates do not comply with the Canadian Institute of Mining, Metallurgy,and Petroleum (CIM) terminology under NI 43-101 guidelines. These estimates are

not mineral resources or mineral reserves and should not be relied upon.

Molybdenum was first reported in the Alice Arm region during World War I, when a

limited quantity of molybdenite was produced from the Tidewater property. At

approximately the same time, molybdenite was reported in the Kitsault and RoundyCreek areas. Serious interest in the Kitsault area began in 1956 when KenncoExplorations (Western) Ltd. (Kennco) examined the property and optioned it the next

year. A number of historic resource estimates were conducted during the 1950s

through the 1980s. These historic resource estimates do not comply with CIM

terminology under NI 43-101.

Kennco began mining the property in 1968, based on their reserve of 36 Mt with an

average grade of 0.138% Mo (0.23% MoS2). Mining was terminated in 1972 due tolow metal prices. Climax Molybdenum Company of British Columbia (CMC)

purchased the property in 1973 and returned it to production in 1981 based on anexpanded reserve of 142 Mt with an average grade of 0.109% Mo (0.182% MoS 2).

Production was again terminated because of low metal prices in 1982. During these

two periods of production approximately 30 Mlbs of molybdenum was produced (BCMINFILE Report number 103P 120). Bell Moly was recognized as a molybdenumproperty in 1965 by Mastedon Highland Bell Mines and Leitch Gold Mines who later

formed Bell Molybdenum Mines that was subsequently purchased by CMC in 1975.

Drilling by both companies identified a resource of 96.4 Mt with an average grade of 0.054% Mo (0.09% MoS2). Drilling by Silurian Chieftain Minerals at Roundy Creek

during the late 1960s resulted in a resource of about 7 Mt with an average grade of 0.066% Mo (0.11% MoS2). Subsequent work by Amax of Canada Ltd. (Amax)resulted in a 1983 resource estimate of 4.24 Mt at an average grade of 0.129% Mo

(0.215% MoS2), found in three separate zones. CMC purchased the property in

1975. CMC transferred the Kitsault property (including Bell Moly and Roundy Creek)to Amax who eventually moved the property to its Alumax division. Alumax was

eventually spun-off to Amax shareholders and later purchased by Alcoa, which wasrenamed ALI.

The Lime Creek Intrusive Complex is composed of at least three stocks and

associated suites of dikes that were emplaced at about 54 Ma, the last two of whichare associated with mineralization in the Kitsault molybdenum deposit.

Compositionally the complex varies from an older diorite in the south to a quartz

monzonite in the central and northeastern portion of the complex. The host rock tothe Lime Creek intrusive complex is the Jurassic Bowser Group, a thick, regional

sequence of fine-grained siltstones and greywackes. The mineralization at Kitsault isrelated in time and space to one of the largest, younger intrusive bodies designated





by Avanti as the Quartz Monzonite Porphyry 1 (QMP1) as well as to younger aplitedikes. The main stage mineralization is cut by a closely-related and largely buried

intrusive body designated by Avanti as the Quartz Monzonite Porphyry ll (QMPll).Mineralization consists of typical stockwork and sheeted, quartz-molybdenite

veinlets±(pyrite and potassium feldspar). Wall rock alteration includes an early and

largely central core of barren quartz veining and earlier feldspar veining andreplacement. This is followed by widespread wallrock silification and sericitization,and late pervasive kaolinization. The overall deposit is annular in plan and cylindrical

or columnar in section.

At Bell Moly, there are two centers of intrusive activity and related mineralizationwhich are hosted by Bowser Lake Group greywacke. The Clary Creek stock consists

of at least five separate intrusions all of which are quartz monzonitic in composition,while the Southwest Zone is primarily a dike swarm composed of quartz monzonite.

Intrusive activity at Bell Moly is of a similar age to that at Kitsault. Quartz-

molybdenite mineralization is closely associated in time and space with the intrusive

activity in the Clary Creek stock and the dikes in the Southwest Zone. Severalsuperimposed ages of mineralization have been defined with an alteration pattern

similar to that at Kitsault.

The Roundy Creek intrusions are also hosted by Bowser Lake Group greywacke andconsist of several 52 Ma quartz monzonite stocks and dikes. Three separate areas

of mineralization are defined by drilling and two adits. The western two areas are the

Sunshine and Sunlight zones and these appear to be well defined by drilling, whilethe more easterly Roundy Zone may still be open at depth.

An initial view of the geology, molybdenum distribution, and geochemistry of all three

mineralized areas indicate that additional exploration targets exist. There is the

probability that additional zones of molybdenum mineralization, and/or expansions of the known deposits is likely.

Avanti had previously signed a Definitive Purchase and Sales Agreement with ALI, to

acquire an undivided, 100% direct interest in the Kitsault mine and surroundingmineral tenures, located in northern British Columbia (BC), subject to a 120-dayconfirmatory due diligence period and regulatory approval (June 20, 2008 Press

Release). The agreement structure required Avanti to pay US$20 M to ALI for a

100% interest in the KM property. ALI has a 1% Net Smelter Royalty (NSR) onfuture production subject to the right, within 90 days from the presentation of a

Feasibility Study suitable for submission to international financial institutions, to electto surrender the NSR in exchange for an additional US$10 M payment payable at

Commercial Production or in Avanti shares at ALI’s election. On October 20, 2008, Avanti announced that they had completed the purchase (on October 17, 2008) of the KM property. The KM property is held by Avanti’s wholly owned subsidiary,

Avanti Kitsault Mine Ltd.





1 . 6 G E O L O G I C A L S E T T I N G

The Kitsault, Roundy Creek, and Bell Moly molybdenum deposits are located within

the western margin of the Bowser basin in the Intermountain tectonic belt, a fewkilometres east of the Coast Range Crystalline Complex.

The area is characterized by intense intrusive activity related to the Coast Range

Crystalline Complex, with younger stocks intruding the sedimentary lithologies andnumerous recent plateau-type lava flows. Intrusive rocks in the Coast RangeCrystalline Complex range in composition from granodiorite to quartz monzonite.

Numerous 50 to 55 Ma granodiorite to quartz monzonite stocks occur in the

sedimentary rocks surrounding Alice Arm which Carter (1981) referred to as the Alice Arm intrusives. Hosts for these stocks are the Lower to Middle Jurassic Hazelton

Formation and Upper Jurassic to Lower Cretaceous Bowser Lake Group. TheHazelton Formation consists of volcanic breccias, tuff, conglomerate, volcaniclastic

sedimentary rocks, and andesite flows, all metamorphosed to greenschist facies.

The Bowser Lake Group consists of interbedded greywacke and argillite with minor conglomerate and limestone metamorphosed to greenschist facies.

Many of the Alice Arm intrusives are the loci for molybdenum mineralization. In

addition to the Kitsault molybdenum deposit, there is intrusive activity and associatedmolybdenum mineralization at Roundy Creek, Bell Moly, Tidewater, and Ajax. All of these resource areas except for Ajax are located in the immediate area of Alice Arm

and all have been the sites of historic exploration and drilling from the 1960s to the1980s with the exception of Kitsault, which was had additional drilling conducted in

2008, and Ajax which had additional drilling conducted in 2007. The Kitsault deposit

is situated about 5 km from ocean access at Alice Arm and lies immediately to thenortheast of the junction of Lime and Patsy Creeks.

Following the emplacement of the Alice Arm intrusives and the related molybdenum

mineralization, a swarm of northeast striking lamprophyre dikes, which are dated at36 ± 1.2 Ma and 34.4 ± 1.2 Ma (Carter, 1974), intruded into the Bowser Lake Group.The youngest igneous event in the region is basaltic plateau-type flows and related

dikes dated at 1.6 ± 0.8 Ma and 0.62 ± 0.6 Ma (Carter, 1974). The entire area was

exposed to glaciation, much (but not all) of which has occurred post-basaltic flows.

Host lithologies for mineralization at Kitsault, Bell Moly, and Roundy Creek are

thermally metamorphosed interbedded argillite and greywacke of the Upper Jurassic

to Lower Cretaceous Bowser Lake Group, and the intrusives of the Early TertiaryLime Creek Intrusive Complex, Clary Creek Stock, and the Roundy Creek intrusive

complex. Intrusive rocks associated with molybdenum mineralization at Kitsault, BellMoly, and Roundy Creek are multiphase diorite, quartz monzonite, and younger

felsic units. Surrounding the intrusives are hornfels aureoles that extend up to 750 m

outward from the intrusive contact. Away from the KM open pit and the adits atRoundy Creek, nearly all surface rock exposures are limited, making an

interpretation of the area’s geology extremely difficult. Principally, the area iscovered by soil, swamp, glacial till, and in places basalt flows. Most of the geological





knowledge for Avanti’s property position comes from the numerous drillholes at thethree deposits.

1 . 7 D E P O S I T T Y P E S

The Kitsault, Bell Moly, and Roundy Creek properties contain intrusive-relatedmolybdenum deposits that, in plan view, are geometrically annular and tend to be

cylindrical-to-tabular and arcuate in a cross-sectional view.

This group of intrusive-related molybdenum deposits differs distinctly from the quartzmonzonite batholithic type, where molybdenum mineralization is related to a series of

sheeted vein systems. These latter deposits are usually more lenticular in shape and

are commonly related to the youngest phases of the batholithic rocks which hostthese deposits.

Although Kitsault shares some similarities with the high-silica, rhyolitic to alkalic

deposits of the western US, with respect to time-space relationships of mineralizationwith its host intrusives, it differs from these deposits in both rock chemistry and in

morphology. This latter group of rhyolitic to alkalic systems is best typified by theClimax and Henderson deposits (Colorado), which are the world’s two largestporphyry molybdenum deposits.

1 . 8 M I N E R A L R E S O U R C E A N D M I N E R A L R E S E R V E E S T I M A T E

1.8.1 M I N E R A L R E S O U R C E E S T I M A T E

The resource estimate has been generated from drillhole sample assay results and is

constrained by the interpolation of a grade probability model which relates to thespatial distribution of molybdenum in the deposit. Interpolation characteristics havebeen defined based on the geology, drillhole spacing, and geostatistical analysis of

the data. The resources have been classified by their proximity to the sample

locations. SRK US has conducted an audit of the resource model and finds it to beacceptable for resource reporting under the CIM “Estimation of Mineral Resources

and Mineral Reserves Best Practice Guidelines”.

A new Lidar survey was conducted during the 2008 field season and that informationhas been utilized to construct a current topographic surface for use as a constraint in

resource estimation.

The mineral resources at the Kitsault deposit have been classified in accordance withthe CIM definition standards for mineral resources and mineral reserves (December

2005). The classification parameters are defined in relation to the distance to sampledata are intended to encompass zones of reasonably continuous mineralization.





During the grade estimation process, distance to closest composite, averagedistance and number of drillholes used to estimate the block were stored in the block

model. Using these values as a basis, blocks were classified as follows:

Measured Mineral Resources – blocks with Mo% grades estimated using a

minimum of 3 drillholes within an average distance of 35 m

Indicated Mineral Resources – blocks in the model estimated using aminimum of three drillholes that are at maximum average distance of 90 m

Inferred Mineral Resources – blocks in the model not meeting the criteria for

indicated resources but are within a maximum distance of 150 m from a

drillhole.

This classification is based on detailed drillhole spacing analysis and assignment of confidence intervals. This analysis has been reviewed by SRK and found to be adefendable basis for resource classification.

Mineral resources are not mineral reserves and do not have demonstrated economicviability. There is no certainty that all or any part of the mineral resource will beconverted into mineral reserves. The effective date of this resource estimate/audit is

March 31, 2009. The mineral resources statement for the KM project is presented inTable 1.1.


March 31, 2009

Resource Classification

Qty(Mt)

Grade Contained Metal

Mo(%)

Ag(g/t)

Pb(%)

WO3

(%)Mo

(Mlb)Ag

(Moz)Pb

(Mlb)WO3

(Mlb)

Measured** 54 0.112 4.54 0.022 0.007 133 8 26 8

Indicated** 153 0.088 5.24 0.025 0.006 297 26 84 20

Measured & Indicated** 207 0.094 5.06 0.024 0.006 430 34 110 28

Inferred** 26 0.069 4.15 0.019 0.005 40 4 11 3

*Mineral resources are not mineral reserves and do not have demonstrated economic viability. All

figures have been rounded to reflect the relative accuracy of the estimates. The cut-off grades are

based on metal price assumptions of US$20.00/lb Mo, and a metallurgical recovery of 89% Mo.Silver, lead, and WO3 were not used in the pit optimization.

** Reported at a cut-off grade of 0.04 % Mo contained within a potentially economic open pit.

The mineral resources are reported at a cut-off grade to reflect the “reasonable

prospects” for economic extraction. SRK considers that portions of the Kitsaultmolybdenum deposit are amenable for open pit extraction, and has not consideredunderground mining methods for deeper portions of the deposit.

Table 1.2 reflects the material present at various cut-off grades.





Table 1.2 Mo Cut-off Grade Sensitivity Analysis within Resource Pit –

Measured and Indicated Resources

Cutoff Grade(Mo%)

Quantity(Mt)

Mo Grade(%)

Contained MetalMo (Mlb)

0.02 315 0.072 5010.025 292 0.076 490

0.03 257 0.083 468

0.035 234 0.088 452

0.04 207 0.094 430

0.045 197 0.097 420

0.05 189 0.099 412

0.055 183 0.100 405

0.06 178 0.102 398

0.065 170 0.103 388

0.07 163 0.105 377

0.075 151 0.108 3570.08 134 0.111 329

Avanti has completed a significant volume of work during the 2008 field season,

which included a 33-hole (10,131 m) infill drilling program. This program focused onboth conversion of inferred to indicated resources as well as the confirmation of thehistorical drilling results. The results of this program have been reviewed; the assay

results from the 2008 drilling compare closely with the assays from the historicdrilling and are overall confirmatory. In addition, the 2008 program successfully

converted inferred resources to indicated and measured resources, and the current

drillhole spacing within the Kitsault deposit appears sufficient to advance the projectto feasibility level studies.

1.8.2 M I N E R A L R E S E R V E E S T I M A T E

The Kitsault Mine Mineral Reserves have been prepared in accordance withNI 43-101 standards and CIM standard definitions. This reserve statement has been

prepared by Mr. Miloje Vicentijevic (P.Eng.) of Wardrop, a QP as defined in NI 43-

101. These reserves are sufficient for close to 15 years of operations at an annualproduction rate of 40,000 t/d. Mineral Reserves are summarized by class inTable 1.3.

The notes accompanying Table 1.3 are an integral part of the Mineral Reserves andshould be read in conjunction with the Mineral Reserve statement.





Table 1.3 Mineral Reserves by Class

By ClassCut-off

Grade (Mo%)Quantity

(Mt)Mo Grade

(Mo%)ContainedMetal (Mlb)

Proven 0.036 55.7 0.109 121.35

0.027 3.8 0.031 2.35Total 59.5 0.104 123.70

Probable 0.036 134.5 0.086 231.08

0.027 21.2 0.031 13.13

Total 155.7 0.079 244.21

Total Proven & Probable 215.3 0.085 367.91

Notes:

1. Reserves calculated in accordance with CIM guidelines.

2. The metal price used for reserve calculation is US$12.51/lb Mo.

3. Metallurgical recovery is 90.6% for Mo.

4. Pit optimization parameters have changed from the time the resource estimate was completed.

As a result, an additional 8.3 million tonnes grading 0.031% Mo of the economic reserveswithin optimized pit was available for the variable cut-off strategy (Note 5 below) making thereserve statement higher than the resource statement (see Tables 1.1 & 1.2)

5. Cut-off grades used were variable, 0.036% Mo and 0.027% Mo.

6. Mining recovery is estimated at 100% and dilution is nil.

7. The waste-to-ore ratio for the deposit is 0.75.

1 . 9 O V E R A L L G E N E R A L A R R A N G E M E N T

Figure 1.4 gives an overview of the arrangement of the project.





Figure 1.4 Overall General Arrangement





1 . 1 0 M I N E R A L P R O C E S S I N G A N D M E T A L L U R G I C A L T E S T I N G

Avanti is proposing the mining and processing of 40,000 t/d of the molybdenum-

bearing resource material. For this study, historical testwork results together with

information supplied by RDi taken from the 2009 testwork program will be used as abasis for the process plant design parameters.

1.10.1 M E T A L L U R G I C A L T E S T W O R K T I ME P E R I O D S

The major testwork and metallurgical reviews, which include laboratory testworkprograms and two plant operation periods, are classified below:

1963 to 1964: testwork focused on mineralogy, flotation, and grindability

1967 to1974: plant operations

1975 to 1979: Feasibility Study

1981 to 1982: plant operations

1983: testwork

2008: Preliminary Economic Assessment

2009: testwork focusing on flotation and grinding under the direction of RDiat SGS Vancouver and Hazen Research Inc., respectively.

1.10.2 M I N E R A L O G I C A L E X A MI N A T I ON

The primary economic mineral in the Kitsault deposit is molybdenite, which iscontained in a stockwork of quartz veinlets. Other minerals of interest that occur inveins throughout the deposit are pyrite and gypsum, followed by minor amounts of

scheelite, galena, sphalerite, chalcopyrite, lead-bismuth sulphosalts, pyrrhotite and

carbonate minerals.

1.10.3 H I S T O R I C A L M E T A L L U R G I C A L R E V I E W

The metallurgical testwork and plant data review will follow the sequence as outlinedabove.

Initial testwork in 1963 returned a poor molybdenum recovery of 65.5% although it

achieved a concentrate grade 53.9% Mo. The best recovery value obtained from thistest program was 80.7% with a final concentrate grade of 52.0% Mo.

A 1964 test program produced a 51.7% Mo concentrate grade with 93.9% recovery.

A different test program in 1964 with a sample feed grade of about 0.19% Moresulted in molybdenum recoveries of between 93.4 to 95.6%, with concentrate

grades varying between 49.9 and 51.7% Mo. This testwork indicated that a relatively





coarse primary grind of 30% plus 105 µm could be anticipated although aconcentrate regrind to 10% plus 74 µm was also required. From the laboratory test

results, a 6,000 tons per day conventional mill was designed. The flowsheetincluded three crushing stages, two rod and ball mill grinding stages, flotation (which

consisted of a rougher stage followed by four stages of cleaning and included

regrinding), and product dewatering and packaging stages.

K I T S A U L T CO N C E N T R A T O R P R O D U C T I O N D A T A 1967 T O 1 9 7 2

The Kitsault concentrator was in operation for five years from 1967 to mid-1972, with

a nominal throughput of 6,000 tons per day. During that time 10,439 t (23 M lb) of molybdenum was recovered from 9.3 Mt of ore, with an average feed grade of 0.123% Mo. The average molybdenum recovery during this period was 89.2%.

P R E - P R O D U C T I O N S T A G E 1 9 7 6 T O 1 9 8 0

A Feasibility Study was conducted in 1976 with the view to re-opening the Kitsaultmine. The plant design was expanded to a throughput capacity of 12,000 tons per

day. The use of existing equipment was a paramount design criterion and thereforedictated some of the process design choices.

Grindability test results gave Bond Ball Mill Index values which varied significantly

between 9.8 and 15.5 kWh/short ton. Using all the available test data, the value of 15.5 kWh/short ton was recommended for the plant design.

A grind size versus flotation recovery test program determined that the optimum

grind size would be 45-50% +150 µm. The laboratory-scale test results indicated

that five stages of cleaning, rather than four, would be required to producemolybdenum concentrate grades which were greater than 54% Mo. As a finalanalysis of the data, the results from the rougher and cleaner tests were evaluated todetermine the overall expected recovery for the flotation process. Based on these

results, and using four stages of cleaning, an overall recovery of 84 to 92% wasanticipated with concentrate grades of not more than 54% Mo. With the addition of the fifth regrind and cleaning stage, recoveries of between 89 and 95% for

molybdenum were indicated with concentrate grades of the order of 58% Mo.

K I T S A U L T CO N C E N T R A T O R P R O D U C T I O N D A T A 1981 T O 1 9 8 2

The Kitsault concentrator was re-opened in April 1981 with a design capacity of

12,000 tons per day. The plant operated for 17 months with a shutdown period of

one month in August 1982, and a reduced work week until the mine ceasedoperations in October 1982. Production data for the Kitsault concentrator indicates

that 4.0 Mt was mined with an average feed grade 0.120% Mo producing a totalamount of 4,100 t (9 M lb) of molybdenum. The molybdenum recovery for the 1981to 1982 period was 85.33% for an average concentrate grade of 54.22% Mo,

although the lead content remained relatively high at 0.032% Pb despite the inclusion





of a lead leach circuit. Of note is that the 2009 study gives a recovery of 90.6% for an average concentrate grade of 52.0% Mo.

S G S L A K E F I E L D P I L O T P L A N T ST U D Y 1 9 8 3

During 1983 a pilot plant test program was undertaken, and poor molybdenumgrades and recoveries were obtained in this study which employed up to seven

stages of cleaning. Final molybdenum concentrate values of about 48% Mo wereobtained.

P R E L I M I N A R Y E C O N O M I C A S S E S S M E N T

SRK conducted a Preliminary Economic Assessment (PEA) in 2008 and utilised

conventional processing facility but which excluded a lead leach circuit. SRK electedto impose a high-lead content smelter penalty in the financial analysis instead. The

present study has included a lead leach circuit in the design which is similar to the

process that was successfully employed by the Kitsault mine during the 1981 and1982 operating years, and which mitigated that the lead content of the final product

was within specification.

L E A D L E A C H I N G T E S T W O R K R E V I E W

The removal of lead from the Kitsault molybdenum flotation concentrates by leaching

was investigated widely in the period from 1968 to 1983 in order to reduce the leadcontent to comply with a product specification level of either <0.10%Pb, or the higher

quality containing <0.02%Pb.

The occurrence of lead in the Kitsault ores was studied and it was concluded thatlead occurred mainly as the mineral galena, and that galena was present as

occluded grains of <10 µm in size within the particles of molybdenite.

The Kitsault plant design in 1981 incorporated a hydrochloric acid leach at elevatedtemperatures which resulted in a leach removal efficiency of about 85% reducing the

lead content from 0.33% to 0.05% Pb which was within the product specificationlimits of < 0.10% Pb

1.10.4 RD I T E S T P R O G R A M

The 2009 RDi testwork program conducted by SGS confirmed the metallurgicalprocess flowsheet that was used in the design of the Kitsault plant. The resultsobtained indicated that a molybdenum recovery of about 90% at a concentrate grade

of up to 54% Mo was attainable upon optimization of the testwork based on theprocess flowsheet as used in the design of the plant.





1 . 1 1 M I L L

The Kitsault concentrator has been designed to process a nominal 40,000 t/d of

molybdenite-bearing ore. The metallurgical processes selected have been designedto produce a saleable high grade molybdenum flotation concentrate containing 52%Mo. Historical plant data from the periods when the mine was operational, namely

1966 to 1967, and 1980 to 1982, as well as metallurgical testwork performed over

the years 1960 to 1990, together with the results of the 2009 testwork, were used inthe selection of the unit processes. In addition, resources set out by Avanti were

also employed. The unit processes employed for the recovery of the mineralmolybdenite will involve conventional size reduction and mineral beneficiationmethods. By-product mineral recovery test results are still being studied and were

excluded from the process design. The tailing disposal design allows for the flotationtailing to be deposited in a conventional tailing pond. Water utilization will be basedon maximizing re-use of process water as water reclaimed from the tailing facility.

Fresh water will only be used for gland water for pumps, reagent preparation, and

cooling of hydraulic units.

1.11.1 M A J O R D E S I G N C R I T E R I A

The concentrator was designed to process 40,000 t/d, equivalent to 14,600,000 t/a.

The molybdenum head grade value used in the design was 0.093% Mo. The major process design criteria parameters used in the design are outlined in Table 1.4. The

grinding data is based on weighted averages of the three rock types being treated,namely monzonite (55%), diorite (25%), and hornfels (20%).

Table 1.4 Major Process Design Criteria

Criteria Unit

Operating Year d 365

Crushing Availability % 80.0

Grinding and Flotation Availability % 92.0

Primary Crushing Rate t/h 2,083.3

Milling and Flotation Process Rate t/h 1,811.6

SAG Mill Feed Size, 80% Passing µm 150,000

SAG Mill Transfer Size, 80% Passing µm 1,500

Ball Mill Circulating Load % 250

Ball Mill Grind Size, 80% Passing µm 250

SAG Breakage Parameter A x b 49.4

SAG Breakage Parameter ta 0.47

Bond Ball Mill Work Index kWh/t 14.58

Bond Abrasion Index g 0.4632

Concentrate Regrind, 80% Passing µm 24

Flotation Concentrate Grade % Mo 52.0

Overall Flotation Recovery % 90.6





1.11.2 P R O C E S S P L A N T D E S C R I P T I O N

C R U S H I N G

A conventional gyratory crusher facility will reduce the size of the rocks in preparationfor the grinding process. The crushing rate will be 2,083 t/h to a product size P80 of

150 mm. The crusher product will be delivered to the crusher ore stockpile.

The open crushed ore stockpile will have a live capacity of 40,000 t. The ore will bereclaimed from this stockpile by apron feeders which will feed the SAG mill using

conveyors. All exterior conveyors will be covered.

G R I N D I N G A N D C L A S S I F I C A T I O N

The grinding circuit will consist of a SAG-ball mill combination circuit. It will be a2-stage operation with the SAG mill in closed circuit with a pebble crusher, and the

SAG mill and ball mill product in closed circuit with the classification cyclones. TheSAG mill will be equipped with a trommel screen to remove pebbles for crushing.

The grinding will be conducted at a nominal rate of 1,812 t/h of new feed material.

The SAG mill will be 10,359 mm in diameter, and have a length of 5,791 mm and willbe equipped with an 11 MW motor. The ball mill will have a diameter of 7,320 mm

and a length of 12,190 mm. The ball mill will also be equipped with an 11 MW motor.

The classification cyclones will produce a product with a cut size P80 of 250 µm.Each cyclone overflow stream will be monitored by a particle size analyzer. The two

product streams will be discharged to the flotation feed conditioning tank ahead of

the flotation process.

F L O T A T I O N C I R C U I T

The milled ore will be subjected to flotation to recover the mineral molybdenite into ahigh-grade molybdenum metal concentrate. The conditioned slurry will overflow theconditioning tank and will be the feed to the molybdenum flotation circuit. The

molybdenum flotation circuit will consist of a rougher and scavenger circuit, with four stages of cleaning including a 1st cleaner scavenger circuit. Three stages of

regrinding have been incorporated into the flotation circuit which will successively

reduce the particle size of the concentrate to a P80 value of 24 µm. Cleaner tailingstreams will be recycled to the preceding stage. Reagents will be added as required,

and stage-wise, if necessary.

C O N C E N T R A T E T H I C K E N I N G

The molybdenum flotation cleaner concentrate will be collected in the concentratethickener. The thickener underflow will be pumped to the lead leach circuit for the

removal of the mineral galena.





L E A D L E A C H C I R C U I T

The lead leach circuit will reduce the content of lead mineral impurities in the

molybdenum concentrate to acceptable limits for sales purposes. The lead leachcircuit design will include a three-tank continuous leach circuit. The concentrate will

be contacted with a hot solution of hydrochloric acid. Steam will be used to maintainthe temperature of the slurry in the leach tanks. On completion of the leach cycle,the leached concentrate slurry will be filtered and washed using a filter press. Thewashed filter cake, which is the leached molybdenum concentrate, will be discharged

onto a conveyor belt, which will feed the concentrate dryer and bagging system.

The filtrate and wash solution from the filter press will be recycled for re-use. Filtratewill also periodically be bled off out of the leach circuit to reduce the build-up of

concentration of the metals. The bleed filtrate will be discharged with the final tailing.

SRK understands that the flow from this stream is relatively low, compared to theoverall discharge into the pond, and the impact on the water quality of the

supernatant is unlikely to be significant enough to warrant the operation of a water treatment plant during the life of the mine. However, additional testwork will be

carried out during the feasibility stage to verify the above assumptions.

C O N C E N T R A T E D R Y I N G

The leached molybdenum concentrate will be filtered, dried, and bagged prior to

dispatch off the mine property. Bulk 1,000 kg bags will be used for the dispatch of the final concentrate product.

T A I L I N G H A N D L I N G A N D D I S P O S A L

The flotation tailing will be discharged to the final tailing pumpbox for delivery to theTailings Management Facility (TMF). Water will be reclaimed from the TMF via apump station with the water being returned to the process water tank for distribution

as required within the process plant. The TMF will be operated to ensure that the

TMF remains flooded to address ARD potential.

R E A G E N T HA N D L I N G A N D S T O R A G E

Various chemical reagents will be added to the grinding and flotation circuit to modifythe mineral particle surfaces and enhance the floatability of the mineral particles for

its collection into a high grade molybdenum concentrate product. These reagentsinclude diesel fuel which will be the collector reagent for molybdenite, Nokes reagent

which will be used as the galena depressant, and DF250 which will be the frother reagent. Flocculant will be used to assist in the concentrate thickener, and lime will

be used for pH control. Hydrochloric acid will be used in the lead leach circuit.





A S S A Y A N D M E T A L L U R G I C A L L A B O R A T O R Y

The assay laboratory will be equipped with the necessary analytical instruments to

provide all the routine assays for the mine samples, the samples from theconcentrator, and the environmental department.

The metallurgical laboratory will undertake all necessary testwork to monitor

metallurgical performance and, more importantly, to improve process flowsheet unitoperations and efficiencies. Standard assay quality control measures will be

employed and analyzed.

W A T E R SU P P L Y

Two separate water supply systems for fresh water and process water will be

provided to support the process operation.

Fresh and potable water will be supplied to a fresh/fire water storage tank from ClaryLake. Reclaimed water will be pumped from the tailing pond to the process water

tank for distribution to the points of usage.

A I R S U P P L Y

Separate air service systems will supply air to the plant areas requiring low pressure

or high pressure air.

1 . 1 2 M I N I N G

Mining will be undertaken using two 18 m3-shovels, one 18 m3 loader, and up tofifteen 177-t haul trucks with related support equipment. Benches will be drilled on

an 8 m by 8 m drill pattern. All blast holes will be sampled and assayed for

molybdenum. The holes will be loaded and shot with a combination of ANFO andemulsion. Benches are 10 m in height and the blast hole drilling will be to a depth of

11.6 m, including sub-drill.

Assay analyses will provide grade control for ore. Haul distances will be shortened

both by the proposed borrow material for starter dam construction and the placementof low-grade material on the top of existing Patsy Dump. Low-grade ore will bestockpiled and processed during the last two years of the operation and higher grade

ore will be sent to the mill. Molybdenum flotation concentrate will be shipped to aroasting facility and subsequently sold to end users.

The mining production schedule is presented in Table 1.5. The final pit is depicted in

Figure 1.5.





Table 1.5 Summarized Production Schedule

Mining

Period

Mine Production

(t)

Low Grade SP

(t)

Waste (t) Total Mined (t) S.R. Mill Produciton

(t)

Grade

%Mo

-1 540,624 8,195,965 8,736,589 15.16 0.067

1 14,600,000 4,185,404 13,413,847 32,199,251 0.71 14,600,000 0.109

2 14,600,000 2,580,297 15,343,253 32,523,550 0.89 14,600,000 0.100

3 14,600,000 1,406,022 16,089,912 32,095,934 1.01 14,600,000 0.103

4 14,600,000 1,235,102 15,550,092 31,385,194 0.98 14,600,000 0.092

5 14,600,000 1,390,588 14,687,240 30,677,829 0.92 14,600,000 0.096

6 14,600,000 1,364,793 14,302,570 30,267,362 0.90 14,600,000 0.086

7 14,600,000 2,629,023 13,261,450 30,490,473 0.77 14,600,000 0.094

8 14,600,000 1,068,413 13,676,918 29,345,331 0.87 14,600,000 0.092

9 14,600,000 1,757,636 9,735,190 26,092,826 0.60 14,600,000 0.085

10 14,600,000 2,174,165 7,805,783 24,579,948 0.47 14,600,000 0.083

11 14,600,000 1,205,514 7,628,557 23,434,071 0.48 14,600,000 0.088

12 14,600,000 2,533,348 8,749,146 25,882,494 0.51 14,600,000 0.086

13 14,600,000 927,156 1,796,306 17,323,462 0.12 14,600,000 0.089

14 463,953 1,915 1,695,122 2,160,990 3.64 14,600,000 0.033

15 10,863,953 0.031

Total 190,263,953 25,000,000 161,931,351 377,195,304 0.75 215,263,953 0.085

Figure 1.5 Final Pit





1 . 1 3 P I T H Y D R O G E O L O G Y

The objective of the groundwater assessment was to characterize the hydrogeology

of the proposed expansion of the Kitsault pit, to provide input to the geotechnicaldesign to estimate inflows, and assess probable inflow geochemistry to the pit. Alladditional hydrogeological issues on site are addressed in other sections of the

report.

The 2009 hydrogeological program was based on data collected from two drillholesdrilled and instrumented in 2008, and two additional drillholes drilled in 2009 in the

northeastern and southern parts of the proposed pit. Continuous packer testing was

carried out in the 2009 holes to provide a profile of hydraulic conductivity for the rockmass in the respective section of the proposed pit wall. Results from the packer

testing indicated that the rock mass tested has low bulk low transmissivity, as isexpected in the intrusive geological setting of the site.

Currently, significant structural features, either intersecting or adjacent to the plannedslopes, that could produce high inflows or residual destabilizing pressures duringexcavation, have not been identified in Avanti’s pit area geological model. This is

supported by the 2009 drilling and testing program. However, this cannot be ruled

out at this time, and will be a focus of further work if additional structuralinterpretation does identify any major features.

Flowing artesian pressures were observed in two drillholes. At this time it is not

known if these pressures would pose a significant risk to pit slope stability, and sowill be assessed in the ongoing feasibility level slope design assessment. If thepressures are deemed to be a risk, a program to delineate and verify ability to

depressurise these parts of the slope will be implemented in the hydrogeology

feasibility program. Depressurization is expected to be controlled with a standardarray of horizontal drainholes drilled into the pit slopes.

Available hydrogeological data were used to develop a simple analytical inflow model

for the Kitsault pit development, with the objective to estimate the potential range of groundwater inflow rates to the pit at various stages of excavation. Results suggest

that inflows will be generally low, and are coincident with the low permeability of thepit lithologies. It is anticipated that excess groundwater inflows to the pit will becollected in sumps and pumped via the site water management system into the

Kitsault TMF.

Water quality of potential inflows to the pit was assessed during two sample roundson drillholes that were screened to depths to coincide with proposed pit walls. Water

quality results indicated that concentrations of dissolved parameters are typical for

such a deposit. Molybdenum concentrations were found to be consistent with thepresence of molybdenum mineralization and are elevated with respect to typical

groundwater concentrations in non-mineralized areas.





1 . 1 4 G E O T E C H N I C A L

A geotechnical pit slope evaluation was undertaken for the Kitsault open pit. The

primary objectives of the evaluation were to evaluate rock mass conditions in thearea of the anticipated open pit and to, consequentially recommend pit slope designcriteria, including interramp slope and bench configuration, to be used for further

mine planning and pit design.

The current ultimate pit follows the recommendations of the pit slope evaluation. Thepit utilizes an interramp slope angle of 51º with double 10-m (20 m-high) benches,

with average sustainable bench widths of 9 m and 70º mean bench face angles.

Potential effects of pore water pressures on global pit slope stability, as well as pitdewatering requirements, should be evaluated further.

1 . 1 5 T A I L I N G , W A S T E R O C K , A N D W A T E R M A N A G E M E N T

The TMF has been designed for secure and permanent subaqueous storage of all

tailing from the proposed mining operations in an impoundment created by a water

retaining embankment constructed with a combination of local borrow materials andwaste rock from the mining operation. The TMF includes a 350 m high water retaining embankment, a tunnel water diversion system, a hydroelectric generating

facility, tailing delivery and distribution pipe works, a seepage monitoring system and

a reclaim system to recycle water to the plant site.

The TMF has been designed to meet all current Canadian Dam Association (CDA)

Dam Safety Guidelines. The Dam Safety Guidelines assign each structure to a

“Dam Class” based on the incremental losses resulting from failure of the dam withrespect to loss of life, environmental and cultural values, and infrastructure and

economic losses. The Dam Class defines the required Inflow Design Flood (IDF)and Earthquake Design Ground Motion (EDGM) for the design of the dam structureand water management systems. The TMF embankment has been classified as

"Very High" under the CDA Dam Safety Guidelines. However, in order to provide

some additional conservatism to the design of the major structures, on account of theoverall height and location upstream of the town of Kitsault, the design parameters

for a Dam Class of "Extreme" have been selected. The corresponding IDF is theProbable Maximum Flood (PMF) and the EDGM is a 10,000 year return period

earthquake event.

The TMF embankment is designed as a water retaining earth/rockfill embankmentthat will be constructed in stages as the elevation of the stored tailing and ponded

water increases with time. The embankment will be constructed initially using borrow

material consisting of overburden material and quarried rock, and waste rock fromthe mining operations as soon as it is available.





The Stage 1 embankment crest elevation has been selected to provide storage for the first 2 years of operation and ongoing raising of the embankment crest will be

carried over the first 8 years of operations. The annual fill placing requirements havebeen designed to utilize available waste rock from the open pit over the first 8 years

of operation. Following completion of the TMF embankment in Year 8, all waste rock

for the remainder of the mining operations will be stored in the upper reaches of theTMF, downstream of the Lime Creek diversion dam and tunnel inlet. This waste willbe fully contained within the TMF during operations. During closure, 25 Mt of this

potentially acid-generating (PAG) waste rock will be relocated to the open pit belowflood elevation, and the remaining PAG waste rock in the TMF will have a cover

applied as part of the reclamation design.

Overall water management is the key to success of all mining waste managementsystems. The TMF is located in the lower reaches of Lime Creek and hence has the

ability to collect all runoff from areas impacted by the mining operation that can then

be recycled for reuse or discharged to lower Lime Creek. The project is located

within an area where a significant overall surplus of water is unavoidable and hencerunoff from upstream areas not impacted by the mining operation will be diverted to

the maximum practicable extent.

Diversion of the major portion of the upstream catchment around the TMF will beachieved using a diversion tunnel starting at the confluence of Lime Creek and Patsy

Creek and located on the left (south) bank of Lime Creek and terminating at a point

downstream of the TMF embankment where the diverted flows will discharge into asurface spillway.

Diversion of runoff to the maximum practicable extent has been achieved by avoiding

impact to Patsy Creek and making provision for the diversion of Patsy Creek around

the open pit along a bench on the wall of the pit.

Tailing will be discharged from the mill as an unthickened slurry via gravity. The

design objective is to discharge the tailing slurry from the embankment face as far as

practicable in order to provide tailings beaches for additional seepage control and tofacilitate ongoing embankment expansion by the centerline construction method.The tailing distribution system has been designed for ease of operations and to

minimize the number of pipeline moves during operations. There will be one tailing

line conveying the full tailing stream. During the final years of operations, the tailinggenerated from the processing of the low-grade stockpile will be discharged to the

open pit, rather than the TMF.

Water for mill processing will be recovered from the TMF supernatant pond usingsubmersible, inclined, multi-stage well pumps in the early years of operation and a

floating reclaim pump-station during the later years. The water is pumped via asingle reclaim pipeline to a head tank at the mill for reuse in the process. A booster pump-station is included in the system due to a significant elevation difference

between the TMF and the mill. The elevation of the TMF pond initially increasesrapidly. This reduces the required pump head throughout the life of the mine. The





pump-stations will be modified as appropriate to accommodate these significantchanges in pond elevation and pumping head.

The water recovery system will also be used to extract clean surplus water from the

TMF. This excess water will be discharged into the Lime Creek diversion tunnel.

Due to the considerable elevation difference between the TMF and the upstream endof the Lime Creek Diversion Tunnel, a parallel booster pump-station is required topump water to this location. The elevation of the TMF pond increases with time

which reduces the required pump head through the life of the mine. Pump-stations

will be modified as appropriate to accommodate significant changes in head.

A single fresh water pipeline connects Clary Lake to a freshwater tank at the mill to

provide clean water for process use, fire water and potable water. A separate

system is designed to supply potable and fire water to the camp, which is located atthe Roundy Creek and Alice Arm confluence.

1 . 1 6 L I M E C R E E K H Y D R O E L E C T R I C P O W E R F A C I L I T Y

Significant water diversion and control infrastructure will be required as part of themine’s redevelopment, presenting a unique opportunity to incorporate a hydroelectricpower plant (HEP). The hydropower development may augment the power supply to

the mine or supply power to BC Hydro grid. An HEP producing 9.8 MW of

hydropower (between 30 and 35 GWh/a) is based on sale of the clean renewableelectricity to BC Hydro.

The HEP will be constructed in Year 2 of mine operations; it will include an intake

structure at the end of the diversion tunnel, and a steel penstock to the powerhouse

situated at the toe of the tailing embankment. The HEP will operate using the LimeCreek Diversion Tunnel until Year 15, after which it will receive flow directly from theTMF. After closure, the annual inflows will be regulated in the TMF to provide a more

constant flow at the powerhouse.

1 . 1 7 I N F R A S T R U C T U R E

The plant and mine facility layouts are located to take advantage of theuncompromising natural terrain and to the extent possible, minimize the impact on

the environment.

Ore from the pit will be trucked up a haul road to the primary crusher. The primarycrusher is located northwest of the ultimate pit. A conveyor material handling system

and services will be constructed to deliver the mill feed for processing from theprimary crusher. The conveyor system will extend from the west side of the primarycrusher and run approximately parallel to haulage and access roads. Pipelines for

tailing disposal to the TMF and for recycling reclaimed water will also be constructed.





Molybdenum concentrates (averaging approximately 70 t/d) produced at the processsite will be transported approximately 15 km by contract trucking firms on new and

upgraded access roads to a port facility south of the existing Kitsault town.Concentrates will be loaded and shipped via ocean transport to North American

and/or overseas roasters.

The Kitsault Project infrastructure will consist of the following: camp/accommodationcomplex, access road from the new port facility to the mine site, a water diversion

channel through the pit, processing plant buildings, maintenance and administration

facilities, TMF, diversion ditches, water diversion tunnel, pipelines for water andtailing, hydro electric scheme, power distribution system and associated substationsand a revised transmission line alignment.

1.17.1 P O W E R SU P P L Y A N D D I S T R I B U T I O N

Power to the mine site will be via a 138 kV overhead line from the Aiyansh

Substation located approximately 42 km away. A new step-down substation locatednear the plant site will reduce the voltage to 13.8 kV which will be used as the plant’s

main distribution voltage level. In addition, the project must also meet the need toprovide a new 138/25 kV substation in order to maintain/provide power to the existing

Kitsault town site at its present 25 kV level and provide the same voltage to the newport and camp facilities.

The mine’s incoming substation includes the utility metering equipment, primary

disconnecting means, two main transformers, and secondary 15 kV switchgear line-up which provide the means to selectively distribute power to the various plant areas.

The power distribution method to load areas will be either by overhead power line for

the more distant loads or by a 15 kV cable system for the loads which are relativelyclose by.

The plant will also include electrical rooms for distribution voltage step-down, Motor Control Centres (MCCs), and emergency generators.

1 . 1 8 A C C E S S R O A D S

There are several road access routes to the mine and plant site, including an access

road from the new camp and port facility to the mine site.

The site can be accessed from Smithers by travelling 110 km northwest onHighway 16 to the Cassiar Highway Junction, and then north 76 km on Highway 37

towards the Cranberry Junction. The westbound Nass Forest Service Road (FSR)

runs 32 km to the Kinsqyish FSR. This road winds 14 km past the Nass Bridge, the

Kinkush-Alice Arm Pass Road junction, several concrete bridges, and thedecommissioned transformer station next to the mine access road.





The site can also be accessed starting at Terrace and travelling 100 km north toNass Camp and then a further 95 km on an upgraded gravel road to site. An

alternative access road route is from Terrace via Highway 37 to Cranberry Junction(175 km) then via a gravel connector road from Cranberry Junction to site

(approximately 106 km).

The re-routed access road from the existing town site of Kitsault heading easttowards the Clary Lake existing road is approximately 7 km in length. The re-routing

of the existing Kitsault town site access road is necessary as the TMF will inundate a

major section of the existing road within the first years of operation.

1 . 1 9 L O G I S T I C S A N D P O R T F A C I L I T Y

Access to the Kitsault property can be achieved by water or air (float plane), and by

vehicle from Terrace or Smithers.

The proposed location for the port and accommodation complex/town site facilities isapproximately 1.6 km south of the existing Kitsault town site (from an existing bridge

crossing) and is in a gently sloping area. Earthmoving in this area should berelatively easy and no problems are anticipated. The existing bridge on this route willrequire upgrading.

The port facility allows for the barging of equipment and supplies to the mine site, aswell as ocean transport of concentrate from the mine site. The port is expected tohandle concentrate throughput of up to two hundred 20 ft containers every two

weeks.

1 . 2 0 E N V I R O N M E N T A L

1.20.1 O V E R V I E W

The Kitsault Mine is a permitted brownfield site with considerable past mining activityand basic infrastructure in place. The Kitsault Project meets the criteria thatconstitute a non-reviewable project under Part 3 of the Reviewable Project

Regulation of the BC Environmental Assessment Act (EAA). Re-opening the minecan therefore be considered a modification of an existing project reviewable under the Mines Act . Avanti may request the BC Environmental Assessment Office (BC

EAO) to “opt-in” the environmental review of the KM project under the EAA in order to harmonize federal and provincial regulatory reviews as well as the requirements of

the Nisga’a Final Agreement (NFA).

Avanti is currently in discussions with the BC Ministry of Energy, Mines andPetroleum Resources (MEMPR), BC EAO, and Nisga’a Lisims Government (NLG) to

confirm the environmental and permitting review process under the Mines Act and

the Northwest Mine Development Review Committee or the EAA. In support of these





discussions, Avanti has obtained confirmation from the BC EAO that Section 51 of the BC EAA applies to the KM project, confirming that the EAA does not technically

apply to the project, as the current Mines Act Permit M-10 predates the EAA.However, Avanti recognizes that the provisions of the NFA must be complied with in

any environmental review and permitting of the Project. The permitting process will

be agreed upon by all federal and provincial agencies as well as the NLG.

As part of Avanti’s efforts to undertake the necessary studies in preparation for

environmental assessment and permitting under the Mines Act or the EAA, the

Company has developed a provisional draft Project Terms of Reference andenvironmental work plan which have guided the biophysical and socioeconomicbaseline studies in 2008 and 2009. Significant work was undertaken in 2009 in

relation to the study.

Studies have established a clear basis for continuing to develop an Application for aMines Act Permit or an EAA review for the Project. While the full environmental

impact assessment has not been completed, the issues identified are typical for alarge open-pit mine in Northwest BC and principally involve water management,

metal leaching/acid rock drainage, and wildlife and fisheries effects mitigation. The

delivery of field studies by Avanti environmental consultants has involved Nisga’aCommercial Group members throughout all phases of field study.

Several environmental studies remain to be completed. Further discussions with

provincial and federal agencies in relation to the environmental review process andreview of studies to date could result in requests to refine studies in conjunction withfinalizing the Project Terms of Reference.

The Kitsault Project environmental assessment consultation program would be

guided by the Northwest Regional Mine Development Review Committee’sconsultation requirements which will be consistent with the guidelines derived from

the EAA, the Act’s Public Consultation Policy Regulation (2002); the Provincial Policy

for Consultation with First Nations (2002) and the Supplementary Guide to

Proponents: BC Environmental Assessment Process. Consultations with NLG,communities and public, will be required during the pre-application and applicationreview phases.

Consultation with Nisga’a Nation is a requirement of the NFA, one of Canada’s firstmodern-day treaties. Nisga’a interests will be reflected in agreements to addressand/or accommodate Nisga’a Nation issues, values, and concerns. The provisions

of the NFA, and specifically as defined in Chapter 2 – General Provisions, and

Chapter 10, Section 8 – Environmental Assessment and Protection, along with theProvincial Policy for Consultation with First Nations (2002), will be followed to fulfill

this requirement.

The next steps in the environmental, permitting and consultation programmes areunderway and involve further engagement of regulatory authorities in the defined

review process, approval of the Project Terms of Reference, completing any





outstanding environmental baseline work, and engaging Nisga’a Nation and publiccommunities in project reviews.

Avanti would develop the Application for the Project approvals with the goal of

submitting the Application for review and approval in Q1 2010 under the Mines Act

review process, and Q2 2010 under the EAA review process, which harmonizesfederal, provincial, and NFA requirements. Under this schedule, Mines Act approvalcould be forthcoming in late 2010, providing the authorization to proceed with

construction of the Kitsault mine concurrent with acquiring other key agency

operational permits.

1.20.2 AC I D R O C K DR A I N A G E P O T E N T I A L A N D M E T A L L E A C H I N G

Initial characterization of the potential for metal leaching and acid rock drainage

(ML/ARD) has occurred as a component of Avanti’s exploration drilling program todefine the resource for the expansion of the existing pit. Every interval was analysed

for sulphur concentrations as part of a multi-element ICP scan producing 3,326results. Every fourth interval was also analysed for carbonate content.

Sulphur concentration determined by the ICP method was used as a surrogate for total sulphur concentration determined by specific methods. Typical sulphur

concentrations were between 1 and 2%. Geological information and subsequent

analysis of a subset of samples showed that calcium sulphate minerals occur insome parts of the proposed open pit resource. This indicates that total sulphur

concentrations will tend to over-estimate acid potential (AP) due to sulphide mineraloxidation where sulphate minerals are present. As a result, Avanti used a regression

equation to estimate sulphide content from sulphur content. The current assessment

is that correction of total sulphur for sulphate content does not significantly affect the

overall prediction of acid rock drainage (ARD) potential for the waste rock but this isbeing investigated further as the ML/ARD characterization of the project proceeds.

Carbonate content was used as a surrogate for acid neutralization potential (NP). Insituations where the main carbonate minerals are composed of calcium and

magnesium carbonates, the surrogate provides a reliable indication of NP becausethese minerals are effective acid consumers. If iron and other heavy metalcarbonates are present, carbonate content over-estimates NP because these

minerals provide limited neutralization. Ongoing mineralogical studies arequantifying carbonate mineralogy to evaluate the suitability of carbonate as asurrogate for NP.

For the purpose of the PFS, the following was assumed:

95% of rock was nominally classified as PAG (NP/AP <2) because thedetailed data to develop a site specific criterion are being obtained and,

based on experience, BC MEMPR will likely consider rock with NP/AP below

this value as PAG for the purpose of assessment and permitting.





Limited previous testing for the site indicated that a site specific criterion

near 1 may be appropriate. For the purpose of defining rock with relatively

lower risk of ML/ARD potential, a placeholder NP/AP criterion of 1.1representing the 50th percentile of the distribution was used for wastemanagement planning.

Management of ARD potential should be considered in future project

designs.

The existing waste rock dumps at the site, some of which originate from mining in the

1960s, are currently not showing ARD though acid generating rock is observed on

surface, and drainage contains elevated sulphate concentrations. The lack of ARDmay reflect a number of factors including composition of the rock and delay of acidicconditions due to slow depletion of carbonate minerals. This evidence implies that

ARD is not likely to start in the mine life and that waste rock and pit water management plans can allow for a lag time of decades when considering

management of PAG rock and ARD.

Metal leaching refers to the release of regulated potential contaminants at levels thatmay not be acceptable for direct discharge to the environment regardless of drainage

pH. Depending on the site specific application of water quality guidelines, a number

of elements may be a concern in non-acidic drainage. Parameters potentially of concern are molybdenum, cadmium, zinc, and sulphate.

1.20.3 T A I L I N G

Tailing samples have been produced from metallurgical testing of three rock-typebased ore composites prepared by Avanti. The three composites produced similar

tailing, implying that rock type was not an important variable. Total sulphur concentrations (dominantly as sulphide) in whole tailing varied from 1.2 to 1.8%.NP/AP varied from 0.7 to 0.9 leading to the conclusion that the tailing have potential

for ARD and need to be appropriately managed. The current strategy would requirea permanently f looded impoundment.

Analysis of metals in the tailing and testwork supernatants showed that leaching of

molybdenum, zinc, cadmium, sulphate, and fluoride may need to be considered for disposal of the tailing.

1 . 2 1 R E C L A M A T I O N A N D C L O S U R E

1.21.1 T A I L I N G A N D W A TE R M A N AG E ME N T F AC I LI TI E S

The Kitsault Project will be developed, operated, and closed with the objective of leaving the property in a condition that will mitigate potential environmental impacts

and restore the land to an agreed to land use and capability. Closure andreclamation activities will be carried out concurrent with mine operation wherever





possible, and final closure and reclamation measures will be implemented at the timeof mine closure.

For the purposes of reclamation planning, the Kitsault Project has been broken down

into the following key reclamation units:

TMF, including the embankment

waste rock dump and low-grade ore stockpile

open pit

mine site facilities (process building, truck shop, conveyors, crusher, andcrushed ore stockpiles)

infrastructure (camp and wharf facility in Alice Arm)

access and haul roads

surface water diversion structures.

Once active mining from the pit has ceased, the low-grade stockpiles will beprocessed for the remainder of Year 14 and a portion of Year 15. The full tailing

stream will be discharged to the pit. Reclaim water will continue to be sourced fromthe TMF, as the settling of tailing solids in the pit will likely not be sufficient to sourcereclaim water from the pit supernatant. During this time period, an estimated 25 Mt

of PAG waste rock will be removed from the south end of the TMF and hauled backto the pit and placed below the ultimate flood elevation (below 515 masl).

The Patsy Creek diversion will be decommissioned in Year 16 and will be redirected

into pit so as to speed up the pit filling. If it is found during decommissioning that the

water from the Patsy Creek diversion is required to dilute the surplus flow from thetailings pond to meet receiving water targets in Lime Creek, an altered strategy to fill

the pit may be required, including allowing the pit to fill naturally while the PatsyCreek diversion remains in place until the pit spills. Upon cessation of tailingsdeposition sometime in Year 15, the Lime Creek Diversion Tunnel will be

decommissioned by sealing off the east portal. A new 300-m long tunnel connecting

the TMF to the end of the existing tunnel will allow surplus flow from the TMF to passinto the hydroelectric scheme, generating power in perpetuity.

Tailings will remain submerged in perpetuity. The flood spillway elevation is set at

470 masl and will be designed to accommodate the PMF event.

An engineered cover and growth medium will be applied to the downstream face of the TMF embankment and the remaining PAG waste rock in the south end of the

TMF. The cover will then be revegetated.

The seepage collection pond at the toe of the TMF will remain active post closure,with the seepage volume being fed into the discharge from the flood spillway. The

seepage quality will continue to be monitored after closure. As previously discussed,

ARD is not likely to start during the life of the mine and it is expected that there will





be a lag time of decades before management of PAG rock and ARD will need to beconsidered. The seepage volume differential compared to the annual surplus from

the TMF is expected to be sufficient to negate the need for immediate water treatment of the seepage water from this source after closure. However, an

allowance to construct and operate a future water treatment plant has been included

in the post closure cost.

Buildings and structures that comprise the mine site facilities (mill, camp, power

station, administration building, maintenance shop, laboratory, site roads, fuel

storage, and explosives storage) will be removed at closure. These facilities will bedemolished. Salvageable items within the building will be removed from site andsold. Hazardous wastes will be removed from site and disposed of in an approved

facility.

All pumps, pipes and other infrastructure related to the tailing, reclaim, and surpluswater systems will be dismantled and removed from site.

Some of the roads on site will be maintained post-closure for ongoing sitemaintenance and monitoring. This will include the main road to the Kitsault town site,the roads accessing the area near the penstock and dam face, and the roadway on

the dam face itself. Roads that can be reclaimed will have the culverts removed,stream crossings re-graded, and their surfaces scarified to encourage vegetation

growth.

1 . 2 2 P R O J E C T E X E C U T I O N P L A N

Upon Mines Act approval and authorization to proceed in Q4 2010, the facility will

take approximately 35 months to construct and be completed in Q3 2013.Construction will be concurrent with acquiring the key agency permits. A high levelproject schedule is shown in Figure 1.6.





Figure 1.6 Project Schedule





1 . 2 3 C A P I T A L A N D O P E R A T I N G C O S T E S T I M A T E S

1.23.1 C A P I T A L C O S T E S T I M A T E

An initial capital of US$641 M (Q4 2009) will be required for the project, based on

capital cost estimates developed by the following consultants:

Wardrop: mine capital costs, mine rock dump costs, process plant and

associated infrastructure costs, power supply costs, and access road costs

KP: tailing management, haul road, construction access roads for TMF and

tunnel diversion, fresh water supply, and water management costs

SRK (Canada): mine closure and reclamation costs

Rescan: associated environmental and permitting costs

WPW: marine port facilities costs.

All currencies in this section are expressed in US dollars, unless otherwise stated.Costs in this report have been converted using a fixed currency exchange rate of

Cdn$1.00 to US$0.92. The expected accuracy range of the capital cost estimate is

±30%.

“Initial capital” has been designated as all capital expenditures required to produce

concentrate. Some capital costs have been relocated to operations and sustaining

capital as directed by Avanti.

A summary of the major capital costs is shown in Table 1.6.





Table 1.6 Capital Cost Summary

Description US$000

Direct Works

Overall Site 37,300

Mining 51,000

Crushing 24,000

Crushed Ore Storage and Reclaim 12,900

Process 106,200

Tailing Managem ent Facility 122,200

Water Management 33,200

Site Services and Site Utilities (Mine Site) 4,500

Ancillary Buildings (Mine Site) 16,700

Plant Mobile Fleet 2,400

Temporary Services (Port Site) 13,300

Port Site Facilities 8,700

Subtotal 432,500

Indirects

Project Indirects 104,000

Owner's Costs 26,000

Contingencies 78,000

Subtotal 208,100

Total Capital Cost 641,000

1.23.2 O P E R A T I N G C OS T E S T I M A T E

The costs in this section are stated in Q4 2009 US dollars, unless otherwise stated.

When it was required, certain costs in this report have been converted using a fixedcurrency exchange rate of Cdn$0.92 to US$1.00 from Avanti. The expectedaccuracy range of the operating cost estimate is ±25%.

The operating costs are defined as the direct operating costs including mining,processing, tailing storage, and general and administrative (G&A).

Table 1.7 Unit Cash Operation Costs (Life of Mine Average)

Description US$

(000s)US$/tMined

US$/tMilled

US$/RecoveredPounds

Mining 657,107 1.63 3.05 1.79

Processing 768,406 1.91 3.58 2.09

G&A 118,825 0.30 0.55 0.32

Power Generation 3,421 0.01 0.02 0.01

Tailings Management 61,085 0.15 0.28 0.17

Totals 1,608,844 4.00 7.47 4.38





Sustaining capital includes all capital expenditures after the process plant has beenput into production.

P R O C E S S O P E R A T I N G CO S T S

The process operating cost for the Kitsault Project was estimated at US$3.58/t milledand includes the cost of power. The estimate was based on an average annual plant

processing rate of 14,600,000 t/a.

Table 1.8 Process Operating Cost Summary

Description Personnel

Annual

Cost(US$)

Unit Cost

(US$/tTreated)

Human Power

Operations Staff 12 1,438,342 0.10

Operations Labour 44 3,079,819 0.21

Maintenance Labour 30 2,672,848 0.18

Met Lab & Quality Control 12 813,659 0.06

Sickness and Vacation Contingency Operations 7 548,864 0.04

Sickness and Vacation Contingency Maintenance 3 267,285 0.02

Sub-total Staff 108 8,820,816 0.60

Supplies (Process Plant)

Operations Supplies 32,039,136 2.19

Maintenance Supplies 795,800 0.05

Power Supply 10,650,225 0.73

Sub-total Supplies (Process Plant) 43,485,161 2.98

Total Process Operating Costs 108 52,305,977 3.58

G&A Staff 27 2,238,209 0.15

G&A Expenses 5,819,000 0.40

G&A Contingency 0 0.00

Total G&A Costs 27 8,057,209 0.55

M I N I N G O P E R A T I N G C O S T S

The operating cost for the Kitsault Project was estimated at US$1.63/t mined or

US$3.05/t milled, which includes the mining contractor costs as well as the cost of

power from the grid line.





Table 1.9 Mining Operating Costs

Description

Annual

AveragePersonnel

Average Annual

Cost(US$)

Unit Cost

(US$/tTreated)

Human Power Operations Staff 21 2,137,692 0.15

Operations Labour 128 9,663,843 0.66

Maintenance Labour 67 5,377,356 0.37

Sub-total Staff 216 17,178,891 1.18

Mine Supplies


Maintenance Supplies 4,605,385 0.32

Power Supply 599,083 0.04

Fuel 11,171,358 0.77

Load/Haul Contract 1,665,631 0.11

Sub-total Mine Supplies/Misc 26,326,034 1.80G&A

G&A Staff 10 1,060,758 0.07


Total G&A 10 1,060,758 0.07

TOTAL MINE OPERATING COST 226 44,565,683 3.05

T A I L I N G O P E R A T I N G C O S T S

The tailing operating costs were estimated at US$0.29/t milled. The estimate was

based on an average annual plant processing rate of 14,600,000 t/a.

Table 1.10 TMF Annual Operating Costs per Tonne Ore

Description

Average

AnnualCost (US$)

Unit

Cost(US$/t)

Human Power 331,521 0.02

Materials 1,511,094 0.10

Equipment 104,631 0.01

Power 1,904,338 0.13

Contingency 384,253 0.03Total Tailing Disposal Costs 4,235,838 0.29





1 . 2 4 F I N A N C I A L A N A L Y S I S

Metal revenues projected in the Avanti Kitsault mine cash flow models were based

on the average metal values indicated in Table 1.11.

Table 1.11 Metal Production from Kitsault Project

Life of Mine

Total Tonnes to Mill (Mt) 215.3

Annual Tonnes to Mill (Mt) 14.6

Average Grades

Molybdenum (Mo%) 0.085

Total Production

Molybdenum (Mlb) 367.91

Average Annual Production

Molybdenum (Mlb) 24.5

A financial model was created utilizing the mine production schedule, the associated

metal grades based on the geological resource estimate, metal recoveries, and

capital and operating costs as set out herein and base case metal prices (based onthe CPM Molybdenum Industry Analysis). The project economics were alsocompleted for the three additional pricing scenarios (Table 1.12) and used for

comparison purposes (Table 1.13).

Table 1.12 Molybdenum Pricing Cases (US$)

Period

CPM Market

Study(Base Case)

CPM Market

Study(Low Case)

CPM Market

Study (HighCase)

Wardrop/

EMCFPrices*

Average Mo Market Price $15.76 $14.07 $16.90 $11.99

Year 1 (Q4 2013) $14.38 $12.88 $16.88

Year 2 $13.56 $12.75 $15.25

Year 3 $14.31 $13.38 $15.13

Year 4 $14.88 $13.88 $15.56

Year 5 $16.13 $14.38 $17.06

Year 6 to End of Mine Life** $16.50 $14.50 $17.50

* Wardrop’s new policy utilizes the Energy & Metals Consensus Forecasts (EMCF) quarterlyreports (The Consensus Economics Inc.) in calculating the Wardrop/EMCF prices. This new

approach is to avoid large fluctuations in metal prices from study to study and to use the long termprice averaged from three EMCF quarterly reports. For this study, if executed between August 31,

2009 and January 1, 2010, the long term metal prices would be derived by averaging the long termprices for previous January, April, and July quarterly reports to drive the Wardrop/EMCF prices.

** The average long term price of Mo computed using CPM projections until 2018 and then

maintained flat for the rest of the mine life.





Modelling at base case metal prices shows that the project could generate a post-taxnet present value (NPV) of US$551 M (discounted at 8%), a 20.6% internal rate of

return (IRR), and a 3.8 year payback period.

Table 1.13 summarizes the key results of the financial model for the evaluated

scenarios. The notes accompanying Table 1.13 are an integral part of the post-taxfinancial analysis.

Table 1.13 Summary of the Post-Tax Financial Analysis

Units

CPM Market Study Wardrop/

EMCFPricesBase Case Low Case High Case

Molybdenum US$/lb 15.76 14.07 16.90 11.99

Exchange Rate US:Cdn 0.92 0.92 0.92 0.92

NPV (at 8%) US$M 551 372 696 155

IRR % 20.6 16.9 24.2 12.1

Cash Cost/accountable lb Mo US$/lb 4.43 4.43 4.43 4.43

Payback Period years 3.8 4.3 3.3 5.2

The financial analysis includes revenue estimated by KP from the run-of-river hydro-

electric project. Wardrop has used KP’s revenue estimate in this financial analysis

but has not independently verified it.

1.24.1 P O S T - T A X AN A L Y S I S

PricewaterhouseCoopers LLP (PWC) performed a due diligence review on a tax

model prepared by Wardrop. PWC have confirmed that the tax model accuratelypredicts the after-tax cash flow.

The post-tax base case financial model used the same inputs as the pre-taxeconomic evaluation with the following key taxation rates:

Federal corporate tax at 15%

Provincial corporate tax at 10%

Provincial net proceeds tax at 2%

Net provincial revenue tax at 13%.

1.24.2 SE N S I T I V I T Y AN A L Y S I S

Table 1.14 displays the sensitivity to pre-tax NPV, analyzed by variations of metalgrades and prices, capital and operating costs, and exchange rate.





Table 1.14 Base Case Sensitivity to Pre-tax NPV (US$000s) at 8% Discount

Rate

-30% -20% -10%BaseCase 10% 20% 30%

Moly Price 108 378 648 919 1,189 1,460 1,730Exchange Rate 2,003 1,551 1,200 919 689 497 335

Moly Head Grade 163 415 667 919 1,171 1,423 1,675

Operating Cost 1,154 1,076 997 919 840 762 684

Capital Cost 1,101 1,040 979 919 858 798 737

Table 1.15 displays the sensitivity to the post-tax NPV, analyzed by variations of metal prices and discount rates.

Table 1.15 Sensitivity to Post-Tax NPV (US$000s)

Mo Price(US$)

Discount Rate (NPV) (US$ M)

0% 5% 8% 10%

8.00 -204 -346 -392 -412

12.00 772 324 156 71

16.00 1,620 893 618 477

20.00 2,476 1,460 1,073 874

24.00 3,335 2,027 1,526 1,269

S U P P L Y A N D D E M A N D B A L A N C E

Increased demand for molybdenum, notably between 2003 and 2007, ushered in asubstantial restructuring of the molybdenum industry. To meet demand, the pace of new supplies also has had to accelerate. As miners sought to develop deposits that

were once uneconomic, operating costs for both primary producers and by-productproducers rose. All molybdenum producers are now subject to operating costs (pro-

rata) that are greater than the long-term average nominal price.

Market analysis reveals that demand for molybdenum may remain fairly priceinelastic and the growth in demand justifies the rising costs due to the following

developments:

Increases in mine supply are in the pipeline but geological, political,environmental, and financial constraints are slowing down this progress.

There are few viable substitutes for molybdenum in its major applications;thus future demand, for the vast majority of end uses, is not likely to be

swayed by strong prices.





A period of tight supply and demand conditions is forecast to unfold in 2011 and2012, pushing real molybdenum prices up to an average of $23.00/lb, over this two

year period. The ramp-up of larger molybdenum projects is estimated to move themarket into a surplus from 2013 through 2017. However, the running balance of

inventories suggests that stock levels will remain fairly low. As such, real

molybdenum prices may decline to average $14.95/lb over this 5-year period. Whilegrowth in secondary supplies from recycled catalysts may help mitigate the pressurefrom market fundamentals, fresh supplies from either by-product or primary

producers is necessary to fill the gap in supply in 2018, the tail-end of CPM’s 10-year projections. In 2018, real molybdenum prices are forecast to average $16.50/lb

(Table 1.13).

1 . 2 5 R E C O M M E N D A T I O N S

The following sections outline areas to investigate for project improvements. To

complete each recommendation, a high-level budgetary estimate is provided inCanadian dollars.

1.25.1 G EOLOGY

It is recommended to advance the project to feasibility-level studies. The resource

model as it stands is sufficient but the secondary metals, if recoverable, maycontribute to project revenues. Therefore, it is recommended to perform an audit on

the current model for lead and silver grades to establish potential by-product

recovery ($10,000).

1.25.2 M I N I N G

Recommendations related to mining are as follows:

Investigation of a higher production rate (greater than the current base caseof 40,000 t/d). Considering the available mineral resources, there is

potential to increase production rate and potentially improve the overall

project economics

Further study of the Patsy Creek Diversion, in the south pit wall, isnecessary in order to confirm geotechnical parameters and pit slope

stability. The study should look at minimizing additional waste stripping

caused by the water diversion. Alternative primary crusher location should be investigated. There may be

potential to move the crusher to the northwest corner of the pit that will

reduce initial capital cost for construction of the access to the crusher; andmine operating and capital costs.

Further study of the transportation system(s) for moving waste from the pit to

the tailing dam for use in embankment construction.





Detailed hydro-geology evaluation of the area is needed in order to improve

the accuracy of pit dewatering design. Vertical and horizontal dewatering

wells have not been included as part of the required PFS activities. It maybe necessary to lower the water table within the pit prior to mining and toprevent water inflow from the nearby creeks and surface. It is

recommended that a water management plan for the location of the pit water discharge be included in future studies.

Considering hydro-geology evaluation, a hydrology assessment is requiredso that the diversion and water management plan will be developed for the

mining area, for both surface and groundwater quantities.

Foundation testing of the waste dump and low grade stock pile is required.

Detailed drill and blasting studies are recommended for more future studies.These will help to determine the penetration rate expected for the selected

drills and the specific rock types in the pit and the most applicable powder factor for the rock types that are present in the Kitsault pit. It is

recommended that a detailed blasting study will include a cost benefitassessment between owner run blasting versus full contractor blastingresponsibility. The most effective system will result in better blasting

efficiency and lower costs.

It is recommended that the optimization of the shovel fleet be completedduring the feasibility phase.

The use of contractor to move the material volumes in the schedule be

reassessed in the next study in line with shovel optimization.

It is recommended that a study to address the optimization of the haulagefleet be carried out during the feasibility phase. There is the potential to use

a larger-sized truck fleet for the current and possibly higher production ratescenarios.

1.25.3 T A I L I N G A N D W A TE R M A N AG E ME N T

Recommendations for the TMF and site water management systems include thefollowing:

additional and complimentary geotechnical site investigations in the footprint

of the TMF main embankment (inclusive of the overflow spillways required

for each stage of embankment)

geotechnical site investigations at the intake structures and alignment for theLime Creek Diversion Tunnel, at the intake structures of the Patsy Creek

Diversion, at access roads and at the construction borrow areas

develop updated sediment control plans for site development and borrow

area excavations





trade-off study for main haul road to TMF to consider incorporating

engineered retained wall along steep sidehill sections to reduce rock

excavation costs

detailed laboratory testing program on samples of the tailing and waste rock

hydrological and water quality analyses of Clary Lake and Roundy Creek

continued hydrological assessment of the Lime Creek and Patsy Creekdrainages

trade-off study for placing the balance of the waste in the TMF versus intothe pit at mine closure.

The following is a list of recommendations for future studies:

All input data used in the PFS represents best estimates/assumptionsavailable at the time of the study and need to be confirmed.

The exact locations and elevations of various system components need tobe confirmed.

The size distribution and rheology for bulk tailings need to be assessed.

Water quality in the TMF surface water pond and suitability for releasewithout treatment needs to be assessed and confirmed.

Seepage discharge that combines seepage through the embankment and

runoff from the surrounding terrain need to be confirmed.

Clary Lake water quality and intake location need to be assessed.

Roundy Creek water quality and flow rate need to be assessed.

As part of the feasibility design, a dam breach and failure runout analysis willneed to be conducted, as per CDA requirements.

The costs for the tailing and water management recommendations fall into two

categories (site investigations and engineering), as follows:

Site investigation and laboratory testing (including drilling contractors,helicopter time, laboratory costs, etc.) are estimated at $600,000.

All engineering work related to elevating the TMF and water management to

a feasibility-level is estimated at $600,000.

1.25.4 P R O C E S S

The 2009 SGS test program must be augmented with additional flotation testwork to

confirm that the target concentrate grade of 52% Mo, and the overall molybdenum

recovery value of 90.6% can be achieved ($150,000).





The removal of lead from the molybdenum concentrate in order to meet marketspecifications and to ensure that there will be no smelter penalties for above-

specification lead content, will require additional testwork and mineralogicalevaluation to determine the nature of the contaminating lead particles ($100,000).

Continued evaluation of bi-product revenue possibilities from the removal of lead andsilver from the ores should be conducted ($100,000).

1.25.5 I N F R A S T R U C T U R E

Wardrop has made the following recommendations for infrastructure:

Geotechnical site investigations should be conducted at the primary crusher,

truck shop, plant site, camp and port to confirm the foundation conditions for these major site structures.

Marine foreshore investigations should be completed to determine optimal

camp and docking facility location ($170,000).

Further investigation of port access to ensure that a handymax vessel couldsafely navigate into Alice Arm. The results of such a study would have to be

confirmed by the BC Coast Pilots. We recommend that Avanti negotiate acontract with a shipping company for delivery of supplies and material andpossible export of concentrate ($20,000).

Proceed with additional engineering work to provide information required to

request initiation of a system impact study by BC Hydro. This study willproduce utility related costs and requirements ($125,000).

Consideration and active design of feasible energy conservation measures

(process design, equipment selection, energy recovery schemes) inconjunction with BC Hydro Power Smart initiatives to take advantage of energy cost reductions. Prudent planning with sufficient lead time and

involvement by BC Hydro experts can create opportunities to reduce therelative magnitude of Tier II energy rates with respect to the overall project($75,000).

Investigate the construction of a roasting facility in Prince Rupert, and the

possibility of transporting the molybdenum concentrate to this facility($150,000).

A risk analysis would need to be carried out at the front end of a future

Feasibility Study ($75,000).





2 . 0 I N T R O D U C T I O N

This NI 43-101 PFS is intended to be used by Avanti to present a current view of the

project’s likely economic outcome. This PFS has been prepared in general

accordance with the guidelines provided in NI 43-101 “Standards of Disclosure for

Mineral Projects”. The intent of this PFS is to provide the reader with a

comprehensive review of the economics of the mining operations and related project

activities as well as to provide recommendations for future work programs. Metric

system (SI) units of measure are used in this report, unless otherwise stated.

This NI 43-101-compliant report has been prepared for Avanti based on work

performed by the following independent consultants:

Wardrop

KP

WPW

SRK

RDi

Rescan.

A summary of the QP responsible for each section of this report is provided in Table

2.1. Certificates of QPs are provided in Appendix A.

Table 2.1 Summary of Qualified Persons

Section Description

Responsibility

Company QP

1.0 Executive Summary Wardrop Frank Grills

2.0 Introduction Wardrop Frank Grills

3.0 Reliance on Other Experts Wardrop Frank Grills

4.0 Property Description and Location Rescan Rolf Schmitt

5.0 Accessibility, Climate, Local Resources,

Infrastructure, and Physiography

Wardrop Frank Grills

6.0 History SRK Jeff Volk

7.0 Geological Setting SRK Jeff Volk

8.0 Deposit Types SRK Jeff Volk

9.0 Mineralization SRK Jeff Volk

10.0 Exploration SRK Jeff Volk

11.0 Drilling SRK Jeff Volk

table continues…





Section Description

Responsibility

Company QP

12.0 Sampling Method and Approach SRK Jeff Volk

13.0 Sample Preparation, Analyses, and Security SRK Jeff Volk

14.0 Data Verification, Quality Control, and Quality Assurance

SRK Jeff Volk

15.0 Adjacent Properties SRK Jeff Volk

16.0 Mineral Resource and Mineral Reserve Estimates SRK/

Wardrop

Jeff Volk/

Miloje Vicentijevic

17.0 Metallurgical Testing Wardrop/

RDi

Andre de Ruijter/

Deepak Malhotra

18.0 Mineral Processing Wardrop Andre de Ruijter

19.0 Other Relevant Data and Information

19.1 Mining Wardrop Miloje Vicentijevic

19.2 Geotechnical – Pit Slopes SRK Mike Levy

19.3 Pit Hydrogeology SRK Michael Royle

19.4 Site Water Management KP Ken Brouwer

19.5 Tailing Management KP Ken Brouwer

19.6 Waste Rock and Tailing Geochemical

characterization

SRK Stephen Day

19.7 Reclamation and Closure SRK Peter Healey

19.8 On-Site Infrastructure Wardrop Frank Grills

19.9 Off-Site Infrastructure Wardrop Frank Grills

19.10 Power Supply and Distribution Wardrop Frank Grills

19.11 Main Access Roads Wardrop Frank Grills

19.12 Logistics Wardrop Frank Grills

19.13 Capital Cost Estimate Wardrop/

KP

Frank Grills/

Miloje Vicentijevic/

Ken Brouwer

19.14 Operating Cost Estimate Wardrop/

KP

Andre de Ruijter/

Miloje Vicentijevic/

Ken Brouwer

19.15 Financial Analysis Wardrop Miloje Vicentijevic

20.0 Project Execution Plan Wardrop Frank Grills

21.0 Environmental Considerations Rescan Rolf Schmitt

22.0 Marketing and Contracts CPM n/a

23.0 Risk and Mitigation All All

24.0 Opportunities and Recommendations All All

25.0 Interpretations and Conclusions All Frank Grills26.0 References n/a n/a





3 . 0 R E L I A N C E O N O T H E R E X P E R T S

Ms. Catherine Virga and Mr. Doug Horn of CPM have been relied upon for the

section on markets and contracts.

Mr. Peter Acton (P.Eng.) of WPW visited the property on July 20 to 21, 2009 and has

been relied upon for matters and costs relating to the marine facilities.

Mr. Tim Johnston of PWC has been relied upon for matters relating to the review of

the tax component of the financial model prepared by Wardrop. The scope of the

review was limited to the following tax components of the model:

Canadian federal corporate income taxes

BC provincial income taxes

BC Mineral Tax.

Mr. Olen Aasen of Blake, Cassels, and Graydon LLP has been relied on for matters

relating to land tenure, title opinion, and related legal matters.

Mr. A.J. Ali of Avanti has been relied on for matters relating to ownership, royalties,

and agreements.

As outlined in Section 2.0, the Kitsault Molybdenum Project PFS report was

completed by a range of independent consulting companies. All sections of thereport have been provided by experts who are QPs.

Wardrop followed standard professional procedures in the preparation this report.

Data used in this report have been verified where possible; Wardrop has no reason

to believe that the data were not collected in a professional manner.

The certificates of the QPs are provided in Appendix A.

The following was taken directly from SRK’s “NI 43-101 Technical Report on

Resources Avanti Mining Inc. Kitsault Molybdenum Property British Columbia,

Canada” (revised signing date – March 3, 2009). This section remains largely

unchanged and is included strictly for completeness and continuity.

The discussion of land tenure in Section 4.0 of this report relies on the information

contained in the title opinion and the information provided by Blake, Cassels &

Graydon LLP (2008). While the authors have carefully reviewed the information

provided with respect to land tenure and the ownership by ALI, and are of the opinion

that the information is be reliable, the authors have not conducted an in-depth

independent review of the land tenure and ownership of the Kitsault property. The





ownership of the Kitsault claims is reported on the British Columbia Energy and

Mines, Mineral Titles Online BC (MTO) website as being Avanti Kitsault Mine Inc., a

wholly owned subsidiary of Avanti. The authors are relying on disclosures by Avanti

on the recent change in ownership and financial agreements, and have not reviewed

any of the pertinent legal documents.

On September 16, 2008, Avanti announced that it had been denied access to the

ghost town of Kitsault by its owner, Kitsault Resort Ltd. (KRL). At the time, Avanti

held a license and permit from the owner (ALI) of a statutory right of way to use the

roads within the town of Kitsault for mine related activities. As a result, Avanti, in its

capacity as a licensee and permittee, commenced an action in British Columbia

Supreme Court seeking: (1) a declaration that it is entitled to use the roads in the

town and (2) a permanent injunction preventing KRL from denying Avanti use of the

right of way. Avanti is now the registered owner of the right of way.

KRL has commenced a counterclaim seeking a declaration that the right of way is

invalid. If the litigation is not resolved in the Avanti’s favor then it will have to utilize

other means of access to Roundy Creek, which are less convenient and will result in

additional costs. The authors were informed by Avanti that there is no other current

or pending litigation that may be material to the property.

The authors are unaware of any outstanding environmental or other liabilities on the

property except that outlined in Section 4.0, and are relying on the opinion of Avanti

to this effect.

Kitsault is a historic mining and exploration area with associated land disturbance.





4 . 0 P R O P E R T Y D E S C R I P T I O N A N D

L O C A T I O N

Sections 4.1 to 4.5 provide information on the land and mineral tenure matters

relating to land tenure, which rely on the title opinion and information provided by

Blake, Cassels, and Graydon LLP (2008). Additional information is provided by

Avanti on ownership, royalties, agreements, and encumbrances.

Historical information on Permit M-10 in Sections 4.6, 4.7, and 4.7.1 are provided by

Avanti and SRK who managed the Kitsault Mine reclamation program under Permit

M-10. This information was previously provided in SRK’s “NI 43-101 Technical

Report on Resources Avanti Mining Inc. Kitsault Molybdenum Property British

Columbia, Canada” (revised signing date – March 3, 2009).

Rescan is responsible for the information and opinions provided in Sections 4.7.2

and 4.8. The remainder of this section was taken from SRK’s “NI 43-101 Technical

Report on Resources Avanti Mining Inc. Kitsault Molybdenum Property British

Columbia, Canada” (revised signing date – March 3, 2009) and remains largely

unchanged and is included strictly for completeness and continuity.

4 . 1 P R O P E R T Y L O C A T I O N

The Kitsault property is located about 140 km north of Prince Rupert, BC, and southof the head of Alice Arm, an inlet of the Pacific Ocean (Figure 4.1). The propertyincludes three known molybdenum deposits – Kitsault, Bell Moly, and Roundy Creek.

The Kitsault mine site is located within NTS maps 103P044 and 043, at

approximately latitude 55°25'19″N and longitude 129°25'10″, and at about 600 melevation. The 8,286 ha of mineral leases and mining claims are fully described in

the following sections. The principal mining feature on the property is the Kitsault

open pit mine (currently not operating), which is shown in Figure 4.2, in relation to themining claims in the area.

The molybdenum deposit at Bell Moly is located at approximately latitude 55°28'N

and longitude 129°20'W at about 750 m elevation. The Roundy Creek molybdenumdeposit is located at about latitude 55°24'49″N and longitude 129°29'32″W atapproximately 320 m elevation (Figure 4.2).





Figure 4.1 Kitsault Location and Regional Geology





Figure 4.2 Claim Map and Location of Molybdenum Deposits on Avanti Claims





4 . 2 U N D E R L Y I N G A G R E E M E N T

Avanti had previously signed a Definitive Purchase and Sales Agreement with

Aluminerie Lauralco, Inc. (ALI), to acquire an undivided, 100% direct interest in the

Kitsault molybdenum mine and surrounding mineral tenures, located in northern BC,subject to a 120 day confirmatory due diligence period and regulatory approval (June

20, 2008 Press Release). The agreement structure requires Avanti to pay US$20 Mto ALI for a 100% interest in the Kitsault property. ALI has a 1% Net Smelter Royaltyon future production subject to the right, within 90 days from the presentation of a

Bankable Feasibility Study, to elect to surrender the Net Smelter Royalty in exchange

for either an additional US$10 M payment (payable at Commercial Production or in Avanti shares at their election). On October 20, 2008, Avanti announced that they

had completed the purchase (on October 17, 2008) of the Kitsault property.

A finder’s fee for 2,000,000 fully-paid and not-assessable Avanti common shares

was payable to a third party on the closing of the Kitsault acquisition. A success feeof 500,000 Avanti common shares was paid to a Financial Advisor. Both of thesefees were paid to the respective parties at the closing of the Kitsault acquisition on

October 17, 2008.

4 . 3 M I N E R A L T I T L E S A N D S U R F A C E R I G H T S

The land status of the mine area is shown in Appendix B of the SRK report publishedon SEDAR (as referenced at the beginning of this section), consisting of a total of 35

mining leases and 196 mineral claims. The majority of leases and claims are locatedunder Provincial Crown lands with four claims located under privately owned lands at

the Kitsault town site along Alice Arm. Surface rights are granted under the mineralleases from the Crown. Mineral claims surface rights are obtained through the

process of converting claims to leases.

4.3.1 SU R F A C E L AN D S

The project also consists of the following surface lands overlapping a number of theleases and claims.

three parcels of land:

Parcel Identifier 015-583-031, District Lot 2656 Cassiar District (2.34 ha)

Parcel Identifier 015-562-531, Block A District Lot 35 Cassiar District(12.1 ha)

Parcel Identifier 015-562-611, Block B District Lot 35 Cassiar District

(0.057 ha).





one Statutory Right of Way for access through Kitsault town site privately

owned lands – Statutory Right of Way No. BX201679 (the ROW) over thefollowing lands:

District Lot 2757 Cassiar District

Block B (Plan 9849) District Lot 63 Cassiar District

Block A District Lot 63 Cassiar District

Block B District Lot 63 Cassiar District

Block A District Lot 64 Cassiar District Except Plan 6531



Lot 1 District Lot 64 Cassiar District Plan 6531.

4 . 4 L O C A T I O N O F M I N E R A L I Z A T I O N

The Kitsault, Bell Moly, and Roundy Creek molybdenum deposits are located on theKitsault land package as shown in Figure 4.2. The existing access roads, power

lines, and waste dumps are also shown in Figure 4.2.

4 . 5 R O Y A L T I E S , A G R E E M E N T S , A N D E N C U M B R A N C E S

The various leases and claims are subject to the encumbrances outlined in this

section.

4.5.1 R O Y A L T I E S

A royalty payable by CMC to Bell Molybdenum Mines Limited (Bell), on each ton of

ore mined and removed from claims 250340, 250341, 250342, 250343, 250344,250345, 250346, 250347, 250390, and 250391(covering the Bell Moly prospectonly). The royalty is as follows:

US$0.10/t of ore mined from the above-listed claims and treated thereon

or shipped to Climax's molybdenum concentration mill at Kitsault, BC or

another place of treatment, provided that for each US$0.25 that the base

price per pound, as defined therein, of molybdenum contained in

concentrate increases above or decreases below US$2.50, the royalty

shall in like manner increase or decrease by US$0.01.

In addition, the entire property is subject to a 9.22% net cash flow interest which Amax Zinc (Newfoundland) Limited (AZN) owns. Net cash flow is only payable after the recovery of all capital costs associated with construction or sustaining capital.

ALI has a 1% Net Smelter Royalty on future production subject to the right, within 90days from the presentation of a Bankable Feasibility Study, to elect to surrender the





Net Smelter Royalty in exchange for either an additional US$10million payment(payable at Commercial Production or in Avanti shares within 30 days after the date

of the election).

4.5.2 E N C U M B R A N C E S F R O M O THER AG R E E M E N T S, L I E N S, A N D D OCUMENTS

In a “Wrap Up Agreement” dated September 29, 2005 between CMC and ALI and

Alumax Inc., ALI covenanted and agreed with CMC that it will not dispose of any of the leases, claims, or lands required for or associated with the completion of the

Work Program, as defined therein, until the completion of such Program.

There are two Claims of Liens in the amounts of Cdn$820,000 and Cdn$40,917.15against the following 11 Leases: 254543, 254544, 254545, 254546, 254547, 254548,

254563, 254569, 254571, 254573, and 254576.

There is a Trust Indenture between Bell and The Canada Trust Company dated July

1, 1980. The Trust Indenture is a charge over the rights of Bell to receive the aboveroyalty payable by CMC to Bell.

4 . 6 E N V I R O N M E N T A L L I A B I L I T I E S

Tailings from the Kitsault Mine were deposited into Alice Arm. The Alice Arm

Tailings Deposit Regulations, made under the federal Fisheries Act , explicitly allowed

for the deposit of mill process effluent from the Kitsault Mine into the waters of Alice Arm, BC. There is no known risk of environmental statutory liability under the

Fisheries Act associated with the tailings deposit into Alice Arm.

Kitsault Mine reclamation activities have been undertaken, pursuant to an AmendedPermit Approving Reclamation Program (Permit M-10). Reclamation commenced in

1996 and was completed in 2006. During this time all buildings and ancillary

infrastructure were decommissioned and removed from the site. The waste rockdumps and low grade mineable resource stockpiles have been re-sloped and re-vegetated. Measures to address acidic drainage in the pit have been implemented in

accordance with the approved reclamation program. A spillway suitable to maintain

the long term integrity of the settling pond at the bottom of the pit has beenconstructed. The remainder of the pit has been reseeded. While the site has been

remediated, there are a number of ongoing environmental commitments associatedwith the site:

water quality monitoring

re-vegetation monitoring

maintenance fertilization and reseeding in the pit

submitting an annual reclamation report to the MEMPR

annual inspections of the diversion ditches and spillway.





The annual cost associated with these activities items is about Cdn$260,000 TheKitsault mine closure and reclamation is regulated by the Kitsault Mine Reclamation

Workplan (SRK, 1997) and is administered by SRK. SRK conducts yearlyassessments of the property to monitor reclamation progress. During the 2007

survey, SRK concluded that re-vegetation of the property was complete and water

quality is within limits for protection of freshwater aquatic life (SRK, 2008).

At Bell Moly the only evidence of past exploration activity are faint traces of drill pads

and access trails. At Roundy Creek, there are two short adits and small associated

dumps. Neither of these two outlying sites has any known environmental liabilities.

A current environmental audit of the Kitsault property has been completed as part of

the due diligence program associated with the purchase.

4 . 7 P E R M I T T I N G

The Kitsault Mine project will require multiple Provincial and Federal permits,approvals, licenses, and authorizations. This section lists the permits already

obtained from ALI on the closing of the Kitsault property acquisition on October 17,2008, and those likely required for exploration and development activities associatedwith the Kitsault Project.

4.7.1 P E R M I T S OB T A I N E D F R O M T H E V E N D O R O F T H E M I N E

The Amended Permit Approving Reclamation Program Permit M-10, issued by the

BC MEMPR, allows the permittee to carry out reclamation activities in respect of the

mine project.

An application dated February 1997 for an amendment to the conditions of the permit

entitled "Kitsault Mine Reclamation Workplan" for the protection and reclamation of

the surface of the land and watercourses affected by the mine was submitted to theChief Inspector of Mines on March 6,1997 in accordance with Section10(6) of the

Mines Act . This application was subsequently approved and the reclamation work

including building demolition, rock dump resloping, and revegetation was carried out

over the period 1997 to 1999. In 2003, a further application was submitted toMEMPR to amend the M-10 permit for remediation of the pit area and the Orange

Pond. This work was approved on November 1, 2004 and carried out in the summer of 2006.

An application dated March 2008 for an amendment to the conditions of the permit

entitled “Closure Management Manual Update 2-Kitsault Mine” for the protection andreclamation of the surface of the land and watercourses affected by the mine was

submitted to the Chief Inspector of Mines on April 1, 2008 in accordance with Section10(6) of the Mines Act . MEMPR approved this amendment on May 14, 2008 subject

to a number of terms and conditions.





These included:

submittal of a reclamation/closure update by 2011

continued monitoring of site drainage chemistry and vegetation

continued monitoring of all ditches, culverts, and the settling pond

submit an annual reclamation report each year by March 31

develop a contingency plan should the 2006 pit remediation work fails

collect and treat mine discharge if site drainages result in applicable water quality exceedances at W-01.

The following permits were also obtained from the Vendor of the mine:

Road Use Permit for Industrial Use of a Forest Service Road No. 08-7876-

01 issued by the BC Ministry of Forests and Range (MFR).

Special Use Permit 9228 issued by the MFR.

4.7.2 P E R M I T S T O B E O B T A I N E D T O C O M M E N C E W O R K A T T H E M I N E

The following permits will need to be obtained in order to commence work at the

Kitsault Mine:

A Permit to Conduct Exploration and Operational Activities – Permit M-10

set out above is a Permit Approving a Reclamation Program and is not apermit to explore or conduct operational activities in respect of the Kitsault

Mine. A new application must be made to the MEMPR by the submission of

a Notice of Work in any of the following circumstances:

to be able to carry out/continue the reclamation program

to carry out exploration work

to operate the mine through an amendment to the mine permit reflecting

final engineering design.

Air and Effluent Discharge Permits – to be obtained from the BC Ministry of

Environment to operate the Kitsault Mine.

The lead environmental assessment agency for review of the project is stillto be confirmed. Avanti has received confirmation that the project does notautomatically trigger the provincial Environmental Assessment Act review, in

lieu of the grandfathering provision of Section 51 of the Environmental

Assessment Act , which applies to Permit M-10. Accordingly, Avanti willdecide whether to request an opt-in to the BC Environmental Assessment

(EA) review process, or pursue assessment as a major mine under the

Mines Act review process administered by the Northwest Regional MineDevelopment Review Committee. Avanti is considering a number of factors

in making this decision, including how each review process is able to





integrate federal and Nisga’a treaties into the review process. TheCanadian Environmental Assessment Agency will coordinate the review of

the federal agencies in terms of their need to screen the project componentsfor any federal triggers or authorizations that may be required to advance

the project.

Forest Service Road or other forest-road access permit as may be requiredfor efficient and authorized access to the Kitsault Mine.

Foreshore Access from Provincial Crown Integrated Land ManagementBureau at Alice Arm for efficient water access as may be required.

Water License from BC Ministry of Environment to divert fresh water for

mine use such as domestic consumption and reagent mixing.

License to Cut from Ministry of Forests for tree removal.

License of Occupation under the Land Act for occupancy of Crown lands

outside of the mining lease.

Each Federal and Provincial permit will have compliance stipulations that requirescrutiny and negotiation that can typically be resolved within 60 days. However, an

EA process can take more than a year to complete. Project delays could occur

because of public opposition, Nisga’a Lisims Government negotiations, inefficienciesin regulator review, or project changes made by the proponent.

4 . 8 C O N S U L T A T I O N W I T H N I S G A ’ A A N D F I R S T N A T I O N S

The Kitsault mine area falls outside of Nisga’a Lands owned by the Nisga’a Nation,under the terms of the NFA, which became effective on May 11, 2000. However, it is

within the Nass Area and the Nass Wildlife Area (as defined in and governed by theNFA), and as such it is subject to the constitutionally protected rights of the Nisga’a

Nation under the terms of the NFA.

The NFA states that if the proposed mining activities at the Kitsault Mine area mayreasonably be expected to have adverse environmental effects on residents of

Nisga’a Lands, Nisga’a Lands, or NFA interests then a specific process for

consultation by the Federal or Provincial government, as the case may be, is set outin the NFA. In addition, when an EA is carried out under Provincial or Federal law,the NFA grants specific rights to the Nisga’a Nation in respect of any environmental

assessment process. The NFA also enumerates various requirements that are

additional to the requirements under EA legislation.

Other than the Nisga’a Nation, Avanti is not aware of any other First Nations that

may have aboriginal rights, interests, or claims relevant to the proposed mining

activities at the Kitsault Mine area. Final confirmation can be obtained from BCMinistry of Aboriginal Relations and Reconciliations’ office responsible for treaty

negotiations.





5 . 0 A C C E S S I B I L I T Y , C L I M A T E , L O C A L

R E S O U R C E S , I N F R A S T R U C T U R E , A N D

P H Y S I O G R A P H Y

The following section was taken directly from SRK’s “NI 43-101 Technical Report on

Resources Avanti Mining Inc. Kitsault Molybdenum Property British Columbia,

Canada” (revised signing date – March 3, 2009). This section remains largely

unchanged and is included strictly for completeness and continuity, except for

Sections 5.4.1 to 5.4.8 that have been updated based on the latest available

information.

5 . 1 C L I M A T E A N D L E N G T H O F O P E R A T I N G S E A S O N

The climate in the Kitsault area is temperate coastal, verging on a rain forest

environment. Prince Rupert is the closest reliable weather station where

Environment Canada keeps records. The coldest month is January with an average

minimum temperature of -2.1°C; the warmest month is August with an average

maximum temperature of 16.7°C. Average precipitation at Alice Arm and Anyox is

about 2 m. At the Kitsault mine site, a majority of the precipitation is snow. When

the mine was operating, snow did not influence operations. The mine can operate

365 d/a, with no days lost to snow stoppages.

5 . 2 P H Y S I O G R A P H Y

The Kitsault property is in upland, hilly plateau country characterized by thick stands

of timber interspersed with small lakes, meadows, and swamps. The dominant

topographic features are a series of eroded basaltic lava flows that commonly form

cliffs up to 100 m high.

Generally, topography rises quickly from tidewater at Alice Arm to an elevation of

600 to 800 m at the plateau. The Kitsault and Bell Moly sites are on the plateau

(Figure 5.1) and the Roundy Creek site is midway up the elevation change from

tidewater to the plateau.

Bedrock is generally blanketed by a few metres of glacial till and is commonly

overlain by a layer of peat bog up to 1 m thick. Outcrop in this area, except for the

basalt cliffs, averages less than 1%.





Figure 5.1 General View of the Kitsault Mine Area





5 . 3 A C C E S S

Kitsault is accessible from Prince Rupert, BC either by floatplane or boat to the

Kitsault town site and then by vehicle on gravel road to the mine site. The Kitsault to

mine site gravel road will be upgraded for mine traffic. Additionally, CMC constructed

a gravel road from pavement near Nass Camp, BC, to the mine site. This road

roughly parallels the powerline from the Aiyansh substation. The road is

approximately 95 km long. While the alternate road had limited maintenance since

then, it is still seasonally accessible and could be utilized as a sole means of access

with full time maintenance. This road is considered secondary access to the mine

but with upgrade and maintenance could be a primary access. This will require

consultation and agreement with the NLG and the MFR.

Access to the Bell Moly portion of the property is by all-terrain vehicle over primitive

trails. Access to Roundy Creek is by gravel road to within a few hundred metres of

the adit and then by foot.

Limited access from the mine site to the new port facility will initially be by the

existing gravel road. However, plant access will be provided by a new and upgraded

gravel access road from the port facility to the mine site which is included in the this

study.

5 . 4 L O C A L R E S O U R C E A N D I N F R A S T R U C T U R E

The Kitsault Mine includes the open pit, reclaimed mill foundations, and several mine

dumps. During the most recent production period, CMC maintained the Kitsault town

site for its employees. The town site was sold by CMC in 2005 to KrishnanSuthtantrinan. The current state of the town is unknown but it could be a possible

housing alternative during renewed mining although this is probably unlikely. This

study has assumed the construction of a separate staff housing facility near Kitsault

or Alice Arm. Mr. Suthtantrinan has denied Avanti access to its Statutory Rights of

Way described in Section 4.3 and Avanti has been forced to take this matter to the

BC Supreme Court for adjudication. Further details about this litigation are disclosed

in Section 1.1.

The town of Alice Arm is on the opposite side of the inlet, where there are limited

facilities available to support mining or exploration. All other supplies are obtained in

Prince Rupert and are transported to the site by water or air, and during summer byvehicle from Terrace or Smithers.

5.4.1 AC C E S S R O A D A N D T R A N S P O R T A T I O N

Primary access to the mine is by water. Barges, boats, and floatplanes can access

the new port facility from which a gravel road extends to the mine site via the existing

Kitsault town site. This 5 km-long gravel road will be upgraded to allow mine traffic





from the port, staff housing facility to the mine, and mill. However, access to the

town site is the subject of the litigation with KRL disclosed in Section 1.1. The road

upgrade costs are imbedded within the mine capital and operating costs.

For the purpose of this PFS, concentrates will be moved by truck to the new port

facility south of Kitsault, BC. A 60 ha probable site, owned by the Crown andavailable for lease. The wharf facilities and the staff housing facility will be located in

this area.

5.4.2 P O W E R SU P P L Y

Power to the mine site will be via a 138 kV overhead line from the Aiyansh

Substation located approximately 42 km away. A new step-down substation located

near the plant site will reduce the voltage to 13.8 kV which will be used as the plant’s

main distribution voltage level. In addition, the project must also meet the need to

provide a new 138/25 kV substation in order to maintain provide power to the Kitsault

town site at its present 25 kV level.

5.4.3 P O R T

An assessment has been prepared for a new port facility for the project. The facility

is located on a site currently owned by the Crown and available for lease and is

approximately 7 km by existing road from the mine site to tidewater. The port facility

will comprise an access trestle, barge loading ramp with counterweighted ramp lifting

towers, stop dolphins, berthing dolphins, and mooring dolphins. The proposed port

facility is located approximately 1.6 km southwest of the town of Kitsault on Alice

Arm.

The design barge considered for the PFS is the Seaspan Series 200, which will visit

the site every two weeks to carry the expected throughput of up to two hundred 20 ft

containers. Given the relatively small number of containers to be exported from the

site to support mine operations, two reach stackers be used to load full and unload

empty containers between the storage yard and the barge. The largest components

to be transported across the loading ramp at the port facilities are expected to be

delivered during construction and start-up of the mine operations, including empty

mobile mining equipment and mining process facility components.

The proposed mine port will be located approximately 140 km from Prince Rupert,

and will take on average 16 hours travel time by barge. Tides rise and fall 4.57 m

(15 ft). The channel depth exceeds 91 m (300 ft).

Fuel will be transported within a fuel barge with reagents carried by liquids tanker

trucks either on top of the fuel barge or the barge carrying containers. Both fuel and

reagents are carried in separate compartments of the same fuel barge.

Refer to Section 19.0 and Appendix B of this report for more details on the port

facility.





5.4.4 BU I L D I N G S A N D AN C I L L A R Y F A C I LI TI E S

There are no buildings or other facilities on the current property. Mill and ancillary

buildings will be constructed as part of the mine infrastructure development.

5.4.5 P E R M A N E N T AC C O M M O D A T I O N F AC I LI T Y

A permanent camp facility will be based on upgrading a section of the construction

camp to provide for 400 personnel at the peak of construction. The facility will be

constructed adjacent to the port facilities. The proposed complex will include a

dormitory style camp with a centralized cafeteria, recreation center, and

administration building.

5.4.6 T A I L I N G S M A N A GE ME N T F A C I LI T Y

There is no TMF currently on the property. During the CMC production period,

tailings were deposited in Alice Arm, which is not allowed under current regulations. A new TMF site will be located in the Lime Creek valley are being investigated, as

are alternative methods of tailings disposal. Section 19.0 of this report provides

further detail.

5.4.7 W A S T E D I S P O S A L AR E A

There are two reclaimed waste dumps on the property that were used during the KEL

and CMC production periods. Access to the plant site will be on the eastern edge of

the Clary dump but the Patsy dump sites will remain relatively undisturbed.

5.4.8 H U M A N R ESOURCES

This part of British Columbia has a mining history and it is likely that sufficient human

resources will be available for any renewed production for the site. The mine

operations, at full production will employ approximately 330 people. Part of the

workforce will come from surrounding communities. The remaining workforce will

commute weekly or bi-monthly from outside the immediate area and live in the

permanent accommodation complex.





6 . 0 H I S T O R Y

Historical mineral resource and reserve estimates presented in this report are based

on information collected and compiled from the 1950s to the early 1980s. Historic

resource estimates do not comply with the CIM terminology under NI 43-101guidelines. The reader is cautioned that these estimates are not mineral resources

or mineral reserves and should not be relied upon.

Attention was first drawn to the molybdenum potential of the Alice Arm’s regionduring World War I when a small quantity of molybdenite was produced from the

Tidewater property (Woodcock and Carter, 1976).

Exposures of silver bearing polymetallic veins southeast of the Kitsault molybdenumdeposit were first staked in 1911 and represent the first known exploration activity in

the mine area (Woodcock and Carter, 1976). Turnbull (1916) first reportedmolybdenite in exposures along Lime Creek, although it was not considered

economically significant at the time.

KEL initially examined claims at Kitsault (originally known as the Lime Creekmolybdenum deposit) in 1956 and optioned the property the following year. The first

recorded drilling activity on the property commenced in 1959 when KEL undertook

exploration for stockworks-style molybdenum mineralization (Hodgson, 1995). In1964, KEL announced a resource for the deposit of 36 Mt averaging 0.15% Mo(0.23% MoS2) (Hodgson, 1995). Open pit mining started in January 1968 and was

suspended in August 1972 due to low molybdenum prices. During that period, about9.3Mt of ore were produced with about 22.9 Mlb of molybdenum recovered(Hodgson, 1995).

In May 1973, CMC purchased the property from KEL, conducted additional drilling

and defined a deposit containing 142 Mt with an average grade of 0.125% Mo(0.182% MoS2) (Steininger, 1980). The property was transferred to ACL in 1979 who

returned the deposit to production in April 1981. Operations were suspended inNovember 1982, again because of low metal prices. During this second productionperiod, about 4.08 Mt of ore and stockpile material were processed and 8.99 Mlb of

saleable molybdenum were recovered (Amax 10K, 1982 and 1983).

Molybdenum was first recognized at what become known as Bell Moly in 1965 whenMastodon Highland Bell Mines, Ltd. and Leitch Gold Mines collected geochemical

samples and eventually staked the property (Steininger and Card, 1979). To fundcontinual exploration to follow up on the initial encouraging drilling results, the twocompanies were re-organized as Bell Molybdenum Mines, Ltd. (Bell). Through 1975,

a total of 36 core holes were drilled for an aggregate of 5,462.6 m. Bell reported a





resource of 23.6 Mt with an average grade of 0.070% Mo (0.117 MoS 2) (Carter,1967).

CMC leased the property in 1975 and conducted drilling programs in 1976 and 1977.

They drilled 17 holes totalling 5,519 m and developed a new resource containing

96.4 Mt with an average grade of 0.054% Mo (0.09% MoS2) (Steininger and Card,1979).

Molybdenite was first recognized at Roundy Creek in the early 1900s. The principalexploration work was conducted by Silurian Chieftain Minerals Company Limited

between 1965 and 1971. That company completed about 9,300 m of diamond coredrilling and 780 m of underground development (Woodcock and Carter, 1976). The

drilling and underground sampling was used to estimate a resource of 7 Mt with an

average grade of 0.066% Mo (0.11% MoS2) (Woodcock and Carter, 1976). CMCpurchased the property in 1975.

At some point in the years between closure of the Kitsault Mine in 1982 and 1993,

the property, including the three molybdenum deposits, was transferred to Amax’s Alumax aluminum division. In late 1993, Amax merged with Cyprus Minerals and aspart of the merger the Alumax division was spun off to the Amax shareholders,

including the Kitsault molybdenum deposits (Alumax 1996 10-K). Alcoa purchased Alumax in 1998 and placed the Kitsault land holdings with its subsidiary, ALI.

CMC and ALI jointly managed maintenance of the town site and reclamation under a

joint management agreement. In 2005, the town site was sold to a private party.

CMC and ALI agreed to a “final reclamation program” which was completed in 2006. After this program, the mining property was transferred 100% to ALI.

6 . 1 H I S T O R I C A L R E S O U R C E E S T I M A T E S

The reported historical “resource” and “reserve” inventories cannot be considered amineral resource or a mineral reserve under CIM guidelines as economic parameters

used to derive the estimates do not reflect accurately the current economics of exploiting these deposits. Furthermore, procedures and data used have not been

reviewed and verified by a Qualified Person and therefore cannot be classified as a

mineral resource under Canadian Securities Administrators NI 43-101 guidelines. Inall cases, insufficient documentation exists that would allow SRK to classify historicreserve and resource estimates into the categories as currently defined by CIM

guidelines. These historic estimates should be considered unclassified mineralized

material.

The authors could not find documentation for the KEL resource estimation

procedures, except for the tonnage and grade which was published in several papers

on the deposit. The most-widely published historic resource is 36 Mt averaging0.15% Mo (0.23% MoS2) (Hodgson, 1995).





CMC produced detailed resource and reserve estimates and a feasibility study for the Kitsault molybdenum deposit (Table 6.1).

Table 6.1 CMC Historic Resource & Reserve Estimate for the Kitsault Deposit*

CategoryCut-off

Grade (Mo%) MtGrade(Mo%)

ContainedMo (Mlb)

1979 Resource** 0.06 129 0.11 312.2

1983 Reserve** 0.06 104.3 0.12 275.4

* Climax, 1978 and Amax, 1985.

** Historic resource estimates do not comply with the CIM terminology under Canadian Securities Administrators NI 43-101 guidelines. The reader is cautioned that these estimates are not mineralresources or mineral reserves and should not be relied upon.

A cross sectional resource was developed for the Clary Creek portion of the BellMoly deposit at the conclusion of the CMC exploration program (Table 6.2)

(Steininger and Card, 1979).

Table 6.2 CMC Historic Resource Estimate for Bell Moly Clary Creek Deposit*

Cut-off Grade

(% MoS2) Mt

Grade

(Mo%)

Contained

Mo (Mlb)

0.05 96.4 0.054 115.1

0.10 13.5 0.084 24.9

* Historic resource estimates do not comply with the CIM terminology under Canadian Securities

Administrators NI 43-101 guidelines. The reader is cautioned that these estimates are not mineralresources or mineral reserves and should not be relied upon.

Woodcock and Carter (1976) reported a historic resource for the three Roundy Creek

deposits, as shown in Table 6.3. This resource estimate shows a combined resourceof 8.38 Mt at 0.099% Mo for a total molybdenum resource of about 16.7 Mlbs.

Table 6.3 Woodcock and Carter Historic Resource Estimate for Roundy Creek

Deposit, 1976*

Zone

Cut-off Grade

(% Mo) Mt

Grade

(% Mo)

Contained

Mo (Mlb)

Roundy Creek unknown 7.0 0.066 10.17

Sunshine unknown 1.35 0.208 6.19

Sunlight unknown 0.035 0.400 0.31

Total 8.38 0.099 16.67


Administrators NI 43-101 guidelines. The reader is cautioned that these estimates are not mineralresources or mineral reserves and should not be relied upon.

The last known resource estimate made for the Roundy Creek deposits was done by

J.W. Mustard of Amax in 1983, after a detailed in-house reconciliation of all historicaldata. The results of this estimate are summarized in Table 6.4. This resource





estimate shows a combined total of 4.24 Mt grading 0.129% Mo contained in threeseparate zones above a 0.06% Mo cut-off, for a total molybdenum resource of

approximately 12.0 Mlb.

Table 6.4 J.W Mustard Roundy Creek Resource Estimate; 1983*

Zone

Cut-off Grade

(% Mo) Mt

Grade

(% Mo)

Contained

Mo (Mlb)

Roundy Creek 0.06 2.74 0.098 5.97

Sunshine 0.06 1.45 0.179 5.73

Sunlight 0.06 0.050 0.320 0.353

Total 0.06 4.24 0.129 12.06


Administrations NI 43-101 guidelines. The reader is cautioned that these estimates are not mineralresources or mineral reserves and should not be relied upon.

6 . 2 H I S T O R I C F E A S I B I L I T Y A N D E C O N O M I C R E V I E W S

In the CMC Feasibility Study (CMC, 1978), a careful comparison was made betweenactual production and the Feasibility Study resource estimates for the Kitsault Mineduring the period between 1968 and 1972. Table 6.5 shows how the block model,

based on KEL blast hole data, compares with the CMC exploration resource model.The model comparison was conducted over a vertical extent of 325 ft, representing

10 mined benches.

Table 6.5 CMC Resource Model vs. Blast Hole Model Comparison*

Cut-off Grade

MoS2

(%)

Blast Hole Model Exploration Model

% Difference

Exploration – Blast Hole

Ore

(k ton)

MoS2

(%)

MoS2

(k ton)

Ore

(k ton)

MoS2

(%)

MoS2

(k ton)

Ore

(k ton)

MoS2

(k ton)

0.01 21,167 0.167 35,349 20,326 0.168 34,148 -3.97 -3.40

0.13 13,790 0.212 29,235 1 5,960 0.188 30,005 +15.74 +2.57

0.15 12,053 0 .222 2 6,758 13,099 0 .198 2 5,936 +8.68 -3.07

0.16 11,143 0 .228 2 5,406 11,416 0 .204 2 3,289 +2.46 -9.09

0.20 7,474 0.252 18,784 5,349 0.233 12,463 -28.24 -33.65

* taken from Kitsault Feasibility Study, Volume 2 (CMC, 1978).

The model estimate compares reasonably well with the blast hole results. Theincrease in tonnage and decrease in grade in the resource model compared to the

blast holes is consistent with smoothing or averaging as the result of more widely

spaced exploration drillhole samples. At higher cut-off grades, the exploration modeltends to under predict both tonnage and grade.

For the four years of full operation from 1968 to 1972, mill recovery averaged over

89% and the milling cost/ton averaged US$0.96.





The CMC feasibility study concluded that using capital costs of approximatelyCdn$150 M in 1978, a 7.3% return on investment (ROI) could be realized at a

molybdenum price of US$4.41/lb assuming the Cdn$ and US$ to be on par. Sincethe Canadian to US$ conversion rate was approximately Cdn$0.91 to US$1.00 at the

time, the expected ROI was 9.4% with capital payback occurring in 7.6 years. It was

on this basis that CMC put the Kitsault mine back into operation in 1981.

Production for the period May 1981 to December 1982 was 9.7 Mlb of Mo from

4.08 Mt of ore from stockpiles and new production (Amax 1981, 1982, 10K) of which

only 8.99 Mlb were reported as saleable due to high lead in the concentrateproduced in 1981. The problem was addressed by the installation of a leach circuitto remove the lead in late 1981.





7 . 0 G E O L O G I C A L S E T T I N G

7 . 1 R E G I O N A L G E O L O G Y

The Kitsault, Roundy Creek, and Bell Moly molybdenum deposits are located within

the western margin of the Bowser basin in the Intermountain tectonic belt, a few

kilometres east of the Coast Range Crystalline Complex (Figure 7.1).

The area is characterized by intense intrusive activity related to the Coast Range

Crystalline Complex, with younger stocks intruding the sedimentary lithologies and

numerous recent plateau-type lava flows. Intrusive rocks in the Coast Range

Crystalline Complex range in composition from granodiorite to quartz monzonite.Numerous 50 to 55 Ma granodiorite to quartz monzonite stocks occur in the

sedimentary rocks surrounding Alice Arm which Carter (1981) referred to as the Alice

Arm intrusives. Hosts for these stocks are the Lower to Middle Jurassic Hazelton

Formation and Upper Jurassic to Lower Cretaceous Bowser Lake Group. The

Hazelton Formation consists of volcanic breccias, tuff, conglomerate, volcaniclastic

sedimentary rocks, and andesite flows, all metamorphosed to greenschist facies.

The Bowser Lake Group consists of interbedded greywacke and argillite with minor

conglomerate and limestone metamorphosed to greenschist facies.

Many of the Alice Arm intrusives are the loci for molybdenum mineralization (Figure

7.1). In addition to the Kitsault molybdenum deposit, there is intrusive activity and

associated molybdenum mineralization at Roundy Creek, Bell Moly, Tidewater, and

Ajax. All of these resource areas except for Ajax are located in the immediate area

of Alice Arm and all have been the sites of historic exploration and drilling from the

1960s to the 1980s with the exception of Kitsault, which had additional drilling

conducted in 2008, and Ajax, which had additional drilling conducted in 2007. The

Kitsault deposit is situated about 5 km from ocean access at Alice Arm and lies

immediately to the northeast of the junction of Lime and Patsy Creeks (Figure 7.1

and Figure 7.2).

Following the emplacement of the Alice Arm intrusives and the related molybdenum

mineralization, a swarm of northeast striking lamprophyre dikes, which are dated at

36 ± 1.2 Ma and 34.4 ± 1.2 Ma (Carter, 1974), was intruded into the Bowser LakeGroup. The youngest igneous event in the region is basaltic plateau-type flows and

related dikes dated at 1.6 ± 0.8 Ma and 0.62 ± 0.6 Ma (Carter, 1974). The entire

area was exposed to glaciation, much of which, but not all, has occurred post-

basaltic flows.





Figure 7.1 Geology of the Lime Creek (Kitsault Area)





Figure 7.2 Geology of the Lime Creek Property





7 . 2 G E N E R A L G E O L O G Y

Host lithologies for mineralization at Kitsault, Bell Moly, and Roundy Creek are

thermally metamorphosed interbedded argillite and greywacke of the Upper Jurassic

to Lower Cretaceous Bowser Lake Group, and the intrusives of the Early Tertiary

Lime Creek Intrusive Complex, Clary Creek Stock, and the Roundy Creek intrusive

complex. Intrusive rocks associated with molybdenum mineralization at Kitsault, Bell

Moly, and Roundy Creek are multiphase diorite, quartz monzonite, and younger

felsic units. Surrounding the intrusives are hornfels aureoles that extend up to 750 m

outward from the intrusive contact (Figure 7.1). Cross-cutting relationships within the

intrusives indicate that multiple mineralizing events are related to the molybdenum

deposits. Away from the Kitsault open pit and the adits at Roundy Creek, surface

rock exposures are limited, making an interpretation of the local geology extremely

difficult. Principally, the area is covered by soil, swamp, glacial till, and in places

basalt flows. Most of the geological knowledge for Avanti’s property position comes

from the numerous drill holes at the three deposits.

7 . 3 K I T S A U L T D E P O S I T G E O L O G Y – G E N E R A L

The Early Tertiary, Lime Creek Central Intrusive Complex is hosted by Upper

Jurassic to Lower Cretaceous Bowser Lake Group sedimentary rocks (Figure 7.2).

The Lime Creek Central Intrusive Complex as determined from the 2008 Avanti

Mining drill program contains at least six separate igneous phases and associated

dikes, with the youngest four related in time and space to the molybdenum

mineralization. All of these four intrusive bodies and the molybdenum mineralization

are cut by widespread, younger, 36 Ma-old lamprophyre dikes, along with the rare

and much younger vesicular basaltic dikes, likely related to the 1.6 Ma-old basalt

flows found capping nearby surrounding topographic highs (Woodcock, 1964; Carter,

1981; Steininger, 1985). The molybdenum mineralization forms a hollow, steeply

dipping, annular cylinder shape that is well-developed on three sides of the margins

of the central Lime Creek Intrusive Complex. The exterior of this annulus it is

localized within, along and slightly exterior of its contact with the hornfelsed Bowser

Basin sediments (Figure 7.2). Interior to the annular and cylindrical molybdenite

mineralization is a barren core. The south side of this intrusive complex is in contact

with another largely unmapped intrusive and in this area, the exterior boundary of the

annular and hollow cylinder is less well constrained by drilling and may extend

across this contact.

In the mine area, the Bowser Lake Group sedimentary rocks are characterized by

interbedded siltstone, argillite, and greywacke. Intrusion of the Lime Creek Central

Igneous Complex and associated mineralization has produced a hornfels aureole up

to 750 m outward from the complex margin (Figure 7.1); this aureole is

superimposed on regional chlorite-sericite-epidote-albite greenschist facies

metamorphism. The hornfels zone has an outer, weakly-developed albite-epidote





facies, a central pale-brown biotite zone, and an inner brown biotite zone (Kamilli,

1977).

Historically, the Lime Creek Intrusive Complex has been described as comprising six

separate intrusive phases (Woodcock, 1964; Steininger, 1985). From oldest to

youngest, these phases are:

1. the East Lobe

2. the Border Stock

3. the Southern Stock

4. the Central Stock

5. the Northeast Porphyry

6. a system of Intramineral Porphyry dikes (Figure 7.2).

The oldest intrusive phase is considered to be the East Lobe. This unit is poorlyexposed, is now covered by mine dumps, and has not been intersected by recorded

drill holes. No radiometric dates exist for this intrusive body. Chronologically

younger in the intrusive sequence is the Border stock (Carter, 1974). Woodcock

(1964) showed that the mineralogical makeup of the Border stock is relatively silicic,

but all subsequent workers have chosen to call this rock type a quartz diorite or

diorite. The largest mass of this intrusive occurs along the west side of the complex

but a compositionally similar intrusive body also occurs along part of the southeast

side of the Central Stock. Texturally, this rock type is medium-grained and

equigranular, and locally shows a distinct foliation defined by the alignment of

abundant biotite. Although the two rock types are reported to be similar, the exact

relationship between the quartz diorite in the southeast part of the Intrusive Complex,

with the adjoining East Lobe body, is unknown. Petrographic studies by Woodcock(1964) describe the Southern Stock as a quartz monzonite. Woodcock (1964)

reported that the probable contact of the Central and Southern Stocks, at the surface

south of Patsy Creek, is marked by extensive development of secondary K-feldspar

alteration but the location of this contact and the relationship with the Central Stock

remains enigmatic. Woodcock also stated that the South Stock is flanked on either

side by more basic intrusives which he described as quartz diorite and for which he

reported a date of 53.5 ± Ma. Avanti’s 2008 geologic work has shown that the

majority of the Central Stock is variably porphyritic quartz monzonite porphyry that

occupies most of the area covered by the Lime Creek Intrusive Complex, and it is the

dominant host rock for the Kitsault molybdenum deposit. The Northeast porphyry,

determined by the 2008 work to be quartz monzonite porphyry, was historicallyrecognized only in drill holes along the northeast, north, and northwest parts of the

deposit; the Intramineral Porphyry dikes have historically been spatially related to this

rock unit.





7 . 4 K I T S A U L T D E P O S I T G E O L O G Y – D E T A I L E D

The Central Stock, lying north of Patsy Creek, has been defined by all investigators

as the central and more leucocratic part of the Lime Creek Intrusive Complex.

Previous petrographic work (Woodcock, 1964) has shown the stock to be quartz

monzonite porphyry in composition. It is the largest intrusive body in the complex,

related in both time and space to molybdenum mineralization. Its texture is variably

porphyritic and seriate, and it is composed primarily of feldspar, quartz, and biotite.

Geologic work conducted by Avanti indicates that there are two phases of this rock

type:

The first phase, designated in this report as QMP1, is pre- or syn-

mineralization in age and is generally equivalent to the well-mineralized

Central Intrusive body of previous investigators.

The second and younger porphyry phase, designated in this report as

QMPll, postdates all economic molybdenum mineralization and in part

includes the historic Northeast Porphyry body, which was considered to be

the youngest major intrusive phase in the Lime Creek Intrusive Complex

(Steininger, 1985).

The differences between QMP1 and QMPll are subtle; these include the presence of

morphologically distinctive quartz and feldspar phenocrysts, which produce a more

obvious porphyritic appearance compared to the QMP1. Where the porphyritic

texture is well developed in the QMPll, especially where the porphyry contains

abundant round quartz phenocrysts, it is invariably barren of molybdenum

mineralization. In drill core, the contact between QMP1 and QMPll is always

obscured by strong silicification making age relationships difficult to determine. The

Northeast Porphyry, now included within the QMPll, was previously interpreted to bepresent only in the subsurface. Avanti’s work indicates that the QMPll also sub crops

in the center or the barren core of the Lime Creek Intrusive Complex and that it

postdates all economic molybdenum mineralization. The distinction between the

QMP1 and QMPll contacts in drill holes within the central area is more difficult to

determine than in drill holes in the northeastern and northern parts of the complex.

Two other significant, pre-molybdenum mineralization rock types are found in the

Central Stock; aplite dikes, (alaskite of earlier investigators), and intramineral

porphyry dikes. Aplite bodies typically range from centimeters to plus meter widths,

and form sheeted zones of dikes, sometimes with cumulative widths of 10m or more.

These dikes generally display a time-space and likely a genetic relationship with the

higher-grade molybdenum mineralization. They cut across all pre-lamprophyre rocktypes except for the QMPll. Intramineral Porphyry dikes have historically been

considered to be appendages of the Northeast Porphyry (now QMPll). Avanti’s 2008

logging of drill core was unable to decipher a distinctive time relationship between

the intramineral porphyry dikes and QMPll, but intramineral porphyry dikes clearly cut

across the QMP1. Most significant molybdenum deposits owe their better grades

and sizes to the superposition of several magmatic-hydrothermal events; intramineral





dikes generally document timing relationships and improve the understanding of

such multiple mineralization cycles. Steininger (SRK 2008 PEA) describes at least

one such brecciated intramineral porphyry dike at Kitsault that contains included

fragments of MoS2-mineralized rock, as well as molybdenite in its own siliceous

matrix. In the 2008 drill core, barren quartz veining, but not molybdenite veining, was

observed to predate the intramineral porphyry dikes.

The quartz monzonite porphyry (QMP1), the dominant host and central to

molybdenum mineralization at Kitsault, has been dated by Carter (1981) at

53.2 ±3 Ma, while the flanking older quartz diorite returned a date of 51.4 ±1.5 Ma.

Both K-Ar dates were obtained from rocks that were somewhat hydrothermally

altered and therefore probably reflect ages of mineralization rather than intrusive

ages. The “Late” porphyry of Carter (1981, Table 3), subsequently called the

Northeast Porphyry by Steininger (1985) and here designated as QMPll, gave a K-Ar

date of 48.3 ±1.6 Ma (Carter, 1981). This younger date supports the conclusion that

this distinctive body is the youngest major intrusive of the Lime Creek Central

Intrusive Complex.

The central QMP1 stock and its flanking quartz diorite in most places display steeply

outward dipping contacts with the enclosing, hornfelsed Bowser Basin sediments at

near surface elevations. The marginal quartz diorite skirts the hornfels contact, along

the west and north sides of the QMP1 stock; it is more abundant in the subsurface

along the eastern and southeastern contact than its surface distribution would

suggest. The QMPll has been demonstrated by the 2008 drilling to reach the surface

as a narrow, circular, sub-cropping plug within the barren core, and it broadens with

depth towards the north and northeast.

A 1970 Kennecott Exploration Services geologic bench map, which predates later

Amax mining, displays quartz monzonite porphyry dikes cutting through themineralized QMP1 quartz monzonite porphyry stock along the north half of the

deposit. These dikes, as mapped in 1970, form a crudely radial pattern around the

central core area. Together with the intramineral porphyry intrusion breccias mapped

by Kennecott (Figure 7.2), these radial dikes may be the wider high-level equivalents

to the narrow intramineral porphyry dikes observed at depth in 2008 drill core.

Although few aplite dikes are observed on the 1970 bench map, earlier investigators

(Woodcock, 1964; Carter, 1964) show concentrations of northerly -trending aplite

dikes and swarms along the west, east and southeast sides of the central QMP1

stock. The 2008 drilling confirmed these earlier observations and also intersected

abundant aplite dikes in the northeast and northwest sectors of the mineralized zone.

Aplite dikes show little continuity between adjacent drill holes, confirming their

irregular shapes as observed on a small scale in drill core. Aplite dikes commonly

show gradations from the typical sucrosic texture to micro-pegmatitic textures and

are generally spatially associated with the margins of the QMP1 stock. However, the

aplite dikes may be interpreted as a high-level expression of an unidentified deeper

intrusive. Aplite dikes commonly contain disseminated molybdenite that appears to

have formed contemporaneously with the silicate matrix of the dikes. Numerous

incipient quartz veins within the dikes, with or without molybdenite, are a common





characteristic of the aplite dikes, suggesting that these dikes represent a continuum

between the igneous and hydrothermal phases of the Kitsault system.

Lamprophyre and basalt dikes and basalt flows postdate and are significantly

younger than the molybdenum mineralization. The dense, very fine-grained, black,

hornblende-biotite lamprophyre porphyry dikes of the area have been dated at 36.5±1.2 Ma (Carter, 1981). In the Kitsault pit, these mafic dikes generally have a

northerly or easterly strike direction and are steeply dipping. These unmineralized

dikes range from 1 m to 10 m wide, and although they are irregularly-developed

across the width of the mineralized zone, they are volumetrically insignificant. One

million year old basalt flows (Carter, 1981) are found between the Kitsault and Bell

Moly deposits to the north, including those, which prominently cap Widdzech

Mountain, situated immediately to the northeast of the deposit. The rare, thin,

vesicular basaltic dikes found in the 2008 drill core are probably related to these

locally extensive and well-preserved basaltic flows.

No major faults were observed in the rock exposures of the existing pit, when it was

cursorily examined in 2008. Very few faults could be projected from hole to hole in

the 2009 compilation of the geological cross-sections. Compilation on a mid-level

plan of the deposit indicates the existence of one post mineral east-west striking fault

on the east side of the mineral zone, and another north-south, post mineral fault on

the north side of the mineral zone, both with ≤30-35 m of offset. These structures

are shown on a plan map discussed in a subsequent section. Earlier mapping by

Carter (1964) indicates the presence of a northeast-trending surface fault with minor

offset at the hornfels/intrusive contact in the northeast corner of the deposit (Figure

7.2). A review of data derived from a 1 m compilation of a Lidar survey carried out

for Avanti Mining in 2008 suggests that the east-west trending branches of Patsy

Creek and the northerly trending portion of the main Lime Creek drainage may

occupy or parallel structural zones with persistent, >1.0 km strike lengths.

7 . 5 R E V I S E D G E O L O G I C I N T E R P R E T A T I O N

Geologic information generated from the logging of all the 2008 core holes and

historic drill hole geological data has been plotted on 15 east-west and 16 north-

south cross-sections spaced at 50 m intervals, covering the entire known mineralized

zone. Cross-sectional data has been projected to the present surface, which is at

approximately the 600 m elevation, and to three deeper levels, at 550, 450, and

350 m elevation plans. Two east-west and one north-south cross-sections, as well

as the surface and 450 m level plans have been included in this document to

illustrate the deposit geology. These plans and sections display both the outline of

various molybdenum grade zones as well as the interpreted distribution of rock

types.

The compilation of the surface plan and the near surface geology on cross-sections

has been hindered by the absence of a current surface geologic map. Some of the

data shown on the Avanti, “projected to surface” geologic plan provided in Figure 7.3,





were obtained from the detailed Kennecott bench map (Giles, 1970). Although these

data reflect the pre-Amax mining surface, and may have been mapped 30-50 m or

more above the current reclaimed pit surface, the strike directions and location of

near vertical features such as the lamprophyre dikes are assumed to be similar and

have been utilized in the construction of the Avanti surface geologic map. Some

additional surface data utilized for the Avanti map was taken from Carter’s 1964 map,as provided in Figure 7.2. The construction and rectification of the east-west and

north-south orthogonal sections and plans has enabled Avanti to develop a three-

dimensional picture of the geology that both confirms historic interpretations and also

offers some new insights to distribution of rock types at the Kitsault deposit. Figure

7.3 to Figure 7.7 demonstrate the general geologic features and distribution of

molybdenum grade as follows:

a surface plan at the ~600 m level (Figure 7.3)

a plan of the 450 m elevation (Figure 7.4)

three cross-sections:

east-west Section 6,141,185ON (Figure 7.5)

east-west Section 6,141, 1900N (Figure 7.6)

north-south Section 473,350E (Figure 7.7).





Figure 7.3 Generalized Geology and Mineralization Projected to Surface from DDHs





Figure 7.4 Generalized Geology – Mineralization Level Plan 450 m





Figure 7.5 Molybdenum Assay/Geology Section 6141850N – Looking North





Figure 7.6 Molybdenum Assay/Geology Section 6141900N – Looking North





Figure 7.7 Molybdenum Assay/Geology Section 473350E – Looking West





Lamprophyre dikes are the youngest post-mineral intrusives as shown on the cross-

sections. Generalized historical surface maps and 2008 pit observations give

reasonable control on the strike and dip of these dikes. Although lamprophyre dikes

are widespread within the deposit, they tend to display poor continuity along strike

and down dip. Few dikes appear to have the vertical continuity and thickness

exhibited by the dike shown on Section 6,141,900N (Figure 7.6).

The next youngest rock unit at Kitsault is the QMPll, which includes the former

Northeast Porphyry. This unit, where it can be definitively traced on plans and on

sections, postdates all significant molybdenum mineralization. Where it carries

>75 ppm Mo values, the mineralization is either due to the presence of late quartz-

carbonate veins within the 3 m assay interval that contain trace to small amounts

molybdenite, or due to the presence of very rare, isolated quartz-molybdenite

veinlets. Most assay intervals within this rock type carry <25 ppm Mo values. The

generalized surface geology plan (Figure 7.3) shows this unit to occupy part of the

central barren core, while the 450 m plan (Figure 7.4) and the sections (Figure 7.5,

Figure 7.6, and Figure 7.7) show it to expand significantly in size with depth, cuttingthrough and underlying the higher grade mineralization in the northeast corner of the

deposit.

The 2008 drilling data suggests that the intramineral porphyry dikes predate most of

the molybdenum mineralization, but they are typically not well mineralized. Previous

reports (Steininger, 1978; 1985) indicate that these dikes postdate what was

historically referred to as the first period, but predate what was interpreted as two

later periods of molybdenum mineralization. However, such conclusions could not

be substantiated from the logging of the 2008 drill holes. Intramineral porphyry dikes

intersected in the 2008 drilling were found to be volumetrically insignificant within the

mineralized zone, and few are shown on the cross-sections. Both Carter’s 1964 map

(Figure 7.2) and the 1970 Kennecott bench map (Giles, 1970), suggests that suchdikes may have been both more abundant and wider at the original pre-mining land

surface. The Kennecott map shows them as radiating out from the area now

identified as the barren core which is occupied in part by the post mineral QMPll.

The current surface of the barren core is probably >50 m below the 1964, or possibly

the 1970, surface.

Aplite dikes (the alaskite dikes of earlier investigators) cut all earlier intrusive units.

The 2008 drill core logging also suggests the aplite dikes cut the intramineral

porphyry dikes, and no aplite dikes have been observed cutting the QMPll. These

dikes are shown on early generalized geologic maps (Carter, 1964) with 15-75 m,

northerly trending strike lengths, and moderate widths of 3-15 mm (Figure 7.2).

However, there is virtually no correlation of these aplite dikes between close spaced

drill holes in the 2009 interpretive cross-sections. This suggests that these dikes are

podiform to irregular in shape and with little continuity along strike or down dip. Their

source area is unknown, although it can be postulated that they are the product of

local residual melt pockets that have been sourced from the QMP1, the dominant

host rock for the Kitsault molybdenite deposit. Aplite dikes can contain abundant

disseminated molybdenite and pockets of high-grade clots of molybdenite as well as





irregular quartz-molybdenite veinlets, which visually appear to originate from within

the aplite bodies. This observation is compatible with the interpretation that the aplite

dikes are a likely source rock for some of the molybdenum mineralization.

The QMP1 and the marginal and older quartz diorite cover most of the area

designated as the historic Lime Creek Intrusive Complex (Steininger, 1985 and SRK,2008). The large, circular QMP1 body, approximately 500 m wide at surface, and on

the 450 m level plan, displays a plug-like morphology on the east-west cross-

sections. The QMP1 and the flanking quartz diorite have outward steeply-dipping

contacts with the enclosing hornfels on the west, north and east flanks of the deposit.

The present mineral zoning on the two sections and the 450 m level plan suggest

that that the QMP1 may have been the primary source for much or most of the

presently defined molybdenum mineralization at Kitsault. To the south, the QMP1

has an unknown relationship and morphology with what has been called the South

stock of Steininger (SRK, 2008) or the Basic stock of W oodcock (1964). Carter

describes this stock as a separate intrusive (1964) but it is not displayed separately

on his surface map (Figure 7.2). No separate and distinct quartz monzoniteintrusive, differing from the dominant QMP1, was observed in the limited amount of

2008 drilling that extends a short distance south of the present deposit. The vertical

extent of the QMP1 body is unknown. Historic drilling has generally tested this plug

down to the 200m elevation. Two historic drill holes of approximately 700 m length

have penetrated the QMP1 to about the 0 m elevation. Variable but generally weak

molybdenum mineralization in the range of 0.03-0.06% Mo and rarely up to +0.10%

Mo levels are observed to the bottom of these holes.

Quartz diorite is the oldest intrusive unit of the Lime Creek Intrusive Complex. On

the surface, the quartz diorite intrusive constitutes the western intrusive-hornfels

contact. It also flanks the east and west contacts of the Basic or South stock

(Woodcock, 1964; Carter, 1964; Steininger, 1985) as displayed in Figure 7.2. Thecontacts between the QMP1 and quartz diorite are typically obscured by alteration

and mineralization, but at core scale, inclusions of Quartz Diorite are observed within

QMP1. Although it is unclear in drill core logging whether the Quartz Diorite is an

early phase of the QMP1 or a distinctly separate intrusive phase, the 2009 geologic

compilation interprets the quartz diorite as a earlier intrusive that has been cut by the

central QMP1 intrusive leaving screens or major inclusions within, and a rind of

quartz diorite around, the main stock of QMP1.

The Cretaceous Bowser Lake Group argillites and graywackes of the Kitsault area

are the oldest rocks in the deposit. Cross-sectional interpretations support that the

outer boundary of the annular molybdenum mineralization on three sides is confined

largely within, coincident with, or extending not far beyond this contact. This

geometry as well as data from the 2008 core logging suggests that the hornfels was

not receptive to mineralization, except where it was heavily cross-cut by aplite dikes.

Minimal bedding attitude information (Carter, 1964) and cursory pit observations

made in 2008 indicate that the Bowser Lake Group sediments in the pit area have a

general N25°-45°E strike and northwesterly dip angles of 20-60° (Figure 7.1 and





Figure 7.2). These trends indicate that the Lime Creek Intrusive Complex is largely

discordant to the bedding in these thermally altered sediments.

Attempts to compile fault structures on cross-sections have not proven to be very

successful. Widespread structural breaks, generally marked by gouge-bearing fault

zones from centimetres to less than a metre in width, are commonly observed inmany of the 2008 drill holes. However, these fault zones are difficult to correlate

between sections. The 450 m level plan displays two interpreted faults:

1. a likely east-west fault with 50 m of offset at the intrusive-sediment contact

on the E side of the deposit

2. a northerly trending fault on the north side of the deposit with up to 30 m of

offset on the same contact.

Cursory, 2008 surface observations of the pit walls identified no major fault zones

within the pit area. Carter’s generalized geologic map of the pre-mining surface

(Figure 7.2) shows a single fault, striking N60°E in the northeast sector of thedeposit, with a 15 m offset along the hornfels -intrusive complex contact. The Avanti

“projected to surface” geology plan (Figure 7.3) shows a probable pre-mineral shear

or fault zone in the southwest corner of the deposit.

7 . 6 B E L L M O L Y P R O S P E C T G E O L O G Y

Molybdenum mineralization at the Bell Moly deposit is related to two intrusive events

that are hosted by the Upper Jurassic Bowser Lake Group rocks (Steininger and

Card, 1979). Surface outcrops of non-basaltic rocks are quite limited, and essentially

all of the geological knowledge of this deposit is based on drill core. Within the Bell

Moly drill area, the principal member of the Bowser Lake Group is massive

greywacke to micro-greywacke, interbedded with minor black slate and argillaceous

mudstone. Intrusive activity produced extensive hornfelsing of the sedimentary

rocks. Two separate zones of hornfels are known: one is closely associated with the

Southwest Zone intrusive. The other, larger zone is associated with the Clary Creek

stock and extends significantly northwest of the Clary Creek stock, suggesting that

additional intrusive rocks may be present at depth to the northwest (Figure 7.8).

Relatively unaltered sedimentary rocks were only encountered at the periphery of the

principal drill area.

Two principal intrusive centers have been identified: the Clary Creek Stock Complex

and the Southwest Zone (Figure 7.8). The Clary Creek stock consists of at least fiveseparate intrusive phases of generally quartz monzonitic composition, while the

Southwest Zone is primarily a dike swarm apparently centered on a small stock, also

of quartz monzonitic composition.





Figure 7.8 Bell Molybdenum Area Geology

The Clary Creek Stock is an east-west elongated elliptical mass, with surface

dimensions of roughly 700 m x 300 m. There is a conspicuous hornfels roof pendant

in the center of the stock. The oldest intrusive phase, predominant in the

northeastern portion of the complex, is quartz monzonite porphyry which displays amafic-rich border approaching granodiorite in composition. Carter (1974) obtained

an age date of 51.7 ± 2.2 Ma for the border phase. A swarm of alaskite dikes appear

to have been intruded into the center of the quartz monzonite porphyry. A quartz-eye

porphyry subsequently intruded into the southwestern part of the complex, its dikes

cross-cutting the previous two intrusives. Carter (1974) obtained a 52.9 ± 2 Ma age

date for this latter intrusive phase. A second generation of quartz-eye porphyry dikes

cut all of the older phases, but no central intrusive body related to this unit has been

identified. The youngest known intrusive, occurring in the southwestern portion of

the complex, is called the Crowded Porphyry. This intrusive unit appears to postdate

molybdenum mineralization.

The Southwest Zone is located about 1 km southwest of the Clary Creek stock

(Figure 7.8). All of the intrusives intersected by drilling are dikes, with the possible

exception of a small stock encountered in one drill hole. Although these dikes have

similar compositions, the varying age relationships with respect to mineralization

suggest a continuum of intrusions throughout development of the molybdenum

mineralization. All of the dikes are sufficiently altered to preclude determination of





original composition, but they appear to be similar to the rhyolitic quartz-eye

porphyries of the Clary Creek stock.

Numerous mafic dikes occur in the drilled portions of the prospect and are probably

related to the regional but very young basaltic intrusive event. These basaltic dikes

have been dated at between 0.62 ± .80 Ma and 1.2± .80 Ma (Carter, 1981). Coredrilling has revealed that lacustrine and glacial deposits lie below many of these

basaltic flows, but on top of the Tertiary or Cretaceous bedrock.

Carter (1967) defined two ages of regional folding, one with a north to northwest axis

and a younger one with an east-northeast axis. Faults, some hosting mafic dikes,

are commonly oriented along the trend of the younger fold direction. In the

immediate area of the Clary Creek stock and Southwest Zone, faults with north-

northwest trends are common.

7 . 7 R O U N D Y C R E E K G E O L O G Y

Molybdenum mineralization at the Roundy Creek deposit is associated with a small

composite quartz monzonite stock that is disrupted by faulting (Woodcock and

Carter, 1976) (Figure 7.9).

Figure 7.9 Geology of the Roundy Creek Area





The central intrusive is a leucocratic quartz-phenocryst-bearing quartz monzonite

porphyry that is locally brecciated. Two phases have been identified: one rich in

biotite xenoliths and one that is banded. A younger biotite-bearing quartz monzonite

is present at the margins and in the core of the intrusive complex. This phase has

been dated at 52.5 + 2 Ma (Carter, 1981). Alaskite dikes cut all of the quartz

monzonite bodies and in turn are cut by narrow dikes of light gray biotite quartzmonzonite. The Roundy Creek Complex intruded a sedimentary rock package

similar to that found at Kitsault, resulting in the development of a 60 m wide hornfels

zone outward of the intrusive margin. Lamprophyre dikes with a general

northeasterly strike cut the entire complex.

7 . 8 R E V I E W O F G E O L O G Y

SRK conducted a detailed review of the geologic interpretive work conducted by

Avanti personnel at Kitsault, and concludes that data collection procedures,

methodology used for geologic interpretation and documentation geologic data is of high quality. The following data was reviewed:

geologic logs and logging procedures

core photographs

interpreted level plans and cross-sections

visual inspection of outcrop in the existing pit.

It was noted that geologic data was not used to constrain grade estimation in the

current resource model but simplified geologic 3D solids were constructed from these

data and utilized for density assignment in the model. Statistical analysis by rock

type confirmed that these data represent a single population, validating grade

estimation across lithologic boundaries.





8 . 0 D E P O S I T T Y P E S

The Kitsault, Bell Moly, and Roundy Creek properties are intrusive-related

molybdenum deposits that, in plan view, are geometrically annular and tend to be

cylindrical-to-tabular and arcuate in a cross-sectional view. Other molybdenum

deposits in North America which have similar intrusive rock types, but dissimilar

morphology to those observed at the Kitsault property, include Ajax (Dak River, BC),

Cannivan Gulch and Bald Butte (Montana), and, to a lesser extent, those at Pine Nut,

Liberty (Hall), and Buckingham (all in Nevada) and Creston (Sonora, Mexico).

This group of intrusive-related molybdenum deposits differs distinctly from the quartz

monzonite batholithic type, where molybdenum mineralization is related to a series of

sheeted vein systems. These latter deposits are usually more lenticular in shape andare commonly related to the youngest phases of the batholithic rocks which host

these deposits. Typical examples of the batholithic type in North America are

Endako and Adanac (BC), Mount Tolman (Washington), and Thompson Creek

(Idaho).

Although Kitsault shares some similarities with the high-silica, rhyolitic to alkalic

deposits of the western US, with respect to time-space relationships of mineralization

with its host intrusives, it differs from these deposits in both rock chemistry and in

morphology. This latter group of rhyolitic to alkalic systems is best typified by the

Climax and Henderson deposits (Colorado), which are the world’s two largest

porphyry molybdenum deposits.





9 . 0 M I N E R A L I Z A T I O N

9 . 1 K I T S A U L T M I N E R A L I Z A T I O N

The 2008 drilling program conducted on the Kitsault deposit demonstrates that the

economically important molybdenum mineralization occurs primarily in two styles.

The first (and perhaps earliest) style is a system of quartz-molybdenite stockworks;

the second mineralization style is an irregularly-developed, sheeted vein zone style

of mineralization, which tends to be best developed in close proximity to the aplite

dikes. Together, these two styles of mineralization define the hollow, cylindrical, and

annular shaped body of mineralization that more or less follows the contact of the

QMP1 and Quartz Diorite with the surrounding hornfelsed sediment as it extends todepth. A third, much younger, and economically insignificant type of molybdenum-

bearing mineralization consists of steeply dipping quartz-carbonate veins that contain

sphalerite, galena, a variety of Pb-Bi sulphosalt minerals and occasional trace

amounts of molybdenite. This very late base metal mineralization is widely

distributed and is found inside of, within, and well outside of the limits of the

potentially economic molybdenum mineralized zone. Within the deposit, this base

metal veining seems to be best developed on the south side of the molybdenite

zone. The quartz-carbonate veins may have some economic significance because

of their persistent, low-level silver content. This base metal mineralization post-dates

the QMPll intrusive, and a few very rare and isolated quartz-molybdenite veinlets

have also been observed in the deeply buried QMPll intrusive.

The economically important molybdenite-bearing veins at Kitsault consist almost

exclusively of milky quartz veins and veinlets, with lesser to minor amounts of fine-

grained molybdenite and only minor amounts of pyrite or other gangue minerals.

These veins exhibit two forms:

1. 0.25 to 1.50 cm-wide quartz-molybdenite veins and veinlets forming

stockworks

2. more widely spaced, sub-parallel, 3-20 cm wide, sheeted, milky quartz veins

containing fine-grained molybdenite.

Coarse-grained molybdenite has been observed rarely in the second type of molybdenite-bearing veins. Past exploration workers at Kitsault termed these latter

veins “ribbon banded” quartz-molybdenite veins (Climax, 1978; Steininger, 1985).

Where these veins are less than 1.5 cm wide, molybdenite is confined to vein

borders. Where the veins are wider, there are multiple epitaxial bands of

molybdenite within the quartz vein as well as on its margin. Most of the ribbon-

banded quartz-molybdenite veins carry little or no interstitial potassium feldspar and





minimal sericite and/or feldspar alteration halos, and are most abundantly developed

in and around the aplite dikes. Disseminated, crystalline, fine-grained molybdenite is

sporadically distributed within the groundmass of all premineral rock types, but

typically is most abundant in the aplite dikes. Many molybdenum deposits are

distinguished by widespread development of molybdenite coated fracture and joint

surfaces, commonly called “moly paint”, but this is not commonly observed atKitsault. One other molybdenite vein style was noted in the 2008 drill holes. This

consists of 0.50 to 2.0 cm pyrite veins containing subordinate molybdenite and trace

amounts of quartz. The molybdenum contribution of this vein type, which is largely

confined to the hornfels, is negligible.

Quartz-carbonate (dolomite+calcite) base metal-sulfide veins, typically ranging from

4 cm to 1 m or more in width, are abundant in virtually all of the core from the 2008

drill holes. These veins have sharp contacts and cut across all rock types of the

Lime Creek Intrusive Complex, including the post-molybdenite-aged QMPll.

However, these veins appear to predate the lamprophyre dikes. Typically, these

veins contain from 2 to 5% total sulfide, although occasionally the total sulfidecontent can be as high as 50% or more. Sulfides minerals include pyrite, an

unusual, honeycomb-textured amber sphalerite, galena, and sectile and soft, grey

lead-bismuth sulfosalts, including akinite, cosalite, and neyite. Neyite (Drummond et

al., 1969) contains essential Ag and Kitsault is the type locality for this mineral. The

Pb-Bi sulphosalt minerals occur as very fine-grained vermiform layers within the

quartz-carbonate gangue, which with or without a hand lens resemble fine-grained

molybdenite. Much care is required to differentiate between the trace molybdenite

occasionally found in these veins and the much more common sulphosalts.

Systematic trace element analysis of the 2008 core has shown that persistent but

trace amounts of anomalous silver values are coincident with the distribution of the

quartz carbonate veining. Other minor gangue minerals identified in these base

metal veins include fluorite and barite. Rare to minor violet-colored, late anhydritebearing veins have also been identified in drill core but no systematic distribution of

this vein type has yet been recognized.

The Kitsault molybdenite deposit contains very little copper based on existing

analyses, and no visible chalcopyrite was noted in the 2008 drill cores. Systematic

multi-element analyses show that background copper levels, outside of zones with

abundant quartz-carbonate veins, are in the 20 to 50 ppm Cu range. Copper values

in drill core, where late quartz-carbonate veining is abundant, are typically in the 50

to 250 ppm Cu range, with very rare instances where values exceed 1,000 ppm Cu.

Unlike some of the other molybdenite deposits in North America to which Kitsault has

been compared, such as Buckingham and Liberty (Hall) in Nevada and the Creston

property in Mexico, the Kitsault molybdenum mineralization is essentially a copper-

free molybdenum system.





9 . 2 M O L Y B D E N U M D I S T R I B U T I O N – P L A N A N D S E C T I O N

As a result of the 2008 drilling program the distribution of molybdenum values at

cutoff grades of 0.05%, 0.10%, 0.15%, and 0.20% Mo, has been examined on east-

west and north-south cross-sections, throughout the entire area of the mineral

deposit. This distribution has also been interpreted by projection to the present

surface (~600 m elevation) and at the 550, 450 and 350 m elevations. Examples of

these interpretations in plan for the near surface mineralization are shown on

Figure 7.3 and at the 450 m elevation on Figure 7.4. Examination of the near surface

mineralization at a cutoff of 0.05% Mo shows it to display a hollow cylindrical or

annular shape with widths from 100 to 150 m on the east, west, and north sides, and

a less well defined zone to the south, where it is at least 300 m wide and may extend

to nearly the southern limits of drilling. At about ~150 m below the present surface

(450 m elevation), as provided in Figure 7.4, the width of the mineralized annulus at

a cutoff of 0.05% Mo compares well to the geometry and extent of the near surface

contour, but the 0.15% Mo contours have narrowed substantially. As observed in

cross-section (Figures 7.5, 7.6, and 7.7), the mineralized annulus displays variablewidths at a 0.05%Mo grade cutoff and extends to at least the 200 m elevation. On

the north-south cross-section (Figure 7.7) the annular mineralization along the north

side or limb of the barren core has a -70° northward dip, and is confined largely

within the intrusive complex. In contrast, the southern limb is much wider and dips

north at about -40° near the surface, steepening with depth. On the east-west cross-

section 6,141,900N (Figure 7.6), the east and the west, carrot-shaped limbs of the

cylindrical mineralization show a more or less vertical inclination. The width of the

eastern limb, at grades above 0.05% Mo, is about 175 m near the surface but thins

substantially below the 400 m level. The width of the western limb is about 130 m

and it persists at that width at depth. The steeply inclined northward dip of the

annular cylinder on the north-south section (Figure 7.7) suggests that the deposit

may have been tilted to the north during post mineralization deformation.

9 . 3 K I T S A U L T G E O C H E M I S T R Y

A wide ranging geochemical study of the areal distribution of multiple elements was

carried out over the Kitsault/Lime Creek deposit and the immediately surrounding

area by Woodcock (1964), early in the history of the exploration of the deposit. This

study was based on a large collection of outcrop samples, and core samples from

the top of the bedrock surface in areas where there was heavy surficial cover. The

samples were taken from within and at some distance outside the 0.75 km2

area

covered by the Lime Creek Intrusive Complex. Many samples were also taken alongtraverses up to 1.75 km outward, in orthogonal directions, from the margin of the

intrusive complex. These surface samples, nearly all of which were reported to

contain fresh sulfide minerals where mineralization or alteration was present, were

analyzed by Kennecott Exploration Services at their Salt Lake laboratory. The

analyses were done by either emission spectrography, x-ray fluorescence, or wet

chemical methods. This work showed that molybdenum values outboard of the east,





north, and west sides of the annular molybdenum deposit from the intrusive-hornfels

contact generally dropped dramatically from economically significant grade values to

150-200 ppm Mo, within a distance of 50 to 75 m outside of this contact. However,

these data also show that strongly anomalous Mo values extend well south of the

southern half of the annular molybdenum deposit as defined by the 0.10% Mo grade

contour. The exterior boundary of the +0.10% Mo zone coincides with the surface of a major drainage known as Patsy Creek. Samples collected in 1963 for this study

from up to 100 to 200 m south of the deposit include a small number of samples

which contain Mo values of 200 ppm to as much as 370 ppm Mo. These were

collected in part from areas around and up to 50 m south of the three southernmost

Kennecott drill holes, collared 150 m south of Patsy Creek (Figure 9.1). Kennecott’s

geochemical data further show that barium values greater than 2,000 ppm are

centered on and tightly confined to the area of the known deposit but that anomalous

barium, at levels of ≥1,800 ppm Ba, also extends some 200 m or more south of

Patsy Creek. Anomalous silver at values of +4.0 ppm, and fluorine values between

1,400 to 1,800 ppm overlie the south half of the molybdenum mineralized zone, but

also extend some 200 m or more south of Patsy Creek and are in part coincidentwith the anomalous Mo in this area. Bismuth in the 10 to 50 ppm Bi range, lead in

the 80 to 160 ppm Pb range, and zinc in the 160 to 300 ppm Zn range also form

coincident, broad halos over the south half of the molybdenum deposit. But

geochemical halos of the these same elements also extend about 100 to 200 m

south of Patsy Creek and, in part, are co-incident with the Mo, Ag, and Ba anomalies

around and just south of the southernmost drill holes. The main Kitsault deposit is

encircled on three sides by the hornfels-intrusive complex contact, but on the south

side, the QMP1 is in contact with the quartz monzonite of the South Stock. An

analysis of the Kennecott multi-element data suggests:

hornfelsed sediments, where present, tightly constrain all hydrothermally

introduced elements of possible interest, particularly Mo, to within 50 to 75 mor less within the Lime Creek Intrusive Complex

Ag, Be, and F, in addition to Mo, are well developed at anomalous levels

within the Mo deposit

anomalous Bi, Pb, and Zn are distributed along and to the south of the

contact between the central QMP1 stock and the South intrusive, south of

the main molybdenum deposit

multi-element Pb, Zn, Bi, Ag, F, Ba anomalies are coincident with scattered

high Mo values around and to the south of the southernmost Kennecott drill

holes, south of Patsy Creek.

It is unknown whether this base-precious metal-molybdenum mineralization

extending well south of the known deposit could potentially overlie a separate, blind,

molybdenum system that has not been fully tested by the existing drilling in this area.





Figure 9.1 2008 DDH Location Map





9 . 4 K I T S A U L T H Y D R O T H E R M A L A N D T H E R M A L A L T E R A T I O N

The Bowser Basin Group argillites and siltstones surrounding the Kitsault-Lime

Creek Intrusive Complex (Figure 7.1) have been thermally metamorphosed or

hornfelsed at distances as far as 750 m away from the intrusive contact (Woodcock,

1964 and Carter 1964, 1981). The hornfels consists predominantly of brown fine-

grained secondary biotite and cherty to massive cryptocrystalline quartz. Biotite

hornfels locally contains small veinlets of epidote and clots of andraditic garnet. The

absence of any well-developed skarn in the hornfels indicates that no significant

calcareous units were present in the Bowser Lake Group sediments adjacent to the

deposit. It is believed that the thermal metamorphism which created the hornfels is

most likely a result of the introduction of the early Quartz Diorite and the associated

East Lobe intrusive. This more mafic and older intrusive complex may have

originally been present in the entire area now occupied by the Central Intrusive

Complex, and therefore had the most intrusive surface area available for affecting

the thermal metamorphism along its contact with the host sedimentary units.

Hydrothermal alteration is superposed upon the thermal alteration and can be

divided into four phases. The most dominant and perhaps the earliest phase is

silica and potassium feldspar introduction associated with 1 to 4 cm-wide barren,

grey to white quartz and associated pink K-feldspar stockwork veining. Quartz is

present only as veins and not as pervasive silica flooding but the K-feldspar can be

found in both habits. This alteration is best developed and most observable within,

and immediately adjacent to, the barren core. However, it extends away from the

core where it is cross-cut by molybdenite-bearing quartz veins, suggesting that the

silica-K-feldspar alteration is early. The second type of alteration is phyllic,

consisting of occasional and poorly-developed envelopes of sericite surrounding

quartz-molybdenite-pyrite veins and veinlets as well as abundant and widely-dispersed disseminated zones of quartz-sericite-pyrite. This more or less circular

and pervasive quartz-sericite-pyrite zone or halo encompasses most of the

molybdenite zone and extends outward from the molybdenite stockwork zone to the

contacts of the QMP1 with the enclosing, hornfelsed, Bowser Lake Group

metasediments. Some of the biotite hornfels has also been bleached and sericitized.

The third alteration type is an erratically developed pyritic halo. At the present land

surface it is poorly developed or absent on the east and north sides of the intrusive

complex, but is moderately well developed within the QMP1 along the south side of

the deposit where it is coincident with the inner annulus of the mineralized zone.

However, the pyrite halo is probably best developed at the surface in the hornfelsed

sediments along the west and northwest sides of the molybdenum mineralizationwhere it overlies and is immediately outboard of the intrusive-sedimentary contact.

Late in the hydrothermal sequence and overprinting the three earlier alteration and

mineralization events is a sporadically distributed but widespread, pervasive, argillic

alteration. This alteration is expressed as kaolinite-illite preferentially replacing

plagioclase feldspar Where well developed, the clay minerals completely replace the

primary plagioclase as well as some of the primary potassium feldspars.

Megascopically, argillic alteration had little if any effect on the surrounding hornfelsed





sediments. This alteration type forms a superimposed, roughly circular pattern,

extending from the outer boundary of the barren core to the contact of the QMP1

and/or quartz diorite intrusives with the hornfelsed sediments. The widespread

pervasive development of finely disseminated calcite and dolomite within the

intrusive rocks appears to be the youngest alteration event. Most of the more

strongly altered intrusive rocks will, when sprayed with hydrochloric acid, exhibitmoderate to strong pinpoint effervescence. This suggests that the very fine grained

free carbonate is probably calcite, which may be an alteration product formed during

the late argillization event.

9 . 5 B E L L M O L Y M I N E R A L I Z A T I O N

Two zones of molybdenum mineralization are documented at the Bell Moly property -

one closely associated with the Clary Creek stock and the second associated with

the intrusives in the Southwest Zone (Steininger and Card, 1979) (Figure 7.8).

Molybdenum occurs most commonly in quartz veinlets, which are up to 0.5 cm wide,and less commonly as disseminated grains in the intrusives. Within quartz veins,

molybdenite occurs in the following, in decreasing order of abundance:

as selvages along vein borders

in sub-parallel bands throughout the veins

as finely divided crystals throughout the veins

as hairline quartz-molybdenite veinlets

as fracture coatings.

Wider-spaced, sheeted quartz-molybdenite veinlets are the most prevalent form of mineralization, with stockworks of quartz-molybdenum veinlets being less common.

High-grade veins are abundant in the drill core in orientations that are sub-parallel to

the core axis. Quartz-molybdenite vein distribution is erratic and results in assay

values in drill core to vary from 0.03% Mo to 0.29% Mo in adjoining 3 m assay

intervals.

Scheelite occurs in quartz veinlets that generally cross-cut molybdenite bearing

veinlets. Pyrrhotite is the most widespread and abundant sulfide in the deposit,

occurring most commonly within hornfels surrounding the molybdenum zone. Pyrite

is commonly associated with pyrrhotite, but has a reverse relationship, increasing in

abundance toward the center of the molybdenum zone, while the pyrrhotite

decreases in abundance in the same direction.

Disseminated pyrrhotite appears to have formed during hornfels formation. The

oldest hydrothermal mineralization is manifested by quartz-albite-pyrrhotite veinlets

that cross-cut and alter hornfels, but are in turn clearly cut by quartz-molybdenite

veinlets. There are at least four stages of quartz-molybdenite veinlets, the earliest of

which are up to 5 cm wide, with bands and clots of pyrrhotite, pyrite, chalcopyrite,





and minor molybdenite. The second molybdenum event is represented by narrow

veinlets with finely divided molybdenite grains. Stockworks and preferred

orientations to the veinlets are common, and these veinlets are probably related to

the first intrusive phase of the Clary Creek stock. Older quartz-eye porphyry appears

to have been the source for the third stage of quartz-molybdenite veinlets. These

veinlets are commonly banded and up to 1 cm wide. The abundance of molybdenitein veinlets decrease with depth, and the veins converge into a silicified zone. The

final stage of molybdenum mineralization is represented by a set of narrow quartz-

molybdenite veins that cross-cut all of the earlier vein sets. Base metal sulfide-

bearing quartz veinlets which cut molybdenum mineralization make up the last stage

of mineralization.

Hydrothermal alteration is developed along all veins margins. There is a silicified

zone underlying the molybdenum zone in the Clary Creek stock, grading upward into

a zone of secondary potassic feldspar, which is more or less coincident with the

molybdenum mineralization. Much of the secondary feldspar can only be observed

at the microscopic scale. Argillization is common outward for the molybdenum zonein areas where veining is less dense. A poorly developed zone of propylitic alteration

is superimposed on the hornfels surrounding the molybdenum deposit.

Mineralization and alteration in the Southwest Zone are similar to those near the

Clary Creek stock but less well understood.

9 . 6 R O U N D Y C R E E K M I N E R A L I Z A T I O N

Three zones of molybdenum mineralization are observed at Roundy Creek,

developed in an eastern zone and two higher-grade zones in the western part of the

intrusive complex (Carter, 1981) (Figure 7.9). The eastern Roundy Zone consists of

quartz-molybdenite veins that commonly have random orientation and molybdenitepaint on fracture surfaces, resulting in a more continuous but lower grade

mineralization than found in the other two zones. This zone is defined by limited

drilling and appears to be open at depth. There are two higher-grade molybdenum

zones in the western portion of the complex the Sunlight Zone and the Sunshine

Zone. Much of the molybdenite in these zones occurs in the alaskite phase of the

intrusive complex (Woodcock and Carter, 1976). Most of the mineralization in the

three zones is closely associated with the introduction of secondary potassium

feldspar or biotite alteration.

The western Sunshine Zone has been drilled and exposed in two adits at the 260 m

and 320 m elevations. The deposit as outlined by the 0.06% Mo grade contour has a

podiform shape and a continuous internal zone of 0.12% Mo. The western Sunshine

Zone has been defined only by drilling. The eastern Sunlight Zone is smaller than

the western Sunshine Zone. This lens-shaped zone, as defined by the 0.06% Mo

contour, contains numerous 3 m intervals of +0.6% Mo. Drilling appears to have fully

delineated the zone. The Roundy Zone is exposed at the surface but is not well

defined by drilling. The zone is lower in molybdenum grade than the Sunlight and

Sunshine Zones and does not appear to contain an internal high-grade component.





1 0 . 0 E X P L O R A T I O N

Because most relevant exploration activities occurred from the late 1950s through

the 1980s, exploration activities are also summarized in the “History” section of this

report (Section 6.0). It is important to understand the results of previous exploration

programs for the optimization of recommendations for additional exploration.

Therefore, an expanded summary of past exploration is provided in the sections

below.

1 0 . 1 K I T S A U L T

Molybdenum mineralization was first recognized at Kitsault in the early 1900s.However, the Kitsault deposit was not seriously explored for molybdenum until 1956,

after a prospector brought the property to the attention of Kennecott Exploration

Company (KEL) (Woodcock and Carter, 1976). KEL optioned the property in 1957

and conducted an initial drilling program. The success of this early work resulted in

KEL purchasing the property in 1959; mine production subsequently commenced in

1967. KEL continued to drill and define the deposit until the mine closed in 1972.

CMC purchased the property from KEL in 1973 and continued drilling the deposit,

recommencing production in 1981. The mine was closed in 1982, again due to low

molybdenum prices; no exploration activities have since been conducted on the

property.

In addition to drilling, KEL and CMC developed an extensive geological and

geochemical database. With the exception of some historical digital assay and drill

hole geologic data compiled by CMC in 1973 to 1979 and recovered by Avanti in

2008 from the Mintec consultants in Tucson, AZ, very little of the hard copy data

relating to the Kitsault deposit has been recovered in the last two years. However,

recent investigations show that much of the original Kennecott data probably resides

in the KEL (now Rio Tinto) archives in Vancouver, BC.

1 0 . 2 B E L L M O L Y

Molybdenum was discovered at Bell Moly in 1965 by Highland Bell Mines, Ltd. and

Leitch Gold Mines. Drilling conducted in 1967 and 1968 by the combined company,

under the name Bell Molybdenum Mines Ltd., lead to the initial definition of a low-

grade molybdenum resource. CMC optioned the property in 1975 and continued

exploration, resulting in the definition of two separate molybdenum zones, which

were too low-grade to be considered for a standalone mining operation. No further

exploration was conducted after the 1978 drilling program, although CMC retained





the property as a possible satellite deposit to the Kitsault operation. An extensive

database was developed, and some of this data was recovered and is now available

for planning additional exploration in the area.

1 0 . 3 R O U N D Y C R E E K

Although molybdenum mineralization at Roundy Creek was documented in the early

1900s, exploration was conducted primarily between 1965 and 1971 (Woodcock and

Carter, 1976). During that period, about 9,300 m of core drilling and 780 m of

underground development were conducted. CMC purchased the property in 1975,

but only mapping and limited sampling were carried out at that time. The property

was subsequently considered too small for a standalone operation, but was

maintained as a possible satellite deposit for the Kitsault operation. Much of the

compiled information from the latest Amax work done in the 1970s on Roundy Creek

was recovered by Avanti in 2008 from a former Climax employee.

1 0 . 4 F U T U R E E X P L O R A T I O N W O R K

10.4.1 K I T S A U L T

In early-to-mid 2008, while Avanti Mining was in the early stage in its re-evaluation of

the three molybdenum deposits, some initial ideas were formulated for the expansion

of the defined molybdenum resource and the possible discovery of previously-

unknown molybdenum zones at Kitsault. Much of the region is covered by glacial till,

swamp, soil, and post-mineralization basalt flows, all of which may conceal additional

areas of molybdenum mineralization.

Several exploration targets of interest were proposed in the July 15, 2008 report on

resources (SRK, 2008). These included:

the exploration of the deep levels of the deposit below and adjacent to the

barren core zone, where limited historic drilling had intercepted

mineralization

areas at depth to the northeast of the present deposit, if the deposit was

tilted as previously postulated

a general search for a blind intrusive which would have been related to the

youngest of the historically postulated three stages of molybdenitemineralization.

During the 2008 Avanti drill program, only a limited amount of drilling tested the

deeper areas of potential mineralization, below and adjacent to the barren core zone.

Most of this drilling was to the depths tested failed to discover significant (+0.10%

Mo) mineralization. Drilling to the northeast and east of the known mineralization,

along and external to the contact of the intrusive complex with the hornfelsed





sediments, suggested that the mineralization in this area has dramatic grade

gradients and does not extend far beyond the hornfels contact. The late quartz

monzonite porphyry QMPll, which was thought prior to 2008 to be a buried and not

necessarily a post ore intrusive, was much better defined by the 2008 drilling. This

recent work reveals that it subcrops in the central core, postdates the known

economic molybdenum mineralization, and is not a likely target for a blindmolybdenum deposit.

Several primary dispersion geochemical studies were completed at Kitsault; the most

significant of these were reported on by Woodcock (1964). The Woodcock data

have been discussed in Section 9.3 (Kitsault Geochemistry) of this report. Several

anomalous elemental distributions encircle or partially encircle the deposit and

extend south of the existing drilling, but within the boundary of the South Stock.

These include Ag, Bi, Pb, Zn, F, Ba and most importantly, Mo. Based on the

knowledge gained from the 2008 drill holes, the first four and possibly five of these

elements are considered to be related entirely to the very late stage post-ore quartz-

carbonate Zn-Pb-Ag-Bi veins that carry rare, trace Mo. The 2008 drill results as wellas the 1964 geochemical studies show that the Ag, Bi, Pb, Zn, F, and Ba anomalies

occur over the south half of the known deposit and extend for up to 200-250 m south

of Patsy Creek, up to 50-100 m south of the southernmost historic drilling. The

molybdenum values in subcrop at the top of the southernmost drill sites range from

60 to 700 ppm Mo and peak at 960 ppm Mo. Approximately 50-100 m further south,

the values range from 20-40 ppm Mo with a peak value of 200 ppm Mo. This entire

area requires further surface sampling and study, followed up with drill testing.

The general lateral and vertical extents of the Kitsault deposit were largely but not

completely defined by the 2008 drill program, and the potential exists to expand the

limits of the known zones to the south and west sides of the deposit. Of particular

note is the northwest sector of the deposit, where abundant and wide dikes of aplitewere encountered in drill hole K08-22 (Figure 9.1); this hole bottomed in +0.15% Mo.

Based on the geology observed in K08-22, a limited amount of shallow to

intermediate depth drilling around the presently known limits of molybdenum

mineralization in the northwest sector of the deposit should be targeted for future

drilling. Molybdenum distribution patterns outlined by the 2008 drilling indicate that

at near-surface to mid-level depths, the southern part of the deposit, under and

south of Patsy Creek warrants further drilling to define lower grade (≥0.05 ≤0.10%

Mo), but still potentially economic zones of mineralization. The prospect for the

existence of another separate molybdenite-bearing igneous body further to the south

of the most southerly drill holes remains a target; however, the likelihood for success

for such a discovery is considered to be low.

10.4.2 BE LL M O LY

At the Bell Moly property, hornfels is closely associated with intrusive activity, and

these intrusives are closely associated with molybdenum mineralization. Hornfels

extends well northwest of the Clary Creek stock, suggesting additional intrusive

activity in that direction and the possibility of additional molybdenum mineralization.





Initially, a geophysical program may be the most prudent method of exploration in

this area, to assist in defining potential drilling targets.

10.4.3 R O U N D Y C R E E K

A drill hole location map, cross-sections displaying down-hole assays, and geological

maps of the two adits were recently acquired by Avanti for the Roundy Creek

property (D. Davidson, personal communication, 2008). These drill results suggest

that while the Sunlight and Sunshine zones appear to be fully delineated by drilling,

the Roundy Zone is still open at depth. A re-evaluation of the existing geologic data

is warranted to determine if these near-surface deposits are upper level expressions

of a deeper mineralized body of significant size. Positive results from such a study

would then require ground follow-up to determine if further drilling is appropriate.





1 1 . 0 D R I L L I N G

The digital drill hole databases for Kitsault and Roundy Creek have been recovered

by Avanti from archived files located at Mintec, Inc. (Mintec) offices in Tucson,

Arizona. Mintec personnel (J. Thornton, personal communication, 2008) indicated to

Avanti that these were the master databases that supported the reserve statements

submitted to the US Securities and Exchange Commission (SEC) by Amax for the

periods 1980 through 1985.

An intensive search was conducted at the former Amax and CMC records located in

the Freeport McMorran Inc. (Freeport) archives in Phoenix, AZ. This effort recovered

a large amount of useful data but the majority of the many documents known to have

existed were not found. Additional investigation for the possible location of any LimeCreek files that might have formerly resided in the archives at the KEL headquarters

office in Salt Lake City, Utah have been non-productive (R. Blair, personal

communication, 2009). However recent discussions with the KEL (now Rio Tinto)

office in Vancouver, BC, indicate that data from the Lime Creek (now Kitsault) project

including drill logs, is stored in the KEL archives in Vancouver, BC, and is available

for purchase. Discussions with Dick Woodcock (R. Woodcock, personal

communication, 2009) and Nick Carter (N. Carter, personal communication, 2009)

resulted in the recovery of valuable early data, including core logging notes on many

of the KEL drill holes (N. Carter, personal communication, 2009). The Carter logging

data was used in the 2009 Avanti geology and mineral zone compilation and

interpretive work.

1 1 . 1 K I T S A U L T H I S T O R I C

Several drilling campaigns were undertaken at Kitsault — initially by KEL, and related

companies, and finally by CMC and ACL — to define the molybdenum deposit. The

original operator, KEL, drilled 98 core holes totalling 18,372.1 m (Amax, 1979). CMC

drilled an additional 38 holes totalling 7,651.3 m. Three holes totalling 1,627.65 m

were also drilled by Amax outside of the mine site and along a proposed tailing

tunnel site. All of the historic drilling conducted on the Kitsault deposit was core

drilling ranging from AX to NQ in diameter. The majority of drilling was conducted

using BQ and NQ core. Most of the early holes were collared inside and angledoutward to encounter the margins of the steeply dipping annular cylinder of

molybdenum mineralization. At the time that Climax ceased operating the Kitsault

mine, the core was stored in a facility near the Kitsault town site. Discussions with

geologists who had access to the KRL town site in 2008 (R. Blair, personal

communication 2008) indicates that all of the core was destroyed in 2007 by KRL.





1 1 . 2 K I T S A U L T – 2 0 0 8 P R O G R A M

In late 2008 Avanti carried out a 33-hole drilling program.

The total completed meterage of the 2008 drill program was 10,130.76 m in 33 drill

holes. The program was completed on November 10, 2008, and all f ive objectives

were accomplished. A plan map showing the traces of all historic drill holes as well

as the 2008 Avanti drill holes is provided in Figure 9.1. All holes were collared at HQ

core size and often were completed using either HQ or NQ core size. All holes were

surveyed for azimuth and dip, at approximately 50-m intervals from the collar to total

depth.

1 1 . 3 B E L L M O L Y

Through 1975, Bell Moly and its predecessor companies (Steininger and Card, 1979)

drilled 36 core holes with an aggregate meterage of 5,462.6 m. CMC leased theproperty in 1975 and implemented drilling in 1976 and 1977. CMC drilled 15 holes

totalling 5,519 m. When CMC terminated its drilling program, the core was stored in

covered shelters on the property. A recent visit to the old field camp location by

Avanti personnel (R. Blair, personal communication, 2008) indicates that a significant

amount of core is still on site but the core racks have collapsed. With careful re-

boxing, much of this core may be recoverable.

1 1 . 4 R O U N D Y C R E E K

Woodcock and Carter (1976) report that 9,300 m of diamond core drilling was

undertaken by several companies on the Roundy Creek property between 1960 and

1969. In the late 1970s, Amax collected the Roundy Creek drill core and stored it

with the Kitsault drill core in a storage shed located adjacent to the Amax office at the

Kitsault town site. The storage shed, along with the drill core samples, were

destroyed by KRL in 2007.

A database was compiled by CMC in 1982 that reconciled collar survey differences

between the different drilling campaigns. This database contains information for 156

drill holes (9,556 m) and was recovered by Avanti from the Mintec office in Tucson.

Additional hard copy maps generated by Amax for the Roundy Creek deposit were

recovered by Avanti in 2008 (D. Davidson, personal communication, 2008; R. Blair,

personal communication, 2008).

1 1 . 5 2 0 0 8 K I T S A U L T D R I L L P R O G R A M O B J E C T I V E S

In July 2008, Avanti Kitsault Mines Ltd. (AKM) carried out a mineral resource study

based on a 3D block model created from the historic Climax digital database





recovered from Mintec in Tucson, AZ. The results of this study were included in the

initial Kitsault Canadian NI 43-101-compliant Report on Resources (SRK, July 17,

2008). On the basis of this study, the decision was taken to make the primary

objective of the 2008 drill program the conversion of the inferred mineralized blocks

within the projected pit, into the indicated category by increasing drill hole density in

key areas. Most of the inferred blocks were located near the bottom of the resourceultimate pit, so that the primary drill targets tended to be near the lower limits of the

known mineralization.

A secondary though equally critical objective of the 2008 drill program was to confirm

assay values in historic drill holes with close-spaced new drilling. Twinning of historic

holes was not possible since former mining of 50 to 80 m, or recent remediation of

the 1982 surface, has removed the collars of all 1960 to 1970s drill holes that were

located in the pit area. Other objectives of the 2008 drill program included:

testing the higher level margins of the proposed pit, especially where drill

hole density was low

testing several areas of deep projected mineralization external to the

interpreted mineralized zones

drilling shallow holes for metallurgical purposes.

Avanti management made the decision that all holes would be drilled parallel with

N-S or E-W cross-sections and at angles to the steeply plunging historic mineralized

zones. With these objectives and criteria, some 33 core holes were proposed. Most

of the holes were targeted at 45-60° angles below the horizontal and planned for

depths ranging from 250 to 450 m. Holes were to be started HQ diameter and

carried to depths of approximately 250 m before reducing to NQ size to complete the

holes. Core sawing rather than mechanical core splitting was selected as thepreferred sampling m ethod.

1 1 . 6 2 0 0 8 K I T S A U L T D R I L L P R O G R A M

Drilling was initiated on September 4, 2008 and concluded on November 10, 2008.

The drill contractor was Driftwood Diamond Drilling Ltd., of Smithers, BC, who used

two drill rigs running on a 24-hour basis. All of the holes were collared within or

immediately adjacent to the historical open pit. Thirty-three diamond drill holes

totalling 10,130.76 m were completed, including the three approximately 25-m holes

drilled for metallurgical testing purposes (Figure 9.1). Core recovery was generally

excellent, averaging approximately 93%, and all holes (except the three short

metallurgical holes) were surveyed down hole every 50 m using a Reflex EZ-SHOT

survey tool. A Reflex ACT core orientation tool was employed in the six holes

selected for geotechnical testing, and was supervised by SRK geotechnical

personnel. After logging the whole core, the Avanti geological team marked the core

into 3 m assay intervals and photographed it. The core was then sawn in half with

one half bagged and sent for sample preparation to the Terrace, BC laboratory and a





subsequent pulp for analysis to the Vancouver, BC, laboratory of ALS Chemex.

Avanti independently prepared refereed molybdenum standards and blank material.

These, along with tags for the intervals selected for re-splitting, were inserted into the

sample stream by Avanti personnel on the Kitsault property site. The 3 m bagged

core samples were submitted to ALS Chemex’s Terrace laboratory, where they were

crushed to 90% -10 mesh, split and pulverized to 80% -200 mesh. The pulpenvelopes were then sent to the ALS Chemex Vancouver laboratory where they

were analyzed for Mo and other elements by means of the +33 element EM-ICP61

method, utilizing a four acid digestion. Details of the sample analysis, standards, and

other QC/AC controls are discussed in a following section of this report. The

remaining half core was stored in the original wooden core boxes and racked in

some thirteen rugged, modular, and weatherproof core storage units at the 2008

camp site.





1 2 . 0 S A M P L I N G M E T H O D A N D A P P R O A C H

There is no reliable information on the methods KEL used to sample and analyze drill

core from the property. Since KEL was part of a major mining company reportedly

operating under industry-acceptable standards, it is assumed that sampling and

analyses met industry standards at the time. This, coupled with the production

reconciliations as conducted during the Amax Feasibility Study, suggests that the

drilling data and assays of that era are reliable. There is no information available on

the preparation and assaying of Bell’s samples. There is no information on how

sampling and analyses were performed at Roundy Creek; therefore, no comment

can be made on the reliability of the reported grades. The sampling and analytical

methods used by CMC at Kitsault are documented and described below.

In addition, Avanti conducted an extensive infill and stepout drilling program in 2008;

the results of this program confirm the results obtained during the historic KEL and

Amax drilling campaigns. This program implemented an extensive quality

assurance/quality control (QA/QC) program which conforms to industry-accepted

practices.

1 2 . 1 F A C T O R S I M P A C T I N G A C C U R A C Y O F R E S U L T S

During the CMC program, in areas where molybdenite paint on fractures or coarse-

grain molybdenite was observed, care was taken to produce an even split of those

intervals. There is also no known geological feature that is preferentially mineralized,

or unmineralized, in the deposits. Based on comparison of field duplicate results at

the time of the historic drilling and as confirmed by the 2008 Avanti drill program,

there is no indication of any bias imparted during the core preparation and splitting

process. Comparisons of the analytical results obtained by CMC from the Kitsault

and Bell Moly deposits and subsequent comparison between these historic data and

the 2008 Avanti program confirm the historic data are both reliable and can be

considered to be representative of mineralization in the Kitsault and Bell Moly

deposits.

1 2 . 2 S A M P L E Q U A L I T Y

Sample preparation during the CMC program was done using a core-splitter on HQ

and NQ core. The entire core hole was split and assayed on typically 3 m intervals,

which is considered an appropriate sample length given the style of mineralization

and mining method. Core recovery was reportedly quite good, and field inspection of

mineralized material in the pit during the site visits confirms overall good rock quality.

Given the generally disseminated nature of mineralization throughout the deposit and





the overall lack of structurally related fracturing and alteration, SRK is of the opinion

that the sample preparation practices during the CMC drilling campaign are

acceptable and have resulted in reliable and representative samples.

Sample preparation utilized during the 2008 Avanti program was consistent with the

earlier CMC program in terms of core preparation and sample length. The resultsfrom the QA/QC program confirm that the data is of high quality. Additionally, these

data confirm the earlier results of both the KEL and CMC drilling programs. SRK is

of the opinion that this confirmation validates the reliability of these historic drill

campaigns, and that the exhaustive data set is acceptable for use in resource

estimation.





1 3 . 0 S A M P L E P R E P A R A T I O N , A N A L Y S E S ,

A N D S E C U R I T Y

1 3 . 1 S A M P L E P R E P A R A T I O N A N D A S S A Y I N G M E T H O D S

13.1.1 H I S T O R I C D R I L L I N G P ROGRAMS

The sample preparation, analytical methods, and laboratory QA/QC measures for the

KEL and Bell core are largely unknown.

The procedures utilized by CMC for sample preparation, analytical methods, andQA/QC are well documented and are described below.

CMC split its core at the splitting facility at Kitsault and Bell Moly, starting at the first

3 m increment within bedrock, through the total depth of the hole. The entire core

hole was split on site using a core splitter, with half retained for geological reference

and the second half sent for assay. CMC bagged each 3 m split half of the core in

canvas bags and shipped the samples to its preparation facility in Smithers, BC.

Once the core samples were processed, the analytical pulps were sealed in

envelopes and not opened again until they reached Skyline Labs, the external

commercial lab in Wheatridge, Colorado. No information is available with regard to

security or the chain of custody during sample preparation and transport.

Each sample was crushed to 100% passing -8 mesh and split in half using a Jones

splitter to produce about 2,000 g sample. The samples were then split in two one

half splits again to produce about 500 g sample. The 500 g sample was then

pulverized to 100% passing -100 mesh, rolled on a mat, and 100 g split-out as an

assay pulp. The assay pulp was shipped to Skyline Lab for analysis.

Gordon Van Sickle, Manager of Skyline Lab at that time (Van Sickle, personal

communication, 2008), reported that the “Molybdenum-Henderson Method” was

used to analyze the pulps. This method was developed by CMC for analysis of its

ore; it involves a three-acid digestion of the pulp followed by the titration of an assay

aliquot and analysis by mass spectrometer. At the time, the laboratories were not

certified by any agency. The Skyline Lab was known in the industry for producing

high-quality analytical results, which was confirmed by the CMC QC/QA program at

Kitsault and Bell Moly (Steininger, personal communication, 2008).

Molybdenum content was reported as percent MoS2, as was the case for all CMC

operating mines assay work (Steininger and Thornton, personal communication,





2008). No information is available with regard to procedures for assaying for

secondary metals

13 .1.2 2 008 AV A N T I D R I L L P R O G R A M

Core was placed in core boxes at the drill site and transported to the core logging

facility by the drill crew where it was systematically logged by the geology staff

almost as soon as it was available. The geologic staff first marked the core for the

3 m assay intervals, recorded the core recovery, rock quality designation (RQD), and

other geotechnical details, and then logged the core for geologic and mineralogical

features. The core was then photographed and transported to the core sawing

facility along with the bags containing the Avanti molybdenum standards and blank

material and the tags for the periodic re-splits carried out on the coarse rejects. The

core sawyers were advised to cut the core in a random manner except where

obvious major veins were present, and then to bisect any such veins. They were

also advised to sample those finely-broken zones that could not be successfully

equally cut with the saw, with a spoon, and to sample any significant gouge zonesthat would wash by sawing, in a similar manner prior to sawing the intact core.

Otherwise, no particular scrutiny that might bias the results is applied to the

alternating halves selected for the assay sample. The core inventory system was

scrupulously maintained. After sawing, the 3 m split sample was bagged and a lab

generated sample ticket was inserted with the sample. Samples were shipped every

day or so as the sample inventory built up.

Samples arrived by truck at the ALS preparation facility in Terrace, BC, in sealed

bags. The chain of custody passed from core sampling on site to the sample

preparation facility in Terrace. Tampering with individual bags or the ties would be

immediately evident when the samples arrived at the sample prep lab. Any

tampering with the larger bags would also be apparent on arrival at the lab.

Documentation was provided so that it would be difficult for a mix up in the samples

to occur either during shipment or at the lab. No breach in the chain of custody has

ever been noted.

All procedures were being carefully attended to and met or exceeded industry

standards for collection, handling, and transport of drill core samples.

S A M P L E PR E P A R A T I O N P R O T O C O L

Samples in plastic bags were dried at 60°C until the desired moisture content was

achieved. The entire sample was crushed to 85% passing 10 mesh (2 mm). Thecrusher was cleaned with high pressure air after every sample. The entire sample

was then run through a Jones or riffle splitter to obtain 500 g. Rejects were retained

and are currently stored at a company warehouse in Terrace, BC.

The 500 g sample was pulverized in a ring-and-puck pulverizer to 95% passing

150 mesh (65 µm). The particle size of the samples was checked by screening





random samples. The pulverizer was cleaned after every sample with high pressure

air.

The 500 g sample was then prepared for assay.

All assaying for samples from Kitsault was provided by ALS Chemex Vancouver, BC,and Acme Laboratories Vancouver, BC. Both ALS and Acme are ISO 9001:2000

certified. All assaying for the 2008 program conducted by ALS utilized a 33 element

four acid ICP-AES method (procedure ME-ICP61). Molybdenum is reported as Mo in

ppm and these data are converted to MoS2 prior to entry in the digital database. The

assay pulps are currently stored at the ALS Chemex facility in Terrace, BC.

1 3 . 2 Q U A L I T Y C O N T R O L A N D Q U A L I T Y A S S U R A N C E

13.2.1 H I S T O R I C D R I L L P R O G R A M S

It is reported that CMC conducted a QA/QC program at Bell Moly and Kitsault that

included insertion of a non-certified standard after every tenth sample and a

duplicate pulp after every fifth sample (Don Davidson, personal communication,

2008). If the assay from the second pulp was not within +10% of the original value,

the entire assay batch was re-analyzed by Skyline and any deviation of a “few

percent” from the accepted value for the standards or duplicates was cause for re-

analysis of the samples in question.

QA/QC was administrated by the Kitsault project geologist. A limited number of pulp

duplicates assays from the CMC Bell Moly drilling program were recently obtained by

Avanti (which are available in Appendix C of the SRK report that was published on

SEDAR, as referenced at the beginning of Section 4.0 in this report). The pulp

duplicate assays confirm the reproducibility of molybdenum assay results for the

Climax drilling at Bell Moly (Figure 13.1). The results of the duplicate assay samples

appear reasonable; however, conclusions cannot be drawn regarding assay quality

at the Kitsault property based on data from the Bell Moly area.





Figure 13.1 Scatterplot of the Bell Moly Duplicate Assay Data

Dr. Steininger was responsible for the QC/QA program during the Climax drilling

program and the protocols and procedures were conveyed verbally to SRK

personnel.

13 .2.2 2 008 AV A N T I D R I L L P R O G R A M

Q C S A M P L E I N S E R T I O N

The sampling staff inserted molybdenum standards every 20th sample. There was

every indication that the procedure was being strictly followed and QC sample

coverage was adequate for the drilling. These standards were prepared by CDN

Laboratories in Delta, BC.

Blank material was inserted at the rate of one in every 40-45 samples. The blank

material was silica sand.

A summary of the performance of the three standards implemented by Avanti are

provided in Figure 13.2 through to Figure 13.4. Three values fell outside of control

limits for Standard AV2. Follow up of the out of control standards with a set of pulp

duplicate assays at Acme Laboratories in Vancouver, BC, showed the ALS Chemex

assays returned consistently higher values. Standards submitted with the Acme

samples returned values well within control limits. Based on these results, all the

samples in hole K08-30 (the hole affected by the out of control AV2 standard assays)

with an original value greater than 300 ppm molybdenum were re-submitted to ALS

Chemex for assay. The difference between original and duplicate assays for hole





K08-30 are shown in the scatter plot below. The duplicate values are consistent with

results from Acme and are significantly lower than the originals. Subsequently, the

original molybdenum values for hole K08-30 in the assay database were replaced by

the duplicate assay values.

Standard AV3 had one out of control result where the value was substantially lower than the lower control limit. In addition, other elements assayed in the sample did

not conform to typical results from the standard material. After review, it was

concluded that a sample had been mislabelled, and there was no problem indicated

in the assay process. Duplicate assays submitted in conjunction with the other AV3

results that exceeded the upper control limit all returned values within 10% of the

original assays. The results suggested the assay process was under control during

the time these samples were undergoing assay. The out of control rate for the AV3

standard was less than 10% of the samples submitted, and it was concluded the AV3

standard results demonstrated assay process control.

Figure 13.2 CDN Standard AV1 Control Chart

604530150

Sequence Number

720

680

640

600

560

520

480

M o l y b d e n u m

( p p m )

AVG = 601.4833

LCL = 510

UCL = 670

Accepted value = 590






806040200

Sequence Number

1400

1300

1200

1100

1000

900

M o l y b d e n u m

( p p m )

AVG = 1186.

LCL = 1000

UCL = 1300



50403020100

Sequence Number

2200

2000

1800

1600

1400

1200

M o l y b d e n u m

( p p m )

AVG = 1792.8125

LCL = 1584

UCL = 1936






S A M P L E B L A N K P E R F O R M A N C E

No molybdenum values in the blank material are exceed the control limit. A graph

showing blank material performance is provided in Figure 13.5.

Figure 13.5 Performance of Kitsault Blank Material

100806040200

Sequence Number

20

16

12

8

4

0

M

o l y b d e n u m

( p p m )

AVG = 2.8462

UCL = 15

C O A R S E DU P L I C A T E SA M P L E P E R F O R M A N C E

Coarse duplicate samples check the adequacy of the sample preparation protocol.

The charts provided in Figure 13.6 and Figure 13.7 show that the pairs fall within

control limits at above the prescribed rate (i.e. 90% within ±30%). This indicates the

sample preparation process is under control and producing samples that are

sufficiently homogeneous to form the basis for resource estimation.





Figure 13.6 Correspondence of Molybdenum in Kitsault Coarse Reject

Duplicates

40003000200010000

Original Molybdenum

0.4

0.2

0.0

-0.2

-0.4

R e l a t i v e D i f f e r e n c e

+30%

-30%

Figure 13.7 Correspondence of Silver in Kitsault Coarse Reject Duplicates

201612840

Original Silver

1.2

0.8

0.4

0.0

-0.4

-0.8

-1.2

R e l a t i v e D i f f e r e n c e

+30%

-30%





P U L P D U P L I C A T E SA M P L E P E R F O R M A N C E

Except for the duplicate results from hole K08-30 lab check pulp duplicates fell within

±10% of the original assay well over ninety percent of the time (Figure 13.8). Pulp

duplicates from hole K08-30 identified an error in the original assays from that hole

and all samples with a significant value were submitted for a second set of assays.Results from this second round of assaying were consistent with the check duplicate

results.

Figure 13.8 Performance of Check Assays from Pulp Duplicates

40003000200010000

Original Mo (ppm)

4000

3000

2000

1000

0

R e - a s s a y M o ( p p m )

Original = Re-assay

-10%

1 3 . 3 I N T E R P R E T A T I O N

13.3.1 H I S T O R I C ASSAY D A TA

The QA/QC programs conducted by CMC as described by CMC personnel

responsible for these programs met or exceeded industry standards at the time

(R. Steininger, personal communication, 2008). Assays from the CMC programproduced results that duplicated those for nearby KEL holes at the Bell Moly

property, indicating that the latter assays were also reliable. However, conclusions

drawn from this comparison cannot be directly applied to the Kitsault assay data,

which is the subject of this resource report. Given the detailed descriptions of the

processes and procedures implemented and the results obtained as provided by

CMC personnel, combined with the close agreement between the 2008 Avanti data





and these historic data, SRK is of the opinion that the data resulting from these

historic drill programs are reliable and suitable for use in resource estimation.

13 .3.2 2 008 K I T S A U L T AS S A Y D A T A

The QA/QC programs implemented by Avanti were executed and monitored in an

industry best practice manner. Results of this program were reviewed in detail by

SRK, and SRK is of the opinion that these data are reliable and suitable for use in

resource estimation.





1 4 . 0 D A T A V E R I F I C A T I O N

1 4 . 1 H I S T O R I C D R I L L P R O G R A M S

As part of the CMC drilling programs, several holes were located sufficiently close to

holes drilled by KEL to confirm the quality of the KEL assays. Specifically, the CMC

cluster of holes 7404-7406 and 7411-7413 were drilled in close approximation to

existing KEL holes, producing similar grade intervals. Other CMC holes were drilled

as in-fills to the KEL pattern and confirmed the sample and distribution of the nearby

KEL holes (Steininger, personal communication, 2008). These data were sufficient

to support CMS’s conclusion that the KEL assay results were reliable.

To further confirm the KEL drillhole data, Avanti conducted additional analysis to

assess the reliability of the pre-CMC drilling data. The drilling results prior to the

purchase of the Kitsault mine by CMC were compared to post-purchase drilling

results using the following procedure:

Nearest neighbour assignments of drillhole composite sample grades were

made to blocks within the resource model within 50 m of a sample.

Blocks having an assignment from both pre- and post-CMC drillholes were

extracted.

The grade distributions from the two sets of drilling were compared by

means of a quantile-quantile (QQ) plot (Figure 14.1).

The QQ plot shows that the grade distributions from the two sets of drilling agree

reasonably well. The correspondence between the distributions indicates the drilling

from both operators is comparable and can be used together when estimating

resources.





Figure 14.1 QQ Plot

SRK also conducted nearest-neighbour analysis, by pairing the historic assay data

with those from the 2008 Avanti program. This analysis showed remarkably close

agreement between the historic data and the 2008 Avanti data out to separation

distances of 200 m; no bias was detected between the two data sets.

It should be noted that BC Moly was in production for several years before being sold

to CMC. The exploration model used by Kennecott was prepared by Kennecott’s

Bear Creek Mining. During its Feasibility Study, Amax conducted a reconciliation of

their resource model to the Kennecott blast hole model of their production, and foundthat they compared well. The results of this comparison are provided in Table 14.1.

Table 14.1 shows how the block model based on KEL blast hole data compares with

the CMC exploration resource model. The model comparison was conducted over a

vertical extent of 325 ft, representing 10 mined benches.

Table 14.1 CMC Resource Model vs. Blast Hole Model Comparison*

Cut-off

Grade

MoS2

(%)

Blast Hole Model Exploration Model

% Difference

Exploration – Blast Hole

Ore

(k ton)

MoS2

(%)

MoS2

(kst)

Ore

(kst)

MoS2

(%)

MoS2

(kst)

Ore

(kst)

MoS2

(kst)

0.01 21,167 0.167 35,349 20,326 0.168 34,148 -3.97 -3.40

0.13 13,790 0.212 29,235 15,960 0.188 30,005 +15.74 +2.57

0.15 12,053 0.222 26,758 13,099 0.198 25,936 +8.68 -3.07

0.16 11,143 0.228 25,406 11,416 0.204 23,289 +2.46 -9.09

0.20 7,474 0.252 18,784 5,349 0.233 12,463 -28.24 -33.65

* Taken from Kitsault Feasibility Study, Volume 2 (CMC, 1978).





As none of the original documentation exists, with the exception of the original

resources and reserve document, SRK has not been able to undertake any

independent verification of the original data for each property except as presented in

electronic data format. The only verification of the Kitsault data is that reported

molybdenum production is in alignment with the historic exploration block model, and

consistent with both tonnage and grade therein. However, as the historic blast hole,resource models, and pre-Amax production records cannot be located, SRK is

unable to confirm the results of this historic comparison.

1 4 . 2 2 0 0 8 A V A N T I D R I L L P R O G R A M S

SRK verified approximately 10% of the database for molybdenum and silver by

comparison of the original lab certificates from ALS Chemex with the digital database

as provided by Avanti, under the direction of the QP. This exercise identified two

data entry errors for molybdenum values out of the random 332 analyses checked by

SRK. These errors were related to precision/rounding, and equate to an error rate of approximately 0.06%. SRK is of the opinion that this error rate is very low, and is not

material with respect the overall integrity of the database. SRK also notes that

molybdenum values below detection (1 ppm Mo) were set to half detection limit

(0.5 ppm), which is normal and customary procedure. However, SRK notes that

silver values below detection limit (0.5 ppm) were set to zero, which is a conservative

approach, and has no material impact on the outcome of grade estimation. Based

on this database verification, SRK is of the opinion that data entry of the original lab

certificates was conducted in an accurate and professional manner.

1 4 . 3 L I M I T A T I O N S

SRK was unable to independently verify the collar, survey and assay information

provided by Avanti for the historical drilling campaigns, due to the fact that none of

the original documentation can be located. It is SRK’s practice to verify a portion of

the database through comparisons of original drilling and assay documents with the

data provided digitally in a database. The lack of any physical documentation for any

of the information normally verified in a resource audit would normally preclude the

possibility of assigning a reasonable level of confidence with respect to a resource

model based upon these data. However, the verifiable production history during two

distinct periods of mining activity and the results of historic mine-model

reconciliations provide the authors with a high degree of comfort with respect to

drillhole assay reliability and accuracy. Security of the original database by CMCpersonnel is well documented, and SRK is of the opinion that the chain of custody for

these data has been maintained. The archived version of this original database has

been provided to SRK by Avanti in what is believed to be its original form. The

testimonials of former CMC employees who supervised data collection and

compilation, the close agreement between the historic resource estimates and the

estimate that forms the basis for this report, and the documented historic production





reconciliations provide SRK with a reasonable confidence level in both the quality of

data and compliance to what are now industry accepted practices.

Statistical and visual comparisons between close spaced historic drilling data and the

2008 Avanti drill data show remarkably close agreement, and the Avanti drilling

adequately confirms the results derived from previous drilling programs. Based onthis assessment, SRK is of the opinion that the consolidated database is reliable and

suitable for use in resource estimation.





1 5 . 0 A D J A C E N T P R O P E R T I E S

There was no additional information included from any adjacent properties used in

the Kitsault property resource estimation.





1 6 . 0 M I N E R A L R E S O U R C E A N D M I N E R A L

R E S E R V E E S T I M A T E

1 6 . 1 M I N E R A L R E S O U R C E S

The mineral resource estimate was prepared by Robert Sim, P.Geo, independent

consultant, under the direction of Bruce Davis, FAusIMM, Avanti’s Manager of Business Development. Grade estimations were conducted using 3D block models

based on geostatistical applications using commercial mine planning software

(MineSight® v4.50). The project limits were based in the UTM coordinate systemusing a nominal block size of 10 x 10 x 10 m. Molybdenum mineralization at Kitsault

occurs in an elliptical pattern related to the contacts of granodiorite and dioritic rocksthat have intruded the host sedimentary rocks. Drillholes tend to be oriented in a

semi-radial distribution in an attempt to intersect the mineralized zone approximately

perpendicular to the intrusive contacts. Drilling tends to be at an average spacing of somewhat less than 60 m throughout the deposit (Kennco and Amex drilled onapproximate 61 m (200 ft) sections in a NW and EW directions; Avanti drillholes in

2008 tended to fill in gaps in that grid ). The majority of drilling occurred during threeseparate periods: prior to mine production in the late 1960s, 1974 to 1978, and again

in 1981 to 1982. Avanti drilled 30 holes to further delineate grade in 2008.

The resource estimate was generated from drillhole sample assay results and the

constrained by the interpolation of a grade probability model which relates to thespatial distribution of molybdenum in the deposit. Interpolation characteristics were

defined based on the geology, drillhole spacing, and geostatistical analysis of thedata. The resources were classified by their proximity to the sample locations. SRKhas conducted an audit of the resource model and finds it to be acceptable for

resource reporting under “CIM Estimation of Mineral Resources and Mineral

Reserves Best Practice Guidelines” (2005).

16.1.1 D R I L L H O L E D A T AB A S E

Drilling activities took place in the 1960s, late 1970s and early 1980s; they were

conducted in imperial units with locations in a local mine grid. All information wasconverted into metric units in the UTM NAD 83 coordinate system. The conversion

from mine grid (metric) to UTM is listed in Table 16.1. Drillholes completed by Avantiin 2008 were located with respect to the UTM coordinate system.





Table 16.1 Grid to UTM Conversion

Direction Mine Grid UTM

X (easting) 0E 459,806.06E

Y (northing) 0N 6,130,750.06N

Z (elevation) 0m 0m

There are a total of 169 drillholes in the Kitsault database with a cumulative length of

33,110.58 m. The 169 holes include 30 holes drilled by Avanti in 2008 (5,459.53 m). Avanti also drilled an additional three holes for metallurgical purposes (the

metallurgical holes were not used in the development of the resource model).Individual holes range from 3-763 m in length and average 196 m in length. The

drilling is located on a series of north-south and east-west oriented cross sections

spaced at approximately 60 m intervals. Drillhole spacing averages somewhat lessthan 60 m throughout most of the deposit area. It is common that several drillholes

are fanned out from a single platform at varying orientations. The distribution of drillholes is shown in plan in Figure 16.1.

There are a total of 12,169 samples in the database of which 8,845 are pre-Avantiand 3,324 of the samples were generated during Avanti’s 2008 drilling program. Pre-

Avanti intervals were analyzed for molybdenum sulphide (MoS2) and values were

converted to Mo% for use in resource modeling (Mo%=MoS2*0.6). Avanti drillsamples were assayed for Mo% directly. Mo% values range from a minimum of 0%

Mo to a maximum of 2.922% Mo. There are 8,573 intervals analyzed for Pb% (0-1.11%), 4,519 intervals tested for Cu% (0-0.39%Cu), 7,016 intervals tested for Ag g/t(0-566 g/t) and 4,519 intervals tested for Fe% (0.1-10.7%). The main potential

economic contributor is molybdenum; however, additional elements were evaluated

and estimated in the model for information purposes.

In addition to the elements listed above, the Avanti drillholes also contain analyses

for S%, Cd, Ni and Zn. Although the sample density for these items is relatively low,

estimates were made in the resource model in order to provide additional information

for metallurgical and environmental purposes.

Individual sample intervals range from a minimum of 0.05 m to a maximum of

27.43 m with an average of 3.06 m.

Some information is available regarding drilling recoveries in the pre-Avanti drilling.

Anecdotal information indicates very good core recovery and no relationship between

core recovery and molybdenum. Core recovery from the Avanti drilling averagedover 97%.

The geologic information is derived primarily through observations during logging andincludes rock type designation.





Figure 16.1 Drillhole Plan





16.1.2 E X P LO R A T OR Y D A TA AN A L Y S I S

Exploratory date analysis (EDA) involves the statistical summarization of the

database in order to quantify the characteristics of the data. One of the mainpurposes of this exercise is to determine if there is evidence of spatial distinctions in

grade which may require the separation and isolation of domains during interpolation.The application of separate domains prevents unwanted mixing of data duringinterpolation and the resulting grade model will better reflect the unique properties of the deposit. However, applying domain boundaries in areas where the data is not

statistically unique may impose a bias in the distribution of grades in the model.

A domain boundary, which segregates the data during interpolation, is typicallyapplied if the average grade in one domain is significantly different from that of

another domain. A boundary may also be applied where there is evidence that thereis a significant change in the grade distribution across the contact.

B A S I C S T A T I S T I C S B Y D O M A I N

The basic statistics for the distribution of molybdenum were generated by rock type

with the results presented using a boxplot in Figure 16.2. There are similar

molybdenum contents in all rock types except the lamprophyre dikes which areslightly lower grade.

Figure 16.2 Kitsault Rock Types

Note: refer to Table 16.2 for a complete list of rock types.

C O N T A C T P R O F I L E S

The nature of grade trends between two rock type domains is evaluated using thecontact profile which graphically displays the average grades at increasing distances

from the contact boundary. Contact profiles which show a marked difference in





grade across a domain boundary, are an indication that the two data sets should beisolated during interpolation. Conversely, if there is a more gradual change in grade

across a contact, the introduction of a “hard” boundary (i.e. segregation duringinterpolation) may result in much different trends in the grade model – in this case

the change in grade between domains in the model is often more abrupt than the

trends seen in the raw data. Finally, a flat contact profile indicates no grade changesacross the boundary. In the case of a flat profile, “hard” or “soft” domain boundarieswill produce similar results in the model.

Contact profiles were generated to evaluate the change in molybdenum gradebetween the main rock types. The results, in Figure 16.3 through Figure 16.8, showeither no change in grade or very transitional changes in molybdenum grades

between rock types.

Figure 16.3 Contact Profile for Kitsault Molybdenum Major 4-NE Porphyry

Contact





Figure 16.4 Contact Profile for Kitsault Molybdenum Diorite-Granodiorite

Contact





Figure 16.6 Contact Profile for Kitsault Molybdenum Hornfels-Granodiorite

Contact





Figure 16.7 Contact Profile for Kitsault Molybdenum Hornfels-Diorite Contact





Figure 16.8 Contact Profile for Kitsault Molybdenum Major 3-Alaskite Contact

C O N C L U S I O N S A N D M O D E L L I N G IM P L I C A T I O N S

The results of the EDA indicate that there are no distinct domains that define unique

grade frequency distributions of molybdenum based on the rock types other than the

lamprophyre dike. The dikes represent an insignificant (~1%) volume of drilling in thedatabase.

S E C O N D A R Y E L E M E N T S

The statistical properties of the minor or secondary elements (lead, silver, copper,

iron, and tungsten) were reviewed by rock type codes and no distinct differences





were identified. None of the other elements used for metallurgical or environmentalpurposes showed any differences by rock type.

SRK conducted an independent statistical analysis of the data provided by Avanti

and concurs with the conclusions reached by Avanti modelling personnel.

H I S T O G R A M S F O R M O L Y B D E N U M

Since the decision was made to use assay results without regard to rock type,histograms for the molybdenum grade in the global dataset for both raw drillholesamples and composite samples are given in Figure 16.9.

Figure 16.9 Histogram for Molybdenum Grade

16.1.3 T O P O G R A P H Y

A new Lidar survey was conducted during the 2008 field season, and was used to

constrain the current resource estimate. SRK conducted visual comparisonsbetween the 2008 drillhole collars and the 2008 topography, and there is goodagreement between the two. Mining was conducted at Kitsault during the 1970s and

1980s, and the 2008 topographic surface reasonably reflects the post-mining surface

observed at site. SRK considers that the topography provided digitally by Avanti isaccurate and appropriate for use in resource estimation.





Figure 16.10 Drillholes – Topographical Surface with 5 m Contour Lines

16.1.4 G EOLOGY

The molybdenum mineralization in the Kitsault deposit is related to the contact zonebetween granodiorite and diorite rock which have intruded into a host rock comprisedof metamorphosed greywacke and argillite referred to as hornfels. Disseminated and

stockwork molybdenum occurs in a diffuse zone along the contact area. There are

several intrusive Alaskite dikes which are also mineralized. Also present are rare

post-mineral lamprophyre dikes. There is up to 10 m of overburden logged in somedrillholes, but essentially all of this material was removed during previous miningactivities.

The original pre-Avanti lithology database is comprised of approximately 85 differentrock codes. These were summarized, based on historical information into six basicrock types listed in Table 16.2. Avanti has retained the basic rock types listed in

Table 16.2 while logging core from their 2008 drilling program. Rock types 1, 2, and3 comprise approximately 90% of the volume of the drillhole intervals in thedatabase. Alaskite dikes represent almost 6% of the data. The remaining 4% of the

drilling is comprised of overburden and weak/unmineralized lamprophyre dikes.





Table 16.2 Simplified Rock Types

Domain Rock Code Comments

Hornfels 1 Pre-intrusive host rocks.

Diorite 2

Granodiorite 3 Main intrusive phase responsible for mineralization.

Alaskite 4 Minor series of dikes. Mineralized.

Intramineral Porphyry 5

Northeast Porphyry 6 Quartz Monzonite Porphyry (post mineralization)

Lamprophyre Dike 7

Overburden 8 Surface soil and gravels.

Fault 9

Mafic Dyke 10 Non-lamprophyre

The distribution of the three main rock types (hornfels, diorite, and granodiorite) was

interpreted on east-west cross sections and linked together to form three dimensionalwireframe models. The distribution of the intrusive phases is shown in Figure 16.11.

These wireframes were used to back-flag model blocks for the assignment of density, and were not utilized in the grade estimation process. SRK has reviewedthese wireframes using visual comparisons in section and plan, and is of the opinion

that they are adequate for assignment of rock densities.

Figure 16.11 Distribution of Intrusive Phases





16.1.5 C O M P O S I T I N G

Compositing of drillhole samples is carried out in order to standardize the database

for further statistical evaluation. This step eliminates any effect related to the samplelength which may exist in the data.

In order to retain the original characteristics of the underlying data, a composite

length was selected which reflects the average original sample length. Thegeneration of longer composites results in some degree of smoothing which could

mask certain features of the data. The mean sample length is 3.1 m and 79% of the

intervals are exactly 3 m in length (actually the most common sample intervals is10 ft, or 3.05 m in length). A standard 3 m composite sample length was generatedfor statistical evaluation and for use in grade estimations in the block model.

Drillhole composites were generated in 3 m downhole lengths, meaning thatcomposites begin at the top of each hole and are generated at 3 m intervals downthe length of the hole. Several holes were randomly selected and the composited

values were checked for accuracy. No errors were found. SRK generated anindependent composite database, and statistical comparisons with the Avanti

composite database show close agreement.

16.1.6 E V A L U A T I O N O F OU T L I E R G R A D E S

Histograms and probability plots of the distribution of modelled elements were

reviewed in order to identify the existence of anomalous outlier grades in thecomposite database. In addition, a decile analysis of the data was also conducted inorder to quantify the distribution of contained metal with respect to the sample

density.

Review of the physical location of these samples show that they tend to occur in asomewhat random pattern throughput the mineralized area and their effects are

limited by the relatively dense drillhole distribution throughout the deposit. Ultimately,

it was decided that the effects of these higher-grade intervals would be limited duringblock grade interpolation using an “outlier limitation”. Outlier samples above a

defined threshold are limited to a maximum distance of influence of 20 m duringblock grade interpolation. The various limitation thresholds and resulting effects aresummarized in Table 16.3.





Table 16.3 Summary of Outlier Limitations

Element Grade Threshold Comments

Molybdenum 0.40% Inside shell Affects 27 comps inside and 11 outside prob shell;overall 1.7% reduction in contained Mo metal in model0.25% Outside shell

WO3 450 ppm 40 comps affected; overall 3.7% reduction in WO3 inmodel

Cu% 0.15% 6 comps affected; -14.3%Cu metal in model

Pb% 0.50% 22 comps affected; -12.0%Pb m etal in model

Fe% n/a Outlier limit not required

Ag g/t 100 g/t 12 comps affected; -9.2% Ag metal in model

Note: metal reduction calculated in measured and indicated class resources.

SRK conducted independent statistical analysis using the raw assay data providedby Avanti, and is in general agreement with the threshold values determined by

Avanti.

16.1.7 SP E C I F I C G R A V I T Y

Specific gravity (SG) determinations were made using the wax seal method and drillcore from the 2008 Avanti drilling. SRK conducted measurements on 13 Quartz

monzonite samples, 1 Diorite sample, and 21 Hornfels samples. RDI Metallurgical

Testing (RDI) measured SG on 55 Quartz monzonite samples, 25 Diorite samples,and 18 Hornfels samples. Based on these measurements SG was assigned by rocktype in model. Assignments were as follows:

Hornfels: 2.70t/m3

Diorite: 2.66t/m3

Quartz Monzonite: 2.63t/m3.

16.1.8 V A R I O G RA M AN A L Y S I S A N D M O D E L L I N G

The degree of spatial variability in a mineral deposit depends on both the distanceand direction between points of comparison. Typically, the variability between

samples increases as the distance between samples also increases. If the degree of

variability is related to the direction of comparison, then the deposit is said to exhibitanisotropic tendencies which can be summarized with the search ellipse. The semi-variogram is a common function used to measure the spatial variability within a

deposit.

The components of the variogram include the nugget, the sill, and the range. Oftensamples compared over very short distances (even samples compared from the

same location) show some degree of variability. As a result, the curve of the

variogram often begins at some point on the Y-axis above the origin – this point iscalled the “nugget”. The nugget is a measure of not only the natural variability of the





data over very short distances but also a measure of the variability which can beintroduced due to errors during sample collection, preparation and assaying.

The amount of variability between samples typically increases as the distance

between the samples becomes greater. Eventually, the degree of variability between

samples reaches a constant, maximum value. This is called the “sill” and thedistance between samples at which this occurs is referred to as the “range”.

The spatial evaluation of the data in this report was conducted using a correlogramrather than the traditional variogram. The correlogram is normalized to the variance

of the data and is less sensitive to outlier values, generally giving better results.

Variograms were generated using the commercial software package Sage 2001©developed by Isaacs & Co. Multidirectional variograms were generated for

composited molybdenum sample data. Note that one variogram was produced fromall of the data to be used both inside and outside of the probability shell during grade

interpolation. This approach is an attempt to retain some degree of continuity of

grade across the boundary when it (locally) exists. The results are summarized inTable 16.4 through Table 16.10.

Table 16.4 Variogram Parameters – Molybdenum, Inside Shell

Zone

Variance Components 1st Structure 2nd Structure

Nugget S1 S2 Range (m) AZ Dip Range (m) AZ Dip

Inside Shell 0.300 0.323 0.377 53 4 2 1,164 310 0

Spherical 33 274 1 275 218 84

11 169 88 100 40 6

Note: correlograms conducted on 3 m DH composite data.

Table 16.5 Variogram Parameters – Molybdenum, Outside Shell

Zone



Outside Shell 0.257 0.257 0.486 208 205 60 3,693 179 83

Spherical 37 106 5 400 178 -7

12 14 29 115 88 0


Table 16.6 Variogram Parameters – WO3

Zone



All Data 0.174 0.534 0.292 155 60 1 962 55 -68

Exponential Practical Range 50 318 85 379 21 -16

9 150 4 209 106 15






Table 16.7 Variogram Parameters – Copper

Zone



All Data 0.676 0.036 0.288 51 4 -1 2,620 126 72

Exponential Practical Range 43 274 -4 1118 1 10

18 287 86 288 88 -14


Table 16.8 Variogram Parameters – Lead

Zone



All Data 0.337 0.584 0.078 40 122 13 1,575 175 74

Exponential Practical Range 21 335 75 604 47 10

4 214 8 308 135 -12


Table 16.9 Variogram Parameters – Iron

Zone



All Data 0.256 0.277 0.466 323 23 -19 3,794 99 60

Exponential Practical Range 302 313 45 886 173 -9

82 98 39 165 78 -28


Table 16.10 Variogram Parameters – Silver

Zone



All Data 0.400 0.469 0.131 69 275 -34 1,312 272 51

Exponential Practical Range 38 360 8 456 88 38

7 258 54 377 359 -2


SRK conducted independent variogram analysis on the provided data set and was

able to closely replicate the results produced by Avanti.

16.1.9 P R O B A B I L I T Y SH E L L

In the absence of a domain boundary or drillholes which sufficiently limit the lateral

extent of the mineralization in the deposit, a probability shell was generated which

represents the area in which molybdenum mineralization is likely to occur. Figure16.12 shows a cumulative log-probability plot of the molybdenum sample data. The





inflection at a grade of 0.05% Mo occurs at the 40th percentile of the distribution andindicates a reasonable point to divide the distribution into less well mineralized and

more strongly mineralized portions. Indicator values were assigned at this gradethreshold (value of “1” assigned to composites >0.05% Mo and “0” to intervals

<0.05% Mo) and an indicator variogram was generated from the indicator values.

Ordinary kriging was used to estimate probability values in blocks. The results werecompared to the original Mo% grades in drilling and it was decided that a 50%probability threshold formed a reasonable division of the data (i.e. there is a >50%

chance that blocks within the shell will exceed 0.05% Mo). The resultant probabilityshell was used as a constraint during grade estimation.

SRK has reviewed the methodology utilized to construct the probability shell and has

inspected it graphically in plan and section. SRK is of the opinion that it provides areasonable boundary to control grade estimation.

Figure 16.12 Cumulative Probability Plot

Due to a lack of drilling at depth, the probability shell was limited to a maximum depthof 250 m. The extent of the probability shell is shown in Figure 16.13.





Figure 16.13 Probability Shell





Composited drillhole samples and blocks in the model were assigned unique codevalues representing inside and outside of the probability shell. These are then

matched during composite retrieval for block grade interpolation.

Search orientation and ranges are based on a 0.05% Mo indicator variogram, with

the variogram parameters provided in Table 16.11. Probabilities were estimatedusing ordinary kriging (OK) using search distances, search orientation, andcomposite selection criteria as provided in Table 16.12.

Table 16.11 0.05% Molybdenum Indicator Variogram Parameters

Zone



All Data 0.150 0.128 0.722 102 32 4 518 140 43

Spherical 70 302 0 272 304 46

35 212 86 128 42 8

Note: indicator variograms conducted on 3 m DH composite data.

Table 16.12 Indicator Interpolation Parameters

Interpolation

Domain

Search Ellipse Range (m) No. of Composites

Other X Y Z

Min/

Block

Max/

Block

Max/

Hole

In/Out >0.05% MoProbability Shell

300 300 300 13 24 6 6 DH/octant

1 6. 1. 10 BL O C K M O D E L L I M I T S

A block model was initialized in MineSight® and the dimensions are provided in

Table 16.13. The selection of a nominal block size measuring 10 x 10 x 10 m is

considered appropriate with respect to the current drillhole spacing as well as theselective mining unit (SMU) size typical of an operation of this type and scale.

Table 16.13 Block Model Limits

Direction Minimum Maximum

Block

Size (m)

No. of

Blocks

East 472,300 474,300 10 200

North 6,140,900 6,142,800 10 190

Elevation -200 1000 10 120

Blocks in the model were coded on a majority basis with the probability shell domain.

During this stage, blocks along a domain boundary are coded if >50% of the blockoccurs within the boundaries of that domain.





The proportion of blocks which occur below the topographic surfaces are alsocalculated and stored within the model as individual percentage items. These values

are utilized as a weighting factor in determining the in-situ resources for the deposit.

1 6. 1. 11 G R A D E E S T I M A T I O N

The block model grades for molybdenum were estimated using OK. The results of

the OK estimation were compared with the Hermitian (Herco) polynomial change of support model (also referred to as the Discrete Gaussian correction). This method is

described in more detail in Section 16.1.12.

The Kitsault OK model was generated with a relatively limited number of samples inorder to match the change of support or Herco grade distribution. This approach

reduces the amount of smoothing (averaging) in the model and, while there may be

some uncertainty on a localized scale, this approach produces reliable estimations of the recoverable grade and tonnage for the overall deposit.

All grade estimations use length weighted composite drillhole sample data. Sample

data is not mixed across the probability shell boundary and the variogramparameters listed in Table 16.4 through Table 16.10 are used both inside and outsidethe probability shell. The interpolation parameters are summarized by domain in the

Table 16.14.

Table 16.14 Interpolation Parameters

Interpolation Domain

Search Ellipse Range (m) No. of Composites

Other X Y Z

Min/

Block

Max/

Block

Max/

Hole

Mo% in >0.05% MoProbability Shell

300 300 100 6 15 5 1 DH/quadrant

Mo% Out Probability

Shell

300 300 100 6 20 5 1 DH/quadrant

WO3 In/Out Shell 300 300 100 6 20 5 1 DH/quadrant

Cu, Pb, Fe, Ag 300 300 100 6 20 5 1 DH/quadrant

Note: probability shell used as hard boundary for Mo and WO3 models only.

Estimates for S, Cd, Ni, Zn, AP were made using ID2; a 300 x 300 x 200 m search;

5 comps per holes; min5/max15 per block; 1DH per quadrant. No shell was used to

constrain the interpolation. NP uses 300 x 300 x 200 m search and max 3

comps/holes, min3/max9 per block, with no quadrant search limit.

The grade of all secondary elements (Cu, Pb, W, and Ag) was estimated using the

inverse distance weighting (ID) estimation method. The probability shell was notused during the estimation of these elements.





1 6. 1. 12 M O D E L V A LI D A TI O N

The results of the modeling process were validated through several methods. These

include a thorough visual review of the model grades in relation to the underlyingdrillhole sample grades, comparisons with the change of support model,

comparisons with other estimation methods, and grade distribution comparisonsusing swath plots.

V I S U A L I N S P E C T I O N

Detailed visual inspection of the block model was conducted in both section and plan

to ensure the desired results following interpolation. This includes confirmation of theproper coding of blocks within the respective domains and below the topographicsurface. The distribution of block grades were also compared relative to the drillhole

samples in order to ensure the proper representation in the model. An example level

plan and cross-section through the block model showing block grades, resource

classification, composite data, and resource pit outline are provided in Figure 16.14and Figure 16.15, respectively.





Figure 16.14 Plan 450 m Elevation





Figure 16.15 Section 6141850N, East-West Vertical Cross Sections





M O D E L C H E C K S F O R C H A N G E O F S U P P O R T

The relative degree of smoothing in the block model estimates were evaluated using

the Discrete Gaussian or Hermitian Polynomial Change of Support method

(described by Journel and Huijbregts, Mining Geostatistics, 1978). With this method,the distribution of the hypothetical block grades can be directly compared to the

estimated (OK) model through the use of pseudo-grade/tonnage curves. Adjustments are made to the block model interpolation parameters until anacceptable match is made with the Herco distribution. In general, the estimated

model should be slightly higher in tonnage and slightly lower in grade when

compared to the Herco distribution at the projected cut-off grade. These differencesaccount for selectivity and other potential ore-handling issues which commonly occur

during mining.

The Herco (Hermitian correction) distribution is derived from the declusteredcomposite grades which were adjusted to account for the change in support as one

goes from smaller drillhole composite samples to the large blocks in the model. Thetransformation results in a less skewed distribution but with the same mean as theoriginal declustered samples.

The distribution for the OK and ID models, shown in Figure 16.16, show a desireddegree of correlation with the Herco results. Note the “bump” in the OK and ID

models at 0.05% Mo corresponding to the threshold grade selected for the

generation of the probability shell. Although this is considered an artificial artifact, itis felt that the use of the probability shell helps control the outer limits of the deposit

where drillholes were terminated somewhat prematurely (some drillholes werestopped less than 5 m after exiting the mineralized zone).





Figure 16.16 Recovered Molybdenum 10 x 10 x 10 m SMU – 2009 Model Inside

Shell

C O M P A R I S O N O F IN T E R P O L A T I O N M E T H O D S

For comparison purposes, additional molybdenum models were generated using

both the ID and nearest neighbour (NN) interpolation methods (the NN model was

made using data composited to 10 m intervals). The results of these models arecompared to the OK models at a series of cut-off grades in the grade/tonnage graphin Figure 16.17. Note that this comparison is limited to blocks classified in the

Indicated category and above the 250 m elevation. Overall, there is very good

correlation between these models. Reproduction of the model using differentmethods tends to increase the confidence in the overall resource.





Figure 16.17 Grade/Tonnage Graph

1 6. 1. 13 SW A T H P LOTS (D R I F T AN A L Y S I S )

A swath plot is a graphical display of the grade distribution derived from a series of bands, or swaths, generated in several directions through the deposit. Gradevariations from the OK model are compared using the swath plot to the distribution

derived from the declustered (NN) grade model.

On a local scale, the NN model does not provide reliable estimations of grade but, ona much large scale, it represents an unbiased estimation of the grade distribution

based on the underlying data. Therefore, if the OK model is unbiased, the grade

trends may show local fluctuations on a swath plot but the overall trend should besimilar to the NN distribution of grade.

Swath plots were generated in three orthogonal directions for distribution of

molybdenum in the Kitsault deposit. Examples in the EW, NS, and vertical directionsare shown in Figure 16.18 through Figure 16.20.





Figure 16.18 Swath Plot East-West 50 m Swaths





Figure 16.19 Swath Plot North-South 50 m Swaths





Figure 16.20 Swath Plot Vertical 20 m Swaths

The results of the ID model were included in the swath plots for comparisonpurposes. There is good correspondence between the models in all of these areas.The degree of smoothing in the OK model is evident in the peaks and valleys shown

in the swath plots. Deviations tend to occur for two reasons. First, reduced

tonnages near the edges of the deposit tend to accentuate the differences in gradebetween models. Second, differences in grade become more apparent in the lower-

grade areas – these typically are the flanks of the deposit where the density of drillingoften decreases.





SRK has reviewed the results of these comparisons, and is of the opinion that allcomparisons show close agreement with the underlying OK model. SRK also

conducted independent comparisons, which included generation of independentinverse distance and nearest neighbour models, using independent grade estimation

parameters, as well as statistical comparisons between the OK block grades and the

underlying composite data. SRK was able to closely reproduce both the tonnageand grades in the global OK model, and is of the opinion that both grade estimationand resource classification methodologies utilized by Avanti are in accordance with

Canadian Securities Administrators’ NI 43-101.

1 6. 1. 14 R E S O U R C E C L A S S I F I C A T I O N

The mineral resources at the Kitsault deposit were classified in accordance with the

CIM definition standards for mineral resources and mineral reserves (December

2005). The classification parameters are defined in relation to the distance to sampledata are intended to encompass zones of reasonably continuous mineralization.

During the grade estimation process, distance to closest composite, average

distance and number of drillholes used to estimate the block were stored in the blockmodel. Using these values as a basis, blocks were classified as follows:

Measured Mineral Resources – blocks with Mo% grades estimated using aminimum of 3 drillholes within an average distance of 35 m

Indicated Mineral Resources – blocks in the model estimated using a

minimum of three drillholes that are at maximum average distance of 90 m

Inferred Mineral Resources – blocks in the model not meeting the criteria for indicated resources but are within a maximum distance of 150 m from a

drillhole.

This classification is based on detailed drillhole spacing analysis and assignment of

confidence intervals, a discussion of which is beyond the scope of this section. SRKhas reviewed this analysis and is of the opinion that it forms a defendable basis for resource classification.

1 6. 1. 15 M I N E R A L R E S O U R C E ST A T E M E N T

The mineral resources for the Kitsault molybdenum deposit were audited by SRK at54 Mt at an average grade of 0.112% molybdenum classified as Measured Mineral

Resources, 153 Mt at an average grade of 0.088% molybdenum classified asIndicated Mineral Resources, and an additional 26 Mt grading an average of 0.069%molybdenum classified as Inferred Mineral resources. Although molybdenum forms

the basis for this resource estimate, silver, lead, and tungsten were also estimatedand are included in the resource tabulation provided in Table 16.15. This resource

estimate was prepared by Avanti personnel.





The mineral resources are reported in accordance with NI 43-101 and wereestimated in conformity with the generally accepted CIM guidelines. Mineral

resources are not mineral reserves and do not have demonstrated economic viability.There is no certainty that all or any part of the mineral resource will be converted into

mineral reserves. The audit of this resource estimate was completed by Jeffrey Volk,

P.Geo., an independent qualified person as this term is defined in NI 43-101. Theeffective date of this resource estimate is March 31, 2009. The mineral resourcesstatement for the Kitsault molybdenum project is presented in Table 16.15.


March 31, 2009

Resource Classification

Qty

(Mt)

Grade Contained Metal

Mo

(%)

Ag

(g/t)

Pb

(%)

WO3

(%)

Mo

(Mlb)

Ag

(Moz)

Pb

(Mlb)

WO3

(Mlb)

Measured** 54 0.112 4.54 0.022 0.007 133 8 26 8

Indicated** 153 0.088 5.24 0.025 0.006 297 26 84 20Measured & Indicated** 207 0.094 5.06 0.024 0.006 430 34 110 28

Inferred** 26 0.069 4.15 0.019 0.005 40 4 11 3

*Mineral resources are not mineral reserves and do not have demonstrated economic viability. All

figures were rounded to reflect the relative accuracy of the estimates. The cut-off grades are basedon metal price assumptions of US$20.00/lb Mo, and a metallurgical recovery of 89% Mo. Silver,lead, and WO3 were not used in the pit optimization.

** Reported at a cut-off grade of 0.04 % Mo contained within a potentially economic open pit.

The mineral resources are reported at a cut-off grade to reflect the “reasonable

prospects” for economic extraction. SRK considers that portions of the Kitsaultmolybdenum deposit are amenable for open pit extraction, and has not considered

underground mining methods for deeper portions of the deposit.

The “reasonable prospects for economic extraction” requirement was tested bydesigning a series of conceptual pit shells using the Lerchs-Grossman optimizingalgorithm. These parameters were selected by SRK to represent an “optimistic”

expectation reflecting the intent that the resource should comprise material that is

potentially economically mineable. The reader is cautioned that the results from thepit optimization are used solely for the purpose of reporting mineral resources that

have “reasonable prospects” for economic extraction by an open pit. After review of several scenarios considering different metal prices, design criteria and operating

costs assumptions, SRK assumed a molybdenum price of US$20.00/lb, a

metallurgical recovery of 89%, and slope angles of 40° in all areas. These

parameters were selected by SRK to represent an “optimistic” expectation reflectingthe intent that the resource should comprise material that is potentially economically

mineable in the future.





1 6. 1. 16 R E S U L T S O F SRK A U D I T

SRK verified and validated this model using a number of different methodologies,

and find the model acceptable as a basis for resource reporting under NI 43-101guidelines. This verification included:

visual comparison of block grades and composite grades in plan and section

statistical comparisons between block and composite grade distributions

independent grade estimation (inverse distance and nearest neighbour models).

All of these independent checks show good agreement between the Avanti model

and all other ancillary checks.

1 6. 1. 17 M I N E R A L R E S O U R C E SE N S I T I V I T Y

In order to assess the impact of cut-off grade on contained metal, tonnage, and

grade were summarized within the 0.04% Mo resource pit above a series of molybdenum cut-offs (Table 16.16 and Table 16.17). As seen in these sensitivities,

the resource is relatively insensitive to cut-off grade in the 0.04 to 0.06% Mo range,

which is likely the cut-off grade range of economic interest.

Table 16.16 Mo Cut-off Grade Sensitivity Analysis within Resource Pit –

Measure and Indicated Resources

Cutoff Grade

(Mo%)

Quantity

(Mt)

Mo Grade

(%)

Contained Metal

Mo (Mlb)

0.02 315 0.072 5010.025 292 0.076 490

0.03 257 0.083 468

0.035 234 0.088 452

0.04 207 0.094 430

0.045 197 0.097 420

0.05 189 0.099 412

0.055 183 0.100 405

0.06 178 0.102 398

0.065 170 0.103 388

0.07 163 0.105 377

0.075 151 0.108 3570.08 134 0.111 329





Table 16.17 Mo Cut-off Grade Sensitivity Analysis within Resource Pit – Inferred

Resources

Cutoff Grade

(Mo%)

Quantity

(Mt)

Mo Grade

(%)

Contained Metal

Mo (Mlb)

0.02 77 0.042 720.025 64 0.047 66

0.03 45 0.054 54

0.035 36 0.060 48

0.04 26 0.069 40

0.045 23 0.072 37

0.05 20 0.076 34

0.055 17 0.080 31

0.06 16 0.082 28

0.065 14 0.084 27

0.07 13 0.085 25

0.075 10 0.090 200.08 7 0.095 15

1 6. 1. 18 D I S C U S S I O N A N D C O N C L U S I O N S

Avanti has completed a significant volume of work during the 2008 field season,which included a 33 hole (10,131 m) infill drilling program. This program wasfocused on both conversion of inferred to indicated resources as well as the

confirmation of the historical drilling results. SRK has reviewed the results of this

program, and is of the opinion that the assay results from the 2008 drilling compare

closely with the assays from the historic drilling and are overall confirmatory. Inaddition, the 2008 program was successful in converting inferred resources toindicated and measured resources, and the current drillhole spacing within the

Kitsault deposit appears sufficient to advance the project to feasibility level studies.

1 6 . 2 M I N E R A L R E S E R V E S

The Kitsault mine Mineral Reserves have been prepared in accordance withNI 43-101 standards and CIM standard definitions. These reserves are sufficient for

close to 15 years of mining at an annual production rate of 40,000 t/d. MineralReserves are summarized by class in Table 16.18.

The notes accompanying Table 16.18 are an integral part of the Mineral Reserves

and should be read in conjunction with the Mineral Reserve statement.





Table 16.18 Mineral Reserves by Class

By Class

Cut-off

Grade (Mo%)

Quantity

(Mt)

Mo Grade

(Mo%)

Contained

Metal (Mlb)

Proven 0.036 55.7 0.109 121.35

0.027 3.8 0.031 2.35Total 59.5 0.104 123.70

Probable 0.036 134.5 0.086 231.08

0.027 21.2 0.031 13.13

Total 155.7 0.079 244.21

Total Proven & Probable 215.3 0.085 367.91

Notes:

1. Reserves calculated in accordance with CIM guidelines.

2. The metal price used for reserve calculation is US$12.51/lb Mo.

3. Metallurgical recovery is 90.6% for Mo.

4. Pit optimization parameters have changed from the time the resource estimate was completed.

As a result, an additional 8.3 Mt grading 0.031% Mo of the economic reserves within theoptimized pit was available for the variable cut-off strategy (see Note 5) making the reserve

statement higher than the resource statement (see Table 16.15 and Table 16.16).

5. Cut-off grades used were variable, 0.036% Mo and 0.027% Mo.

6. Mining recovery is estimated at 100% and dilution is nil.

7. The waste-to-ore ratio for the deposit is 0.75.





1 7 . 0 M E T A L L U R G I C A L T E S T I N G

1 7 . 1 I N T R O D U C T I O N

Wardrop conducted a review of the historical metallurgical testwork, reports, and

plant operation data related to the Kitsault deposit. Avanti obtained this information

from the previous owners and provided this to Wardrop for use in this study.

Units of measure are stated in full throughout this section for clarity as a result of the

use of imperial measures in many of the historical documents.

Avanti is proposing the mining and processing of 40,000 tonnes per day of amolybdenum-bearing resource material located in the Cassiar land district of BC.

For this study, historical testwork results together with information by Resource

Development Inc. (RDi) based in Denver, Colorado, and taken from the 2009 SGS

testwork program, will be used as a basis for the design parameters. This 2009 SGS

test data is the latest and most comprehensive testwork conducted on samples taken

from the resource material that has been slated for treatment. The 2009 SGS

metallurgical test program was completed early October 2009 and the results have

been incorporated as Section 17.5.2 in this report.

The proposed Kitsault Project plans to mine three distinct mineralized zones

identified geologically as monzonite, diorite and hornfels rock types.

The molybdenum in the feed material is present as the mineral molybdenite. Minor

lead, silver, and tungsten components are also present in the ore.

The initial feed to the plant is planned to have a nominal head grade of 0.093% Mo,

the molybdenum recovery has been estimated to be 90.6%, and the grade of

flotation concentrate will be 52% Mo.

1 7 . 2 M E T A L L U R G I C A L T E S T W O R K T I M E P E R I O D S

The major testwork and metallurgical reviews are classified in the following periods:

1963 to 1964: Testwork focused on mineralogy, flotation and grindability by

Lakefield Research and other laboratories

1967 to1974: Plant operations by British Columbia Molybdenum Ltd. (BC

Molybdenum)

1975 to 1979: Feasibility Study by Kilborn Engineering Ltd. (Kilborn)





1981 to 1982: Plant operations by AMAX of Canada

1983: By-product evaluation at SGS Lakefield

2008: PEA by SRK US

2009: Testwork focusing on flotation and grinding under the direction of RDiat SGS Vancouver and Hazen Research respectively. The test report

detailing this information was provided by RDi.

Metallurgical test programs have been conducted at several laboratories throughout

the different project phases dating back to 1963. Operational plant data was

obtained from Avanti for the two periods when the mine was in production. This

included the operating period from 1967 to 1972 and from 1981 to 1982.

The Kitsault concentrator operated from 1967 to 1972 as a 6,000 tons per day

throughput facility. Plant operating data was generated during this period and is

reviewed in a subsequent section.

The Climax Molybdenum Company conducted a Feasibility Study in 1978 at the

AMAX laboratory in Golden, Colorado. Additional grindability studies were

concurrently conducted by the Allis-Chalmers equipment company.

Following the 1978 Feasibility Study, a revised plant design that included doubling

the throughput of the plant, was implemented. The Kitsault concentrator was

subsequently returned to operation in April 1981 with a throughput of 12,000 tons per

day. The plant operated until November 1982 when it was shut down as a result of

low metal prices.

Following the shutting down of the plant in 1982, a by-product study program was

conducted which continued into 1983. The plant and the town site were maintainedby AMAX until 1994 at which time the Kitsault concentrator was completely

dismantled, salvaged and sold, and the site restored.

In the 2008 PEA, SRK advocated a relatively conventional processing facility with a

throughput of 40,000 tonnes per day.

For the 2009 PFS, a new testwork program was implemented by Avanti under the

direction by RDi. Flotation process optimization tests are currently still being

performed by SGS Vancouver, while grindability evaluation testwork was conducted

by Hazen Research Inc, with a review of the test results by Contract Support

Services Inc., Red Bluff, California.

Flotation test results have been evaluated by RDi, and plant design

recommendations have been made based on the results obtained. The conceptual

evaluation of by-product recovery methods was also explored by the 2009 SGS test

program.





1 7 . 3 M I N E R A L O G I C A L E X A M I N A T I O N

Mineralogical evaluations have been done at various times throughout the history of

this deposit. The following description of the mineralogy is a summary as outlined in

a report issued in 1981 (Steininger, 1981).

The primary economic mineral in the Kitsault deposit is molybdenite which is

contained in a stockwork of quartz. Other minerals of interest that occur in veins

throughout the deposit are pyrite followed by minor amounts of scheelite, galena,

sphalerite, chalcopyrite, various lead-bismuth sulphosalts, pyrrhotite, tetrahedrite and

carbonate minerals.

The molybdenite occurs within the quartz veins as individual disseminated grains,

ribbons, fracture coatings, and rosettes. The majority of the molybdenite is present

as individual grains of less than 0.05 mm in diameter making up between 80 and

90% of the total molybdenite present. The next most abundant occurrence of

molybdenite is as ribbons of up to 2 mm in width, followed by molybdenite coatingson fracture surfaces. Other occurrences of molybdenite are as rosettes of up to

5 mm in diameter and as rare clots of up to 20 mm wide. Molybdenite disseminated

throughout the matrix of breccia dikes is the rarest of the molybdenite occurrences.

Since the recovery of by-products may be developed at a later stage of the project, a

synopsis of the occurrence of these minerals of potential interest is included in this

review.

Scheelite commonly occurs in the quartz-pyrite veins. The scheelite grains are

generally less than 2 mm across with an approximate average size of 1 mm. The

scheelite grains are erratically disseminated throughout the quartz-pyrite veins. The

scheelite is apparently free of molybdenum based on fluorescence analysis tests.

The majority of the base metal mineralization is associated with the polymetallic

veins with the notable exception of galena, which was found to be associated with

molybdenite. Chemical and microprobe examination indicates that the majority of the

galena has a grain size of approximately 10 µm. As much as 20% of the galena

occurs as inclusions encapsulated within the molybdenite lamella with the remaining

galena occurring as attachments situated at the edges of the molybdenite crystals.

Silver appears to be directly associated with the lead probabaly as solid solution with

the galena, and possibly in tetrahedrite.

The polymetallic veins, which occur throughout the deposit, contain chalcopyrite,

tetrahedrite, pyrite, sphalerite, galena, lead-bismuth sulphosalts, molybdenite,

fluorite, and carbonate.

The predominant oxide gangue minerals are quartz, fluorite, gypsum-anhydrite, and

carbonate.





1 7 . 4 H I S T O R I C A L M E T A L L U R G I C A L R E V I E W

The metallurgical testwork and plant data review will follow the sequence as

previously outlined.

The initial plant design of 6,000 tons per day of production was completed by Wright

Engineers. The initial BC Molybdenum mill (formerly Kennco Explorations (Western)

Limited, a subsidiary of Kennecott Copper) began plant operations in 1967.

17.4.1 BC M O L Y B D E N U M P E R I O D – 1963 T H R O U G H T O 1976

Wardrop received reports and data for the BC Molybdenum period from Avanti,

which was compiled by Climax Molybdenum Limited. This information was used in

determining the scope of amenability testing and concentrator design criteria

recommendations put forward by the Extractive Metallurgy Laboratory for the

expansion in 1979. The information was compiled into seven sections as follows:

Section 1: Introduction

Section 2: Pre-Plant Startup Testing

Section 3: Concentrator Operations Reports

Section 4: Post-Plant Startup Testing

Section 5: The Lead Problem and Lead Removal Testing

Section 6: Analytical Reports and Procedures

Section 7: Proposed Expansions Reports Prior to 1976.

Sections 2, 3, and 5 are of particular metallurgical interest and will be discussed in

this section. Sections 1 and 4 were not available.

SE C T I O N 1: IN T R O D U C T I O N

The information in this section is not available as it is missing from the database.

SE C T I O N 2: PR E-PL A N T START -U P R EPORTS 1963 TO 1964

During the period of 1963 to1964 a number of test programs were facilitated. The

following are summaries of four reports found in Section 2.

The Mineral Processing division of the Canadian Department of Mines and Technical

Surveys was tasked with completing a metallurgical investigation on a sample of

molybdenite ore submitted by Wright Engineers. The testing was begun in early

1963 and the final report was issued in mid-1964. Testing focussed on determining

the grade and recovery of molybdenite that could be obtained from the sample.

Results indicated that a poor recovery of 65.5% was achieved at a reasonable





concentrate grade of 53.9% Mo. This poor result was attributed to finely intergrown

gangue in the final concentrate. The highest recovery value obtained was 80.7%

with a final concentrate grade of 52% Mo. It was found that appreciable amounts of

impurities were present in the final concentrate including lead, copper, and bismuth

sulphide minerals. The primary grind size was at a P80 of 70% minus 74 µm. The

final concentrate size was not analyzed, although a regrind time of 20 minutes wasstated to have been used on the rougher concentrate but this was not correlated with

a product particle size.

Mineralogical investigations were carried out by the Canadian Department of Mines

and Technical Survey on samples supplied by Wright Engineers and consisted of

about 0.05 kg of broken drill core sample as well as three molybdenum concentrates

samples. The principal metallic minerals were molybdenite and pyrite, and the main

non-metallic component was quartz. The molybdenum concentrates of varying

grades were found to contain some quartz with trace amounts of pyrite as the

gangue minerals.

In 1964, Britton Engineers, located in Vancouver, BC, were tasked with a test

program to determine the open cycle rougher flotation recovery at various grind

sizes, to obtain grindability work index, investigate regrind size, and determine

cleaner flotation recovery. Parameters included reagent and pH requirements at a

targeted final concentrate grade of 50.9% Mo. Observations on settling

characteristics and filtration behaviour were also requested. The composite feed

sample assayed 0.19% Mo and a series of rougher flotation tests at various grinds

were completed. The final grind parameter chosen for testing was 44% minus 74 µm

for the primary grind followed by a flotation and two stages of regrinding in the

cleaner circuit. Only material in the plus 44 µm fraction was reground. The overall

final concentrate particle size was 75% minus 74 µm. Reagent dosages were not

varied as results were found to be favourable with the initial reagent types anddosages chosen for the program. Following the rougher flotation testing and using

the results from the best case, cleaning tests were then conducted. The best overall

test gave a result of 51.7% Mo concentrate grade with 93.9% recovery. Locked

cycle tests were not done to confirm the results obtained. However, Britton

Engineering anticipated that a further recovery increase to 95% or higher could be

realized with the re-circulation of the cleaner tailings and finer grinding in the cleaner

circuit. During flotation, sodium cyanide was used as a depressant in each section of

the cleaner circuit. Also completed was a specific gravity determination of the ore

which was reported as 2.65. Work index determinations were carried out on a

standardized mill the average result of 12.9 kWh/short ton being reported for the

Bond Work Index. Only general observations on the need for flocculant prior to

thickening and filtration were reported. It was stated that the flocculant added to the

final concentrate hindered the rate of filtration.

The Western Mining Division Research Department of Kennecott Copper

Corporation was requested to complete independent amenability testing on sample

material taken from Alice Arm (Kitsault) in 1964. The test report included the

amenability test results along with the projections for anticipated grade and recovery





obtainable at full scale operation together with a flowsheet based on the test results.

The feed grade of the samples tested was between 0.18% Mo and 0.20% Mo. A

recovery of 93.4 to 95.6% was obtained and the final concentrate met the

specifications of a marketable grade which was specified as having less than 1.25%

copper contained in the concentrate. Concentrate grades obtained were between

49.9 and 51.7% Mo and these results were given as the projected attainable grade.The results were taken from the locked cycle test data generated from testwork

conducted on two composite samples. Tests indicated that a relatively coarse

primary grind of 30% plus 105 µm could be anticipated; however, the need for

regrinding to 10% plus 74 µm was necessary to make the saleable concentrate

grade requirement for copper. It was not possible to make a concentrate grade of

53.9% Mo containing a low level of copper impurity with the sample tested. An

indicated Bond Work index number of 15.0 kWh/ton was reported based on the use

of a modified test program. From the laboratory tests results obtained, a 6,000 ton

per day conventional mill was designed. The flowsheet included three stages of

crushing, two stages of rod and ball mill grinding, a flotation circuit which consisted of

a rougher stage followed by four stages of cleaning and including regrinding, productas well as dewatering and product packaging and shipping.

SE C T I O N 3: CO N C E N T R A T O R O P E R A T I O N S R EPORTS

Reports included in this volume were mainly of the concentrate production that was

produced in 1971 and 1972. Some general observations of limited information

provided is that the primary grind was typically 44 to 50% plus 150µm and that lead

had become a concern in the concentrate necessitating the need for a lead leach

circuit to be installed. The year-to-date November 1971 concentrate grade was

55.3% Mo, and year-to-date March 1972 concentrate grade was 54.0% Mo.

SE C T I O N 4: PO S T PL A N T ST A R T -U P T E S T I N G

The information in this section is not available as it is missing from the database.

SE C T I O N 5: TH E LE AD PR O B L E M A N D L EA D R EMOVAL T E S T I N G

This section includes all the reference material dedicated to the study of the

occurrence of lead in the molybdenum flotation concentrate and the subsequent

requirement for lead removal from the concentrate. This section includes a collection

of 32 articles related to the presence of lead in the molybdenum flotation

concentrates, and the teswork conducted. This information was compiled during

plant operations from 1967 and continued through to the mine closure in 1972. No

plant design changes were undertaken during this phase of the operation. The

report entitled “Lead at Kitsault” that reviewed all the BC Molybdenum investigations,

concluded with the statement that it was anticipated that lead would be a continuing

problem throughout the life of the mine (Holland, 1976). A complete summary of the

lead reports is located in Section 17.4.7.





SE C T I O N 6: A N A L Y T I C A L R E P O R T S A N D PR O C E D U R E S

This section is a compilation of the various assay procedures used in the BC

Molybdenum era. Memorandums are also included regarding the molybdenum

concentrate grade analysis with inter-laboratory comparison studies, together with

methodology for the oil and moisture determinations required on the molybdenumconcentrate samples.

SE C T I O N 7: PR O P O S E D EX P A N S I O N S R E P O R T S PR I O R T O 1976

Expansion studies were investigated from 1968 through to 1975 with various

expansion ideas put forward.

17.4.2 K I T S A U L T C O N C E N T R A T O R P R O D U C T I O N D A T A 1967 TO 1972

The Kitsault concentrator was in operation for five years from 1967 to mid-1972.

During that time 10,439 t of molybdenum was recovered from 9.3 Mt of ore whichhad an average feed grade of 0.123% Mo. The average recovery during that period

was 89.2%. Table 17.1 contains the production information from the period of

operation of the mine between 1967 and 1972, and this information was sourced

from the1976 Kilborn Feasibility Study. Prior to closure, lower grade material was

mined which may have allowed for an increase in the throughput in the mill.

Table 17.1 Production Summary Data – Kitsault Concentrator 1967 to 1972

Concentrator

Data Units

Production Year

1967(1)

1968 1969 1970 1971 1972(2)

1967

to 1972

Ore Milled t n/a 2,002,937 2,137,794 2,443,255 2,246,348 473,210 9,303,544

Feed Grade % Mo n/a 0.132 0.130 0.122 0.112 0.117 0.123

Mo Recovery % n/a 88.3 88.7 90.7 89.4 85.9 89.2

Mo Production kg n/a 2,311,086 2,528,000 2,788,443 2,318,811 492,733 10,439,073

(1)Data not available – it appears to be included in the 1968 data.

(2)4 months of production.

17.4.3 SE C O N D P R E - P R O D U C T I O N P E R I O D F R O M 1976 TO 1980

Testwork from this period, together with BC Molybdenum production plant data from

the plant operating period of 1967 to 1972, was used as part of the design basis for changes to the facility prior to the re-opening in 1981 as a 12,000 tons

(10,886 tonnes) per day throughput plant.

Kilborn undertook a Feasibility Study in 1976 with the view to re-opening the Kitsault

mine as an economically viable venture. In order to do this, the plant design was

expanded to a throughput capacity of 12,000 tons per day. The use of existing

equipment was a paramount design criterion and therefore dictated some of the





process design choices. Another contributing factor in the plant design was the

availability of production plant data from the operating mine life from 1967 to 1972 as

well as test data generated in further studies. Much of the actual test data generated

was included in appendices located in Volume 3 of the Feasibility Study, which is not

available from the Avanti information database. However, reports and summary

reports that were included in other parts of the Kilborn Feasibility Study will be usedas a summary of the data.

Volume 1 of the Kilborn Feasibility Study includes the metallurgical evaluation based

on the test results up to and including report updates issued in 1978. The major

conclusions and design is based on the 1976 pilot plant testing program carried out

at the AMAX Extractive Metallurgy Laboratory. These test results are summarized

and included in the report (Enochs, 1977b). This report was used as a reference to

the choices made in the overall plant design and was an updated revision of the

earlier report issued in 1977 (Enochs, 1977a).

The 1976 pilot plant test program conducted by AMAX was a study to confirm the

various operating requirements by establishing adequate crushing and milling

capacities. It was also used to better define flotation reagent requirements and to

characterize the leach process for lead removal. The technical report was based on

the results from the pilot plant study and was subsequently compiled in 1977 and is

referred to as the overall guide to the design of the Kitsault plant (Olin, 1977). The

report alludes to appendices which included the actual test data obtained during the

pilot plant study, but this information from these appendices is no longer available.

Concurrently with the pilot plant study, a grindabililty study was also undertaken at

Allis-Chalmers to determine the work indices of the different rock types in the ore

body.

Specific aspects of the 1976 pilot plant test program, as reported in the Kilborn

Feasibility Study, will now be presented.

H E A D AN A L Y S I S

Table 17.2 summarizes the assay results for the head grade analysis of samples

used in the AMAX test program. Lots 1 and 2 are surface samples from the West Pit

and Lots 3 and 4 are surface samples from the East Pit. The sample size used in

preparation for the pilot plant study was approximately 2 tons per lot but no

information was provided as to the sampling method employed.





Table 17.2 Ore Head Analysis – AMAX 1976

Lot

Head Analysis

% Mo % Pb % Cu % WO3

1 0.143 0.074 0.016 0.014

2 0.267 0.175 0.032 0.011

Average West Pit 0.205 0.125 0.024 0.013

3 0.064 0.008 0.015 0.008

4 0.189 0.018 0.015 0.005

Average East Pit 0.127 0.013 0.015 0.007

It is apparent that the head grades of both pit samples are higher than the grade

estimated for the present (2009) design, namely 0.093% Mo. The variation in head

grade for molybdenum as well as lead and copper is apparent, while the head grade

for the lead from the West Pit is notable.

G R I N D A B I L I T Y

Allis-Chalmers used the classical Bond grindability procedure to determine the work

and abrasion indices of various ore types from the Kitsault orebody. A summary

table of the results was included in the Kilborn Feasibility Study and has been

replicated in Table 17.3.

Table 17.3 Summary of Work and Abrasion Indices – Allis-Chalmers 1977

Sample Identification

Impact

Index

Rod Mill

Work Index

(1)Ball Mill

Work Index

(1)Abrasion

Index

Specific

DensityHornfels East Pit 8.0 18.2 15.5 0.447 2.75

Hornfels West Pit 8.8 17.1 13.5 0.323 2.71

Hornfels North Pit 12.6 17.7 14.3 0.432 2.70

Granodiorite 6.5 12.7 11.4 0.360 2.67

Diorite-Granodiorite West Pit 6.6 9.7 9.8 0.211 2.69

Granodiorite North Pit 6.2 9.6 10.2 0.194 2.64

(1)Units are kWh/short ton.

The AMAX test program of 1976 also included grindability test results conducted on

the four sample lots used in the pilot plant study. The work indices were not

classically derived Bond grindability indices but were based on a proceduredeveloped by AMAX and reported in “Grindability of Kitsault Ore Types” February 28,

1977, but the source file is no longer available and therefore the methodology cannot

be evaluated. However, a summary of the results as reported is presented in Table

17.4. The variation between the different rock types and samples tested is seen to

be significant.





Table 17.4 Summary of Grindability Studies – AMAX 1976

Sample Identification

Work Index

(kWh/ton)

Abrasion

Index

Lot 1 13.42 0.60

Lot 2 15.84 0.64

Average West Pit 14.63 0.62

Lot 3 22.78 0.72

Lot 4 17.87 0.69

Average East Pit 20.33 0.71

Based on the results obtained, a Bond Work Index value of 18.0 kWh/short ton was

recommended at the time of writing of the 1976 report which was prior to the results

from Allis-Chalmers being available. As the type of Bond Work Index value was not

specified in the report, it will therefore be assumed to be a Ball Mill Bond Work Index

based on the inference.

In the 1977 report issued by Wright Engineering for the Climax Molybdenum

Feasibility Study and using all the available test data, the following grinding index

design criteria were used. For comparison purposes a conversion of the grinding

index numbers to metric units has been included.

Table 17.5 Grinding Design Criteria – Climax Molybdenum Feasibility Study

1977

Item Imperial

Units

Metric

Units

Impact Index 12.6 kWh/ton 14.2 kWh/tonneRod Mill Index 18.2 kWh/ton 20.1 kWh/tonne

Ball Mill Index 15.5 kWh/ton 17.1 kWh/tonne

Abrasion Index 0.447 g

The numbers are virtually identical to those obtained by Allis-Chalmers for the

hornfels East Pit Sample and it can only be surmised that the plant was designed

with using the hardest and most abrasive numbers generated during the testing

program.

Grinding Testwork Conclus ions

As indicated by testwork and issued design memorandum, the grinding index varies

greatly with different rock types and samples tested, and Bond Ball Work Index

values have varied from 12.8 to 18.0 kWh/ton (14.1 to 19.8 kWh/tonne). The Allis-

Chalmers work in particular highlights the variations which ranged from 9.8 to

15.5 kWh/ton (10.8 to 17.1 kWh/tonne), and even the highest Allis-Chalmers number

is less that the recommended 18.0 kWh/ton (19.8 kWh/tonne). The inferred





18.0 kWh/ton was likely related to the results obtained during the 1976 AMAX

grindability testing, which used a comparative test procedure and gave results as

high as 22.8 kWh/ton (25.1 kW/tonne).

G R I N D S I Z E V E R S U S FL O T A T I O N R E C O V E R Y T ESTWORK

A total of f ive primary grind sizes per sample lot, as described earlier, were evaluated

in the laboratory to determine the optimal grind-recovery relationship prior to the pilot

plant testing. As illustrated in Figure 17.1, depending on origin of the feed sample, a

rougher flotation recovery of between 91% and 97% was obtainable at a grind of

45% plus 100 mesh (150 µm). The results as shown are at a 20:1 ratio of

concentration. The above results have been reported verbatim from the 1977 report

since the actual test data and the complete material balances are no longer available

(Olin, 1977).

Figure 17.1 Molybdenum and Lead Recovery during Rougher Flotation as a

Function of Primary Grind – Climax 1977





Once the optimum grind size of 45 to 50% plus 150 µm was chosen, a further four

flotation tests were performed. The overall results are shown in Table 17.6 and

although the complete material balances are no longer available, Figure 17.2 shows

the results obtained from these tests.

Table 17.6 Metallurgical Summary of Laboratory Batch Flotation Tests –Olin 1977

Lot

Wt

(%)

Calculated

Feed Grade

Rougher

Concentrate Grade

Distribution

in Concentrate

Mo (%) Pb (%) Mo (%) Pb (%) Mo (%) Pb (%)

1 3.2 0.149 0.078 4.37 0.329 95.3 13.6

2 6.2 0.285 0.165 4.48 0.526 97.2 19.7

3 6.8 0.074 0.007 0.99 0.038 90.9 35.5

4 6.2 0.221 0.017 3.32 0.068 93.8 24.5

Figure 17.2 Mo and Pb Recovery during Rougher Flotation as a Function of

Flotation Time at a Grind of 45 to 50% plus 150 µm – Olin 1977





The cumulative rougher grade ranged between 1.0% and 4.4% Mo at the 45% grind

with 12 minutes of flotation depending on the sample tested.

Following these preliminary rougher flotation tests, a pilot plant was operated in order

to produce sufficient rougher concentrate which could then be subsequently cleaned

in the laboratory to produce a high grade concentrate. The metallurgical summaryfor the rougher pilot plant flotation test results is shown in Table 17.7. It is not clear if

the samples created in the rougher portion of testing were kept separate throughout

the cleaner circuit testing and therefore an arithmetic average of the molybdenum

recoveries in the rougher circuit will be used for comparative purposes.

Table 17.7 Metallurgical Summary of Pilot Plant Rougher Flotation – Olin 1977

Lot

Wt

(%)

Calculated

Feed Grade

Rougher

Concentrate Grade

Distribution

in Concentrate

Mo (%) Pb (%) Mo (%) Pb (%) Mo (%) Pb (%)

1 11.1 0.126 0.038 1.07 0.138 93.2 39.81 7.6 0.141 0.070 1.79 0.420 96.2 46.0

2 3.1 0.206 0.116 6.41 0.780 96.7 20.8

2 4.0 0.204 0.110 4.94 0.571 97.2 209

3 1.7 0.058 0.015 3.07 0.176 92.9 20.0

3 3.7 0.061 0.007 1.53 0.038 92.3 19.6

4 4.4 0.179 0.010 3.86 0.058 95.2 24.8

4 4.2 0.191 0.014 4.26 0.081 94.6 24.5

The cleaner portion of this testing employed alternative operating conditions than

those typical of the BC Molybdenum era with the addition of depressants to the

regrind mill in the form of Nokes reagent and “N” silicate and the addition of the

frother and collector to the flotation cell. BC Molybdenum practise in the regrind

circuit had kerosene addition to the mills followed by lead depressant additions to the

cyclone feed sumps which was felt by the author of the report to inhibit the lead

depressant capabilities through fouling the galena surface with oil. This change in

reagent addition points combined with the fifth regrind may have contributed to the

results which showed extremely low lead content in the final concentrate prior to lead

leaching. Recirculation of middlings products was not incorporated during testing.

Initially four cleaner tests were run to evaluate the number of grinding stages

required, and to confirm the reagent scheme employed. After the four tests, the

remainder of the sample was processed in the laboratory. The concentrate producedin the final series of tests was used in downstream lead removal leach testwork. The

laboratory test results indicated that five stages of cleaning, rather than four, would

be required to produce concentrate grades of greater than say, 54% Mo. A summary

of the cleaner results obtained is shown in Table 17.8.





Table 17.8 Summary of Regrind-Cleaner Flotation Upgrading of Pilot Plant

Concentrates – Olin 1977

Lot

Regrind &

Cleaning

Stages

Wt

(%)

Calculated

Rougher Grade

Cleaner

Concentrate Grade

Distribution

in Concentrate

Mo (%) Pb (%) Mo (%) Pb (%) Mo (%) Pb (%)

1 4 3.3 1.97 0.202 54.24 0.09 92.1 1.4

1 5 3.0 1.85 0.227 59.16 0.03 97.5 0.4

2 4 9.0 5.03 0.532 53.94 0.14 96.7 2.3

2 5 10.4 6.20 0.585 58.80 0.05 98.4 0.9

3 4 3.0 1.79 0.084 54.42 0.06 90.9 2.3

3 5 3.4 2.07 0.070 58.37 0.03 96.3 1.4

4 4 7.5 3.63 0.083 46.87 0.07 96.5 6.4

4 5 6.8 4.08 0.082 58.23 0.03 96.7 2.5

The difference between the Mo and Pb results obtained is of great interest,particularly the Mo concentrate grade and recovery values for four and five regrind

stages. In the absence of specific detail, and in comparing these results, it has been

assumed that the overall tests procedures have been the same. The final

concentrates molybdenum grade after 5 stages of regrind and cleaning shows

grades between 58.2 and 59.2% Mo.

The associated Pb values in the tests which included five stages of cleaning ranged

between 0.03 and 0.05% Pb. Figure 17.3 shows the typical lead rejection and the

molybdenum grade achieved during the cleaning process employed for all four

samples. The graph chosen for illustrative purposes is that of the Lot 1 sample. The

graph clearly indicates that a large amount of lead is rejected at the fifth cleaner

stage. The actual regrind size achieved at each stage was not stated.





Figure 17.3 Lead Removal from Molybdenite Concentrate during Each Stage of

Beneficiation – Olin 1977

As a final analysis of the data, the results from the rougher and cleaner tests were

combined to determine the overall expected recovery for the flotation process. As

shown in Table 17.9, the overall recoveries are dependent on both the number of

cleaning stages as well as the sample source. Based on the test results and using

four stages of cleaning, an overall recovery of 84.2 to 92.3% is expected with

concentrate grades of not more than 54.4% Mo. With the addition of the fifth regrind

and cleaning stage, recoveries of between 89.2 and 95.4% for molybdenum were

shown with concentrate grades exceeding 58% Mo.





Table 17.9 Combined Rougher and Cleaner Flotation Results

Lot

Regrind &

Cleaning

Stages

Average Recovery to

Rougher Concentrate

Recovery to

Cleaner Concentrate

Overall

Calculated

Recovery

Mo (%)Mo (%) Pb (%) Mo (%) Pb (%)

1 4 94.7 42.9 92.1 1.4 87.2

1 5 94.7 42.9 97.5 0.4 92.3

2 4 97.0 20.9 96.7 2.3 93.8

2 5 97.0 20.9 98.4 0.9 95.4

3 4 92.6 19.8 90.9 2.3 84.2

3 5 92.6 19.8 96.3 1.4 89.2

4 4 94.9 24.7 96.5 6.4 91.6

4 5 94.9 24.7 96.7 2.5 91.8

It is of note that each of the samples showed the same tendency with the increase in

the number of regrind and cleaner stages, namely a higher molybdenum recovery

and a higher concentrate grade. Similarly, the extent of lead rejection in the fifth

cleaner stage, as compared with four cleaner stages only, is dramatic with a final

lead grade of between 0.03 to 0.05% Pb in the final concentrate.

The very high values of molybdenum concentrate grades obtained in this series of

tests are discussed in Section 17.4.8.

LE A D L E A C H I N G

Effective lead reduction to the saleable concentrate specification of <0.02% was

found to be dependent on the use of depressants in all stages of grinding andflotation, followed by leaching with hydrochloric acid. In order to reduce the lead

impurity of the molybdenum concentrate a hot hydrochloric acid leach was

implemented. The lead content was reduced to less than 0.02% Pb within 2 hours of

leaching. A minimum leach time of 2 hours was chosen in the plant design criteria.

It must be noted that the tests done were on a combination of the fifth cleaner

concentrates generated in the flotation tests. The feed to the lead leach circuit was

very low on average at 0.03% Pb.

B Y- P R O D U C T S

The AMAX study indicated that tungsten, lead, and silver minerals occur in theKitsault orebody in sufficient amounts to warrant further investigation. However,

although the production of by-products was tested, it was concluded that further,

more detailed investigations would be required to implement this portion of the

circuit.





17.4.4 K I T S A U L T C O N C E N T R A T O R P R O D U C T I O N D A T A – 1981 TO 1982

The Kitsault concentrator was re-opened in April 1981 under the ownership of AMAX

Inc. with a design capacity of 12,000 tons per day. The plant operated for 17 months

with a shutdown period of one month in August 1982, and a reduced work week for

the remainder of the operating life until October 1982. Production data for theKitsault concentrator has been summarized in Table 17.10. The information in the

table is the adjusted production data as reported during the time of operation.

Table 17.10 Production Summary — Kitsault Concentrator 1981-1982

Concentrator Data Units 1981(1)

1982(2)

Ore Milled t 1,785,819 2,228,861

Mill Ore Feed Grade

Mo % 0.118 0.121

Pb % 0.017 0.018

Cu % n/a 0.005Mo Recovery % 82.5 87.6

Mo Production as Mo kg 1,738,494 2,361,709

Mo Production as MoS2 t 3,178.8 4,377.6

Mo Concentrate Grade

Mo % 54.69 53.87

Pb % n/a -

Mo Packaged kg 1,738,494 2,361,709.2

Packed Concentrate Grade

Mo % 54.69 53.95

Pb % n/a 0.092

Cu % n/a 0.077Mill Availability % 70.4 90.5

Source: Kitsault Concentrator Production Data 1981-1982(1)

Production April to December.(2)

Production January to July, and September to October at reduced capacity.

Also available from this production period is the actual monthly concentrator

production data from 1982. This information is presented in Table 17.11 and was

extracted from the actual plant data. Table 17.11 also gives the figures for the

concentrate material which was dispatched, leached, which by-passed the leaching

circuit, and the amount of material which had to be reprocessed in the leach circuit

since it did not meet specification.





Table 17.11 Monthly Production Data – Kitsault Concentrator 1982

Month

Feed Grade

Flotation

Concentrate Grade

Packed

Concent rat e Grade Tot al Dr y Weight (l bs Mo)

% Mo % Pb % Cu % Mo % Pb % Cu % Mo % Pb % Cu Packed Leached Bypassed

Reprocessed

1981 Material

January 0.093* 0.017 0.007 53.05 0.089 0.110 54.20 0.023 0.066 327,736.6 301,310.4 26,426.2 0.0

February 0.104 0 .120 0.010 51.36 0.078 0.076 53.47 0.027 0.068 445,856.5 445,856.5 0.0 0.0

March 0.113 0.017 0.007 52.74 0.080 0.070 52.74 0.049 0.071 507,727.5 416,248.2 91,479.3 5,049.0

April 0.102 0.022 0.006 53.40 0.107 0.089 53.70 0.036 0.010 631,811.1 582,655.3 49,155.8 105,724.2

May 0. 130* 0 .012 0. 004 5 6. 08 0. 073 0. 067 5 5. 77 0 .039 0. 118 761, 890. 5 548, 454. 8 213, 435. 7 113, 992. 0

June 0. 154 0. 008 0 .008 56. 40 0 .060 0 .084 54. 94 0 .036 0 .127 1 ,042, 411. 3 683, 518. 5 358, 892. 8 104, 371. 0

July 0.126 0.020 0.004 5 3.96 0.097 0.053 5 3.79 0 .047 0.117 694,346.3 539,043.5 155,302.8 68,738.4

August 0.166 0.019 0.007 52.05 0.101 0.128 52.13 0.081 0.144 329,202.6 253,024.5 76,178.1 40,657.6

Septem ber 0.117 0.028 0.005 5 1.69 0.170 0.072 5 2.47 0 .041 0.106 372,760.0 372,760.0 0.0 74,735.8

October 0.126 0.028 0.007 54.00 0.115 0.032 53.61 0.032 0.115 628,940.2 628,940.2 0.0 120,245.4

Total/Average 0.121 0.018 0.005 53.89 0.092 0.076 53.87 0.038 0.081 5,742,682.6 4,771,811.9 970,870.7 633,513.4

*changed in table where transpositions from original production data were noted.





The following aspects of the plant data are of interest:

The head grade of the 2009 design is substantially lower at 0.093% Mo than

the feed grade for both periods of production, namely 0.123% Mo (for 1967

to 1972) and 0.120% Mo (for 1981 to 1982).

The molybdenum recovery for the 1981 to 1982 period was 85.33% for an

average concentrate grade of 54.22% Mo. The 2009 study gives a recovery

of 90.6% for an average concentrate grade of 52.0% Mo.

The lead content of the molybdenum concentrate in 1981 was 0.092% Pb

and this was reduced by leaching to 0.032% Pb which is still higher than the

limit of 0.02% Pb required in the specifications.

Overall recovery for the year of 1982 averaged 87.6% and the overall

average molybdenum concentrate grade for the year was 53.89% Mo. It

should be noted that in 7 months of the 10 months of production, a

concentrate grade of less than 54% Mo was obtained.

Also available from the plant data were the typical reagent consumptions required

during the 1982 production year, and this is shown in Table 17.12.

Table 17.12 Reagent Consumptions – Production Data 1981 to 1982

Plant Data 1982 kg/t

Collector Oil (Unspecified) 0.142

Frother (Terpene, Dowfroth 250, and Polypropylene Glycol) 0.046

Phosphorus Pentasulphide* 0.077

Hydrochloric Acid 0.046

Sodium Hydroxide* 0.111

Lime 0.111

Flocculant (Separan NP-10) 0.001

* Phosphorous Pentasulphide and Sodium Hydroxide reagents are mixed to create the Nokes

reagent required for the depression of lead and copper.

Metallurgical laboratory and plant tests were carried out throughout the 1981 to 1982

production era. Although no test data from that period remains, there are some

generalized references made to testing in the mill’s month-end reports.

17 .4.5 SG S L A K E F I E LD P I L O T P L A N T ST U D Y – 1983

During 1983 a pilot plant test program was undertaken at SGS Lakefield Research

using Kitsault sample material with the objective to develop a flowsheet and reagent

scheme what would produce an acceptable molybdenite concentrate grade and

recovery. An additional objective was to produce a saleable silver-lead concentrate.

Bench scale testing had been completed at Kitsault in early 1983 after the cessation

of plant operations, but there was insufficient concentrate generated in order to

continue with the work.





As a continuation to the bench-scale testing completed at Kitsault, two ore types

were tested at SGS Lakefield. The first type was identified as hornfels and the

second sample was identified as Intrusive. The Intrusive sample was not identified

further geologically and it is not clear whether it is a monzonite type or diorite-type

sample.

Poor molybdenum grades and recoveries were obtained in this study which

employed seven stages of cleaning in an attempt to obtain a reasonable

molybdenum concentrate grade. Testing appears to have only been done on the

rougher concentrate with 80.4% recovery for the hornfels and 83.2% recovery for the

Intrusive samples. The final molybdenum concentrate values of not more than 48%

Mo were realized during testing. The upgrading of the sulphide concentrate

produced as a by-product was investigated without promising results. The majority

of the lead and silver was lost in the cleaner circuit to the final tailings. An adaptation

whereby the Nokes depressant was removed from the flotation circuit did improve

the results and produced a lead, silver, copper concentrate, but no further testwork

was undertaken.

1 7. 4. 6 S RK P EA – 2 00 8

For the 2008 PEA, SRK proposed a conventional crushing, grinding and flotation

circuit, with tailings discharged to the tailings facility via thickeners. However, SRK

excluded a lead leach circuit from the design of the ore processing facility opting

instead to applying a smelter penalty for the treatment of concentrates having a lead

content which exceeded the market specification value.

17.4.7 L E A D L E A C H I N G T E S T W O R K R E V I E W

IN T R O D U C T I O N

The specification for molybdenum concentrate for use in steelmaking requires a lead

content of < 0.02%Pb. Concentrates with lead contents higher than 0.25% Pb are

not saleable. A penalty will be incurred for lead concentration between 0.02 and

0.10% Pb, and addition penalties will be charged for lead contents between 0.10 and

0.25%Pb.

The removal of lead from the Kitsault molybdenum flotation concentrates by leaching

was investigated widely in the period from 1968 to 1983 in order to reduce the lead

content to a level either < 0.10%Pb, or <0.02%Pb. The following sections present

the conclusions and findings from the testwork data reviewed from the various test

programs.





BC M O L Y B D E N U M PE R I O D 1960 T O 1980

Mineralogical Invest igat ions

The occurrence of lead in the Kitsault ores was studied and discussed in the 1968

technical report (Last et al., 1968). Their major conclusions were the following:

Lead occurred as the mineral galena (PbS) in variable amounts in the Alice

Arm deposit.

Free galena grains in the recovered molybdenum concentrate were finer

than 400 mesh (37 µm).

Subsequent investigations confirmed that lead occurred as galena in the

molybdenum concentrate. Furthermore, it was observed that part of the galena was

occluded within the particles of molybdenite so that the galena could not be

contacted by the leach solution. It was estimated that about 20% of the galena

particles were occluded within molybdenite grains and were therefore not leached(Stephens, 1968).

It was subsequently confirmed that galena was present as occluded grains of

<10 µm within molybdenite. In addition, it was found that a variety of lead sulphosalt

minerals containing varying amounts of copper, arsenic, antimony, bismuth, and

silver were also present although these did appear to form a minor proportion of the

total occurrence of lead (Steininger, 1981).

Laboratory Leaching Tests

Lead leaching using ferric chloride at atmospheric pressure was investigated at theKennecott Research Center (Prater, 1968). The results showed that the lead content

was reduced from 0.10 to 0.022% Pb after a 2-hour leach at 85°C using a 10% ferric

chloride solution.

The BC Molybdenum Limited Laboratory examined the lead leaching process at

ambient temperature using hydrochloric acid, nitric acid, and Syntex L (a dispersant)

as leachant solutions. Optimum results were obtained with a 3-hour leach using 5%

hydrochloric acid, 5% nitric acid, and 0.01% Syntex L. Under these conditions, the

lead grade in the molybdenite concentrate was reduced from 0.079 to 0.012% Pb

(Lane, 1968).

A sand milling process was tested in 1970 to prove that it could meet the specifiedlead grade. The sand-milling process refers to a process where the concentrate is

mixed with coarse sands (+420 µm) and hydrochloric acid in a vessel equipped with

an agitator. The test results showed that the lead grade was reduced significantly

from 0.32 to 0.017 Pb% (Vreugde, 1970).





However, subsequent testwork in1971, using the same process, failed to produce a

concentrate product which was within the specified lead content of <0.02% Pb (Ford,

1971).

In 1977, AMAX Extractive Metallurgy Laboratory carried out further lead leaching

testwork using hydrochloric acid at high temperature as a part of its metallurgicalinvestigation. Two types of molybdenite concentrate samples were tested, one was

produced at Kitsault during the 1967 to 1974 plant operations and the other was

prepared at the AMAX laboratory in Golden using Kitsault ore from four ore lots.

BC Molybdenum Concentrate Sample

The leach tests were carried out at 80°C and 20% solids pulp density on flotation

concentrates produced while the BC Molybdenum plant at Kitsault was in operation.

The effects of varied leach time and leach acid strength were investigated.

Table 17.13 Lead Leaching of BC Molybdenum Kitsault Mill Concentrate

Test

No.

Sample

ID

HCl

Strength

(%)

Leach

Time

(h)

Feed

Grade

(% Pb)

Residue

Grade

(% Pb)

Lead

Rejection

(%)

1 Low Lead 2.5 1.0 0.060 0.030 50.0

2 Low Lead 5.0 1.0 0.060 0.028 53.3

3 Low Lead 10.0 0.5 0.056 0.027 51.8

4 Low Lead 10.0 1.0 0.056 0.025 55.3

5 Low Lead 10.0 1.5 0.056 0.027 51.8

6* Low Lead 10.0 1.5 0.027 0.024 11.1

7 High Lead 2.5 1.0 0.185 0.039 78.9

8 High Lead 5.0 1.0 0.185 0.034 81.6

9 High Lead 10.0 0.5 0.190 0.028 85.3

10 High Lead 10.0 1.0 0.190 0.027 85.8

11 High Lead 10.0 1.5 0.190 0.025 86.8

12* High Lead 10.0 1.5 0.025 0.024 4.0

Average 0.1228 0.029 68.1

*The concentrates from Tests 5 and 11 were leached for an additional 1.5 hours with results

tabulated and reported as results for Tests 6 and 12 respectively.

Although up to 87% of the lead was leached as shown in Table 17.13, the lead

content in all the leach residues was still higher than the test objective value of

0.02% Pb which is the steel industry standard specification value.

AMAX Laboratory Sample

Four molybdenite concentrate samples were produced in the AMAX Extractive

Metallurgy Laboratory for treatment in the lead leaching circuit. Two sets of leaching

tests were conducted. The first set used a leaching time of 2 hours at 20% solids





heated to 80°C. The second set of tests used a leach time of 20 hours at 33% solids

heated to 85°C. All the tests used a hydrochloric acid solution strength of 5%.

Table 17.14 Lead Leaching from AMAX Laboratory Concentrate Samples

Sample

Feed Grade

(% Pb)

Residue Grade (% Pb) Lead Rejection (%)

2 h Leach 20 h Leach 2 h Leach 20 h Leach

1 0.032 0.011 0.009 65.6 71.9

2 0.055 0.014 0.012 74.5 78.2

3 0.029 0.009 0.008 69.0 72.4

4 0.038 0.013 0.013 65.8 65.8

Average 0.039 0.012 0.011 68.7 72.1

As shown in Table 17.14, all the tests attained the target lead grade of <0.02% Pb.

Also, the lead rejection difference between the 2 hour and 20 hour tests is relatively

small. However, the lead leach feed grades of the four samples tested was lowvarying between 0.029 and 0.055% Pb which is significantly lower than the feed

grades of the BC Molybdenum samples tested. This is partially attributable to the

change in the flotation test procedure namely, the addition of depressants to the

regrind stages as well as the inclusion of a further regrind step. The maximum lead

rejection value in this set of tests was only 78% which is lower than the 87%

achieved in the previous tests. However, the percentage of lead rejection is relative

to the feed to the leach and is not considered a significant factor. Of major

importance is the amount of lead remaining in the leached concentrate. When

comparing all the results and the variables that influence the amount of lead in the

leached concentrate, the most significant factor is the fineness of the grind of the

flotation concentrate.

Pi lot P lant Lead Leaching Tests – BC Molybdenum Ki tsaul t P lant S i te

Ambient Temperature Lead Leach Tests

In July 1969, a batch lead leaching pilot plant was set up at the BC Molybdenum

Kitsault Plant Site. The objective was to determine the economic feasibility of

producing a molybdenum concentrate containing <0.10% Pb. The leach tests used a

5% hydrochloric acid solution at ambient temperature and an average pulp density of

41% solids. After approximately 10 hours of leaching time, the lead content in the

samples was reduced, on average, to 0.042% Pb from feed grades which rangedfrom 0.34% Pb to 0.08% Pb.

Subsequent to the initial leach tests, the inclusion of nitric acid as an oxidising agent

in conjunction with hydrochloric acid to reduce hydrochloric acid consumption was

tested to follow the apparently successful laboratory tests conducted in 1968.

However, this procedural change resulted in the generation of nitric acid fumes and





excessive frothing problems which resulted in the termination of testing for safety

reasons.

High Temperature Lead Leach Tests

Laboratory tests had effectively shown that the use of elevated temperature during

leaching would reduce hydrochloric acid consumption and leach times. In 1971,

modifications were made to the existing pilot plant in order to carry out leach tests at

an elevated temperature of 60°C. The targeted lead content in the residue was

0.10% Pb. Initially, the pilot tests were unsuccessful and lead content in the residue

fluctuated widely with most values reported being higher than 0.10% Pb. It was

determined that the addition of lime as a flocculant aid to the final concentrate was

essentially neutralizing the acid being used in the hot leach. The operational

procedure was modified to remove the lime and the lead values in the concentrate

were reduced to range between 0.042 to 0.077% Pb during further leach tests. The

average leaching time was reduced significantly to approximately 4.3 hours with a

decrease in filtration time also realized.

AMAX PR O D U C T I O N PE R I O D 1981 TO 1982

Leach Plant Product ion Reports

A lead leach plant was in operation at the Kitsault plant during 1981 through to the

plant closure in 1982, although there is very limited information remaining from this

period. Three leach plant production reports from 1981 and three from 1982 were

reviewed. In general, during 1982, leach batches were approximately 2.4 tonnes

each with a total of 6 to 8 leaches conducted per day. Blending of leached and

unleached concentrate was introduced in March 1982 to allow for concentrateproduction that would meet the copper and lead specifications required for shipment

Various operational problems were experienced in the leach plant facility. Some of

the following constraints may be relevant to the proposed plant design:

an inadequate waste liquid handling system

a high lead concentration in the leaching plant feed material, which was 2 to

3 times higher than the designed feed value

high volumes of hot wash water handling

various mechanical problems which resulted in the bypassing of themolybdenum concentrate for packaging without first leaching the

concentrate

the filter press capacity and washing appeared to be under-designed and

was a bottle neck.





SGS Lakef ie ld Pi lot P lant Study 1983

In the 1983 pilot testwork carried out by SGS Lakefield, the lead leaching of

molybdenite concentrates was tested on a laboratory scale. Concentrate samples

created in the study were leached at 80°C for 2 hours in a 5% hydrochloric acid

solution with a 30% solids density. The leach residue was then filtered and washedwith hot water. The results obtained are shown in Table 17.15.

Table 17.15 Lead Leach Test Results — SGS Lakefield 1983

Test No.

Feed Grade Residue Grade Mo

Upgrade

(%)

Lead

Rejection

(%)Mo (%) Pb (%) Mo (%) Pb (%)

L-26 46.2 0.24 48.1 0.008 4.1 96.7

L-27 49.4 0.16 49.7 0.009 0.6 94.4

L-28 45.5 0.39 46.7 0.048 2.6 87.7

Average 47.0 0.26 48.2 0.022 2.5 92.9

The SGS Lakefield test results showed that lead rejection values of between 88 and

97% could be achieved although the highest lead content samples could not achieve

the lead objective of <0.02% Pb. The results also indicated that a molybdenum

upgrade of about 2.5% on average could be anticipated as a result of the lead leach

process.

LE A D L E A C H L I Q U O R TR E A T M E N T

A method was proposed in 1969 to recover the lead and silver from the waste acid

solution produced in the lead leaching process (Lane, 1969). The report assumedthat the lead leaching was performed with concentrated hydrochloric acid (37%

concentration), and consequently an almost unchanged acid concentration was

expected for the leach liquors. The lead and silver recovery was performed by

diluting the leach liquor with water. Lead chloride and silver chloride were

precipitated as a result of the lower solubility of the compounds in the diluted leach

liquor. In theory, when the waste leach liquor was diluted to a 9% acid strength

(concentration), approximately 84% of the lead in solution could be recovered as a

lead-chloride precipitate. However, no results were reported, and this inadequately

described lead recovery system cannot be used for the design of the lead leach

circuit.

Aluminium precipitation of the dissolved heavy metals in the leach plant liquors was

tested on site after the closure of plant operations (Kivari, 1983). The test results

have been summarized in Table 17.16. Although the precipitation efficiencies for

lead, copper, and silver were very high, the target concentration of 1 ppm was not

achieved for lead, copper, cadmium, and zinc. Re-dissolution of the precipitated

metals was given as the reason for not achieving the 1 ppm concentration objective.

No analyses were conducted for bismuth and iron.





Table 17.16 Metal Concentration of Leach Plant Liquor Before and After

Precipitation

Metal

ConcentrationPrecipitation

Efficiency

(%)

Pre-precipitation

(ppm)

Post-precipitation

(ppm)

Pb 780 33 95.6

Cu 135.0 1.8 98.7

Ag 16.0 0.29 98.2

Cd 4.1 2.9 29.2

Zn 91 89 2.2

A further test was subsequently conducted at Kitsault where leach plant liquors were

passed through aluminium weighing dishes packed in columns (Cooper, 1983). The

averaged test results are given in Table 17.17, and were promising in terms of lead,

silver, bismuth, and copper removal, but for the other heavy metals no significantprecipitation was observed. These results were similar to the precipitation results

previously obtained.

Table 17.17 Summary of Leach Liquor Heavy Metal Concentration

Metal

Concentration Precipitation

Efficiency

(%)Pre-precipitation

(ppm)

Post-precipitation

(ppm)

Pb 500 2 99.6

Cu 151.0 0.7 99.5

Ag 6.40 0.56 91.3

Bi 39 5 87.2

Cd 2.38 1.20 49.6

Zn 76 60 21.1

Mo 310 270 12.9

Fe 195 185 5.1

No further testwork related to the possible recovery of heavy metals from the lead

leach solutions appears to have been subsequently conducted. Also, no subsequent

method for the recovery of the individual metals from the precipitate was proposed.

LE A C H R E V I E W C O N C L U S I O N S

The lead leach circuit has been designed on a conceptual basis only. Although the

lead leach process essentially follows the process used by the Kitsault mine when it

was operational in 1981 to 1982, the efficacy of the process remains questionable.

As indicated by the historical testwork, a target lead grade of 0.02% Pb does not





appear to be attainable under anticipated lead contents of up to 2% in the flotation

concentrate.

Also, very little work has been undertaken on the subsequent recovery of heavy

metals from the leach liquors. Although a number of these metals, namely lead,

copper, bismuth, and silver have been shown to be potentially recoverable byprecipitation with aluminium, this process has not been quantified. The subsequent

separation and purification of each of these metals has not been demonstrated.

For purposes of the plant design, the process will be shown to include a

neutralization and filtration stage. The precipitate will be recovered by using a filter

press and the filtrate, consisting of neutralized solution, will be discharged to the

tailings pond. The conceptual lead leach circuit design has allowed for the treatment

of wash liquids and leach liquors with an adequate filtration capacity for the leached

concentrate and the residue. However, the detailed design can only be

accomplished when related testwork become available.

17.4.8 M O L Y B D E N U M AN A L Y S E S

Some of the molybdenum assay results given in the testwork reported by BC

Molybdenum in Table 17.8 are seen to be very high when compared with the

theoretical maximum amount of molybdenum in molybdenite being 59.94% Mo. A

note in the Kitsault plant monthly report dated March 1982 stated that the method of

calculating the MoS2 recovery as used by BC Molybdenum resulted a bias in the

recovery of +1.9%. This bias was attributed to the method then used by BC

Molybdenum of pulverizing oven-dried samples prior to assaying for molybdenum

oxide. No further elaboration was made regarding the extent of this error, and

whether it was a standard error or dependent on the grade of the sample. Despite

this apparent procedural analytical error, the overall conclusions drawn from the

analysis of the historical testwork remain valid.

The accuracy of the molybdenum assay results as conducted by the Kitsault mine

assay laboratory was checked during October 1982, and was reported in the month-

end report dated November 2, 1982. Concentrates from Kitsault had been analyzed

by three other laboratories and the results obtained have been detailed in Table

17.18.

Table 17.18 Comparison of Molybdenum Analyses

Laboratory

Grade (% Mo)

Difference(% Mo)Other Lab Kitsault

Brenda 53.79 54.10 +0.31

Gibralter 52.05 51.93 -0.12

Endako 54.20 54.01 -0.19

Average 53.35 53.35 0





It is interesting to note that the average Kitsault analysis value of 53.35% Mo is

identical to the average of the results obtained by the other three laboratories.

However, in the absence of a detailed statistical review, no additional information can

be obtained from this exercise, although it was used by Kitsault to help identify

possible sources of error related to their analytical procedures.

No comments could be found in the Kitsault reports regarding the comparative

accuracy and compatibility of results obtained using the atomic absorption

spectrophotometric method, or the colorimetric method, or standard titration

methods.

17.4.9 R O C K T Y PE D I S T R I B U T I O N

The 2009 SGS test program has utilized the proportion of the three major rock type

constituents as given in the process design criteria in order to prepare the composite

sample used in the testwork. However, the current (2009) Mine Plan has given a

different distribution of the rock types which will be mined and treated in the plant.Table 17.19 compares the two rock type distributions.

Table 17.19 Rock Type Distribution

Rock Type

Distribution (%)

Mine Plan SGS Testwork

Monzonite 44.8 55.0

Diorite 33.3 25.0

Hornfels 21.9 20.0

Total 100.0 100.0

Although there are significant differences in the amount of monzonite and diorite

slated for treatment, compared with the 2009 SGS testwork composite sample

distribution, there are no changes recommended for the design of the plant at this

stage of the project.

1 7 . 5 R D I R E V I E W

17.5.1 H I S T O R I C A L R E V I E W

The Kitsault mine was developed and operated by BC Molybdenum, a subsidiary of

Kennecott Copper Corporation, from 1967 through to mid-1972 when it closed for

economic reasons. AMAX acquired the property in late 1973 and operated it from

1981 to 1982 when production was again terminated due to low molybdenum prices.

An NI 43-101 PEA Study was completed for Avanti by SRK on November 3, 2008.

The recent 2009 testwork program has been designed to further the understanding of





the of the treatment characteristics of the Kitsault ore and support a pre-feasibility

level study.

In its two production campaigns of 1967 to 1972 and 1981 to 1982, the Kitsault

concentrator processed approximately 13.4 million tons of mineable ore for the

production of approximately 30 million pounds of Mo. The previous Kitsaultconcentrator was completely disassembled, salvaged, and sold after closure in 1982.

Considerable data still exists from the previous campaigns to facilitate the design of a

modern processing facility, and this was evaluated by Wardrop in the earlier part of

this report.

17.5.2 M E T A L L U R G I C A L T E S T I N G

The current metallurgical testwork was undertaken on the samples from the 2008

drilling program. The primary objectives of the testwork were

to develop and confirm a process flowsheet that would produce amarketable-grade molybdenite concentrate

to develop design criteria for the PFS including definition of the grinding

characteristics and the flotation circuit including the regrinding stages.

The grindability studies were performed at the Hazen laboratory, and the flowsheet

development studies were performed by SGS Vancouver.

SA M P L E SE L E C T I O N

The deposit has three major rock types with the projected life of mine (LOM)

distribution as indicated, namely, quartz monzonite (55%), hornfels (20%) and diorite(25%). The selection of samples for the grindability and flotation testing were based

on the following criteria:

The grade of an individual sample used for preparing the rock type

composite must meet or exceed the cut-off grade of 0.036% Mo

Rock types were limited to the three main rock types, namely hornfels,

diorite, and quartz monzonite.

These criteria resulted in the preparation of composite rock samples for the

metallurgical study as indicated in Table 17.20, and are considered to be

representative of the deposit.





Table 17.20 Composite Sample Details

Rock Type

No. Sample

Intervals

Total

Metres

No. of

Holes

Average

Grade (% Mo)

Hornfels 285 846 15 0.097

Diorite 476 1,419 11 0.102Monzonite 721 2,157 16 0.108

C O M M I N U T I O N ST U D I E S

Comminution testing was undertaken on each of the three rock types using the full

JKTech tests, including:

Drop-weight testing

Semi-autogenous mill comminution (SMC) testing

Bond Rod Mill Work Index (RWi) testing

Bond Ball Mill Work Index (BWi)

Bond Crushing Index (CWi) testing

Bond Abrasion Index (Ai) testing.

The test data obtained has been summarized in Tables 17.21, 17.22 and 17.23.

Table 17.21 Summary of Hazen Drop Weight Breakage Evaluation

Parameter Diorite Hornfels

Quartz

Monzonite

SG (by weighing in water and air) 2.64 2.70 2.64

JKSimMet Parameters

A (maximum breakage) 62.61 61.90 71.91

B (relation between energy and impact breakage) 0.94 0.51 0.72

A x b (overall AG-SAG hardness) 58.6 31.3 51.8

Ta (abrasion parameter) 0.54 0.26 0.51





Table 17.22 Summary of Hazen SMC Breakage Evaluation

Parameter Diorite Hornfels

Quartz

Monzonite

SG (by weight in water and air) 2.6 2.7 2.6

SMCT Parameters A (maximum breakage) 68.2 68.8 69.0

b (relation between energy and impact breakage) 0.9 0.5 0.8

A x b (overall AG-SAG hardness) 58.0 33.0 58.0

SMC Test DWi (kWh/m3) 4.5 8.2 4.5

DWi (%) 33.0 80.0 33.0

Mia (kWh/t) 14.8 22.6 14.6

Mih (kWh/t) 10.1 17.4 10.0

Ta 0.7 0.4 0.7

Table 17.23 Summary of Hazen BWi, RWi, CWi, and Ai Evaluation

Sample

BWi

(kWh/t)

RWi

(kWh/t)

CWi

(kWh/t) Ai (g)

Diorite 14.3 13.3 8.47 0.3846

Hornfels 15.7 20.1 8.26 0.4659

Quartz Monzonite 14.3 14.3 7.72 0.4979

The test data was analyzed by Contract Support Services Inc. (CSS) which provides

consulting services on behalf of JKSimMet for companies wishing to develop

computer simulations for grinding circuits employing autogenous or semi-autogenous

grinding (SAG). CSS used the basic test data developed at the Hazen facility to

determine the optimal grinding mill sizes and circuit configurations and conditions for

treating each of the three rock types individually, and as a blend of rock types at

varying mill feed rates. The blend of rock types used the LOM average of 55%

quartz monzonite, 25% diorite, and 20% hornfels. CSS evaluated two basic grinding

circuit configurations for each rock type, and the blend of rock types. CSS also

modelled a SAG mill/ball mill circuit without pebble crushing, and also a SAG mill/ball

mill circuit with pebble crushing (SABC). The simulations indicate the following:

It was determined that while both grinding options could successfully be

used, the provision of pebble crushing with the SABC circuit allowed for the

use of the smallest grinding mills and the least power consumption. Hence,the capital cost is the lowest using the SABC circuit.

Each circuit configuration was modeled for mill feed rates of 20,000, 30,000,

and 40,000 t/d. It was ultimately determined to size the plant for 40,000 t/d

on a blended feed of rock types. The LOM average blend was used to size

the grinding mills.





CSS recommended a single grinding line consisting of one each 34 ft x 19 ft

(10.36 m x 5.79 m) SAG mill with a 12 MW motor and one each 22 ft x 36.5 ft

(6.71 m x 11.12 m) ball mill with a 9 MW motor for the 40,000 t/d feed rate. A SAG

mill pebble crusher is required for this case.

For purposes of the plant design, Wardrop reviewed the CSS data and consideredthat the ball mill was undersized. As a result, the subsequently revised grinding

circuit design consists of a single SAG mill, 10.36 m diameter by 5.79 m in length

(34 ft x 19 ft) with an 11 MW motor, and a single ball mill, 7.32 m diameter by

12.12 m in length (24 ft x 40 ft), also with an 11 MW motor.

FL O W S H E E T D E V E L O P M E N T ST U D I E S

The flowsheet development studies, undertaken by SGS Vancouver, consisted of

optimization of the rougher flotation, cleaner flotation, cleaner scavenger flotation

(with and without regrind), and several stages of cleaner flotation to produce a

marketable grade molybdenite concentrate. The flotation concentrate may need tobe leached with hydrochloric acid (HCl) to reduce the lead content of the

concentrate. The metallurgical leaching testwork is summarized in the below section

titled ‘Molybdenum Concentrate Leach Tests’. The conceptual flowsheet evaluated

in the open-circuit flotation testing and locked cycle flotation testing is given in Figure

17.4.





Figure 17.4 Flowsheet Evaluated in the 2009 SGS Test Program

Rougher Flotat ion Tests

The rougher flotation test results, given in Table 17.24, indicate that a simple reagent

suite consisting of diesel fuel and methyl isobutyl carbonyl (MIBC) frother will recover

91 to 94% of the molybdenite in the rougher concentrate at a primary grind size P80

of approximately 250 µm. The flotation time was maintained at 19 minutes in the

laboratory testing. Most of the tests depicted in Table 17.24 were on 2 kg samplecharges. However, each rock type was also tested in the one cubic foot (28.3 L)

flotation cell. Two one cubic foot tests were conducted on each rock type. One test

was for a rougher flotation time of 18 minutes, while the other test extended the

flotation time to 24 minutes.





Table 17.24 Rougher Flotation Test Results

Test No.

Grind

P80 µm pH

Reagents (g/t)Flotation

Time (min)

ConcentrateTails

(% Mo)

Calc. Head

(% Mo)Lime Diesel Fuel MIBC Nokes Wt. (%) Mo Rec. Mo Grade

Hornfels

1 249 8 .7 30 200 30.5 80 10 2.55 86.1 3.02 0.013 0.0902 205 8.5 10 200 32 80 10 2.40 85.7 3.31 0.014 0.093

3 150 8.6 60 200 32 80 10 2.18 85.8 3.61 0.013 0.092

10 200 10.0 450 200 20 80 13 4.44 89.9 1.85 0.010 0.092

11 250 8.1 0 200 25 80 19 9.09 91.1 0.93 0.009 0.093

14 250 8.0 0 200 25 0 19 7.04 88.5 1.18 0.012 0.094

17 (1 ft3) 249 8.3 0 200 20 0 18 8.63 92.8 0.98 0.007 0.091

20 (1 ft3) 250 8.6 0 200 30 0 24 8.63 92.8 0.98 0.007 0.091

Diorite

4 235 8.6 100 200 33 80 10 2.00 88.5 3.99 0.011 0.090

5 204 8.9 70 200 33 80 10 3.21 92.6 2.73 0.007 0.095

6 151 8.6 70 200 31 80 10 3.14 92.4 2.80 0.008 0.095

12 250 7.8 0 200 30 80 19 7.46 92.8 1.23 0.008 0.104

15 250 7.8 0 200 30 0 19 6.73 93.2 1.44 0.019 0.103

18 (1 ft

3

) 255 8.3 0 200 20 0 18 5.71 82.8 1.50 0.019 0.10122 (1 ft

3) 250 8.1 0 200 30 0 24 5.13 93.1 1.83 0.007 0.101

Monzonite

7 234 8.5 88 200 28 80 10 5.25 93.3 1.90 0.008 0.107

8 190 8 .8 59 200 38.5 80 10 3.76 94.1 2.74 0.007 0.109

9 140 8.8 48 200 31 80 10 3.74 94.4 2.71 0.006 0.107

13 250 8.3 0 200 30 80 19 10.02 93.7 0.98 0.007 0.105

16 250 8.4 0 200 30 0 19 5.96 92.3 1.82 0.010 0.118

19 (1 ft3) 251 8.9 0 200 23 0 18 4.65 82.8 1.95 0.020 0.109

23 (1 ft3) 250 8.4 0 200 30 0 24 5.08 94.6 2.06 0.006 0.111





The flotation time had to be increased to 24 minutes when using the one cubic foot

(28.3 L) laboratory flotation cell to obtain a similar molybdenite recovery as was

achieved using the smaller cell in 19 minutes (Table 17.24). The split between the

rougher and scavenger concentrate for the three rock types is given in Table 17.25.

These data indicate that the rougher flotation time should be 8 minutes and the

scavenger time should be 16 minutes.

Table 17.25 10 kg Rougher/Scavenger Flotation Tests

Test No. Sample

Wt.

(%)

Mo

Recovery

(%)

Rougher/Scavenger

Concentrate

Grade (% Mo)

Hornfels (H)

20 Rougher Conc. 3.31 86.1 2.37

Scavenger Conc. 5.32 6.60 0.11

Diorite (D)

22 Rougher Conc. 2.49 88.3 3.57


Quartz Monzonite (Qtz)

23 Rougher Conc. 2.00 83.5 4.62


Blend (55% Qtz, 25% D, 20% H)

Rougher Conc. 2.40 85.22 3.40


Firs t C leaner Flotat ion Tests

The rougher flotation concentrate was subjected to first cleaner flotation and timed

concentrates were collected. The test results summarized in Table 17.26 indicate

that it is possible to obtain a reasonably good stage concentrate recovery and stage

concentrate grade with a flotation time of three minutes.

Table 17.26 First Cleaner Flotation – Three Minutes

Test No. Sample

Wt.

(%)

Rec.

(% Mo)

Conc.

Grade

(% Mo)

Tailings

Assay

(% Mo)

Calc. Rougher

Conc. Grade

(% Mo)

24 Hornfels 26.5 92.2 11.62 0.35 3.34

25 Diorite 21.5 88.3 14.70 0.53 3.5626 Quartz Monzonite 30.2 93.1 12.00 0.38 3.90

Comp. Blend 27.3 91.72 12.60 0.41 3.70

(1)Blend = 55% quartz monzonite + 25% diorite + 20% hornfels

(2)0.66% based on plant feed.





Primary Regrind Tests

The primary regrind refers to the regrinding of 1st cleaner stage tailings and the

scavenger concentrate. These streams are combined and reground prior to re-

floating in a 1st cleaner scavenger stage. The 1st cleaner scavenger concentrate is

re-circulated to the 1st cleaner stage while the 1st cleaner scavenger tailings arecombined with the scavenger tailings as final tailings. Table 17.27 indicates the

flotation recovery and concentrate grades at three different laboratory regrind times.

Table 17.27 Combined Reground Cleaner Tails and Scavenger Concentrate

Test

No.

Grind

Time

(min)

Concentrate

(Wt. %)

Molybdenum

Recovery

(%)

Concentrate

Grade

(% Mo)

Tailings

Grade

(% Mo)

Calc.

Feed

(% Mo)

Hornfels

27 7.5 12.7 79.3 2.6 0.10 0.41

28 15 13.3 84.9 2.7 0.07 0.41

29 20 11.0 83.7 2.8 0.07 0.37

Diorite

30 7.5 18.0 83.9 2.5 0.10 0.53

31 15 17.1 85.5 2.7 0.10 0.55

32 20 17.0 87.3 2.7 0.08 0.53

Quartz Monzonite

33 7.5 12.5 81.6 1.7 0.05 0.26

34 15 12.5 82.4 2.1 0.06 0.31

35 20 12.0 82.8 2.5 0.07 0.36

The three rock types behaved very similarly and a 15 minute primary regrindappeared to be satisfactory. Based on the above data, a 15 minute regrind time was

selected for the remaining tests, although this was not correlated with an actual

particle size analysis.

Open Circui t C leaner Tests

The above information was used to design open circuit flotation tests consisting of

four cleaning stages with a 15 minute primary regrind, a 4 minute regrind of the 1st

cleaner concentrate and with and without a second regrind of the 3rd cleaner

concentrate. These data are provided in Table 17.28.





Table 17.28 Open Circuit 4-Stage Cleaner Tests with 1st Cleaner Concentrate

Regrind, and With and Without 3rd Cleaner Concentrate Regrind

Sample

Test

No.

3rd Cleaner

Regrind

4th Cleaner Concentrate

Grade (% Mo) Recovery (%)

Hornfels 36 No 47.1 88.4

38 Yes 48.7 91.3

Diorite 41 No 50.0 84.7

43 Yes 54.3 84.1

Quartz Monzonite 46 No 52.0 89.1

48 Yes 52.8 92.3

This data indicates the need for regrinding the 3rd cleaner concentrate prior to the

fourth stage of flotation. All three rock types responded approximately the same. In

all cases, the concentrate grade increased with the additional stage of concentrate

regrinding. The increase in grade for the diorite composite was substantial, althoughthe recovery did go down slightly. Since the rock types have similar treatment

characteristics, it was decided to conduct the complete open circuit flotation tests on

the Master Composite. Table 17.29 indicates the open circuit test results of test 54.

This test resulted in a final product 4th cleaner concentrate grade of 55.10% Mo at

an overall molybdenum recovery of 78.1%.

Table 17.29 Open Circuit Flotation on the Master Composite

Test Product

Wt.

(%)

Assay

(% Mo)

Distribution

(%)

Mo 4th Cleaner Concentrate 9.5 55.10 78.1

Mo 3rd Cleaner Concentrate 12.2 46.90 85.4

Mo 2nd Cleaner Concentrate 16.5 36.20 89.2

Mo 1st Cleaner Concentrate 34.3 18.30 93.4

Rougher Concentrate 100.0 6.70 100.0

Locked Cycle Tests

Two locked cycle flotation tests were conducted on the Master Composite. The test

data for the two tests are presented in Table 17.30 and Table 17.31.





Table 17.30 Locked Cycle Flotation Test (LCT1) on the Master Composite

Product Wt. (%)

Assays Distribution

Mo (%) Pb (%) Ag (g/ t) Cu (%) S (%) Mo (%) Pb (%) Ag (%) Cu (%) S (%)

4th Cleaner Conc. 0.14 54.4 2.336 606 0.423 35.83 82.2 19.5 28.1 12.5 3.2

Cleaner Scavenger Tail 2.14 0.46 0.256 45.9 0.063 2.14 10.3 31.6 31.3 27.6 2.8Rougher Tail 97.72 0.01 0.009 1.3 0.003 1.55 7.5 48.9 40.6 59.9 93.9

Combined Tail 99.86 0.02 0.014 2.2 0.004 1.55 17.8 80.5 71.9 87.5 96.8

Calc. Head 100.0 0.10 0.017 3.1 0.005 1.60 100 100 100 100 100

Table 17.31 Locked Cycle Flotation Test (LCT2) on the Master Composite

Product Wt. (%)

Assays Distribution

Mo (%) Pb (%) Ag (g/ t) Cu (%) S (%) Mo (%) Pb (%) Ag (%) Cu (%) S (%)

4th Cleaner Conc. 0.22 42.4 5.680 1153 0.900 29.6 86.1 51.4 55.4 26.0 4.5

Cleaner Scavenger Tail 3.26 0.25 0.141 25.8 0.057 2.27 7.3 18.6 18.2 24.0 5.0

Rougher Tail 96.52 0.01 0.008 1.3 0.004 1.38 6.6 30.1 26.4 50.0 90.5

Combined Tail 99.78 0.02 0.012 2.1 0.006 1.40 13.9 48.6 44.6 74.0 95.5

Calc. Head 100 0.11 0.25 4.6 0.008 1.47 100 100 100 100 100





LCT1 was able to achieve the target molybdenum concentrate grade of

approximately 54% Mo. However, the molybdenum recovery was unacceptably low

at only approximately 82.2%. The two places for losses of molybdenum to occur are

in the scavenger tailings and the 1st cleaner scavenger tailings. The scavenger

tailing accounted for only 7.5% of the lost molybdenum and the 1st cleaner

scavenger tailings accounted for an additional 10.3% for a total loss of recovery of 17.8% molybdenum. The 1st cleaner scavenger tailings assayed 0.46% Mo. This is

approximately four times the molybdenum grade achieved in the open circuit tests.

The circuit conditions were changed for LCT2. The regrind time and flotation time for

the 1st cleaner scavenger was increased for LCT2. This decreased the grade of the

1st cleaner scavenger tailings from 0.46 to 0.22% Mo and reduced the molybdenum

loss in this stream from 10.3 to 7.3%. However, the final product molybdenum

concentrate grade was unacceptably low at 42.4% Mo.

Additional locked cycle flotation tests must be done to optimize both the final product

4th cleaner concentrate grade and recovery. Both LCT1 and LCT2 were run with six

cycles, and the average of the last three cycles was used to project the finalconcentrate grade and recovery. The new locked cycle tests should be run for more

cycles, possibly for up to nine cycles. RDi believes that once the locked cycle test

conditions are optimized, the data will reflect a marketable concentrate grade of

approximately 54% Mo at an overall molybdenum recovery of 88 to 90%. Most

primary molybdenum operations achieve molybdenum recoveries of approximately

95 to 97% in the cleaning circuit. The test data indicate a rougher flotation recovery

of approximately 93% Mo. The combined recovery, then, should be between 88 and

90%. The optimizing of the process parameters in the 1st cleaner scavenger circuit

should be sufficient to achieve the target recovery. The future test work should also

evaluate the benefits of regrinding the 4th cleaner concentrates and adding a 5th

cleaner stage of flotation.

Molybdenum Concentrate Leach Tests

Copper, lead, and silver are all relatively high in the final product cleaner

concentrate. The copper assay in the final concentrates for LCT1 and LCT2 was

0.423 and 0.900% Cu respectively. Copper would not appear to present problems

since higher copper contents are generally accepted in molybdenum concentrates.

Lead appears to pose a problem in meeting marketable concentrate grades. The

former 1981/1982 plant operation included a hydrochloric acid leach of the final

molybdenum concentrate in order to reduce the lead concentration.

One hydrochloric acid leach test was conducted on the combined 4th cleaner concentrates from the open circuit tests using all three rock types. The combined

concentrate was leached at 35% solids with 5% HCl solution at 80ºC for 8 hours.

The lead assay in the combined cleaner concentrates was 0.60% Pb and this was

reduced to 0.11% Pb in the leached residue. The molybdenum assay in the

combined cleaner concentrates was 48.6% Mo and this increased to 50.4% Mo in

the leached residue. Previous leach test work at SGS on the molybdenum

concentrate generated in the pilot plant indicated the lead in the leached residue to





be less than 0.05% Pb. Lead removal in the leach circuit was also acceptable in the

commercial circuit at Kitsault.

However, the results obtained from the SGS locked cycle tests have indicated high

levels of lead in the final concentrate, namely up to 5.68% Pb in the final concentrate.

In order to reduce the lead content of the final concentrate to the market specificationof <0.1% Pb, this would require a lead removal efficiency in excess of 95%, a figure

not attained in the historical testwork.

A possible explanation for the high lead content in the concentrate is that the rougher

concentrate is not reground but feeds the 1st cleaner stage directly, and coarse

particles which contain galena only see the first regrinding stage when they are

refloated into the 1st cleaner concentrate. It then follows that any galena liberated

during each of the regrinding stages has no ready access to being discarded with the

tailings. Together with the results obtained in the testwork conducted in 1977, this

may imply that the existing regrind philosophy may not be correct and that

successive stages of regrinds with every cleaner stage should be designed. It is

therefore imperative that additional flotation tests which include regrinding stages

and which follow the lead behaviour in the flotation circuit be conducted.

Additional lead leach tests should also be conducted in the next phase of testing to

optimize the process parameters, and to ensure that the lead content be within

market specified values in order that no smelter penalties be imposed.

Variabi l i ty Test ing

A total of 27 samples of quartz monzonite, diorite, and hornfels were tested to

determine the variability of samples of varying feed grade and from a variety of

locations within the orebody. Samples were selected to cover a range of feed gradesfrom 0.049% Mo to 0.279% Mo based on the calculated head grade from each test.

The samples were also selected from areas of varying depths within the orebody.

Only rougher flotation tests were run using the standard 19 minute grinding time

used from the previous test program. All of the tests used the standard 200 g/t of

diesel oil and MIBC. The data suggests that the variability in response was not due

to the varying feed grades. However, the samples were of varying hardness since

the standard 19 minute grind yielded a wide range of P80 values. The flotation

response was related to liberation size. The complete results can be found in the

SGS report provided in Appendix C, which details the variability rougher kinetic test

results.

The data indicates that there are no areas in the mine that are likely to create

problems in the plant. As long as the target rougher flotation grind of approximately

80% passing 250 µm is achieved, all of the ore should respond favourably in the

plant.





Pyri te and Sulphide Flotat ion f rom Final Tai l ings

The combined tailings (rougher tailing plus 1st cleaner scavenger tailings) from LCT1

were used to conduct two flotation tests to determine the feasibility of floating pyrite

from the tailings stream (test 56) and floating all of the contained sulphides (test 57)

from the tailings stream. Both tests used potassium amyl xanthate (PAX) as thecollector reagent. Test 56 used 2 minutes of flotation time. Test 57 used 7 minutes.

The tests are summarized in Table 17.32.

These tests demonstrate the feasibility of removing a substantial amount of sulphides

(82 to 86%) from the bulk of the tailings stream making it possible to deposit the high

sulphide tailing in a separate area of the tailings facility in which it can remain in a

sub-aqueous environment. It may also be possible to produce a saleable by-product

but that would require further testwork.





Table 17.32 Pyrite and Sulfide Flotation Test Data

Test Product Wt. (%)

Assays Distribution

Pb (%) Ag (g/ t) Fe (%) S (%) Pb (%) Ag (%) Fe (%) S (%)

56 Pyrite Rougher Concentrate 4.12 0.40 58 29.24 30.6 74.1 75.7 45.2 82.4

Pyrite Rougher Tails 95.88 0.006 0.8 1.52 0.28 25.9 24.3 54.8 17.6Calculated Head 100.0 0.02 3.2 2.66 1.53 100.0 100.0 100.0 100.0

57 Pyrite Rougher Concentrate 4.87 0.32 55 23.44 24.6 84.5 93.4 44.8 86.3

Pyrite Rougher Tails 95.13 0.003 0.2 1.48 0.2 15.5 6.6 55.2 13.7

Calculated Head 100.0 0.02 2.9 2.55 1.39 100.0 100.0 100.0 100.0





Potent ia l By-product Recovery f rom the Rougher/Scavenger Tai l ings

Nine tests were conducted on the cleaner/scavenger concentrate tailings to

determine whether there is the potential to recover by-product Pb, Zn, Ag, Bi, and

WO3 by gravity or flotation. Table 17.33 indicates the results of each of the three

rock types to gravity concentration using a Knelson centrifugal concentrator. By-

product silver recovery from the 1st cleaner scavenger tailings yielded very poor

results.

Table 17.33 By-product Recovery by Gravity Concentration

Sample

Test

No.

Feed

(g/t Ag)

Concentrate RecoveryConcentrate

Grade (g/t Ag)Wt. (%) Ag (%)

Hornfels 37 6.0 38.5 28.2 4.40

Diorite 42 32.1 29.0 29.8 33.00

Quartz Monzonite 47 9.13 24,8 22.5 8.30

By-product recovery by flotation was investigated in six tests. Each of the three rock

types were tested using either SIPX or AP3477 collector reagent. Table 17.34 and

Table 17.35 indicate the results of the by-product flotation tests using either SIPX

(Table 17.34) or AP3477 (Table 17.35). The tests were not optimized and the results

are similar. However, AP3477 appears to be the better reagent for the diorite.

Table 17.34 By-product Recovery using AP3477 for Flotation

Sample TestNo.

Feed

Grade(g/t Ag) AP3477(g/t)

Flotation Conc.

Recovery Flotation

Conc. Grade(g/t Ag)Wt. (%) Ag (%)

Hornfels 39 15.7 50 15.77 81.8 82.30

Diorite 44 49.9 50 10.34 43.7 211.15

Quartz Monz. 49 14.61 50 15.59 78.0 73.13

Table 17.35 By-product Recovery using SIPX for Flotation

Sample

Test

No.

Feed

Grade

(g/t Ag)

SIPX

(g/t)

Flotation Conc.

Recovery Flotation

Conc. Grade

(g/t Ag)Wt. (%) Ag (%)

Hornfels 40 10.4 50 14.49 79.4 57.00

Diorite 45 56.76 50 6.66 16.6 141.68

Quartz Monz. 50 16.6 50 20.57 82.7 66.67

The diorite sample composite had significantly higher silver grades than the other

two rock types. The silver assay in the cleaner scavenger tailing averages 46.3 g/t





Variability testing of the three primary ore types using the process flowsheet

as established from the testwork in order to confirm the results of the

flotation procedure established.

Pilot plant testing of the process flowsheet developed is required in order to

confirm the results obtained from the locked cycle tests. This will alsogenerate environmental samples, larger concentrate samples, and will

provide the quantities of tailings samples required for additional appropriate

testing. These samples would also be used for settling and filtration tests,

including flocculant screening. The concentrate sample produced will be

used for minor/impurity element determinations which are required to

characterize the presence of potential smelter penalty elements.

Definition of the lead leach circuit is required, including the treatment of

leach liquors. The nature of the contaminating lead particles should be

characterized.

Determination of the amount and type of deleterious and smelter penalty

elements present in the final concentrate.

Standard settling tests for the determination of a thickener area for the

concentrate products.

Standard filtration testing of the final concentrate produced.

The use of site water in the flotation evaluation tests is also recommended.

High Pressure Grinding Rolls (HPGR) testwork is recommended on each of

the major rock types comprising the Kitsault deposit.





1 8 . 0 M I N E R A L P R O C E S S I N G

1 8 . 1 I N T R O D U C T I O N

The Kitsault concentrator has been designed to process a nominal 40,000 t/d of

molybdenite-bearing ore. The concentrator will be designed to produce a marketable

molybdenum flotation concentrate containing 52% Mo at an overall recovery of

90.6%.

1 8 . 2 S U M M A R Y

The unit processes selected were based on historical plant data from the periods

when the mine was operational, namely 1966 to 1967, and 1980 to 1982, as well as

metallurgical testwork performed over the years 1960 to 1990 by Climax, amongst

others. In addition, the results of metallurgical testing performed at SGS Canada Inc.

(SGS) in Vancouver, BC, during 2009, was also used. Resources set out by Avanti

were also employed.

The metallurgical processes selected have been designed to produce a saleable

high grade molybdenum flotation concentrate containing 52% Mo. The unit

processes employed for the recovery of the mineral molybdenite will involve

conventional size reduction and mineral beneficiation methods. Since by-product

mineral recovery test results have been inconclusive to date, this has been excluded

from the process design. The tailings disposal design allows for the flotation tailings

to be deposited in a conventional tailings pond. Water utilization will be based on

maximizing re-use of process water as water reclaimed from the tailings facility.

Fresh water will only be used for gland service, reagent preparation, and cooling of

hydraulic units.

The process plant will consist of the following operations and facilities:

run-of-mine (ROM) ore hopper

primary crushing

conveyors

crushed ore stockpile

crushed ore reclaim

a SAG and ball mill grinding circuit incorporating cyclones for classification

SAG mill pebble crushing circuit





molybdenum rougher flotation

molybdenum scavenger flotation, thickening, and regrinding

molybdenum cleaner flotation stages including regrinding

molybdenum concentrate thickening concentrate leaching, filtration, drying, and bagging

concentrate dispatch to off-site roaster facility

scavenger flotation tailings to tailings pond

tailings pond reclaim water system.

A simplified process flowsheet is given below as Figure 18.1.

The detailed process flowsheets are located in Appendix D as Drawings A00-09-01

to A00-09-10, and the simplified flowsheet as Drawing A00-09-020.





Figure 18.1 Simplified Process Flowsheet





1 8 . 3 M A J O R D E S I G N C R I T E R I A

The concentrator was designed to process 40,000 t/d, equivalent to

14,600,000 t/year. The molybdenum head grade value used in the design was

0.093% Mo. The major process design criteria parameters used in the design are

outlined in Table 18.1. The grinding data is based on weighted averages of the three

rock types being treated, namely monzonite (55%), diorite (25%), and hornfels

(20%).

Table 18.1 Major Process Design Criteria

Criteria Unit

Operating Year d 365

Crushing Availability % 80.0

Grinding and Flotation Availabili ty % 92.0

Primary Crushing Rate t/h 2,083.3Milling and Flotation Process Rate t/h 1,811.6

SAG Mill Feed Size, 80% Passing µm 150,000

SAG Mill Transfer Size, 80% Passing µm 1,500

Ball Mill Circulating Load % 250

Ball Mill Grind Size, 80% Passing µm 250

SAG Breakage Parameter A x b 49.4

SAG Breakage Parameter ta 0.47

Bond Ball Mill Work Index kWh/t 14.58

Bond Abrasion Index g 0.4632

Concentrate Regrind, 80% Passing µm 24

Flotation Concentrate Grade % Mo 52.0Overall Flotation Recovery % 90.6

The complete process design criteria, together with the material balance, are detailed

in Appendix D.

The grinding mills were sized based on the SAG ore breakage parameters provided

by Hazen Research Inc. (Hazen) and evaluated by CSS. The Bond Work Index data

was also provided by Hazen. Wardrop conducted its own evaluation of the CSS and

modified the ball mill size accordingly. The Isa-type regrind mills were sized by

Xstrata plc (Xstrata) based on treatment rates supplied by Wardrop, and utilizing

Xstrata experience.

The flotation cells were sized based on flotation times as provided by Avanti from the

current testwork program being undertaken by SGS. Typical scale-up factors have

been applied.





1 8 . 4 P L A N T D E S I G N

18.4.1 O P E R A T I N G SC H E D U L E A N D AV A I L A B I L I T Y

The complete comminution and processing system will be designed on the basis of two 12-hour shifts per day, for 365 d/year. The plant feed rate will be 40,000 t/d.

The gyratory crusher availability will be 80%, and the grinding and flotation circuit

availability will be based on a running time of 92%, both based on 365 working days

per year. A stockpile with a live capacity of 40,000 t will take up the production surge

between the crushing and grinding throughput rates. These running times will allow

sufficient downtime for the scheduled maintenance of the crushing equipment, and

the equipment in the plant.

1 8 . 5 P R O C E S S P L A N T D E S C R I P T I O N

18.5.1 P R I M A R Y C R U S H I N G

A conventional gyratory crusher facility will be designed to crush ROM ore to reduce

the size of the rocks in preparation for the grinding process. The design crushing

rate will be 2,083.3 t/h.

The major equipment and facilities in this area includes:

crusher feed hopper

rock breaker

gyratory crusher; 1,372 mm x 1,905 mm

crusher discharge conveyor apron feeder; 1,829 mm x 8,000 mm

crushed ore stockpile; 40,000 t live capacity

stockpile reclaim apron feeders; 1,220 mm x 6,000 mm, four

conveyor belts, and belt tear detectors

dust collection system.

The ROM ore will be delivered to the crusher feed hopper by 177 t capacity dump

trucks. The ore will be crushed by the primary gyratory crusher, which will be choke-

fed from the feed hopper. A rock breaker will reduce the size of the oversize materialin the feed hopper. The crusher feed hopper will have a live capacity of 400 t,

sufficient to handle the load from two dump trucks.

The gyratory crusher will reduce the size of the feed rock to a P 80 of 150 mm with the

maximum lump size being 250 mm. The crusher product will discharge onto a

conveyor via an apron feeder to regulate the conveyor feed. The crusher discharge





conveyor will deliver the crusher product to the crusher ore stockpile via the stockpile

feed conveyor. Both conveyor belts will be 1,524 mm wide.

The live capacity of the crushed ore stockpile will be 40,000 t, equivalent to the

production rate for one day.

The crusher facility and belt transfer points will be equipped with a dust collection

system to control fugitive dust that will be generated during the crushing and

conveying operations.

18.5.2 C R U S H E D ORE ST O C K P I L E A N D R E C L A I M

The crushed ore stockpile will have a live capacity of 40,000 t. The ore will be

reclaimed from this stockpile by apron feeders at the nominal rate of 1,811.6 t/h. The

apron feeders will feed a 1,524 mm wide SAG mill feed conveyor, which in turn feeds

the SAG mill. The conveyor belt will be equipped with a belt scale to monitor and

control the feed to the SAG mill.

The crushed ore stockpile reclaim feeders will be equipped with a dust collection

system to control fugitive dust that will be generated during conveyor loading and the

transportation of the ore to the SAG mill.

18.5.3 G R I N D I N G A N D C L A S S I F I C A T I O N

The grinding circuit will consist of a SAG-ball mill combination circuit. It will be a

2-stage operation with the SAG mill in closed circuit with a pebble crusher, and the

SAG mill and ball mill product in closed circuit with the classification cyclones. The

SAG mill will be equipped with a trommel screen to remove pebbles for crushing.

The grinding will be conducted at a nominal rate of 1,811.6 t/h of new feed material.

The grinding and classification circuit will include the following equipment:

SAG mill feed conveyor

SAG mill feed conveyor belt weigh scale

SAG mill: 10,359 mm diameter x 5,791 mm long, 11 MW motor

ball mill: 7,320 mm diameter x 12,190 mm long, 11 MW motor, variable

frequency drive (VFD)

belt magnet and metal detector

pebble crusher: 3,500 mm diameter, 600 kW motor

mill discharge distribution box

two mill discharge pumpboxes

two sets – cyclone cluster feed slurry pumps

two sets – classification cyclone clusters





two particle size analyzers

two mass flow meter systems.

The ore from the crushed ore stockpile will be reclaimed under controlled feed rate

conditions using apron feeders. Up to three apron feeders will be operational at anytime, with the fourth unit being a standby unit. The apron feeders will discharge the

material from the stockpile onto a conveyor belt feeding the SAG mill. A belt scale

will control the feed rate to the SAG mill. Water will be added to the SAG mill feed

material to assist the grinding process at 72% solids.

The SAG mill will be 10,359 mm in diameter, and have a length of 5,791 mm. It will

be equipped with an 11 MW motor. The SAG mill will operate at a critical speed of

78%.

The SAG mill discharge will have a grate opening size of 65 mm. A trommel screen

with 13 mm apertures will remove the screen oversize material. The underflow will

be discharged to the mill discharge distribution box where it will be combined with theball mill discharge material. The trammel screen oversize material will be washed

with sprays and conveyed via transfer conveyors to a pebble crusher. A belt magnet

with a metal detector will ensure that no tramp metal is fed with the pebbles to the

crusher. This 3,500 mm cone crusher will crush the pebbles to a P80 of 13 mm. It

will be equipped with a 600 kW motor, and will crush a pebble circulating load of

25%, equivalent to 452.9 t/h. The crushed material will be returned to the conveyor

belt feeding the SAG mill for further grinding. A crusher bypass facility will be

incorporated in the system to cater for the event where the pebble crusher may have

been stopped for maintenance, but the grinding circuit is still operational.

The discharge from the ball mill will also be discharged into the common mill

discharge distribution box which will split the combined slurry into two streams, each

feeding a mill discharge pumpbox. The slurry in each of the two mill discharge

pumpboxes will be pumped to a cyclone cluster for classification. The cut size for the

cyclones will be at a size of P80 of 250 µm, and the circulating load will be 250%.

The cyclone underflow from each cyclone cluster will be returned to the ball mill as

feed material, with the option to return the underflow to the SAG mill via a grinding

circuit distribution box. The cyclone feed pumps will be equipped with VFD motors.

Ten cyclones will be installed in each cluster, of which nine will be operating and one

will be a standby unit. The overall milling rate will be 1,811.6 t/h and this will

constitute the feed rate to the molybdenum flotation circuit. Dilution water will be

added to the grinding circuit to maintain the density at 70% solids.

The ball mill will have a diameter of 7,320 mm and a length of 12,190 mm. It will be

equipped with an 11 MW VFD motor. The ball mill will operate at a critical speed of

76%.





Each cyclone overflow stream will be monitored by a particle size analyzer. The pulp

density will be 36% solids. The two product streams will be discharged to the

flotation feed conditioning tank ahead of the flotation process

Provision will be made for the addition of lime to the SAG mill feed for the adjustment

of the pH of the slurry in the grinding circuit prior to the flotation process if required.Similarly, diesel fuel collector reagent will be added to the SAG mill feed. The

addition of Nokes reagent to the SAG mill feed material will be provided for as an

option in the event that lead sulphide mineral depression will be required.

Grinding media will be added to the mills in order to maintain the grinding efficiency.

Steel balls will be added periodically, to each mill, using a ball charging kibble for

each mill. The SAG mill grinding media will be 125 mm in diameter, and the ball mill

will use 75 mm diameter steel balls.

18.5.4 F L O T A T I O N C I R C U I T

The milled ore will be subjected to flotation to recover the mineral molybdenite into a

high-grade molybdenum metal concentrate. Conventional tank cells for the rougher

and scavenger stages, and mechanical and column cells for the cleaner stages, will

be utilized.

The molybdenum flotation circuit will include the following equipment:

conditioning tank with agitator: 4,600 mm diameter x 5,200 mm

flotation reagent addition facilities

five rougher conventional flotation tank cells: 160 m3

nine scavenger conventional flotation tank cells: 160 m3

scavenger regrind thickener: 14 m diameter

scavenger concentrate Isa-mill regrind mill: M3000 unit

variable arrangement rougher/scavenger concentrate launder system

four 1st cleaner conventional flotation tank cells: 5 m3

four 1st cleaner scavenger conventional flotation tank cells: 10 m3

1st cleaner Isa-mill regrind mill: M500 unit

1st cleaner regrind cyclone feed pumps

1st cleaner classification cyclone cluster

four 2nd cleaner conventional mechanical flotation cells: 1.5 m3

3rd cleaner column flotation cell: 1,200 mm diameter x 4,800 mm

3rd cleaner Isa-mill regrind mill: M500 unit

3rd cleaner regrind cyclone feed pumps





3rd cleaner classification cyclone cluster

4th cleaner column flotation cell: 1,000 mm diameter x 4,000 mm

pumpboxes and standpipes

slurry and solution pumps particle-size monitors

sampling system.

The overflow from the two streams from the classification cyclones in the grinding

circuit will be the feed to the molybdenum rougher flotation circuit. The slurry

streams will be combined in the flotation conditioning tank for a design feed rate of

1811.6 t/h. The slurry will be conditioned for 1 minute at a slurry density of 36%

solids. Diesel fuel as collector reagent, and Dowfroth 250 (DF250) frother, will be

added to the conditioning tank. Provision will be made for the staged addition of the

reagents to the rougher and scavenger flotation circuits.

Although nominally having a conditioning time of 1 minute, the conditioning tank will

mainly serve to combine and thoroughly mix the two cyclone overflow product

streams, and to distribute the flotation reagents more evenly in the slurry prior to the

rougher stage.

The conditioned slurry will overflow into the rougher flotation bank of cells. The

molybdenum rougher concentrate will be recovered and discharged to the 1st

cleaner flotation circuit via the rougher concentrate standpipe. A relatively short

rougher flotation residence time will be used in order to produce a high grade

rougher concentrate which will be discharged to the 1st cleaner stage for further

upgrading without regrinding.

The rougher stage tailings will feed the scavenger flotation stage. Tailings from the

1st cleaner scavenger stage may be returned to the head of the scavenger circuit to

join the rougher tailings stream. The scavenger flotation circuit will have a longer

residence time of 18 minutes to maximize the recovery of molybdenum. The

scavenger concentrate will be discharged to a thickener where it will be combined

with the 1st cleaner stage tailings. The thickener will dewater the two flotation

products to produce an underflow with a slurry density of 60% solids which will be the

feed to the first regrinding stage. This scavenger concentrate regrind mill will be an

M3000 Isa-mill. The scavenger regrind mill will reduce the particle size from a feed

P80 value of about 250 µm, to a product P80 value of 120 µm.

The tailings from the scavenger circuit will constitute the final plant tailings together

with the 1st cleaner scavenger tailings. This will be discharged to the tailings dam

via the final tailings pumpbox.

The scavenger regrind thickener overflow will be re-combined with the scavenger

regrind mill product for delivery to the 1st cleaner scavenger flotation circuit. The

concentrate from the 1st cleaner scavenger cells will be pumped to the head of the





1st cleaner flotation stage to be combined with the rougher concentrate and the 2nd

cleaner tailings streams. The tailings from the 1st cleaner scavenger stage will either

be discharged to the final tailings pumpbox as final tailings or, depending on the

grade, will be returned to the head of the scavenger flotation circuit.

To completely liberate fine sized grains of molybdenite from the gangue constituentsand to enhance the upgrading of the molybdenum, a series of three regrinding

stages will be incorporated in the scavenger and cleaner flotation circuits. As

mentioned above, the scavenger concentrate will be reground in the scavenger

concentrate regrind mill. Two further regrind stages will be incorporated into the

cleaner flotation circuit. The concentrate from the 1st cleaner stage will be reground

in the 1st cleaner regrind mill. This will be an M500 Isa-mill and will reduce the

particle size from a P80 feed size of about 120 µm to a product size P 80 of 50 µm.

The concentrate from the 3rd cleaner column cell will be reground in the 3rd cleaner

regrind mill. This will also be an M500 Isa-mill and will reduce the particle size from a

P80 feed of about 50 µm to a product size of 24 µm. Both M500 Isa-mill regrind

circuits will incorporate densification cyclones to increase the feed density to the millas cyclone underflow, with the mill product being re-combined with the cyclone

overflow for delivery to the following stage of flotation.

The feed material to the 1st cleaner circuit will be the rougher concentrate, the

tailings from the 2nd cleaner stage, and the 1st cleaner scavenger concentrate. The

concentrate from the 1st cleaner stage will be the feed for the 1st cleaner regrind mill

circuit. The reground 1st cleaner concentrate material will feed the 2nd cleaner

flotation circuit. Diesel fuel collector and DF250 frother reagents will be added to the

1st cleaner circuit and the 1st cleaner scavenger circuit as required. The Nokes

reagent will also be added to both circuits if lead mineral depression is required.

The 2nd cleaner flotation stage will be fed by the reground 1st cleaner concentrate,together with the tailings from the 3rd cleaner stage. The 2nd cleaner flotation

concentrate will be directed to the 3rd cleaner column flotation cell via a standpipe.

The 2nd cleaner tailings will be discharged into a pumpbox for transfer to the 1st

cleaner flotation circuit. Diesel fuel collector will be added to this stage by addition

into the 1st cleaner concentrate pumpbox before the regrinding stage. Diesel fuel

can also be added directly to the 2nd cleaner flotation stage if required. Frother

reagent will be added directly to the 2nd cleaner stage flotation cells.

The 3rd cleaner flotation stage will consist of a column cell equipped with a recycle

pump for agitation and air dispersion. The 3rd stage feed will be made up of the

concentrate from the 2nd cleaner stage and tailings from the 4th cleaner stage.

Flotation concentrate from the 3rd cleaner column cell will be reground in the 3rd

cleaner regrind mill circuit. Tailings from the 3rd cleaner column cell will be pumped

to the 2nd cleaner stage via a pumpbox. Provision will be made for diesel fuel

collector reagent to be added directly to the 3rd and also the 4th cleaner column

cells, if required. DF250 frother will be added to both column cells.





The concentrate from the 3rd cleaner column cell will feed the 4th cleaner flotation

stage via the 3rd cleaner regrind mill circuit. The 4th cleaner concentrate produced

will be the final molybdenum concentrate and will feed directly to the molybdenum

concentrate thickener. Tailings from the 4th cleaner column cell will be returned to

feed the 3rd cleaner column cell together with the 2nd cleaner concentrate.

Particle size analysers will be installed to monitor the primary grinding circuit cyclone

overflow products from each classification stream, and to control the primary grinding

circuit. Particle size analyzers will also be installed in each of the three regrind

circuits to monitor the particle size of the respective reground products. Samplers

will be installed to automatically sample the following streams:

flotation feed

final tailings after the scavenger stage

scavenger regrind product

1st cleaner scavenger tailings

2nd cleaner feed

2nd cleaner concentrate

3rd cleaner tailings

4th cleaner concentrate.

Conventional flotation tank cells will be used for the rougher, the scavenger, the 1st

cleaner, and the 1st cleaner scavenger stages. Conventional mechanical cells will

be used in the 2nd cleaner flotation stage. Column cells will be used for the 3rd

cleaner and 4th cleaner flotation stages.

Molybdenum concentrate thickener overflow water will be sent directly to the process

water tank for re-use in the flotation process as dilution water in launders, sprays and

make-up water.

18.5.5 M O L Y B D E N U M C O N C E N T R A T E T H I C K E N I N G

The molybdenum flotation cleaner concentrate will be collected in the concentrate

thickener. The thickener underflow will be pumped to the lead leach circuit for the

removal of lead contaminants. The molybdenum concentrate thickener circuit will

have the following equipment:

concentrate thickener: 4.8 m diameter

overflow standpipe with pumps

concentrate thickener underflow slurry pumps.

The 4th cleaner product will be the final molybdenum flotation concentrate having a

grade of 52% Mo. This will be discharged to the concentrate thickener to increase





the product density prior to the leaching stage to reduce the amount of lead impurity

in the molybdenum concentrate. The concentrate feed density to the concentrate

thickener will be 15% solids, and the thickener underflow density will be 60% solids.

The thickened slurry will be pumped to the first of three leach tanks in the lead leach

circuit. The concentrate thickener will have a diameter of 4.8 m. Flocculant will be

added to facilitate settling of the molybdenum concentrate.

The concentrate thickener overflow will be collected in the thickener overflow

standpipe. The solution will be pumped from the standpipe to the process water tank

for re-use within the grinding and flotation circuit.

18.5.6 L E A D L E A C H C I R C U I T

The lead leach circuit will reduce the concentration of lead impurities in the

molybdenum concentrate to acceptable limits for sales purposes. The equipment

used in the lead leach circuit will be the following:

three lead leach tanks with agitators: 2,400 mm diameter x 2,700 mm

acid mixing tank with agitator: 2,200 mm diameter x 2,400 mm

acid mixing pumps.

A lead leach circuit has been incorporated in the design of the plant. During the time

when the mine was operating between 1981 and 1982, the molybdenum concentrate

was found to contain unacceptably high lead concentrations. A lead leach facility

was subsequently incorporated into the process plant flowsheet. The lead leach

circuit involved the leaching of the concentrate with 5% hydrochloric acid solution

maintained at the relatively high temperature of 85°C. The leaching time required

was determined to be 2 hours and leaching was conducted on a batch basis. Theleached concentrate was washed and filtered, then dried and bagged for shipping off

the property. The acid was recycled and brought up to the required strength again

before being re-used in the leaching circuit. The resulting molybdenum concentrate

was within specification with respect to lead, while the copper and zinc contents were

also reduced in the leach process. An amount of up to 90% of the lead present in

the concentrate was removed during the leaching step.

The present lead leach circuit design will be the following. A three-tank continuous

leach circuit has been designed. Concentrate will be pumped to the first of three

leach tanks using the concentrate thickener underflow pumps. The slurry density will

be 60% solids. The concentrate will be contacted with a hot solution of 5% strength

hydrochloric acid which will be pumped from the acid mixing tank. The slurry density

in the leach tanks will be 33% solids. The slurry will be kept in suspension with the

leach tank agitators. Steam will be used to maintain the temperature of the slurry in

the leach tanks. Each leach tank has been designed for a residence time, or

leaching time, of 2 hours, for a total nominal leaching time of 6 hours. The slurry will

overflow to the next leach tank in line. On completion of the leach cycle, the leached

concentrate slurry will be filtered and washed using a filter press. The filter press will





be fed by the filter press feed pumps directly from the final leach tank, which will

serve as a filter feed surge tank. The washed filter cake, which is the leached

molybdenum concentrate, will be discharged onto a conveyor belt, which will feed the

concentrate dryer and bagging system.

The filtrate and wash solution from the filter press will be discharged to the acidmixing tank for re-use in the leach circuit after adjusting the acid concentration to 5%

acid strength with concentrated hydrochloric acid. Filtrate will also periodically be

bled off out of the leach circuit to reduce the build-up of concentration of the metals.

This filtrate will then be discharged to the final tailings pumpbox for mixing and

disposal with the flotation tailings.

It is possible that the lead leach circuit may not be required to operate continuously.

If a low lead content molybdenum concentrate is produced, possibly as a result of the

particular ore zone having a low lead concentration, or that the Nokes reagent is

successful at reducing the amount of lead present in the molybdenum concentrate,

the lead leach circuit may not be required to operate continuously.

18.5.7 M O L Y B D E N U M C O N C E N T R A T E D R Y I N G

The leached molybdenum concentrate will be filtered, dried, and bagged prior to

dispatch off the mine property. The equipment used for the filtration, drying and

bagging of the concentrate will be the following:

concentrate filter press feed pumps

concentrate filter press: 1,000 mm x 1,000 mm

concentrate dryer: holoflite

concentrate storage bin: 20 t capacity

concentrate bagging and dispatch facilities

concentrate dust collection system.

After the leaching step, the molybdenum concentrate will be filtered using a pressure

filter. The filter feed pumps will deliver the slurry to the concentrate filter press.

Since filtration with a filter press unit will be a batch process, the third leach tank will

also act as a surge tank for the filtration operation. The filter press will dewater the

concentrate to produce a final concentrate with a moisture content of about 8%.

Compressed air will be used for the core blow of the concentrate filter cake. A hotwater wash will wash out the remaining entrained hydrochloric acid. After the air

blow and wash cycles, the filter cake will be dropped out of the filter press onto a

conveyor belt which will transfer the filter cake to the concentrate dryer feed hopper.

The feed to the concentrate dryer will be regulated in order that a dry molybdenum

concentrate will be obtained. A holoflite dryer will be installed for this purpose. The

dry concentrate will be transferred to a dry concentrate storage bin which will feed





the bagging system. Bulk 1,000 kg bags will be used for the dispatch of the final

concentrate product.

The filtrate from the filtration step will be returned to the acid mixing tank for re-use of

the unused acid, while the bleed solution will be discharged to the tailings pond.

A dust collection system will ensure that any dust generated in the molybdenum

concentrate handling area will be collected and recycled.

18.5.8 T A I L I N G S H A ND LI N G AN D D I S P O S A L

The flotation tailings will be deposited in the tailings pond. The tailings handling

circuit will have the following equipment:

final tailings pumpbox

final tailings pumps.

The scavenger circuit flotation tailings will be the final plant tailings. The tailings from

the 1st cleaner scavenger flotation stage will also be considered to be final tailings

and will be discharged to the final tailings pumpbox, unless sampling of this stream

indicates that it should be recycled to the scavenger circuit. The final tailings will be

discharged into the final tailings pumpbox and will be pumped directly to the tailings

pond.

Water will be reclaimed from the tailings pond via a reclaim water pump station

system with the water being returned to the process water tank for distribution as

required within the process plant.

18.5.9 R E A G E N T H A ND LI N G AN D STORAGE

Various chemical reagents will be added to the process slurry stream to facilitate the

flotation process.

The preparation and distribution of the various reagents will require the following

equipment:

bulk handling system

mixing and holding tanks

metering pumps

transfer pump

flocculant preparation facility

lime slaking and distribution facility

eye-wash and safety showers





applicable safety equipment.

Various chemical reagents will be added to the grinding and flotation circuit to modify

the mineral particle surfaces and enhance the floatability of the mineral particles for

its collection into a high grade concentrate product. Fresh water will be used in the

making up and/or the dilution of the various reagents that will be supplied in solidform, or which require dilution prior to the addition to the slurry. These solutions will

be added to the points of addition of the flotation circuit using metering pumps.

Diesel fuel will be added as the collector reagent. It will be delivered in bulk via a

trailer in 30,000 L batches. It will be transferred from the trailer to the holding tank.

The diesel fuel will not be diluted and will be pumped directly from the holding tank to

the various points of addition using metered pumps.

The Nokes reagent will be delivered as a 20% concentration strength in bulk totes. It

will be pumped directly from the tote to the point of addition without dilution.

The frother reagent DF250 will be delivered in bulk via a trailer in 20,000 L batches.

It will be transferred directly from the trailer to the holding tank. The DF250 will not

be diluted and will be pumped directly from the holding tank to the various points of

addition using metered pumps.

Flocculant will be prepared in the standard manner as a dilute solution with 0.30 %

solution strength.

Lime will be delivered in bulk by 40 t trailer and will be off-loaded pneumatically into a

lime silo. The lime will then be prepared in a lime slaking system as a 20%

concentration slurry. This lime slurry will be pumped to the points of addition using a

closed loop system. The valves will be controlled by pH monitors which will controlthe amount of lime added.

Concentrated hydrochloric acid will be delivered in bulk containers. It will be pumped

from the container to the lead leach circuit acid mixing tank where it will be diluted to

the required strength.

To ensure containment in the event of an accidental spill, the reagent preparation

and storage facility will be located within a containment area designed to

accommodate 110% of the content of the largest tank. In addition, each reagent will

be prepared in its own bunded area in order to limit spillage and facilitate its return to

its respective mix tank in the event of spillage occurring. The storage tanks will be

equipped with level indicators and instrumentation to ensure that spills do not occur during normal operation, or solution make-up, or transfer.

Appropriate ventilation, fire and safety protection, and Material Safety Data Sheet

(MSDS) stations will be provided at the facility. Each reagent line and addition point

will be labelled in accordance with Workplace Hazardous Materials Information

Systems (WHMIS) standards. All operational personnel will receive WHMIS training,





along with additional training for the safe handling and use of the reagents, and

spillage clean-up.

1 8. 5. 10 AS S A Y A N D M E T A L L U R G I C A L L A BO RA TOR Y

The assay laboratory will be equipped with the necessary analytical instruments to

provide all routine assays for the mine samples, the the samples generated by the

concentrator, and the environmental department. The most important of these

instruments includes:

atomic absorption spectrophotometer (AAS)

inductively-coupled plasma (ICP) mass spectrometer

x-ray fluorescence spectrometer (XRF)

Leco furnace.

The metallurgical laboratory will undertake all necessary testwork to monitor metallurgical performance and, more importantly, to improve process flowsheet unit

operations and efficiencies. The laboratory will be equipped with laboratory-sized

crushers, ball and stirred mills, particle size analysis sieves, flotation machines with

various cell sizes, filtration devices, balances, and pH meters.

1 8. 5. 11 W A T E R SU P P L Y

Two separate water supply systems for fresh water and process water will be

provided to support the operation.

FRESH WATER SUPPLY S YSTEM

Fresh and potable water will be supplied to a fresh/fire water storage tank from the

catchment dams and/or wells. Fresh water will be used primarily for the following:

firewater for emergency use

cooling water for mill motors, and crusher and mill lubrication systems

gland service for the slurry pumps

reagent make-up

boiler for hot water and steam production potable water supply.

The fresh/fire water tank will be equipped with a standpipe which will ensure that the

tank contains at least 40 m3, equivalent to a 2-hour supply of firewater, for any

emergency. The fire water distribution system will direct the fire water to various

parts of the plant.





The potable water from the fresh water source will be treated and stored in the

potable water storage tank prior to delivery to various service points.

PROCESS WATER SUPPLY S YSTEM

Process water generated in the flotation circuits as concentrate thickener overflow

solution will be returned to the process water tank. The scavenger concentrate

thickener overflow will be re-used in the regrind circuit but any excess will be

delivered to the process water tank for re-use in the plant. Excess water will be

discharged to the final tailings pumpbox, if necessary. Reclaimed water will be

pumped from the tailings pond to the process water tank for distribution to the points

of usage.

1 8. 5. 12 A I R SU P P L Y

Separate air service systems will supply air to the following areas:

Flotation: low pressure air for flotation cells will be provided by air blowers.

Filtration: high pressure air for the concentrate filter press will be provided by

the plant compressors.

Crushing: high pressure air for the dust suppression system and other

services will be provided by a separate air compressor.

Instrumentation: instrument air will come from the plant air compressors and

will be dried and stored in a dedicated air receiver.





1 9 . 0 O T H E R R E L E V A N T D A T A A N D

I N F O R M A T I O N

1 9 . 1 M I N I N G

The Kitsault property is located at about 140 km north of Prince Rupert, BC, and

south of the head of Alice Arm, an inlet of the Pacific Ocean. The project siteelevation is situated at about 600 masl.

The deposit will be mined by open pit mining to provide a mill feed at a nominal rate

of 40,000 t/d (14.6 Mt/a). At the mine’s operating peak, 29.0 Mt/a of material will bemined annually, with a LOM waste-to-ore ratio of about 0.75:1.

The operation will have a primary equipment fleet consisting of two 269 mm blastholedrills, two 18 m3 electric hydraulic shovels, one 18 m3 front end loader, andseventeen 177-t haul trucks.

The primary equipment will be supported by track- and rubber-tired dozers, motor graders, a compactor, a water truck, a small excavator, and other auxiliaryequipment.

The ore and waste material will be mined in 10-m benches to achieve 20-m double

benches. The overall mining sequence was developed through a series of four mining phases, or pushbacks. The sequence was designed to:

gain early access to the higher grade ore to facilitate early capital recoverywhile providing waste rock for the tailing embankment during construction

minimize waste stripping by careful positioning of haul roads and by

maintaining smooth waste to ore ratio

develop an ore blending strategy to maximize the NPV by mining higher grade ore during the initial years and stockpiling low-grade ore for later processing

integrate the stripping and delivery of non-potentially acid-generating(NPAG) and PAG waste rock for integration with the construction of the

tailing embankment and inter-mixture with the TMF.

integrate the diversion of the Patsy Creek along the south wall of the open

pit.

The general arrangement drawing for the mine, primary crusher, concentrator,ancillary structures, and overall tailing facility is presented in Figure 19.1.





Figure 19.1 General Arrangement





19.1.1 P I T O P T I M I Z A T I O N

G E O L O G Y R E S O U R C E B LOCK MO D E L

The resource model “Kitsault model3 ASCII May 29, 2009” was used in the pitoptimization. The size of a block is 10 m x 10 m x 10 m. The total extent of themodel is outlined in Table 19.1.

Table 19.1 Block Model Limits

Coordinates Y X Z

Minimum 6140100 471610 -210

Maximum 6143500 474970 1160

The model was validated by comparing the measured and indicated ore resourcesand the volumes for the different rock types with the May 7 model. The validationconfirmed that the May 29, 2009 model closely matches the previous resource

model.

G E O T E C H N I C A L SL O P E R E C O M M E N D A T I O N S

The pit slope parameters as applied in pit optimisation and mine design are based on

the recommendations of SRK Consulting Inc. (SRK).

SRK recommended an inter-ramp slope angle (ISA) of 51°, 10-m double bench(20 m high), and average sustainable bench widths of 9 m and 70° mean bench face

angles. The recommendation is based on SRK’s experience that ISAs that yield a30% failure probability for slopes with low-to-moderate failure consequences, andabout 10% failure probability for high failure consequences, are appropriate for most

open pit mines.

SC A L E O F PR O D U C T I O N – BA S E C AS E

The objective of the preliminary optimization is to select the production rate option

that will provide the best economic return. The results of this run will be used insubsequent optimization of the ore cut-off grade to further increase the NPV. Four

production rates, ranging from 25,000 t/d to 40,000 t/d at 5,000 t/d increment were

analyzed.

Base Case – Opt imizat ion Parameters

Model data from the 3D resource model kits_model.txt were imported andmanipulated in Gemcom Surpac™ (Surpac) mine planning software, and thenexported to the Gemcom Whittle™ v4.1.3 (Whittle) pit optimization software. Only





the measured (Class 1) and indicated (Class 2) resource categories of the resourcemodel were used. The metal values for inferred (Class 3) resource were categorized

as waste.

Wardrop calculated the long-term metal price and the exchange rate; Avanti provided

the treatment and concentrate freight charge projections. A projected metal price of Cdn$14.79/lb (US$12.51/lb) Mo was used. The currency exchange rate wasforecast to be US$1.00:Cdn$1.182. There was no value attributed to secondary

metals for the optimization analysis. It should be noted that the pit optimization was

carried out in Canadian dollars.

Operating costs for mining, processing, and G&A were estimated for each of the

production rate options. Different rates of production from recent mining studies

provided the reference costs to extrapolate the costs for each of the four options. Additional data from the 2008 mining reference guide and SRK’s PEA costs providedfurther rationale to the extrapolated costs. As shown in Table 19.2, the projected

cost for the highest production rate option (Option 1 at 40,000 t/d) is a mining cost of $1.56/t mined, a processing cost of $4.87/t milled, and a G&A cost of $0.72/t milled.

The projected cost for the lowest production rate option (Option 4 at 25,000 t/d) is a

mining cost of $1.74/t mined, a processing cost of $5.50/t milled, and a G&A cost of $0.90/t milled. A process recovery of 90% was assumed for all options.

Table 19.2 shows the preliminary technical and financial parameters for the

optimization.

Table 19.2 Whittle Base Optimization Parameters

Optimization Parameters Units Option 1 Option 2 Option 3 Option 4

Daily Mill Capacity t/d 40,000 35,000 30,000 25,000Million Tonnes per Year Mt/a 14.60 12.78 10.95 9.13

Reference Mining Cost $/t mined 1.56 1.61 1.65 1.74

Mining Recovery 1 1 1 1

Mining Dilution 1 1 1 1

Ore Processing Cost $/t milled 4.87 5.21 5.41 5.50

G&A Costs $/t milled 0.72 0.77 0.84 0.90

Closure/Environmental Allocation $/t milled 0.19 0.19 0.19 0.19

Metal Price (Mo) US$/lb 12.51 12.51 12.51 12.51

Selling Cost (Mo) US$/lb 1.22 1.22 1.22 1.22

Exchange Rate (Cdn$:US$) 1.182 1.182 1.182 1.182

Mill Recovery (Mo) % 90 90 90 90Freight

Ocean/Trans./Stevedoring US$/dmt 79.1 79.1 79.1 79.1

table continues…





Optimization Parameters Units Option 1 Option 2 Option 3 Option 4

Smelter Terms

Concentrate Mo Grade % 53 53 53 53

Mo Deduction % Mo 1.0 1.0 1.0 1.0

Mo Roasting Charge US$/lb Mo 1.00 1.00 1.00 1.00Marketing (% of Concentrate Value) % 1.00 1.00 1.00 1.00

Insurance (% of Concentrate Value) % 0.15 0.15 0.15 0.15

Moisture Content % 10 10 10 10

Specific Gravity – Ore t/m3

as per block modelSpecific Gravity – Waste t/m

3

Overall Pit Slope(s) ° 47 47 47 47

Discount Rate % 8 8 8 8

Alcoa Royalty % 1 1 1 1

Impact & Benefits – Royalty % 0.50 0.50 0.50 0.50

Initial Capital Cost Estimate (000 Cdn$) 552,000 494,000 434,000 392,000

Initial Capital Cost Estimate (000 US$) 467,005 417,936 367,174 331,641

The same reference mining studies were used to estimate the initial capital costs for

the mining equipment and pre-production work for each production option.

An overall slope angle of 47° was used in the pit optimization. The slope includes aprovision for a haul road access to the pit bottom.

Base Case – Prel iminary Opt imizat ion Resul ts

Whittle uses the Lerchs-Grossman algorithm to maximize the total NPV of an oredeposit in order to develop a discounted optimal pit outline. The Whittle optimization

was carried out on a 3D resource block model using the technical and financial

parameters in Table 19.2 to calculate the positive economic blocks.

As the metal prices are stepped from low to high values, a set of nested pit outlineswith corresponding values and tonnages was generated, as outlined in Table 19.3.

The smaller values produce the inner pit shells to indicate the best positions for the

initial mining sequence.

Figure 19.2 shows a representation of the tonnages and values indicated in Table

19.3, and illustrates where the potential final pit shells are in the graph. Pit 16 was

selected as the optimal pit shell to be used for the production rate optimization for Option 1. The pit shell reported a total of 190.75 Mt of ore at 0.084% Mo and awaste-to-ore ratio of 0.39 to 1.0.

An optimization was performed for all other options and production rates, using thesame approach; optimal pit shells were subsequently selected. Based on eachoption’s optimal pit shell, an appropriate optimal production schedule was generated





from the Whittle scheduling option consisting of Option 1 at 40,000 t/d, Option 2 at35,000 t/d, Option 3 at 30,000 t/d, and Option 4 at 25,000 t/d.

A base case financial analysis was conducted for each of the options with the results

as shown in Table 19.5 and with the details shown in Appendix E. The analysis

indicated that Option 1, at 40,000 t/d, generated the highest NPV atCdn$1.037 billion. Option 4, at 25,000 t/d, generated the lowest NPV at Cdn$803 M,for a variance of up to Cdn$234 M.

Table 19.3 Nested Pits – Base Optimization

FinalPit

Discounted Cash Flow for Different Pit Sizes

Best Case(Cdn$) Input (t)

StripRatio

Grade

Input(Mo %)

Mine

Life(years)

1 (239,357,556) 12,134,650 0.38 0.134 1

2 (158,456,166) 16,089,136 0 .39 0.13 1

3 (80,584,305) 20,068,293 0 .38 0.127 1

4 (6,026,691) 24,520,719 0.39 0.123 2

5 223,929,047 40,316,488 0.36 0.113 3

6 338,728,928 49,643,212 0.35 0.109 3

7 482,517,778 63,579,023 0.33 0.104 4

8 671,370,516 83,665,001 0.31 0.099 6

9 760,816,243 94,553,660 0.31 0.097 6

10 856,165,597 108,481,229 0.31 0.094 7

11 920,956,731 119,056,596 0.31 0.093 8

12 965,942,807 127,152,914 0.32 0.091 9

13 1,008,222,874 135,283,035 0.32 0.09 9

14 1,126,407,989 161,929,171 0.35 0.087 11

15 1,161,317,230 170,948,244 0.36 0.086 12

16 1,227,158,931 190,753,945 0.39 0.084 13

17 1,257,771,180 201,067,226 0.4 0.083 14

18 1,275,966,536 207,145,334 0.42 0.082 14

19 1,292,319,583 213,727,074 0.44 0.082 15

20 1,301,543,327 217,605,425 0.44 0.081 15

21 1,316,615,722 224,335,097 0.46 0.081 15

22 1,326,811,665 229,445,138 0.46 0.08 16

23 1,349,398,285 239,812,658 0.51 0.08 16

24 1,355,960,790 243,555,991 0.52 0.079 17

25 1,362,090,737 247,254,470 0.53 0.079 17

26 1,371,028,501 251,602,457 0.56 0.079 17

27 1,375,927,363 254,824,937 0.57 0.079 17

28 1,379,889,801 257,754,517 0.58 0.078 18

29 1,384,784,881 261,814,710 0.58 0.078 18

table continues…





Final

Pit


Best Case

(Cdn$) Input (t)

Strip

Ratio

Grade

Input(Mo %)

Mine

Life(years)

30 1,389,255,859 265,080,780 0.6 0.078 1831 1,391,692,712 267,221,630 0.6 0.078 18

32 1,394,787,084 270,025,270 0.61 0.077 18

33 1,398,483,049 273,486,450 0.62 0.077 19

34 1,402,365,356 277,828,070 0.64 0.077 19

35 1,408,087,494 283,901,990 0.67 0.076 19

36 1,410,224,701 287,164,740 0.68 0.076 20

37 1,412,398,119 290,833,650 0.69 0.076 20

38 1,414,940,219 295,488,990 0.71 0.075 20

39 1,419,250,376 306,029,790 0.76 0.074 21

40 1,422,287,397 316,233,890 0.77 0.073 22

41 1,423,029,710 319,182,240 0.78 0.073 2242 1,423,886,994 323,510,780 0.81 0.073 22

43 1,424,264,741 326,225,140 0.83 0.072 22

44 1,424,408,627 328,801,480 0.83 0.072 23

45 1,424,427,617 332,654,120 0.85 0.072 23

46 1,424,387,478 333,578,660 0.85 0.072 23

47 1,424,119,733 336,421,760 0.87 0.071 23

48 1,423,869,801 338,529,490 0.88 0.071 23

49 1,423,419,238 340,766,550 0.91 0.071 23

50 1,422,523,078 344,335,430 0.93 0.071 24

51 1,421,963,156 346,287,100 0.94 0.071 24

52 1,419,531,826 351,570,130 0.98 0.07 2453 1,417,785,110 354,071,090 1 0.07 24

54 1,415,333,246 357,131,130 1.04 0.07 24

55 1,413,295,418 359,166,650 1.07 0.069 25

56 1,408,909,600 362,694,320 1.13 0.069 25

57 1,405,494,693 364,778,580 1.17 0.069 25





Figure 19.2 Potential Final Pit Shells in the Base Optimization

F I N A L C A SE – CU T -O F F O P T I M I Z A T I O N

To meet project strategic objectives and establish tailing dam waste requirements,

and maximum tailing storage capacity, a further optimization was conducted on thecut-off grade. An overall cut-off grade of 0.036% Mo was subsequently establishedfor all options. Table 19.4 shows the set of nested pits with corresponding values

and tonnages and illustrated in Figure 19.3. Pit 24 was selected as the optimal pit

shell to be used for the final production rate optimization. The pit reported a total of 190.6 Mt of ore at 0.094% Mo and a waste-to-ore ratio of 0.94 to 1.0.

An optimization was performed for all other options and production rates, using the

same approach; optimal pit shells were subsequently selected. Four optimalproduction schedules with corresponding capital and operating costs were developed

for each option. All optimizations were carried out in Canadian dollars. A financialanalysis was repeated for each option with the results as shown in Table 19.5 (the

complete analyses are provided in Appendix E).

As in the base optimization, similar economic results were obtained with Option 1, at40,000 t/d, which again provided the highest NPV at Cdn$1.264 billion. Option 4 at

25,000 t/d again provided the lowest NPV at Cdn$990 M, for an NPV variance of

Cdn$274 M between Options 1 and 4.





Table 19.4 Nested Pits – Optimized Cut-off Scenarios

Final

Pit


Best

Case

(Cdn $) Input (t) Waste (t)

Strip

Ratio

Mine

Life

(years)

1 (239,299,851) 10,802,805 5,899,385 0.55 1

2 (160,287,902) 14,128,564 8,166,186 0.58 1

3 (82,110,760) 17,617,388 10,145,872 0.58 1

4 (5,928,103) 21,323,070 12,804,270 0.6 1

5 228,033,950 34,829,341 19,809,789 0.57 2

6 343,382,643 42,309,845 24,820,095 0.59 3

7 492,696,578 53,208,654 31,236,376 0.59 4

8 691,166,314 70,510,287 38,784,473 0.55 5

9 786,889,287 79,903,089 43,749,721 0.55 5

10 888,431,605 91,325,439 50,481,871 0.55 6

11 958,043,545 99,547,532 56,896,268 0.57 7

12 1,006,256,339 105,712,506 61,664,484 0.58 7

13 1,052,459,542 112,104,240 66,856,170 0.6 8

14 1,184,380,641 132,792,559 86,299,451 0.65 9

15 1,223,749,825 139,448,013 93,667,167 0.67 10

16 1,299,231,179 153,674,759 111,572,251 0.73 11

17 1,333,151,467 160,884,191 121,281,989 0.75 11

18 1,355,368,395 165,318,239 129,320,581 0.78 11

19 1,374,414,369 169,599,159 137,101,621 0.81 12

20 1,385,353,337 172,474,509 141,237,231 0.82 12

21 1,402,873,704 177,101,518 149,373,862 0.84 12

22 1,415,185,774 180,537,545 155,306,215 0.86 12

23 1,442,666,665 188,025,045 174,496,805 0.93 13

24 1,450,673,850 190,552,564 180,065,546 0.94 13

25 1,458,632,525 193,046,003 185,717,247 0.96 13

26 1,470,056,874 196,463,730 196,498,500 1 13

27 1,476,063,328 198,734,750 201,587,870 1.01 14

28 1,480,700,514 200,648,970 205,859,900 1.03 14

29 1,486,547,309 203,318,960 211,435,350 1.04 14

30 1,492,027,747 205,564,150 218,652,660 1.06 14

31 1,494,955,919 207,088,560 221,263,050 1.07 14

32 1,498,749,514 209,008,270 226,000,710 1.08 14

33 1,503,363,013 211,396,190 232,991,780 1.1 14

34 1,507,455,020 213,759,790 240,240,210 1.12 15

35 1,513,864,630 217,891,900 254,137,700 1.17 15

36 1,516,132,279 219,528,830 260,284,120 1.19 15

37 1,517,254,641 220,501,960 263,472,820 1.19 15

38 1,520,617,752 223,753,730 274,967,900 1.23 15

table continues…





Table 19.5 Comparison of Base Case versus Final Case

Items Units Option 1 Option 2 Option 3 Option 4

Discount Rate % 8.0 8.0 8.0 8.0

Initial Capital Cdn$ M 552.0 494.0 434.0 392.0

Base Case - Preliminary Optimization

Operating Cost Cdn$/t milled 8.44 9.00 9.53 9.89

Mining Cost Cdn$/t mined 1.91 1.99 2.11 2.26

Ore t 190,753,945 190,109,744 195,180,967 192,651,904

Waste t 74,493,064 80,274,445 90,247,734 89,474,585

Stripping Ratio t/t 0.39 0.42 0.46 0.46

Mo Grade % Mo 0.084 0.085 0.085 0.085

Pre-Tax NPV Cdn$ 000 1,037,197 981,658 898,764 803,296

Mine Life years 14.0 15.0 18.0 22.0

Final Case – Cut-off Optimization


Mining Cost Cdn$/t mined 1.84 1.95 1.99 2.15

Ore t 190,552,564 191,946,897 181,054,239 189,279,493

Waste t 180,065,547 196,453,804 180,658,611 207,813,037

Stripping Ratio t/t 0.94 1.03 1.00 1.10

Mo Grade % Mo 0.093 0.094 0.096 0.096

Pre-Tax NPV Cdn$ 000 1,264,449 1,181,839 1,090,786 990,987

Mine Life years 14.0 15.0 17.0 21.0

Case Differences: Final – Base


Mining Cost Cdn$/t mined (0.07) (0.04) (0.12) (0.11)

Ore t (201,381) 1,837,153 (14,126,728) (3,372,411)

Waste t 105,572,483 116,179,359 90,410,877 118,338,452

Mo Grade % Mo 0.009 0.009 0.011 0.011

Pre-Tax NPV Cdn$ 000 227,252 200,181 192,022 187,691





Figure 19.4 Base Case and Optimized Cut-off – NPV Analysis

Table 19.5 shows that further optimization of the cut-off grade resulted in thefollowing:

Although waste stripping increased by 105.6 Mt from Base Case Option 1 to

Final Case Option 1, the %Mo head grade increased by 10.7% from 0.084%Mo to 0.093% Mo. The ore tonnage changed very little but the increase inhead grade resulted in additional pre-tax (Federal and Provincial taxes) NPV

of about Cdn$227.3 M.

On the basis of NPV analysis (Table 19.5 and Figure 19.4), the Final Case

Option 1 at 40,000 t/d provided the highest NPV, and therefore is the optimaloption for the project. A Decision Analysis (Figure 19.5) was also

undertaken to determine how each option ranks with respect to satisfyingmajor variables such as waste rock availability for the tailing dam, LOM, and

NPV. The results of the Decision Analysis determined that Option 1 ranked

the highest.

There is sufficient space within the permit area to accommodate theplacement of tailing, low-grade ore, waste dump, and the associated

infrastructure. In addition, a higher pre-production stripping and ore

stockpiling in the pre-feasibility study optimization stage will be required toensure full production in Year 1. The possibility of a higher production

throughput should be considered a project opportunity to be pursued in thenext phase of study.





Figure 19.5 Decision Analysis

ST R A T E G I C PL A N N I N G A N D O P T I M I Z A T I O N – PR E-F E A S I B I L I T Y ST U D Y

A production rate of 40,000 t/d (Option 1) has been analyzed and the mine operatingcosts updated for the PFS. Table 19.6 outlines design parameters that were used for

the purposes of this study.

Table 19.6 Design Parameters for Pit Optimization

Design Parameters Units

Daily Mill Capacity t/d 40,000

Mt/a 14.60

Reference Mining Cost $/t mined 1.84

Mining Recovery 1

Mining Dilution 1

Coast Adjustment Factor (CAF) n/a

Ore Processing Cost $/t milled 4.87

G&A Costs $/t milled 0.72

Closure & Environ. Costs Allocation $/t milled 0.19

Selected Method Cut-off

Metal Price (Moly) US$/lb 12.51

Selling Cost (Moly) US$/lb 1.22

Exchange Rate (CDN$ : US$) 1.182

Mill Recovery (Moly) % 89%

Specific Gravity – Ore t/m3 as per block model

Specific Gravity – Waste t/m3 as per block model

Overall Pit Slope(s) degrees 47

Discount Rate % 8%

INITIAL CAPEX 000 Cdn$ 552,000

INITIAL CAPEX 000 US$ 467,005





D I S C O U N T E D O P E N P I T VA L U E S

A molybdenum price of US$12.51/lb was used for all years of the mine life. Using a

mill capacity of 14.6 Mt/a, discounted cash flow analyses were performed for all theoptimized pit shells. Three scenarios were assumed for the analyses: the best case

scenario, the specified case scenario, and the worst case scenario. The best caseassumed that it is possible to mine nested pit shells from the smallest pit thatcontains the highest grades of ore to the largest pit that contains the lowest grades of ore in the right sequence. The specified case assumes mining in four phases

(pushbacks) and fixes lead between them. The worst case assumes that it is notpossible to get into lower benches before pushing back the top bench to its final limit.

Table 19.7 summarizes the results of the analysis.

Table 19.7 Discounted Open Pit Cash Flow for Different Sizes

Discounted Cash Flow

Fin al Be st Case Sp e ci fie d C ase W ors t Case Ore Waste Min e li fe

pi t Cd n $ Cdn $ Cdn $ ton ne to nn e years

1 (434,651,421) (434,651, 421) (434,651,421) 3,659,571 1,477,569 0

2 (288,724,905) (288,724, 905) (288,724,905) 9,546,568 3,592,612 1

3 (136,956,666) (136,958, 374) (136,958,374) 16, 342,550 7,574,950 1

4 82,728,843 81,923,447 77,422,181 29,102,556 12,824,804 2

5 284,857,189 279,965,117 271,732,630 42,513,905 20,007,085 3

6 555,105,856 548,226,197 526,836,501 65,297,761 30,525,199 4

7 712,834,603 699,447,391 668,602,179 81,292,388 41,001,992 6

8 803,778,708 788,114,588 746,936,163 92,365,410 48,392,800 6

9 905,834,841 885,337,742 832,488,953 106,338,314 59,736,366 7

10 985,221,097 959,425,126 898,418,663 119,528,650 71,271,780 8

11 1,066,668,263 1,032 ,126, 147 961, 001,213 134,268,807 86,723,733 9

12 1, 128,152, 628 1,088 ,107, 149 1,004, 068, 575 147, 488,424 102, 657,986 10

13 1, 163,631, 581 1,118 ,754, 910 1,025, 631, 003 155, 509,173 115, 319,467 11

14 1, 191,997, 897 1,141 ,639, 124 1,038, 480, 748 162, 904,297 128, 314,123 11

15 1, 210,830, 977 1,155 ,711, 103 1,045, 514, 441 168, 400,917 138, 337,513 12

16 1, 226,181, 821 1,165 ,680, 196 1,048, 760, 794 173, 670,253 147, 938,657 12

17 1, 236,102, 882 1,171 ,966, 540 1,050, 720, 190 177, 293,413 155, 796,777 12

18 1, 242,750, 065 1,175 ,409, 251 1,050, 132, 527 180, 266,743 160, 952,767 12

19 1, 263, 469, 312 1,182 ,913, 227 1,043, 270, 623 189, 753,923 185, 036,527 13

20 1, 268,237, 533 1,186 ,005, 175 1,041, 468, 228 192, 524,153 190, 543,697 13

21 1, 276,337, 460 1,189 ,405, 187 1,035, 151, 092 197, 548,043 204, 193,947 14

22 1, 279,860, 147 1,189 ,722, 363 1,029, 893, 375 200, 094,623 212, 102,057 14

23 1, 282,339, 328 1,189 ,979, 682 1,026, 206, 035 202, 465,013 218, 516,287 14

24 1, 282,817, 873 1,189 ,974, 658 1,025, 215, 418 203, 196,403 219, 745,457 14

25 1, 284,122, 859 1,188 ,986, 607 1,020, 142, 218 204, 841,653 225, 543,647 14

26 1, 285,258, 741 1,188 ,181, 014 1,015, 501, 954 206, 778,553 231, 750,657 14

27 1, 286,024, 603 1,186 ,598, 194 1,010, 102, 786 208, 591,893 238, 077,327 14

28 1, 286,480, 806 1,183 ,914, 108 1,001, 882, 365 210, 732,093 246, 498,617 14

29 1,286,643,209 1,181 ,479, 578 994, 938,633 212,400,283 253,290,117 15

30 1,286,606,792 1,180 ,679, 899 993, 037,571 212,783,353 255,000,607 15

31 1,286,325,012 1,178 ,636, 261 988, 217,549 213,652,773 258,896,407 15

32 1,285,909,781 1,176 ,472, 873 982, 950,255 214,423,963 263,248,547 15

33 1,285,401,208 1,173 ,999, 199 977, 665,046 215,309,093 267,186,667 15





Figure 19.6 shows the results of this analysis graphically. All of the pit shells outlinedin Table 19.7 but last three (Pit 1, Pit 2, and Pit 3) provide a positive discounted pit

value. The maximum NPV for the Best case scenario occurs for Pit 29, for theSpecified case scenario occurs for Pit 23, and for the Worst case scenario occurs for

Pit 17.

Figure 19.6 Nested Pits – Discounted Cash Flow for Different Pit Sizes

Figure 19.6 shows that, after Pit 19, there is no significant increase in NPV. Since

the TMF storage capacity is limited to approximately 190 Mt, the pit will be available

to accept discharge tailing after mining activities have concluded. Pit 19 is chosenfor further detail pit design.

19.1.2 O P E N P I T D E S I G N

M I N E D E S I G N PA R A M E T E R S

Pit Opt imizat ion Shel ls

As the optimized pits are used as the basis for the design of the pit phases, the

mining width module of Whittle was a necessary tool to apply in approximating

practical operating width between phases. An optimal mining schedule wasdeveloped during the final case cut-off optimization that simulates the sequence of

mining the selected pit shells from initial to final pushbacks. Four pit shells wereselected to guide the phase designs.





Geotechnical S lope Recommendat ions

As stated in Section 19.2.6, SRK recommended a 51° ISA, 20-m double benches, 9-m sustainable bench widths, and 70° bench face angles. The phase design closelyfollows the slope recommendations.

Haul Road and Mining Width

A preliminary equipment analysis was conducted to determine the sizes of shovels

and trucks to be used in the mining operations. The operating width for the trucks

will define the haul road width for the phase design. The operating range of theshovel in conjunction with the haul truck will define the width of each pushback as

well as the number of pushbacks.

The equipment analysis indicated that either a fleet of 22 m3 electric cable shovelsand 177-t trucks, or a fleet of 18 m3 electric hydraulic shovels and 177-t trucks will

provide the lowest present value costs to haul ore to the primary crusher, low-gradeore to the stockpile, and waste rock to the waste dump and tailing dam. The fleet of

electric hydraulic shovels and 177-t trucks was selected to be more suitable to mine

the 10-m benches.

The 177-t truck modelled in the equipment analysis is the Cat 789C haul truck. The

effective width of the haul road was estimated to be 3.5 times the operating width

(7.6 m) of the Cat 789C truck. With the addition of a 2.5-m safety berm and 1 mdrainage ditch, the total road width was established at about 30 m. Roads weredesigned with a maximum gradient of 10%.

A minimum operating width of 60 m was used to design the phases to accommodate

double-side loading of the 177-t truck with an 18 m3

electric hydraulic shovel tomaximize loading productivity.

W A S T E D U MP DE S I G N C R I T E R I A

The proposed waste dump will be situated in the upper reaches of the TMF,

downstream of the diversion dam and tunnel inlet, and west of the open pit. As the

waste rock is expected to consist of about 70% PAG rock, the waste dump will beconstructed in 5-m lifts (similar to the assumptions made in SRK’s 2008 PEA) toimprove compaction. The compaction will minimize surface water percolating

through the dump, and therefore will minimize acid generation. The final overall

slope will be established at 3H:1V, or about 18.4° with an 8-m berm for every lift tofacilitate future reclamation of the waste dump. The waste rock will be dumped at

the angle of repose of about 37°.





LO W- G R A D E O R E ST O C K P I L E C R I T E R I A

The proposed low-grade ore stockpile (LGO stockpile) will be situated north of the

open pit in the reclaimed Clary Dump area (approximately 15 Mt) and east of theopen pit on the top of existing Patsy Dump area (approximately 10 Mt). The

proposed design criteria will be similar to the waste dump design: the stockpile willbe constructed in 5-m lifts at a final overall slope at 3H:1V, or about 18.4° with an8-m berm for every lift. The low grade ore will be dumped at the angle of repose of about 37°.

PH A S E D E V E L O P M E N T A N D M I N E A B L E R E S E R V E S

The initial phases are designed to mine the highest grade ore in order to:

generate higher revenues during the first few years of mill production to

facilitate early recovery of capital

minimize stripping to reduce initial mining capital and maintain a better balance in the truck haul fleet

integrate the stripping and delivery of NPAG and PAG waste rock with thetailing facility

integrate the diversion of the Patsy Creek along the south wall of the openpit.

Table 19.8 shows a summary of the mineable reserves for all phases. Phase

developments 1 through 4 are provided in Appendix F.

Phase 1

Phase 1 was designed to develop the haul roads that will connect with the access tothe primary crusher, LGO stockpile to the east and tailing dam to the northwest. The

phase will be developed down to 490 m elevation to mine a total ore of 32.3 Mt at0.100% Mo, and a waste-to-ore ratio of 0.49/1.0. A total of 48.3 Mt of material will be

mined. Phase 1 will contribute the highest ore grade and represents about 15.0% of

the total ore.

Phase 2

Phase 2 was designed to expand the pit further to the north and south. The haul

roads to the primary crusher, LGO stockpile and tailing dam will remain the same.The phase will be developed down to 440 m elevation to mine a total ore of 47.3 Mt

at 0.091% Mo and a waste-to-ore ratio of 0.27/1.0. A total of 59.9 Mt of material will

be mined. Phase 2 will contribute the second highest grade ore and representsabout 22% of the total ore.





Phase 3

Phase 3 was designed to reach the ultimate pit limit to the northeast, east, and southside of the pit. It also established a 30 m width bench to accommodate a 1%gradient diversion channel in the south wall to divert the Patsy Creek away from the

pit. The phase will be developed down to 320 m elevation to the south to mine atotal ore of 97.6 Mt at 0.080% Mo and a waste-to-ore ratio of 0.95/1.0. A total of 190.5 Mt of material will be mined. The pit is at its largest in Phase 3, which

represents about 45% of the total ore.

Phase 4

Phase 4 is the final phase; it was designed to reach the ultimate pit limit to thenorthwest. It will be developed down to 230 m elevation to mine a total ore of

38.1 Mt at 0.077% Mo and a waste-to-ore ratio of 1.06/1.0. A total of 78.6 Mt of

material will be mined. Phase 4 will represent about 18.0% of the total ore.





Table 19.8 Mineable Ore Reserves

Phase

Ore to Plant≥0.036% Mo

Ore to LGO

Stockpile Total Ore Total Waste (t)Total

Material (t)

Waste

to OreRatiotonnes % Mo tonnes % Mo tonnes % Mo NPAG PAG Total

1 27,162428 0.115 5,133,464 0.030 32,295,892 0.102 6,850,853 9,117,650 15,968,502 48,264,394 0.492 43,129,297 0.097 4,175,026 0.030 47,304,323 0.091 9,649,457 2,916,853 12,566,311 59,870,634 027

3 87,376, 367 0. 086 10, 186, 960 0. 032 97,563, 327 0.080 51, 308, 620 41, 594, 646 92, 903, 265 190,466, 592 0. 95

4 32,595, 861 0. 086 5, 504, 550 0. 031 38,100, 411 0.078 14, 777, 727 28,715, 546 40, 493, 273 78, 593, 684 1. 06

Total 190,263,953 0.093 25,000,000 0.031 215,263,953 0.085 82,586,656 79,344,695 161,931,351 377,195,304 0.75





M I N E R A L R E S E R V E S

The total ore reserves from the four designed phases constitute the reserves for the

ultimate pit (Table 19.8). The ultimate pit contains 215.3 Mt of ore at 0.085% Mo, an

overall waste-to-ore ratio of about 0.75/1.0, and a total of 377.2 Mt of material.

U L T I M A T E P I T D I M E N S I O N S

The ultimate pit (Figure 19.7) has the following dimensions:

N-S direction 1,095.0 m

E-W direction 1,218.0 m

final ramp exit 570 m

pit bottom elevation 230 m

maximum high wall 780 m

maximum depth 550 m.





Figure 19.7 Final Pit





19.1.3 M I N E P R O D U C T I O N SC H E D U L E

G E N E R A L SC H E D U L I N G C R I T E R I A

The mine will deliver ore to the mill at the rate of 40,000 t/d or 14.6 Mt/a. With themill scheduled to achieve full production in Year 1, an adequate quantity of wastemust be stripped during pre-production periods to ensure that adequate ore is

available for the mill.

With the selection of a conventional slurry tailing disposal to the north side of theopen pit, the material specification will require the construction of an earth/rockfill

embankment. The starter embankment is scheduled for completion duringconstruction and assumed to have a tailing storage capacity of two years. It wasassumed that about 59% of the waste rock from the pit over the first 8 years of

operation will be used in the TMF.

With the exception of Years 1 to 3, part of the waste rock will be delivered to theWaste Dump located west of the open pit. Commencing in Year 9, all waste rock will

be hauled to the Dump. It is estimated that 70% of the total waste dumped in thedump will consist of PAG rock.

TA I L I N G MA N A G E M E N T FA C I L I T Y

Waste rock requirements for the tailing embankment at the Kitsault TMF wereestimated by KP for the starter dam to provide storage for the first two years of

milling operations. This rock will be found from the borrow areas close to the TMF.The embankment crest will be progressively raised to final elevation for the first 8

years of operations, using waste rock from the pit.

O R E A N D W A S T E AN N U A L PR O D U C T I O N

An annual mine production schedule was developed to ensure a continuous40,000 t/d (14.6 Mt/a) ore delivery to the mill. Annual tonnage and head grades were

calculated from bench average grades within each mining phase after the inclusionof mining dilution. Pre-production activities will begin in Year -1 to strip the wasterock and to prepare the mine for full mill production in Year 1.

Over the life of the mine, a total of 190.3 Mt of ore at 0.093% Mo will be delivered to

the mill, and 25.0 Mt of low-grade ore at 0.030 % Mo will be delivered to the LGOstockpile, 96.1 Mt of waste rock (63.1 Mt NPAG and 33.0 Mt PAG) to the TMF, and

65.8 Mt of waste rock (19.4 Mt NPAG and 46.4 Mt PAG) to the West Dump, for atotal material mined of about 377.2 Mt. The overall waste-to-ore ratio is 0.75:1.0.

The 25.0 Mt of low-grade ore will be reclaimed from the stockpile and delivered to the

mill for processing. The summarized production schedule is shown in Table 19.9.





Table 19.9 Summarized Production Schedule

MiningPeriod

Mine

Production(t)

LGO

Stockpile(t) Waste (t)

TotalMined (t)

StripRatio

Mill

Production(t)

Grade

(%Mo)

-1 540,624 8,195,965 8,736,589 15.16 0.0671 14,600,000 4,185,404 13,413,847 32,199,251 0.71 14,600,000 0.109

2 14,600,000 2,580,297 15,343,253 32,523,550 0.89 14,600,000 0.100

3 14,600,000 1,406,022 16,089,912 32,095,934 1.01 14,600,000 0.103

4 14,600,000 1,235,102 15,550,092 31,385,194 0.98 14,600,000 0.092

5 14,600,000 1,390,588 14,687,240 30,677,829 0.92 14,600,000 0.096

6 14,600,000 1,364,793 14,302,570 30,267,362 0.90 14,600,000 0.086

7 14,600,000 2,629,023 13,261,450 30,490,473 0.77 14,600,000 0.094

8 14,600,000 1,068,413 13,676,918 29,345,331 0.87 14,600,000 0.092

9 14,600,000 1,757,636 9,735,190 26,092,826 0.60 14,600,000 0.085

10 14,600,000 2,174,165 7,805,783 24,579,948 0.47 14,600,000 0.083

11 14,600,000 1,205,514 7,628,557 23,434,071 0.48 14,600,000 0.08812 14,600,000 2,533,348 8,749,146 25,882,494 0.51 14,600,000 0.086

13 14,600,000 927,156 1,796,306 17,323,462 0.12 14,600,000 0.089

14 463,953 1,915 1,695,122 2,160,990 3.64 14,600,000 0.033

15 10,863,953 0.031

Total 190,263,953 25,000,000 161,931,151 377,195,304 0.75 215,263,953 0.085

PR E- P R O D U C T I O N ST R I P P I N G

Prior to and during Year -1, the haul roads from the in-pit junction to the tailing dam,

West Dump, LGO stockpile area and primary crusher will be constructed. Thestripping of the waste rock will be coordinated with the haul road construction so asto provide base rocks for the roads. The approximate primary haul road distancesfrom the in-pit junction are 4,400 m to the tailing embankment, 600 m to the Waste

Dump, 2,500 m to the LGO stockpile adjacent to the plant, and 2,000 m to the

primary crusher.

Pre-production stripping activities will focus along the north and northeast sections of

the pit in order to finalize all the ex-pit haul roads. The stripping will strip benches

from a maximum elevation of 780 m down to 630 m.

In Year -1, about 8.2 Mt of waste rock will be hauled to the Waste Dump, on the west

side of the pit. This waste rock will consist of 3.8 Mt NPAG rock and 4.4 Mt PAGrock, for a NPAG/PAG ratio of 0.87/1. About 0.7 Mt of low-grade ore will be sent tothe Patsy Creek LGO stockpile. The extent of pre-production stripping is illustrated

in Figure 19.8.

It should be noted that Figure 19.8 through to Figure 19.15 show only the mine,waste, and low-grade stockpile status plans.





F I V E- Y E A R M I N I N G PL A N S

A set of mining plans show the annual development of the open pit from Years 1 to 5

as illustrated in Figure 19.9 through to Figure 19.13.

During the inclusive period, the mine is scheduled to deliver 73.0 Mt of ore at 0.100%Mo to the mill, stockpile 10.8 Mt of low-grade ore at 0.031% Mo and strip 75.1 Mt of

waste rock for a waste-to-ore ratio of about 0.90/1.0. About 64.2 Mt of waste(41.4 Mt NPAG and 22.8 Mt PAG) will be delivered to the TMF and about 10.9 Mt

(1.6 Mt NPAG and 9.3 Mt PAG) to the West Dump. A total of 158.9 Mt of material

will be moved.

The ore will be mined down to 470 m bench and the ultimate wall will be reached at

the north, east, and south sections of the pit. The Waste Dump on the west side of

the pit will be filled up to 540 m and the Patsy Creek LGO stockpile up to 765 melevation.

YE A R 9 M I N I N G PLA N

The mining status as of Year 9 is illustrated in Figure 19.14 along with the integrateddevelopment of the TMF, Waste Dump, and LGO stockpiles.

During the inclusive period from Years 6 to 9, the mine is scheduled to deliver a totalof 58.4 Mt of ore at 0.089% Mo to the mill, stockpile 6.8 Mt of low-grade ore at0.032% Mo and strip 51.0 Mt of waste rock for a waste-to-ore ratio of about 0.78/1.0.

About 32.0 Mt of the waste (21.8 Mt NPAG and 10.2 Mt PAG) will be delivered to theTMF and 19.0 Mt (4.0 Mt NPAG and 15.0 Mt PAG) of waste rock to the Waste Dump.

A total of 116.2 Mt of material will be moved.

The ore will be mined down to the 420 m bench and the ultimate wall will be reachedat the west section of the pit. The Waste Dump will be pushed further to the west at

540 m and the LGO stockpile up to 750 m elevation towards the northeast.

YE A R 14 – F I N A L YE A R M I N I N G PLA N LGO S T O C K P I L E

The mining status as of Year 14 is illustrated in Figure 19.15 along with theintegrated development of the TMF, Waste Dump, and LGO stockpile.

From Years 10 to 14, the mine was scheduled to deliver a total of 58.9 Mt of ore at

0.087% Mo to the mill, stockpile 6.8 Mt of low-grade ore at 0.031% Mo and strip27.7 Mt of waste rock for a waste-to-ore ratio of about 0.42/1.0. All of the waste

(10.1 Mt NPAG and 17.6 Mt PAG) was delivered to the Waste Dump on the Westside of the pit.

The ore was mined down to 230 m bench. The Waste Dump was filled up to 605 m

and the LGO stockpile by the plant up to 795 m elevation.





Figure 19.8 Mining Status as of Year 0

N






N





19.1.4 E Q U I P M E N T AN A L Y S I S A N D SE L E C T I O N

SU M M A R Y O F PR E V I O U S WO R K

SRK’s 2008 PEA selected 142-t Cat 785D haul trucks, a 21-m3

hydraulic front shovel(Hitachi EX3600), and a 199 mm diesel rotary drill as primary equipment for themining operations. The PEA also provided two 17-m3 front end loaders (Cat 994F)

for loading flexibility in the pit and other auxiliary pit functions.

The PEA assumed that the mine is scheduled to operate 24 h/d, 350 d/a, with4 operation crews working 12-h shifts on a 7 day rotation. A total of 15 days was

assumed to be lost due to weather, shutdowns, and miscellaneous other delays.Mechanical availability was estimated to typically range from 85% to 90% for themajor equipment.

As the mill and tailing dam configurations have yet to be finalized, haul profiles were

based on a final in-pit haul of 4,000 m at a 10% ramp gradient, a pit exit to theprimary crusher haul of 1,000 m, and a pit exit to the tailing dam of 3,000 m with a

225 m rise in elevation.

SC O P E O F WO RK

The work involved the analysis and selection of primary loading and hauling

equipment over a range of equipment sizes considering anticipated multiple workingareas and long truck hauls to the tailing dam, waste dump, and the LGO stockpile.

The work consisted of estimating:

shovel-truck fleet productivities for different fleet sizes based on haul profilemeasurements and haul cycle calculations

fleet operating hours based on a preliminary production schedule from the

pit optimization

fleet capital and operating costs based on available cost data

present value cost in Canadian dollars for the shovel-truck fleet for each of the equipment sizes and selecting the fleet with lowest present value costs.

EQ U I P M E N T SE L E C T I O N C R I T E R I A

Production rate is important to determine the number of loading units required for continuous supply of ore to the mill. With larger shovels, fewer units will be needed

to meet overall production but have a higher probability of having insignificantproduction during mechanical breakdowns resulting in low truck utilization. On the

other hand, having smaller loading units will require more operating areas, larger

facilities, and more personnel to perform maintenance resulting in increasedoperating and capital expenses.





The equipment analysis and selection was guided by the following criteria:

Analysis and selection was based on the preliminary production schedule as

generated from the Whittle optimization using the new geology resourcemodel; and production rate of 40,000 t/d.

Trucks were matched with loaders relative size (shovel dumping height vs.hauler size) and number of bucket loads to fill the truck. Experience

indicates that between three to five shovel passes are the most efficient inloading a truck. However, productivity is also affected when the mining

operation is shovel or truck restricted as waiting delays could vary

drastically. The fleet was narrowed down to 15.0 m3 to 26.0 m3 shovel sizesand 136-t to 232-t haul trucks, as shown in Appendix G.

In general, electric cable shovels are more productive and have a lower

operating cost compared to an equivalent hydraulic shovel. When operatingon wider pushbacks and properly developed mine plans, an electric shovel

has a higher production rate, less maintenance time, and lower power costs.

However, the capital cost of electric shovels is higher.

In terms of unit cost, the best option for any mine is to purchase the largestelectric cable shovel and matching truck fleet it could afford. The larger the

units the lower the unit costs tend to be. The limit on the size of this fleet

would be defined by the mine life and capital costs since capital costs tendto increase rapidly as the unit size increases. With higher capital costs more

time will be required to attain the benefit of the lower operating costs. For the cable electric shovel, only the 22 m3 size of shovel was considered asthe next size would be too large for the anticipated pit production. For the

electric hydraulic shovels, 15 m3 to 26 m3 sizes were considered.

Equipment capital and operating costs were based on the 2008 MiningReference Guide.

BC Hydro’s power rates are lower compared to the cost of diesel fuel;

therefore, the use of diesel driven shovels was not considered in the

equipment analysis.

The front-end loader was not included in the analysis as it is a common

equipment item in any fleet as a back-up to the selected primary shovel.

LO A D I N G PA R A M E T E R S A N D T RUCK MA T C H CA L C U L A T I O N S

The mine is scheduled to operate 24 h/d, 365 d/a, and two 12 h shifts per day. Anaverage mechanical availability of 85% was assumed for the cable electric shoveland 82% for the electric hydraulic shovel.

Operating delays such as weather, meal breaks, shift changes, blasting, and

equipment standby are common to all fleets. After allowing for mechanicalavailability, utilization of available time was estimated at 70% for the cable shovel





and 69% for the hydraulic shovel. It was assumed that the truck availability to ashovel is only 80%, in order to obtain a conservative estimate of shovel productivity.

A total of 12 different fleets were evaluated and are as follows:

Fleet 1: 22 m3

electric cable shovel + 232-t haul truck

Fleet 2: 22 m3 electric cable shovel + 177-t haul truck

Fleet 3: 22 m3 electric cable shovel + 136-t haul truck

Fleet 4: 26 m3 electric hydraulic shovel + 232-t haul truck








Fleet 12: 15 m3 electric hydraulic shovel + 136-t haul truck.

PR O D U C T I O N C R I T E R I A A P P L I C A B L E T O EA C H FLEET

Prel iminary Whi t t le Product ion Schedule

The preliminary production schedule produced from Whittle was used to referencethe mined-out materials from the open pit and the corresponding delivery of thosematerials to various destinations. Delivery from the ore to the mill is scheduled for 14 years, for a total ore production of 170.8 Mt. About 6.8 Mt of low-grade ore will

stockpiled, reclaimed, and delivered to the mill. A total of about 202.4 Mt of waste

was stripped with 65% of the material assumed to be hauled to the tailing dam and35% to the Patsy Dump. The maximum total amount of moved material is scheduled

to be approximately 31 Mt/a.

Fleet Avai labi l i ty , Ut i l izat ion, and Shovel Product ion

Electric cable shovels are assigned a mechanical availability that is 2% higher thanelectric hydraulic shovels. The shovel and truck availability has been progressively

scaled down to reflect equipment ageing. Utilization of available hours is higher for

trucks compared to shovel to reflect more shovel delays waiting for trucks.





Cycle Time Calculat ion

A bench-by-bench schedule was developed for the ore, low-grade ore, and wastematerials for each phase, using the Whittle scheduler. The bench centroid for eachof the materials in each phase was estimated. Haul profiles and distances to

respective destinations were then manually measured using MineSight® Acadfunctions. Existing site roads to each material destination were used whereapplicable for the haul profiles in conjunction with the in-pit haul roads using the PEA

open-pit design (which was the only design available).

The information was then inputted into Caterpillar’s Fleet Production and Cost Analysis (FPC) Version 3.05A to calculate the truck cycle time to each of the material

destinations. The shovel loading time was derived from the loading parameters for each of the shovel sizes. Cycle times were calculated for each of the shovel and

truck sizes.

The total average cycle time from the FPC program was manually inputted into the

spreadsheet for each year and for each of the material destinations. The total truckoperating hours calculated for each of the material destinations were accumulated to

estimate the truck fleet productivity and subsequently, number of trucks required for each period. The net operating hours for the shovels were calculated based onestimating the probability of a number of operational shovels in each period.

Equipment Cost Est imat ions

The 2008 Mining Reference Guide was used as the basis for estimating the capital

and operating costs for each of the shovel and truck fleets with the exception of the

cost of fuel, truck tires, and electrical power. These costs were obtained from the

parameters used in the pit optimization. The total capital and operating costs wereaccumulated for each year. The present value of the total annual costs was thencalculated for each of the fleet as the basis for the selection of shovel and trucksizes.

EQ U I P M E N T SE L E C T I O N R E S U L T S

Table 19.9 shows the ranking results for the 12 equipment fleet based on present

value costs. Fleet 2 (two 22.0 m3 BE 295HR electric cable shovels and 177 t trucks),

as the fleet with the lowest present value costs of $189.5 M, compared to Fleet 8(three 15.0 m3 PC 3000 electric hydraulic shovels and 232 t trucks), as the fleet with

the highest present value costs of $355.9 M. Fleet 8 was selected as the primary pitequipment and as the basis in deriving the size of the auxiliary equipment. Although

the PC4000 electric hydraulic shovel is 3% higher in present value cost, it matcheswell with the selected bench height and provides more operational flexibility.





Table 19.10 Ranking of Equipment Fleet

RankingFleetNo. Description

NPV (8%)Cdn$ M

1 2 two BE 295HR electric cable shovels + 177-t trucks 189.51

2 8 two PC4000 electric hydraulic shovels + 177-t trucks 195.933 5 two RH200 electric hydraulic shovels + 177-t trucks 217.40

4 11 four PC3000 electric hydraulic shovels+ 177-t trucks 218.03

5 9 two PC4000 electric hydraulic shovels + 136-t trucks 223.83

6 3 two BE295 electric cable shovels+ 136-t trucks 225.07

7 4 three RH200 electric hydraulic shovels+ 232-t trucks 246.60

8 6 three RH200 electric hydraulic shovels+ 136-t trucks 248.70

9 12 four PC3000 electric hydraulic shovels+ 136-t trucks 258.00

10 1 two BE 295HR electric cable shovels + 232-t trucks 326.77

11 7 Two PC4000 electric hydraulic shovels+232-t trucks 336.75

12 10 three PC3000 electric hydraulic shovels+ 232-t trucks 355.85

19.1.5 M I N I N G E Q U I P M E N T F L E E T P R O D U C T I V I T I E S

G E N E R A L C OMMENTS

Large-scale mining equipment was selected to match the mine production schedule,which is based on 365 d/a. Crews will work in two 12-h shifts, 4 days on and 4 daysoff. Equipment selection, sizing, and fleet requirements were based on expected

operating conditions, long haulage profiles, production cycle times, mechanical

availability, and overall utilization. To determine the number of units for each

equipment type (drills, shovels, haulers, etc.), annual operating hours werecalculated and compared to the available annual equipment hours.

Mobile mine support equipment, such as front-end loaders, track and rubber-tireddozers, graders, water, lube, and fuel trucks were matched with the major mining

units. Auxiliary equipments were assigned to haul road maintenance, snow removal,and mechanical and electrical servicing of the mining fleet.

Equipment additions and replacements were determined for each major and auxiliary

support mining unit.

M I N E EQ U I P M E N T O P E R A T I N G SC H E D U L E

The equipment calendar operating schedule is shown in Table 19.11.





Table 19.11 Mine Equipment Operating Schedule

Loading Parameters UnitsOperating

Time

Calendar Days d/a 365

Scheduled Outages (incl. Holidays) d/a 0Work Days d/a 365

Shifts per Day shifts/d 2

Hours per Day h/d 12

Total Hours h/a 8,760

19.1.6 D R I L L I N G

B L A S T H O L E DR I L L – N E T PR O D U C T I V E OP E R A T I N G T I ME

The initial drill requirements will consist of two blasthole drills capable of drilling

269 mm diameter blastholes. An 8.0 m x 8.0 m pattern has been selected in drilling

the ore and waste rock.

The mechanical availability of the blasthole drill has been applied on a sliding scale

based on the ageing of the unit. The overall mechanical availability averages to

about 82.8%. The use of available hours was assumed to be 75.0% on a constantannual basis. The estimated effective utilization of the drills over the life of the mine

is about 60%.

B L A S T H O L E DR I L L PR O D U C T I V I T Y

Drill productivities were based on a bit penetration rate of 0.35 m/min in both ore andwaste rock. Total drill time per hole including penetration and move time is shown inTable 19.12.

A diesel powered hydraulic percussion track drill will be used for secondary blastingof oversize material and for bench pioneering.





Table 19.12 Blasthole Drill Productivity

Item Description UnitsOre and

Waste Rock

Hole Diameter mm 269

Bench Height m 10.0

Subgrade m 1.6

Hole Length m 11.6

Penetration Rate m/min 0.35

- Penetration Time per Hole min 33.1

- Move/Align Time min 2.0

- Hole Collaring min 1.0

- Grade Control Sampling min 1.5

Total Time Per Hole min 37.6

Holes Per Hour holes 1.59

Average Drilling Rate m/h 18.5

Spacing & Burden m 8.0

Material Weight t/m3 2.60

Rock Mass per Hole t 1664

Availability % 82

Use of Availability % 75

Effective Utilisation % 61.5

Maximum Operating Hours h 5,387

Hourly Maximum Drill Productivity t/h 2,652

Yearly Maximum Drill Productivity t/a 14,289,000

19.1.7 BL A S T I N G

G E N E R A L B L A S T I N G CO N D I T I O N S F O R PR O D U C T I O N H O L E S

Overall explosive consumption has been based on 30% wet holes using 70% ANFOand 30% Emulsion. Drillhole liners will be used in wet holes where practical.

An explosive supplier will erect on-site a bulk explosives plant and equipment, bulkproduct storage facility, and explosives magazines. The supplier will be contractedto supply, deliver, and load explosives into the blastholes. The supplier will also be

contracted to provide the blasting crew and will charge monthly fees for the plant

supervisors. The drill and blast foreman will oversee the contractor’s blasting crewwho will prime, stem, and tie-in blastholes. The contractor will also dewater wet

blastholes.

Table 19.13 shows the blasting parameters.





Table 19.13 Blasting Parameters for 269 mm Production Blastholes

Blasting Parameters UnitOre and

Waste Rock

Burden & Spacing m 8.0

Hole Diameter mm 269.0Bench Height m 10.0

Subgrade m 1.6

Rock Mass per Hole t 1,664.0

Explosives Bulk Density t/m3

0.58

% of Hole Charged % 60.0

Column Charge per Hole kg 231.0

19.1.8 L O A D I N G

G E N E R A L LO A D I N G CO N D I T I O N S

The loading fleet will consist of two 18 m 3 electric hydraulic shovels and one 18 m3

front end loader. The hydraulic shovels are matched with 177 t trucks to load in five

passes in handling ore and waste rock materials. Each of the shovels is assigned adigging cycle of 35 seconds.

The front end loader will load residual materials from the electric shovels and to

perform various functions in the pit and ore stockpile areas. The loader is matchedwith 177-t trucks to load in 5 passes in handling ore and waste rock materials. It has

a digging cycle of 42 seconds.

SH O V E L /L O A D E R LO A D I N G PR O D U C T I V I T Y

The estimated average loading productivity for the shovels and front end loader

loading the 177-t haulers is shown in Table 19.14.

Table 19.14 Average Productivity of Shovels in Ore and Waste Rock

Loading Parameters UnitsElectric

Hydraulic Shovel*Front

End Loader*

Total Hours h/a 8,760 8,760

Mechanical Availability % 75 75

Available Hours h/a 6,570 6,570

Standby/Idle % 3 3

Gross Operating Hours h 6,373 6,373

Total Operating Delays h/shift 2.45 2.45

Total Operating Delays h/a 1,787 1,787

table continues…





Loading Parameters Units

ElectricHydraulic Shovel*

FrontEnd Loader*

Net Operating Hours h/a 4,586 4,586

Use of Availability % 72 70

Effective Utilization % 52 52Bucket Capacity (Heaped) lcm 18.0 18.0

Dry Material Weight dmt/lcm 2.65 2.65

Swell Factor % 30 30

Dry Material Weight dmt/lcm 2.04 2.04

Moisture Content % 3.0 3.0

Wet Material Weight wmt/lcm 2.10 2.10

Bucket Fill Factor % 95 90

Effective Bucket Capacity lcm 17.1 16.2

Tonnes/Pass wmt 35.9 34.0

Truck Capacity wmt 177 177

Average Passes/Truck passes 4.9 5.2Truck Load for Productivity wmt 177.0 177.0

Truck Spot Time seconds 30 30

First Bucket Cycle Time seconds 35 42

Subsequent Bucket Cycle Time seconds 35 42

Load Time/Truck min 3.37 4.14

Maximum Productivity trucks/h 17.8 14.5

wmt/h 3,148 2,566

Truck Availability to Shovel % 80 80

Production Hours h/a 3,669 3,669

Annual Production wmt/a 11,551,720 9,413,536

Productivity wmt/noh 2,519 2,052Truck Specifications

Gross Vehicle Weight kg 317,316 317,316

Empty Vehicle Weight kg 135,955 135,955

Truck Rated Payload t 177.0 177.0

Truck Body Capacity lcm 105 105

* based on loading 177-t trucks.

Note: Equipment Specification Basis:Electric Hydraulic Shovel – PC4000

Front End Loader – Cat 994FHaul Truck – Cat 789C Mechanical Truck

19.1.9 T R U C K H AU LA GE

G E N E R A L H A U L I N G CO N D I T I O N S

The 177-t hauler was selected to match the 18 m3 hydraulic shovel, and 18 m3 front

end loader in determining the number of trucks required for each operating year.





The mine plan is based on 177-t haul trucks delivering 40,000 t/d ore to the primarycrusher. Low-grade ore will be delivered to the LGO stockpiles while the waste rock

will be hauled to the TMF and Waste Dump on the west side of the pit.

As shown in Table 19.9, none of the waste rock will be hauled to the TMF to build the

initial tailing embankment during the pre-production period as the rock will beborrowed from near sources and provide tailing storage for the first two years of milloperations. A total of 96.1 Mt of waste rock from the pit will be dumped in the TMF.

A total of about 190.1 Mt of ore will be hauled directly to the crusher. The 25 Mt of

ore will be hauled to the LGO stockpiles and subsequently re-claimed upon depletionof the ore from the pit and hauled to the primary crusher. 65.8 Mt of waste rock will

be hauled to the Waste Dump on the West side of the pit.

The approximate locations of the different facilities are shown in Table 19.15.

Table 19.15 Facility Locations

Easting(m)

Northing(m)

Top Elevation(m)

Distance to

Pit Junction(Elev. 570 m)

Primary Crusher 473,550 6,142,816 722 2,100

LGO Stockpile 473,900 6,144,000 760 2,500

TMF 471,800 6,144,000 360 4,400

Waste Dump 472,500 6,142,300 540 800

The haul distance from Pit Junction will vary with the bench centroid. In Phase 4, the

length of haul from the pit junction down to bench 240 m is estimated to be 3,700 m.In Phase 3, the length of haul down to bench 330 m is estimated to be 3,040 m. All

measured distances are contained in Appendix H.

H A U L TR U C K PR O D U C T I V I T Y

From computer haul simulations, total cycle times and annual hauler hours were

calculated for delivering annual tonnage requirements to the designated dumpinglocations. The following sections outline information that was used in the simulationanalysis.

Product ion Schedule by Phases and by Bench

A detailed production schedule was developed showing the mill ore, low-grade ore,

waste rock mined, and material by destination. The detailed schedule was used toestimate the location of the centroid of each type of material mined in each period,phase, and bench. The centroid was estimated to be the mid-point of total tonnages

mined between the top and bottom benches for each phase and each period when





the material is mined out. From this centroid, the ore and waste materials are hauledto the respective destinations (Appendix H).

Truck Haulage Prof i les

Haul profiles are laid out from the designed haul ramps for Phases 1, 2, 3, and 4

using MineSight®. From the pit junction, haul profiles were laid for mill ore destinedto the primary crusher, low-grade ore to the LGO stockpile, waste rock to the TMF

and lastly, waste rock to West Dump. It was assumed that existing roads to the

proposed tailing dam, LGO stockpile, and West Dump will be upgraded to 30 m widthto accommodate the 177 t haul truck.

The X, Y, and Z location for each material centroid is then tied to the designed phase

ramps to complete haul road link to each of the material destinations. Haul profileswere then developed to define the haul distances and gradients from period andphase centroid to each destination (Appendix H).

Truck Simulat ion wi th FPC

Caterpillar’s Truck Simulation Program FPC was used to calculate truck cycle times

for each of the material destination based on the Cat 789C truck rimpull performance

curves corresponding to specific haulage profiles. Each of the measured haulprofiles and estimated rolling resistances were manually inputted into the program.

Higher rolling resistances and lower speeds were estimated in areas such as thewaste dump, tailing embankment, and low-grade stockpiles.

The program calculated the haul and return cycle times. The dump and manoeuvre

delay (estimated at 30 seconds) and the shovel load time were added to the traveltimes to complete the total haul cycle (Appendix H).

Total Cyc le Time by Per iod by Phase

Using the estimated cycle time by phase by centroid for each material, the weighted

average cycle times by phase and by destination was calculated for the respective

period (Appendix I).

Typical haulage simulation parameters for the Cat 789C 177-t haulers are shown in

Table 19.16.





Table 19.16 Cat 789C Haul Truck Productivity

Haulage Simulation Parameters – Cat 789C Hauler

Truck Size – Nominal 177 t

Engine Gross Power Rating 1,320 kW

Empty Vehicle Weight 135,955 kg

Tire Size 37.00-R57 (E4)

Rolling Resistance

General Applications 3%

At Dam, Dumps 4% - 5%

Rock

Moisture 3.0%

Fill factor 95%

Payload (dmt) 177.0

Net Operating Time per Shift

Spot and Load times (min) Shovel Loader

Mill Ore to Primary Crusher 3.37 4.14

Low-grade Ore to LGO Stockpile 3.37 4.14

W aste Rock to Tailing Embankm ent 3.37 4.14

Waste Rock to West Dump 3.37 4.14

Turn and Dump Times (min) Shovel Loader

Primary Crusher 0.5 0.5

LGO Stockpile 0.5 0.5

Patsy Dump 0.5 0.5

Tailing Embankment 0.5 0.5

Speed Limit (km/h) Grade (%) Loaded Empty

+3 to 10 30 30

-3 to 3 40 40

-10 to -3 30 30

1 9. 1. 10 P I T D R A I N A G E A N D D E W A T E R I N G

The open pit drainage and dewatering system will consist of diversion ditches and in-pit collection sumps for controlled removal of both precipitation and groundwater

inflows. A series of pumps will be installed to transfer water from active mining

bench sumps to a piping link with the surface diversion ditches.

SU R F A C E WA T E R MA N A G E M E N T

Surface water runoff will be captured and collected in peripheral diversion ditches

outside of the ultimate pit limits. The diversion ditches will be constructed during thepre-production periods to prevent surface water from precipitation from infiltrating intothe open pit operations.





IN- P I T D E W A T E R I N G A N D P I P E L I N E S

The in-pit pumping requirements will vary annually throughout each season, and will

increase as the catchment area increases with successive pushbacks to ultimatewalls. Water reporting to in-pit sumps will be discharged through pipelines to the

surface diversion ditches.

A HDPE pipeline will be laid out up the pit walls and be assembled with steelstrapping, concrete/timber sleepers, and rock anchors.

1 9 . 2 G E O T E C H N I C A L E N G I N E E R I N G – P I T S L O P E S

SRK Consulting (US) Inc. conducted a pre-feasibility geotechnical pit slopeevaluation (SRK, 2009) for the Kitsault project open pit. The primary objectives of

the evaluation were to evaluate rock mass conditions in the area of the anticipated

open pit and to consequentially recommend pit slope design criteria, includinginterramp slope and bench configuration to be used for further mine planning and pitdesign.

19.2.1 G E O T E C H N I C A L D A TA C O L L E C T I O N

The geotechnical data collection program was developed with the primary objective

of rock mass characterization to support development of a geotechnical model

suitable for pit slope stability evaluation. Information such as geologic contacts,profiles of rock strength, and characterization of discontinuities were recorded duringthe program.

G E O T E C H N I C A L C OR E LOGGING

Geotechnical logging and discontinuity orientation of the core recovered from sixdrillholes were conducted in the fall of 2008 for the pre-feasibility investigation. Five

of the holes (i.e. K-08-04, K-08-09, K-08-12, K-08-14, and K-08-16) were drilled to

coincide with holes planned for the Avanti resource drilling program. Based on the

current understanding of the deposit, those particular five holes were selected toprovide the best coverage possible of rock likely to form the Kitsault pit walls. Sinceno further resource drilling was planned in the area of the anticipated western pit, an

additional hole (K-08-06) was drilled specifically to examine rock expected to

comprise that wall segment.

Based on results of the pre-feasibility pit slope evaluation (SRK, 2009), two additional

geotechnical drillholes (K-09-07 and K-09-12) were drilled in the summer of 2009 to

obtain geotechnical data sufficient for a feasibility-level pit slope evaluation. Collar coordinates as well as drillhole azimuth and inclination (below horizontal) of the eight

geotechnical drillholes are summarized in Table 19.17.





Table 19.17 Summary of Drillholes Orientated & Logged for Geotechnical Data

DDH ID

Collar LocationAzimuth

(°)Inclination

(°)Length

(m)Northing Easting Elevation

K-08-04 6141730.0 473100.3 560.2 185 -58 300.5

K-08-14 6141850.0 473570.0 594.2 089 -43 349.6

K-08-09 6141934.7 473743.9 672.5 277 -53 433.4

K-08-12 6141980.0 473300.0 594.2 002 -43 315.8

K-08-16 6141600.0 473580.0 593.6 086 -46 289.9

K-08-06 6141851.0 472989.0 583.0 275 -60 401.4

K-09-07 6141945.8 473534.5 595.7 043 -57 400.2

K-09-12 6141612.4 473249.3 548.5 180 -57 459.6

The geotechnical core logging program was targeted at definition of the MRMR

(Laubscher, 1990) and Geologic Strength Index (GSI) classification systems.

Dependant on block size and discontinuity condition, the GSI is a direct input to theHoek-Brown (Hoek et al., 2002) rock mass failure criterion. Specific geotechnical

parameters that were logged included general lithology, total core recovery, RQD,rock weathering and relative intact strength indices, discontinuity type, frequency,

and characteristics (roughness, shape, and condition) and discontinuity orientation.

As is common with most resource core drilling programs, a 10 ft-long standard corebarrel was used for sample recovery for the six 2008 geotechnical holes; 5 and

10 ft-long triple tube systems were used for the two 2009 geotechnical coreholes.

C O R E OR I E N T A T I O N

Orientation of discontinuities in each run was attempted by use of an ACT core

orientation system. The ACT system is an electronic, accelerometer-based devicemanufactured by Reflex Instruments. The depth of intercept and the angles of thediscontinuities relative to the core axis and perpendicular to the core axis, (alpha and

beta angles, respectively) were measured during logging to enable the calculation of

the true dip direction and dip.

In the 8 coreholes logged, a combined total of 5,054 discontinuities (61%) were able

to be oriented out of the total 8,285 logged. In addition to the oriented core

measurements, structural features or set members observed during the September 2008 site visit were also included in the analyses.

G E O T E C H N I C A L OB S E R V A T I O N S O F EX I S T I N G P I T

During a site visit by SRK between September 8 and September 11, 2008,geotechnical observations of the existing pit wall conditions and performance were

made and noted. The outer walls of the currently exposed pit consist primarily of ahornfels unit cut by relatively small intrusive bodies and lamprophyre dikes. The





existing outer pit walls are comprised of up to approximately six 10 m-high benchesseparated by catch benches, resulting in interramp slope angles of approximately 43°

to 45° over a total vertical height of 60 m. A relatively low slope comprised of one totwo benches is exposed in the interior, intrusive portion of the pit.

Based on the field observations, both the outer, hornfels slopes and the inner,intrusive slopes are in good condition, showing only minor raveling and very fewobservable rock displacements. The displacements observed included relatively

limited plane shear and bench scale wedge failures which were noted particularly in

the outer, north to northeast pit walls, and which most likely occurred duringexcavation when the pit was last active 26 years ago. No major fault structures wereobserved in the pit walls during the SRK site visit; however, some small scale, high

angle faulting, was evident in the north pit wall.

Discontinuity orientation data from the recent core drilling program was compared tostructural observations of existing pit walls. Observations of the existing east and

northeast pit walls correspond well with the southwest dipping set seen in theoriented core data. However, members of the northwest dipping set, believed to be

the most likely to adversely impact bench stability, were not observed in existing pit

walls due to the lack of significant rock exposures on the southeast side of the pit.

LA B O R A T O R Y TE S T I N G

Geomechanical testing was conducted at The University of Arizona Rock Mechanics

Laboratory in Tucson, Arizona. The overall laboratory program included direct shear,unconfined compressive strength (UCS) and triaxial compressive strength (TCS)

testing as well as measurements of unit weight and elastic properties. A total of 35

laboratory tests were conducted on samples of core from the six 2008 pre-feasibility

geotechnical borings. Additional laboratory testing will be conducted on samplesfrom the two 2009 geotechnical drillholes during the feasibility-level pit slopeevaluation.

UCS testing was conducted on 20 samples selected to represent the range of the

rock conditions observed in the coreholes. Elastic properties (Young’s Modulus andPoisson’s Ratio) were measured for 6 of the 20 UCS samples. Valid tests produced:

UCS values ranging from 41.9 to 238.4 MPa, with a mean of 104.7 MPa

Young’s Moduli ranging from 13.7 to 69.4 GPa, with a mean value of

43.0 GPa

Poisson’s Ratios ranging from 0.179 to 0.283, with a mean value of 0.228.

TCS test results were used for direct determination of the Hoek-Brown (Hoek et al.,2002) material coefficient (mi). TCS tests were conducted on eight samples with

confining pressures selected up to approximately one-half of the UCS values as

suggested by Hoek (1997). The eight samples were tested with confining pressures





(1) ranging between 3.4 and 20.7 MPa resulting in compressive strengths (3)

between 63.9 to 321.9 MPa.

Discontinuity shear strengths were measured by direct shear testing seven samples

of core that contained natural fractures. Results ranged between 26° and 44° with a

mean value of 35°.

19.2.2 G E O T E C H N I C A L M O D E L

A geotechnical model was developed to provide a framework for slope stabilitymodeling by mathematically simulating site geotechnical conditions and the responseto stress changes resulting from the proposed open pit excavation. Rock mass

characterization, along with the available information on site geology, formed thebasis for defining the individual Kitsault domains (i.e. the Hornfels [country rock]Domain and the Intrusives Domain). The demarcation between the domains was

assumed to be as defined in dxf wireframes provided by Avanti. SRK understands

that the geologic wireframes provided were obtained from resource modelling andthat they reflect the most current information on rock type boundary definition.

The two geotechnical domains are generally comprised of relatively competent rockwith intermittent zones of weaker material. The weaker zones typically correspond to

intervals of increased fracturing, weathering, and/or alteration, including minor fault

zones and surface weathering. As such, intermittent weaker rock zones represent arelatively small portion and are not considered to be representative of the rock massconditions throughout the deposit.

SRK modeled the rock mass shear strength/normal stress relationships using theGeneralized Hoek-Brown criterion. The Generalized Hoek-Brown criterion defines

curvilinear shear strength envelopes that are considered effective representations of intact rock and heavily jointed rock mass behaviour. Primary input parameters for

the Generalized Hoek-Brown jointed rock mass criterion include the GeologicalStrength Index (GSI), an intact material constant (m i) and a disturbance factor (D), as

defined by Hoek et al. (2002).

19.2.3 D A T A AN A L Y S I S

Evaluation of the field and laboratory data collection programs indicated a highdegree of variation in rock strength and geologic structure at Kitsault. This naturalvariability in rock strength and structure suggests that a probability-based method of

analysis would be most appropriate and would yield less conservative slope anglesthan would the selection of a unique, potentially over-conservative value as is typicalto strictly deterministic analyses.

Model parameters were characterized by statistical distributions of values having acentral tendency and some variation around that central tendency, rather than by asingle, unique value. The variations of geomechanical properties were represented





by distributions that are randomly sampled during each of many calculations used inthe analyses.

19.2.4 I N T E R R A M P /O V E R A L L ST A B I L I T Y AN A L Y S E S

The mathematical geotechnical model was input into the commercially availablegeotechnical modeling software package Slide v.5.039, developed by Rocscience

Inc. (2003). Slide is a two-dimensional, limit equilibrium slope stability analysisprogram that analyzes slope stability by various methods of slices. Spencer’s

Method was selected for the limit equilibrium analyses of this evaluation due to its

consideration of both force and moment equilibrium.

The two vertical profiles considered most critical and representative of conditions

were selected for analysis based on the ultimate pit configuration and geologic solids

provided by Avanti. One vertical profile was through the northeast pit wall and theother, through the south pit wall. SRK analyzed the sections both with and without

pseudostatic loading to examine not only static slope conditions but also sloperesponse to earthquake-induced seismic loading. Dynamic loading from earthquake

ground accelerations was simulated using the pseudostatic analysis previouslymentioned, and a horizontal Peak Ground Acceleration (PGA) expressed as a

coefficient (or percent) of gravity. According to the 2005 National Building Code of Canada, the coefficient of peak horizontal acceleration that corresponds to a 10%

probability of exceedance in 50 years is 0.070 gravity (g) for the Kitsault site (Institute

for Research in Construction, National Research Council of Canada, 2005).

When incorporating a PGA value as input into a slope stability model, it is common

practice to reduce the PGA by a factor of 0.5 according to research conducted by the

US Army Corp of Engineers (Hynes-Griffen and Franklin, 1984). In summary, this

reduction in horizontal acceleration is justified for earth and rock structures for thefollowing reasons:

realization that damage to earth structures results from sustained ground

accelerations, typically less than half of the PGA, which is an instantaneousacceleration

consideration that earth and rock structures effectively attenuateearthquake-induced accelerations.

ST A T I C A N A L Y S I S

Results of the northeast section analysis indicate a probability of failure (POF) of about 3% and the south section yielded a 0% POF for static loading conditions. It

should be noted that while a 0% POF does demonstrate a very low likelihood of

slope instability, it does not imply that slope instability is impossible. The 0% simplymeans that for the samples drawn from the strength distributions defined, no validslip surface had a safety factor less than 1.0. Results such as these simply indicate

that slope stability will be governed primarily by bench configuration and not by





interramp/overall slope angles. Furthermore, mean static safety factors of 1.5 and1.6 resulted from the analyses of the northeast and south sections, respectively.

These results are considered well within the range of acceptable risk of instabilitywhen compared to the industry standard static safety factor of 1.3. Results such as

these simply indicate that slope stability will be governed primarily by bench

configuration and not by interramp/overall slope angles.

PS E U D O S T A T I C AN A L Y S I S

Pseudostatic stability modelling of the northeast and south sections yielded POFs of

7% and less than 1%, respectively. Mean safety factors of 1.3 and 1.5 werecalculated for the Hornfels and Intrusives Domains, respectively, which areacceptable when compared to the industry standard pseudostatic safety factor of 1.1.

19.2.5 BE N C H ST A B I L I T Y C O N S I D E R A T I O N S

For pre-feasibility-level pit slope evaluations, interramp/overall slope angles are of primary interest and full bench design is not typically undertaken until higher level

investigations are undertaken. A full bench design or “backbreak” analysis is nottypically undertaken at the pre-feasibility-level and is therefore not included in this

scope of work.

Rock structural conditions were evaluated relative to current ultimate pit wallorientations to identify structural features that could potentially control achievable

bench face angles. In general, two joint sets were identified that have potential tocreate bench scale instability – one with a moderate northwest dip and the other dipping moderately steep to the southwest. The set orientations are consistent in

both the Intrusive and Hornfels domains.

The northwest dipping set is the most prominent, with a mean dip angle of about 56°and a mean dip direction of about 305° (azimuth). The southwest dipping set has a

mean dip of about 59° and a mean dip direction of 230°. Two high anglediscontinuity sets striking north-northwest and northeast were also identified but arenot anticipated have significant impacts on achievable bench configurations.

Based on the current data, the highest risk pit wall orientations for bench instabilitywill be those with a dip direction of about 290° to 320° (striking 200° to 230°).Consequentially, the most sensitive area of the pit to bench scale instability resulting

from members of this set will be the southeast corner, labelled as SE Sector on

Figure 19.16. Given the relatively narrow portion of the current ultimate pit plan withwalls oriented in this direction (dip direction between 290° and 320° azimuth), the

potential for bench scale instability within this sector is not anticipated to have major impacts on pit slope design.

The southwest dipping set, as previously described, may adversely impact the

stability of benches with a dip direction of about 215° to 245° (strike of 125° to 155°),labelled as NE Sector on Figure 19.16. This set does have minor potential for plane





shear type instability along the northeast pit wall; however, existing data indicate thata majority of structures in this set will dip sufficiently steep such that their impact on

achievable bench face angles will be relatively minor.

Figure 19.16 Sectors of Potential Bench Scale Instability

The two additional, 2009 geotechnical drillholes (K-09-07 and K-09-12) were

designed specifically to target the northwest and southwest dipping discontinuity

sets. The data acquired from these two holes will be used to evaluate potentialstructural impacts on bench design by each of these sets during the feasibility-levelpit slope evaluation, currently in progress.

19.2.6 SL O P E D E S I G N R E C O M M E N D A T I O N S

It is recommended that further mine planning for the Kitsault pit utilize an interrampslope angle of 51° with double 10 m (20 m high) benches, average sustainable

bench widths of 9 m and 70° mean bench face angles. While an interramp angle of

greater than 51° would be analytically indicated based solely on the interramp

analyses, such a configuration would likely produce average bench widthsinsufficient to catch rockfall from higher slopes and to permit some maintenance, if necessary.

Limited information on groundwater levels and flow rates were available in the area

of the proposed pit at the time of the pre-feasibility pit slope design (SRK, 2009).Conservative estimates were made for slope stability modelling; however, it is





important to note that these are only estimates which should be confirmed for further analyses. Potential effects of pore water pressures on global pit slope stability as

well as pit dewatering requirements should also be further evaluated.

SRK has recently completed the field portion of a hydrogeological characterization

program, as described in Section 19.3, using the two recent geotechnical drillholesK-09-07 and K-09-12. Additional time will be required to complete the data analysisportion of the program. Results of the hydrogeological evaluation along with

additional geotechnical data from K-09-07 and K-09-12 will be incorporated into the

feasibility-level pit slope analyses.

1 9 . 3 P I T H Y D R O G E O L O G Y

SRK Consulting (Canada) Inc. was approached by Avanti in July 2008 to implement

a geotechnical and hydrogeological program for the proposed expansion of the

Kitsault pit.

To further the hydrogeological understanding of the open pit in support of Avanti’s

PFS, SRK’s scope was to carry out field investigations designed to characterizehydrogeological conditions, and to provide appropriate analysis of such data. Thiswork has been integrated with ongoing feasibility-level geotechnical pit slope design,

to be reported separately. The geotechnical PFS assessment was presented in the

2008 SRK report Geotechnical Pit Slope Design, Kitsault Project .

This section presents SRK’s hydrogeological assessment of the Kitsault deposits. A

full account of the pit hydrogeological assessment (SRK 2009 report, Kitsault Pre

Feasibility, Pit Hydrogeology) can be found in Appendix J. The following objectives

have been satisfied:

provide hydrogeological input for use in the pit slope design for the proposed

pit and, if necessary, propose a suitable dewatering or depressurizationstrategy

estimate groundwater inflow to the pit

provide a prediction of inflow geochemistry.

To achieve the 2009 PFS objectives, SRK designed and managed an integratedhydrogeological/geotechnical drilling program of 860 m in two drillholes in the

northeastern and southern parts of the proposed pit. Groundwater conditions were

investigated to augment the hydrogeological conceptual model for the site and add tothe data collected during the 2008 investigation program. The updated

hydrogeological understanding was used to estimate pore pressures as related to pitslope stability design, groundwater inflow rates, and inflow water quality. The focusof this program was to characterise hydrogeological conditions surrounding and/or

influencing the open pit mining development.





The dominant lithologies (a quartz monzonite and diorite core, hosting molybdenummineralization, with a hornfels aureole extending outwards from the mineralized

rocks) resulting in what is expected to be a low primary porosity and limitedsecondary porosity (fracture flow) hydrogeological system.

To test this, packer tests were completed over fixed spacing and structure-specificintervals in the two drillholes to provide information on the spatial distribution of hydraulic conductivity.

Packer testing to date indicates that the rock mass across the site does have low

bulk low transmissivity, with no anomalously high conductivity zones encountered.Some fractures zones were observed, but associated clay alteration within these

fractures results in low transmissivity values similar to the general rock mass. The

data do not suggest the presence of significantly high transmissive structures thatcould cause high inflows, nor significantly lower permeability lithologies or structuralfeatures that would have the potential to cause compartmentalization of groundwater

during pit slope dewatering. However, due to the low number of drillholes tested,these cannot be ruled out, and should be a focus of further testing if additional work

identifies areas of hydrogeological concern.

Flowing artesian pressures (piezometric levels above local ground surface) havebeen observed at site in holes K08-23, K09-07-GT, and K09-12-GT. Holes

(K09-07-GT and K09-12-GT) were equipped with an array of downhole vibrating wire

transducers and dataloggers to record changes in pressure over time. Shut-in water pressures of up to 190 kPa (27 psi) recorded during testwork. Such pressures aretypical of an environment with steep topography and are not expected to pose a

significant risk to slope stability. However, this needs to be verified as part of the

ongoing feasibility-level slope design assessment. If the pressures are deemed to be

a risk, a program to delineate and verify ability to depressurize these parts of theslope will be implemented in the hydrogeology feasibility program. Depressurizationis expected to be controlled with a standard array of horizontal drainholes drilled into

the pit slopes.

Available hydrogeological data were used to develop simple analytical inflow modelsfor the Kitsault pit development. The model objective was to estimate the potential

range of inflow rates to the pit at various stages of excavation.

Modelling results suggest that flows are generally low, and are coincident with thelow permeability of the pit lithologies. A summary of the best judgement (BEJ) and

geomean groundwater inflow rates for various pit excavation stages are indicated in

Table 19.18.





Table 19.18 Summary of BEJ and Geomean Groundwater Inflow Rates

Pit

Excavation

Stage

Mean Groundwater Inflow (m

3 /d)

BEJ

Geomean

of Data

5 Year 70 75

10 Year 25 245

14 Year 25 655

Although site water management is beyond SRK’s scope, it is anticipated that excess

groundwater inflows to the pit will be collected in sumps and pumped via the site

water management system into the TMF.

Water quality of potential inflows to the pit was assessed during two sample rounds

on wells K08-23 and K08-24. The wells are screened over 100 m to 250 m and167 m to 400 m respectively. Water quality results indicated near neutral pH toslightly alkaline in both wells. In the October 2008 samples, elevated iron andmanganese concentrations suggested chemically-reducing conditions. In terms of

other metals, molybdenum concentrations were consistent with the presence of

molybdenum mineralization and are elevated with respect to typical groundwater concentrations in non-mineralized areas. Concentrations of other elements were

considered typical for such a deposit.

1 9 . 4 S I T E W A T E R M A N A G E M E N T

19.4.1 G E N E R A L

The Kitsault Project water management plan has been developed based on the

layout of the project facilities, process requirements, topography of the project area,and hydrometeorology. The objective of the water management plan is to keep

clean water clean and to direct all impacted water to the TMF. Water will be divertedaway from the mine site through the Lime Creek Diversion Tunnel, the Patsy CreekDiversion, and several smaller diversion ditches. Surplus water from the TMF will be

treated as required and released to the environment into Lime Creek. A generallayout of the mine site including catchment areas is shown in Figure 19.17.





Figure 19.17 General Layout of the Kitsault Mine Site





19.4.2 H Y D R O L O G Y A N D F LOOD F L O W S

FL O O D M ODELLING

Return period flood modelling for the Kitsault Mine was primarily conducted usinginformation provided in “Streamflow in the Skeena Region” (Obedkoff, 2001). Thisreport includes analysis of Lime Creek hydrology, including 10-year return period

peak instantaneous discharge estimates and scaling relations to calculate variousother return period discharges. Water Survey of Canada (WSC) streamflow datafrom Lime Creek, Patsy Creek, and several other regional stations were also used in

the analysis to develop an instantaneous to daily discharge scaling curve. Using thiscurve and a daily to 24 hour scaling ratio adopted from the Washington RainfallFrequency Atlas, peak 24 hour return period discharges and their associated

volumes were derived.

Further to the return period flood modelling, an estimate of the probable maximumflood (PMF) was derived for design of the TMF. The PMF was estimated using

HydroCAD and the input of the probable maximum precipitation (PMP). The PMPwas calculated using daily precipitation data collected by Environment Canada at

Nass Camp (1011973), which were converted to 24 hour data using the Washington

Rainfall Frequency Atlas scaling ration, and applying the PMP equations presentedin the Rainfall Frequency Atlas of Canada (RFAC). The PMP estimate was adjusted

for the difference in elevation between the regional station and Kitsault Mine, andwas assessed for validity using regional precipitation means and standard deviations

presented in the RFAC. The resultant 24 hour PMP for Kitsault Mine is 592 mm.

This value was then input to HydroCAD, along with numerous assumptions andcalculations, to derive a hydrograph for the Kitsault PMF. The PMF was assumed to

occur in early winter in the presence of a substantial snowpack, the melt of whichwas also modelled in HydroCAD. The PMF was calculated to have a maximuminstantaneous discharge of 1,257 m3/s and have a total volume of 27.0 M m3. This

flood was then routed through the TMF, and based on the available storage capacity

and an optimized outlet size. The maximum instantaneous outflow from the TMFassociated with the PMF is 199 m 3/s.

EX T R E M E PR E C I P I T A T I O N MO D E L L I N G

Extreme precipitation estimates were derived from the Nass Camp precipitationrecord. The mean and standard deviation of the Nass Camp maximum annual

precipitation record were computed and applied to a Gumbel distribution to modeldifferent return period precipitation events. The results were adjusted from daily to24 hour values, and then were scaled to account for the difference in elevationbetween the station location and the Kitsault plant site.





19.4.3 D I V E R S I O N ST R U C T U R E S A N D L I ME C R E E K H Y D R O E L E C T R I C P R O J E C T

Patsy Creek will be diverted around the open pit. The Patsy Creek Diversion will

extend around the east and south side of the open pit along a 30 m-wide bench thatwill be developed during initial open pit development in about Year 3. Flood routing

capacity will be available for events of at least a 1-in-200 year return period along anexcavated diversion channel on the bench. These flows will be conveyed towardsthe Lime Creek Diversion Tunnel intake pond. Several smaller diversion ditches willalso be included around the mine site to capture and direct upslope runoff from

coming in contact with the open pit and other mine components. All diversionstructures have been sized to transfer flows from a 200 year peak instantaneous

flood event.

A water diversion tunnel will be used to divert clean flows from the headwaters of Patsy and Lime Creeks around the TMF. The diversion tunnel will comprise an

unlined drill and blast tunnel which will be 5 m in diameter, approximately 3,100 m

long, and will have a 1% slope. A plan and profile of the tunnel can be seen onFigure 19.18.

The diversion structure at the intake of the tunnel will be sized to allow the 1:200 year

instantaneous peak flood event from both Lime Creek and Patsy Creek to be routedthrough the tunnel, and the intake dam will be approximately 35 m long and 10 mhigh, as shown on Figure 19.19. Access to the diversion structure and upstream

portal will be via a 1,400 m-long spur road on the eastern bank of Lime Creek, below

the pit access haul road, as shown on Figure 19.20.





Figure 19.18 Water Diversion Tunnel – Plan and Profile





Figure 19.19 Lime Creek Diversion Dam – Plan and Cross-section





Figure 19.20 Construction Access Road to Upstream End of Tunnel and Lime Creek Diversion Dam





The road that will connect the old Amax mill site to the proposed water diversionstructure is in an area of steep terrain. The road alignment follows along the base of

an existing waste dump slope and will require a bridge crossing over Lime Creek toaccess the upstream portal area. When PAG waste rock is placed in the south end

of the TMF, the road access to the upstream end of the tunnel may need to be

modified. Access to the downstream tunnel portal will be by way of a 5 km-longaccess road from the port site. This access road will serve as a main haul road andquarry access for the TMF starter embankment construction. The final 1,300 m will

only be used to access the tunnel portal and intake structure. This access road is inmoderately steep challenging terrain and is anticipated to be constructed in cut and

fill on rock.

The design at the downstream end allows water to be discharged from the tunnel intoan intake pond where it will be directed through a penstock to a powerhouse. A

hydroelectric power plant (HEP) producing 9.8 MW of hydropower (between 30 and

35 GWh/a) is based on sale of the clean renewable electricity to BC Hydro. The

HEP will be constructed in Year 2 of mine operations; it will include an intakestructure at the end of the diversion tunnel, and a steel penstock to the powerhouse

situated at the toe of the tailing embankment. The HEP will operate using the Lime

Creek Diversion Tunnel until Year 15, after which it will receive flow directly from theTMF. After closure, the annual inflows will be regulated in the TMF to provide a more

constant flow at the powerhouse (KP, 2009a). Further details on the HEP areprovided in Appendix K.

Storm water flows and excess water not required for hydropower production will be

separated at the intake and discharged down a spillway routed through the rockquarry excavation along on the south side of the valley below the tailing

embankment.

19.4.4 M I N E S I TE W A T ER M A N A G E MEN T

All water in contact with the mine facilities including the waste dumps, LGO stockpile,

and TMF will be collected and conveyed to the TMF pond; excess water will be

released to the environment. This includes the:

pit dewatering system, which collects direct precipitation, runoff, and

groundwater from the open pits and pumps the water to the south end of the

TMF

TMF main embankment seepage collection pond, which collects all seepage

and runoff from the main embankment downstream of the TMF and releasesthe water into Lower Lime Creek with water from the tunnel diversion

collection ditches, which divert potentially impacted water to the TMF or

sediment control ponds

sediment control ponds, which collect and treat contact water for sediment

for release to the environment.





All contact water facilities have been sized for the 200 year peak instantaneous floodevent.

19.4.5 W A T E R ST O R A G E A N D R ELEASE

The TMF will store an operating pond size of approximately 8 M m 3 and has enoughcapacity to store the 200 year peak instantaneous flood event. The main

embankment will be raised in three stages. The spillway will be located on thesouthwest side of the embankment and release water to the left bank of the Lower

Lime Creek downstream of the hydroelectric intake from the Lime Creek Diversion

Tunnel. The Stage 1 and 2 spillways have been sized for the 200 year peakinstantaneous flood event and the Stage 3 spillway has been sized for the PMF, withthe TMF attenuating the 200 year flood event.

On an annual basis, surplus water above the assumed pond capacity of 8 M m 3 atthe TMF will be pumped using the reclaim pumping system into a pipeline that runs

south along the main haul road and flows into the Lime Creek Diversion Tunnel.Preliminary results indicate that the surplus supernatant water will be suitable to

discharge once mixed with the diverted flows from Patsy and Lime creeks.

19.4.6 W A T E R B A LA N C E

A site-wide operational water balance was completed to aid in water managementand to determine if the mine site is expected to be in surplus or deficit conditions.

The water balance was conducted on a deterministic annual basis and theconceptual model of the balance is shown graphically in Figure 19.21.

Figure 19.21 Conceptual Water Balance Model

Mill water requirements will be supplied from the TMF pond and slurry water willreport to the TMF during all mining years with the exception of Years 14 and 15

where it will report to the open pit. Inputs to the water balance include net direct and





runoff precipitation, seepage losses, mill requirements, and water lost to tailing voidspaces. Input values are provided in Table 19.19 and catchment areas and runoff

coefficients are shown in Table 19.20 and Figure 19.17.

Table 19.19 Water Balance Input Values

Note: the combined output of mill requirements from the TMF and the open pit equals the total mill

requirements.





Table 19.20 Water Balance Catchment Areas

The water balance indicates that the mine site is in surplus conditions during allyears of operations with the exception of the final years, when tailing is deposited in

the open pit. Surplus water to be discharged to the environment from the TMF is in

the order of 12.5 M m3/a. Table 19.21 displays site wide annual water balance flowsfrom Year 1 to closure.





Table 19.21 Site-wide Annual Water Balance Flows – Year 1 to Closure

Notes:

1. annual precipitation = 1,985 mm; annual evaporation = 385 mm.

2. open pit runoff includes direct precipitation on open pit surface and undisturbed catchment upstream of open pit.

3. TMF embankment runoff includes area downstream of embankment crest.

4. TMF assumed to have a maximum capacity of 8 M m3

/a.5. Year 13 includes approximately 13 months of production; in Year 14, the tailing is deposited in the open pit.

6. no water treatment is required in Year 14 because tailings and surplus water are deposited in the open pit.

7. water balance results are based on an earlier waste dump and low grade ore stockpile configuration.





19.4.7 C L O S U R E

Once active mining from the pit has ceased, the low grade stockpile will be

processed for the remainder of Year 14 and a small portion of Year 15. The full

tailing stream will be discharged to the pit. Reclaim water will continue to be sourcedfrom the TMF, as the settling of tailing solids in the pit will likely not be sufficient to

source reclaim water from the pit supernatant. During this time period, an estimated25 Mt of PAG waste rock will be removed from the south end of the TMF and hauledback to the pit and placed below the ultimate flood elevation (below 515 masl).

The Patsy Creek diversion will be decommissioned so as to speed up the pit filling. If the water from the Patsy Creek diversion is required to mix with the annual dischargevolume from the site surplus, an altered strategy to fill the pit may be required,

including allowing the pit to fill naturally while the Patsy Creek diversion remains in

place until the pit spills naturally. Upon cessation of tailing deposition in Year 15, theLime Creek Diversion Tunnel will be decommissioned to the point near the discharge

of the tunnel. A 300 m tunnel connecting the TMF pond to the end of the existingtunnel will allow surplus flow from the TMF to pass into the hydroelectric scheme,generating power continuously.

The TMF pond elevation will be regulated and it is expected to fluctuate between462 masl and 466 masl throughout the year. The flood spillway elevation is set at

470 masl, to pass the PMF.

The downstream face of the TMF embankment and the remaining 25 Mt of PAGwaste rock in the south end of the TMF will be covered with growth medium, and

revegetated.

The seepage collection system at the toe of the TMF will remain active, with theseepage volume being fed into the discharge from the final spillway. The seepage

quality will be monitored, but the volume differential compared to the annual surplus

from the TMF is expected to be sufficient to negate the need for water treatment of the seepage water from this source in closure.

All pumps, pipes and other infrastructure related to the tailing, reclaim, and seepage

pumping systems will be dismantled and removed from site.

1 9 . 5 T A I L I N G M A N A G E M E N T

19.5.1 I N T R O D U C T I O N

A trade-off study was completed in early 2009 of different tailing disposal options.Six tailing facility options located at three different sites were identified, with the

location situated in Lower Lime Creek (Site 6) selected as the preferred tailing

disposal option. The trade-off study, and all subsequent pre-feasibility design for





waste and water management, can be found in the report titled, "Pre-FeasibilityStudy - Tailings Management Facility and Site Water Management", Ref. No. VA101-

343/03-3 (KP 2009b).

19.5.2 T A I L I N G M A N A G E ME N T F A C I LI T Y D E S I G N

D E S I G N BA S I S

The TMF has been designed for secure and permanent storage of all tailing from the

proposed mining operations in an impoundment created by a water retainingembankment constructed with a combination of local borrow materials and waste

rock from the mining operation. The TMF includes:

a 350 m high water retaining embankment

a tunnel diversion system that incorporates a hydroelectric generating facility

tailing delivery and distribution pipeworks

a seepage recovery system

a reclaim system to recycle water to the plant site

a surplus water discharge system to discharge excess water to the Lime

Creek Tunnel Diversion.

The TMF has been designed to meet all current CDA Dam Safety Guidelines. The

Dam Safety Guidelines assign each structure to a “Dam Class” based on theincremental losses resulting from failure of the dam with respect to loss of life,

environmental and cultural values, and infrastructure and economic losses. The

Dam Class defines the required Inflow Design Flood (IDF) and Maximum DesignEarthquake (MDE) for the design of the dam structure and water management

systems.

The TMF embankment has been classified as “Very High” under the CDA DamSafety Guidelines. However, in order to provide some additional conservatism to the

design of the major structures, on account of the overall height and location

upstream of the town of Kitsault, the design parameters for a Dam Class of “Extreme” have been selected. The corresponding IDF is the PMF and the MDE is a

10,000 year return period event.

The seismic stability assessment of the TMF has included estimation of seismically

induced deformations of the dam from the Operating Basis Earthquake (OBE) andthe MDE events. The OBE has been defined as the 1-in-475 year earthquake with amean ground acceleration of 0.08 g and design earthquake magnitude of 7.0. The

MDE corresponds to the 1-in-10,000 year earthquake with a mean ground

acceleration of 0.31 g and design earthquake magnitude of 7.5. The TMF would beexpected to function in a normal manner after the OBE. Limited deformation of thetailing embankment is acceptable under seismic loading from the MDE, provided that





the overall stability and integrity of the TMF is maintained and that there is no releaseof stored tailing or water (ICOLD, 1995). Some remediation of the embankment may

be required following the MDE. The overall design basis for the TMF is summarizedin Table 19.22. The depth-capacity relationship for the facility is shown in Figure

19.22.

Table 19.22 Overall Design Basis for the TMF – Page 1 of 3

table continues…





Table 19.22 (con’t) Overall Design Basis for the TMF – Page 2 of 3

table continues…





Table 19.22 (con’t) Overall Design Basis for the TMF – Page 3 of 3





Figure 19.22 Depth-Capacity Relationship for the TMF

Notes:

1. areas and volumes were calculated using closed polygons in AutoCAD with 5 m contour

intervals from LiDAR topography.

2. the tailing capacity does not include the reduction in volume due to the starter dam

embankment.

LA Y O U T A N D OP E R A T I N G ST R A T E G Y

Embankment

The TMF embankment is designed as a water-retaining earth/rockfill embankmentthat will be constructed in stages as the elevation of the stored tailing and ponded

water increases with time. The embankment will be constructed initially using borrow

material consisting of overburden material and quarried rock, and waste rock fromthe mining operations as soon as it is available.

The filling curve for the TMF at the design throughput of 40,000 t/d is shown in Figure

19.23. The Stage 1 embankment crest elevation has been selected to providestorage for the first two years of operation and ongoing raising of the embankment

crest will be carried over the first 8 years of operation. The material requirements for the various stages are summarized in Table 19.23, and the annual fill placing

requirements have been designed to utilize available waste rock from the open pit

over the first 8 years of operation, as shown in Table 19.24.





Figure 19.23 Filling Curve for the TMF

Notes:

1. Filling schedule is based on a throughput of 40,000 t/d for the LOM (14 years).

2. Ultimate tailing elevation assumes disposal of 190 Mt of tailing at 1.4 t/m3

(136 M m3).

3. Elevation based on tailing + 3 months of total water in slurry volume (143 M m3).

4. Elevation based on tailing + 3 months of total water in slurry + Q200 24 h flood volume

(143 M m3).

5. Freeboard consists of 2 m contingency, 2 m depth for 25 m-wide spillway and 1 m for wave.

6. Tailing assumed to have 35% solids by weight.

Table 19.23 Material Requirements for the Various Stages of the TMF

Notes:

1. Upstream fill zone assumed to have a 2:1 slope.2. Embankment volumes calculated from AutoCAD takeoff.





Table 19.24 Annual Fill Placing Requirements

Notes:

1. TMF requirements are based on embankment volumes per stage with values for in-between years interpolated to even-out haulage workload.

2. Mine waste production tonnages are from Wardrop (November 6, 2009).

3. Numbers in bold are significant years for staging

4. Upstream fill zone assumed to have a 2:1 slope.

5. Includes two 4 m-wide filter zones (upstream and downstream of core zone) plus two 2 m-wide filter zones on the upstream face of Stage 1.

6. Core zone is 8 m wide.7. Waste rock density = 2.3 t/m3.





Stage 1 of the embankment is designed as a membrane-faced, water-retainingstructure with Stages 2 and 3 designed as centreline raises with a vertical low

permeability till core and adjacent filter zones. Seepage control is provided by theupstream membrane and vertical till core that will be tied in to the bedrock

foundations and a grout curtain at the abutment contacts.

Water Management

Overall water management is the key to the success of all mining waste

management systems. The TMF is located in the lower reaches of Lime Creek andhence has the ability to collect all runoff from areas impacted by the miningoperation, which can then be recycled for reuse or treated prior to discharge to Lower

Lime Creek. The project is located within an area where a significant overall surplus

of water is unavoidable and hence runoff from upstream areas not impacted by themining operation will be diverted to the maximum practicable extent.

Diversion of the major portion of the upstream catchment around the TMF will beachieved using a diversion tunnel starting at the confluence of Lime Creek and PatsyCreek and located on the left (south) bank of Lime Creek and terminating at a point

downstream of the TMF embankment where the diverted flows will discharge into asurface spillway.

Diversion of runoff to the maximum practicable extent has been achieved by avoiding

any impact to Patsy Creek and making provision for the diversion of Patsy Creekaround the open pit on a bench on the south wall of the pit.

Water accumulating in the TMF surface pond will be recycled for reuse in the

process as required. Surplus water will be discharged into the diversion tunnel

where it will mix with runoff from upstream areas not impacted by the miningoperation and ultimately be discharged to lower Lime Creek.

The crest elevations of the TMF embankment raises have been designed to providesufficient freeboard to contain the runoff from a 1-in-200 year storm assuming all thediversion systems fail. Notwithstanding this, each embankment crest will include an

emergency spillway so that the safety of the dam can never be compromised. Thefinal crest will contain a spillway capable of handling the runoff from a PMF eventover the entire upstream catchment.

Tai l ing Dis t r ibut ion

Tailing will be discharged from the mill as an unthickened slurry. The design

objective is to discharge the tailing slurry from the embankment face as far aspracticable in order to provide additional seepage control and keep the surface pond

remote from the embankment. The tailing distribution system has been designed for

ease of operation and to minimize the number of pipeline moves during operations.





PAG Waste Storage

The waste rock excavated during mining operations will contain both PAG waste andNPAG waste. Construction of the TMF embankment requires the use of all of themine waste from the first 3 years of operations and much of the waste rock for Years

4 to 8 of operation; hence, the TMF embankment will likely contain both PAG andNPAG waste rock.

Construction of the TMF embankment requires placing of large tonnages of waste

rock in the embankment shell zones. The waste rock will be placed in layers ranging

from 1 m thickness in the upstream half of the embankment to 10 m thick in thedownstream shell zone. Some limited material selection and placing will be possible

during construction, and every effort will be made to contain the PAG waste materialwithin the upstream half of the embankment. Seepage through the embankment will

be controlled with the provision of drainage layers that will minimize the contact

between seepage and the waste rock embankment zones. During construction of Stage 3, which comprises the final downstream slope of the embankment, provisionscould be made to include a low permeability capping layer to minimize infiltration of

precipitation into the embankment zone and reduce the supply of oxygen.

Following completion of the TMF embankment in Year 8, all waste rock for theremainder of the mining operations will be stored in the upper reaches of the TMF,

downstream of the diversion dam and tunnel inlet. This waste will be fully contained

within the TMF during operations and options exist at closure to relocate any PAGwaste above the final water level in the TMF into the mined out open pit, which will

be flooded. The base case will allow for relocation of half of the total waste rockvolume from this location to the pit at closure. Refined waste characterizationstudies during the feasibility stage will dictate how much, if any, must be relocated.

19.5.3 E M B A N K M E N T C O N S T R U C T I O N

ST A G E 1 C O N S T R U C T I O N

The TMF embankment is located in the steep-sided Lime Creek Valley and access

for initial construction is a major consideration. Initial access to the valley bottom canmost easily be achieved on the left (south) bank and this requires the construction of

an access road from the port site up the Roundy Creek Valley, then north across analluvial terrace and up along the left bank of the Lime Creek Valley. Steep access

roads, suitable for 6-wheel drive haul trucks, can be constructed to the upstream and

downstream toes of the initial embankment, as shown in Figure 19.24 and Figure19.25. The access roads pass through the three identified sources of borrowmaterial and a potential rock quarry downstream of the embankment. A second

potential rock quarry is located on the right bank of Lime Creek just upstream of the

embankment and can be accessed from the existing Kitsault town site access road.This rock quarry will only come into play once the embankment is constructed up to





an elevation where access to the embankment crest can be achieved from the rightbank.

Initial construction of the embankment will require the use of borrow materials from

the borrow areas located along the access road (Figure 19.24) and from the rock

quarries located downstream and upstream of the embankment (Figure 19.25).These borrow sources are required because the volume of waste rock from the openpit is not sufficient prior to start-up to justify the expense of constructing the haul

road. The volume of borrow material required to achieve this is approximately

16 M m3, as shown in Table 19.23.

Stream control during construction will be achieved with the construction of a

diversion conduit along one side of the river valley under the entire impacted length

of the creek, as shown in Figure 19.26.

At the upstream end, the diversion conduit will be concrete encased where it passes

through the liner system on the embankment face and will have provision for the

installation of a concrete plug. The diversion conduit has been sized to pass thepeak flow from a 1-in-10 year storm from the entire catchment upstream of the dam.This condition will exist for a short period of time when construction of the

embankment is commencing and, as the height of the embankment increases, someflood attenuation capacity will develop allowing greater storms to be passed without

the embankment overtopping. The diversion conduit will be plugged as soon as

impounding of water is required to provide enough water for start-up.

A portion of the mine waste haul road will be constructed from the open pit to theTMF embankment during the pre-strip time period using the waste rock coming from

the pit, as shown in Figure 19.27. The haul road will run from the exit point from the

open pit along the alignment of an existing road to the right bank of the TMFembankment. The balance of the road will be constructed during the first year of

operations as more waste is released from the pit.





Figure 19.24 Construction Access Roads to Borrow Areas





Figure 19.25 Construction Access Roads to TMF Embankment and Downstream Portal of Diversion Tunnel





Figure 19.26 Main Embankment Construction Stream Diversion





Figure 19.27 Haul Road from Open Pit to TMF Embankment – Stage 1





Seepage control through the Stage 1 embankment will be achieved with an upstreamgeomembrane liner system. A series of benches are provided on the upstream face

to allow for liner system installation on the lower sections prior to completion of theentire Stage 1 embankment. An overlying concrete-filled geocell system is also

included on top of the geomembrane system in the lowest section, where the valley

sides are very steep, to protect the geomembrane from damage by falling rocks. Atthe abutment contact between the geomembrane liner and the bedrock, a groutcurtain will be installed to reduce seepage through the bedrock abutments.

Construction of the Stage 1 embankment will include placing of fill to the ultimatedownstream toe of the embankment in order to allow for construction of the seepagecollection/monitoring system and the hydroelectric facility powerhouse.

The crest elevations of the TMF embankment raises have been designed to providesufficient freeboard to contain the runoff from a 1-in-200 year storm assuming all thediversion systems fail. Notwithstanding this, the Stage 1 embankment crest includes

an emergency spillway on the left abutment so that the safety of the dam can never be compromised.

ST A G E 2 A ND 3 C O N S T R U C T I O N

Placing of mine waste rock in the TMF embankment will be a continuous process for

the first 8 years of operation. Construction will involve two major lifts from thedownstream toe to a new crest elevation, and material hauling will be achieved by

the inclusion of a ramp system on the downstream face. Spreading and compactionof the waste rock will be in lifts with the thickness of the lifts increasing towards the

final downstream face.

Seepage control through the Stage 2 and 3 embankments will be achieved with avertical low permeability till core contiguous with the geomembrane liner system

above the Stage 1 embankment crest. At the abutment contact between the till core

and the bedrock, a grout curtain will be installed to reduce seepage through thebedrock abutments.

The crest elevations of the TMF embankment raises have been designed to provide

sufficient freeboard to contain the runoff from a 1-in-200 year storm assuming all thediversion systems fail. Notwithstanding this, the Stage 2 embankment crest will

include an emergency spillway on the left abutment so that the safety of the dam cannever be compromised. The Stage 3 embankment crest will include an interim

spillway during operations and this spillway will be modified at closure for the finaldesign criteria of a PMF over the entire catchment area.

SE E P A G E A N D ST A B I L I T Y AN A L Y S E S

Preliminary seepage and stability analyses have been carried out based on the

information gathered from the site investigation program, assumed fill materialparameters, and the selected earthquake design basis. The analyses indicate that





the total seepage through the final embankment will be in the order of 12 L/s, whichwill be collected by the seepage recovery system. Stability analyses indicate that the

embankment will meet all required factors of safety for stability and that predictedembankment settlement, when subjected to an MDE event, will be in the order of

0.5 m, which is acceptable.

19.5.4 T A I L I N G , R E C L A I M A N D F R E S H W A T E R SY S T E M S

Four major external pipelines manage the process streams associated with the

project. This section describes the design considerations and assumptions used to

establish pipeline sizes and material selection, pumping requirements and equipmentselection for these major pipelines. These are:

bulk tailing pipeline

reclaim water pipeline

surplus water pipeline fresh water pipelines.

In general, pipelines are constructed using high density polyethylene (HDPE) pipes

to the greatest extent possible, with steel pipes used only where line pressuresexceed the capacity of HDPE pipes, or where steel pipes are more economical.Wherever possible, pipelines are designed to be free-draining with drainage outlets

at low points, as required, and with air valves or vents at high points.

The pipelines are typically laid on grade alongside existing, or purpose built accessroads and are protected from vehicle traffic by concrete, earth barriers, or local

burial. The thermal expansion and contraction of pipelines is managed with localpipeline anchorage, burial or cover, and corridors of dimensions sufficient toaccommodate pipe snaking, as required for HDPE pipes. Local select fill cover is

used at appropriate locations to reduce potential damage from rock-fall. Pipelinecrossings of access roads are made in buried culverts, to allow for inspection and

replacement without need for road closure.

PR O J E C T D E V E L O P M E N T

The tailing system uses one TMF over the life of the mine. The requirements of the

system change with the development of the TMF embankment. Subsequently, the

design of the system is divided into four periods to facilitate the development of adetailed cost estimate for the life of the system. Table 19.25 outlines the four periods.





Table 19.25 Tailing System Design Periods

Year/Period Stage

TMF

EmbankmentElevation

End of Year -1 Stage 1: Start-up 380 maslEnd of Year 3 Stage 2 435 masl

End of Year 8 Stage 3 475 masl

End of Year 15 Stage 4: Closure 475 masl

S I T E A N D PR O C E S S D E S C R I P T I O N

Kitsault is an open pit mine located near the headwaters of Alice Arm Inlet,approximately 120 km north-west of Terrace, BC. The elevation for various systemcomponents varies between approximately 100 and 1100 masl. Ore from the open

pit is processed in the mill, located to the north of the pit. A single TMF is used for the bulk tailing storage over the life of the mine. Four major external pipelines

manage the process streams associated with the project. One pipe carries the bulk

tailing from the mill to the TMF for storage. A reclaim line returns water pumped fromthe TMF back to the mill for reuse. Seepage water is collected in a seepage pond

downstream of the TMF-confining embankment and is released into Lower LimeCreek with the surplus water exiting the diversion tunnel. Surplus water is pumped

from the TMF surface water pond to the Lime Creek Diversion Tunnel and is

discharged back to the environment. Fresh water for the mill is pumped from ClaryLake to a fresh water holding tank at the mill. A separate fresh water system is usedto supply the camp with potable water from Roundy Creek.

The design data and the block diagrams for various process streams are presentedin Table 19.26 and Figure 19.28, respectively. The site plan and pipeline routes for the mine development through various stages are shown in Figure 19.29 through

Figure 19.33, with details shown on Figure 19.34.





Table 19.26 Design Data

Note: data represents best estimates/assumptions for the PFS.





Figure 19.28 Block Diagrams for the Various Process Streams





Figure 19.29 Tailing, Reclaim, and Water Management Pipelines – Stage 1 (Year -1)





Figure 19.30 Tailing, Reclaim, and Water Management Pipelines – Stage 2 (Year 3)





Figure 19.33 Camp Potable Water Supply





Figure 19.34 Typical Pipework Details for TMF





PR O C E S S B A T T E R Y L I M I T S

The process battery limits are detailed in Table 19.27.

Table 19.27 Process Battery Limits

Process Stream Battery Limits

Bulk Tailing The design includes the bulk tailing handling system from downstream of the

discharge valve out of the bulk tailing box, to the discharge points along theTMF.

Reclaim Water The design includes the reclaim water handling system from the TMF

supernatant pond to the discharge connection into the plant reclaim water head tank at tank invert elevation.

Surplus Water The design includes the surplus water handling system from the TMFsupernatant pond to the Lime Creek Diversion Tunnel.

Fresh Water The design includes the fresh water handling system from the intake at Clary

Lake to the fresh water tank at the process plant. A separate fresh water

system supplies the camp with potable water and the design includes thehandling system from an intake at Roundy Creek to a water treatment plant atthe camp and includes a water holding tank.

B U LK TA I L I N G

Bulk tailing are carried in a single pipe from the mill to the TMF for storage. The flowis by gravity. It is anticipated that the difference in elevation provides sufficient head,throughout the mine life, for discharge into the TMF. This pipeline can be divided

into two components: the delivery pipeline from the mill to the rock haul road at the

TMF, and the deposition pipeline along the rock haul road and the TMFembankment.

Bulk Tai l ing Character is t ics

The size distribution and rheology for the bulk tailing are not available at this time.

Based on similar systems, it is assumed that the tailing slurry behaves as a two-phase settling slurry. The slurry concentration is assumed to be 36.4% dry byweight, with the solids density of 2.7 t/m3. The total tailing production is 14.6 Mt/a,

with solids amounting to 1,804 t/h. The design flow rate for the bulk tailing pipeline is

3,804 m3/h.

Del ivery Pipel ine

The delivery pipeline carries bulk tailing from the discharge valve out of the bulktailing box at the mill (elevation 755 masl) to the rock haul road (Stage 1 elevation440 masl) that connects the open pit with the TMF embankment. A considerable

elevation difference of 315 m exists between the mill and the rock haul road and the

flow is by gravity throughout. The current design utilizes existing or already planned





with each stage, and varies from approximately 2,200 m long with a head drop of 60 m in Stage 1, to approximately 2,700 m long with an essentially horizontal

discharge section in Stage 3.

Factors considered in the deposition pipeline layout design, pipe material and pipe

size selection include:

The deposition pipeline must remain in service and be regularly relocatedand reused in conjunction with ongoing embankment raises.

The pipeline has the same pressure rating and jointing methods for theentire length to simplify construction, relocation, and maintenance.

The availability of the pipe material in appropriate sizes.

The abrasivity of the tailing.

The higher operating velocities experienced in a system operating under gravity flow conditions on steeper sections.

Minimum deposition velocities experienced in a full pipe system operatingunder pressure for less steep and horizontal sections. For bulk tailing, aminimum design velocity of 2 m/s is assumed.

Based on these selection parameters, a 30” DR 9 HDPE pipe is consideredconservative and suitable for the bulk tailing deposition pipeline, with gravity flowalong the steeper sections, and pressure flow along the horizontal or near horizontal

sections. Initially, discharges will be established at each end and at the centre of the

embankment crest, with large diameter single point off-takes. Additional dischargesmay be installed at the junction with the waste rock haul road and along this road,

and also on the embankment crest as it is lengthened through different stages.

Tailing beaches are developed and maintained against the TMF embankment toensure the establishment of significant seepage paths between the surface water

pond and the embankment.

Each discharge point consists of a lined steel tee, two knife gate valves, andassociated discharge piping. Two valves are required to open the discharge point

and to isolate the downstream pipeline. This will prevent solids build-up in anyisolated downstream section of the line and eliminate the risk of the tailing freezing in

cold weather. The role of the discharge piping is to direct the tailing flow as required.

To prevent erosion of the natural ground, this piping runs from the tee to just abovethe tailing beach, and is shortened as the beach rises. To prevent negativepressures in the discharge pipe, a vent is needed at the top of the line.

R E C L A I M W ATER

Reclaim water is pumped from the TMF surface water pond to a reclaim water holding tank at the mill for reuse in the process. The elevation of the surface water





pond increases rapidly in the initial years of operation, which significantly impacts thepumping requirements and the design of the system.

Reclaim Water Pipel ine

The reclaim water pipeline is a single line that runs from the TMF surface water pond

(elevation varies) to the reclaim water holding tank at the mill (elevation 755 masl).The proposed reclaim water intake is located approximately halfway between the

open pit and the TMF embankment on the west side of the TMF pond. The water

surface elevation changes from about 300 masl during the initial years, to about470 masl by the end of the operations. In the initial years, the exposed side walls of the TMF are extremely steep and limit the choices for a pumping system. Water is

pumped from the pond up to a booster pump station located near the rock haul road.

From there, the pipeline follows this road northward to the intersection with the millaccess road. At this intersection, the reclaim line joins the bulk tailing pipeline, and

both follow the same route to the mill. The total length of the reclaim water pipeline

is approximately 4,300 m with a design flow of 3,600 m3

/h.

In addition to reclaim water, surplus water from the TMF pond is withdrawn at the

same location and is also pumped to the booster pump station. The pipeline fromthe TMF pond to the booster station and the pumping system in the TMF pond areshared between the two streams. Surplus water adds 1,440 m3/h to the pumping

requirements from the pond. The surplus water pipeline from the booster station to

the discharge point is described in more detail later in this section.

Factors considered in the reclaim water pumping system, pipeline layout design, pipe

material and pipe size selection includes:

The pipeline is placed in permanent locations as much as reasonablypossible and along existing road structures. The construction of new roads

is minimized.

The local topography is challenging with very steep canyon walls at the TMFfor the initial stages.

Sections of the pipeline should have the same rating and coupling system to

simplify construction and maintenance.

Pumping requirements and design pressures decrease significantly with

rapid water surface rise in the TMF.

Based on these selection parameters, a 30” steel pipe is considered suitable to bringthe water from the TMF water surface pond to the reclaim water holding tank at the

mill. Due to considerable head requirements, pumping is in two stages. The first

pumping stage, from the TMF pond to a fixed booster pump, will requiremodifications to accommodate for the rapidly raising water level in the pond. Thesecond stage is from the booster pump to the mill and is permanent for the life of the

project.





Due to the significant variation in pumping head requirements, the system used for the first pumping stage for water withdrawal from the TMF pond may need to be

changed during the life of the project. In the initial years, an array of multi-stagesubmersible pumps, laid in large diameter steel pipe shrouds is used. These pipes

are set at a steep angle against the rock face. Each pump is supported on a sled,

which enables the pumps to be winched to a higher elevation and kept clear of tailing. At least one spare standby pump is required to allow system modificationsand, when required, to accommodate reducing head requirements. As the water

level rises and the steep cliffs are flooded, the withdrawal pumping system could bereplaced with barges equipped with pumps.

The booster pump station contains separate pumps for delivering reclaim water to

the mill and for conveying surplus water to the diversion tunnel. The reclaim water requirements in terms of flow rate, pumping head, and pipeline length do not change

for the life of the project between the booster pump station and the mill.

SU R P L U S W ATER

Surplus water from the TMF surface water pond is discharged to the Lime CreekDiversion Tunnel and back to the environment. The pipeline and the pumping

system from the TMF to the booster pump station are shared between the reclaimand the surplus water systems. This section describes the surplus water pipeline

from the booster station to the discharge point.

Surplus Water Pipel ine

The surplus water pipeline is a single line that runs from the reclaim water booster

pump station to the Lime Creek Diversion Tunnel (elevation 500 masl). Thediversion tunnel starts a short distance upstream of the south end of the TMF pond atthe Patsy Creek and the Lime Creek confluence. The surplus water pipeline is

approximately 2,400 m long and follows the rock haul road southwards. The design

flow rate is 1,440 m3/h.

Factors considered in the surplus water pipeline layout design, pipe material and

pipe size selection include:

The pipeline is placed in permanent locations as much as reasonablypossible and along existing road structures.

Sections of the pipeline should have the same rating and jointing methods tosimplify construction and maintenance.

Pumping requirements and design pressures remain constant for the life of the project.

Based on these selection parameters, a 22” DR 13.5 HDPE pipe is considered

suitable to transfer the water from the reclaim water booster station to the dischargepoint at the diversion tunnel. Due to significantly different flow rate and pumping





head requirements from the reclaim water line, a separate pumping system is used inthe booster station for surplus water.

The TMF surface water pond chemistry has not been determined up to this time.

Preliminary mixing studies by SRK Canada indicate that water discharged from the

supernatant pond would be sufficiently diluted after mixing with water from Lime andPatsy Creeks in the diversion tunnel, and that water treatment would not benecessary. Future studies are needed to assess if water treatment is essential prior

to discharge back to the environment.

SE E P A G E WA T E R

A seepage water pond is located at the toe of the TMF embankment to intercept any

seepage water from the TMF and also to collect local runoff. The water from the

seepage pond is released into Lower Lime Creek with the diverted water from theLime Creek Diversion System. If the water quality is not suitable for release, a

pumping system could be installed to pump the water back to the TMF.

FR E S H W A T E R

Two separate fresh water systems are designed to provide clean water for various

purposes of the project. One of the systems provides water for process use, fire and

potable water in the mill, while the other system provides potable and fire water for the camp.

Fresh Water for the Mi l l

Clary Lake was identified as the most appropriate fresh water source for the mill,primarily based on the challenging topography for pipelines from other potential fresh

water sources, as well as the length of new roads that would need to be built. A

single pipeline connecting Clary Lake to a fresh water tank at the mill is used toprovide clean water for process use, fire water, and potable water. The water is

pumped from an intake structure at the lake (elevation 650 masl) to a fresh water holding tank at the mill (765 masl). The total length of the pipeline is about 5,100 m,while the design flow rate is 30 m3/h. Clary Lake water quality has not been

determined at this time and the level of treatment for potable water will be assessed

in future studies.

Factors considered in the fresh water pipeline layout design, pipe material and pipe

size selection include:

The pipeline is placed along existing road structures as much as reasonablypossible. The construction of new roads is minimized.

The terrain topography is limiting for some of the options.





The pipeline should have the same pressure rating and jointing methods for

the entire length to simplify construction and maintenance.

Pumping requirements and design pressures.

Based on these selection parameters, a 6” DR 6.3 HDPE pipe is considered suitableto bring fresh water from the lake to the mill. The pipeline follows the existing roadthat leads to the town of Kitsault, up to the intersection with the planned plant access

road. From this point, the pipeline follows the plant access road to the mill and runsin parallel with the bulk tailing and reclaim water pipelines. An intake structure with

screens and a pumping station are required at Clary Lake. The design must

consider freezing conditions in cold winter months. A drain valve is planned at a lowpoint in the line to allow for pipe draining and to prevent freezing.

Fresh Water for the Camp

A 600-person camp is located on the Roundy Creek delta at Alice Arm inlet.

Considering the long distance between the mill and the camp, a separate fresh water system from the Clary Lake system was identified as most appropriate to supply

potable and fire water for the camp. The water is withdrawn from Roundy Creekapproximately 1,000 m upstream from the camp. Gravity flow brings the water from

the intake structure (elevation 75 masl) to a holding tank (elevation 50 masl), andfurther down to a water treatment plant at the camp (elevation 15 masl). The totallength of the pipeline is about 1,000 m, while the design flow rate is 4 m 3/h. Flow

rate and water quality for Roundy Creek have not been determined at this time.

Future studies are needed to assess if there is sufficient flow year-round and thenecessary level of treatment for potable water.

Factors considered in the fresh water pipeline layout design, pipe material and pipesize selection include:

The pipeline is placed along existing and/or planned road structures as

much as reasonably possible. The construction of new roads is minimized.

Runoff from the road surfaces impacts water quality and is a limiting factor for the intake location.

Sediment transport in Roundy Creek is likely considerable.

The pipeline has the same pressure rating and jointing methods for theentire length to simplify construction and maintenance.

The holding tank has sufficient capacity to satisfy needs for potable water for the camp and contain the necessary volume of water for fire fighting.

Based on these selection parameters, a 3” DR 15.5 HDPE pipe is considered

suitable to bring fresh water from the creek to the camp. For its entire length, thepipeline follows the planned road that connects the camp with the mill. The intake

must include a sediment control structure and be designed for freezing winter conditions. A holding tank is planned approximately 300 m from the intake and is at





a lower elevation to allow for filling under gravity. The tank is 8 m in diameter and5 m high to satisfy the capacity requirements. From the holding tank, the pipeline

continues along the road to the camp where it splits into two branches. One branchis for fire fighting, while the other branch supplies potable water for the camp and

goes through a water treatment plant. With an elevation difference of about 35 m

between the tank and the camp, the pressure requirements for both fire fighting andthe distribution system are likely satisfied.

O V E R V I E W

The following is an overview of the PFS pipeline design for Kitsault tailing and water management systems:

Bulk tailing will be transported by gravity from the mill to the TMF in a 30”

HDPE pipeline.

Reclaim water will be transported by pumping from the TMF surface water pond to the mill in a 30” steel pipeline.

Surplus water will be transported by pumping from the TMF surface water

pond to the Lime Creek Diversion Tunnel in a 22” HDPE pipeline.

Seepage water will be collected at the toe of the TMF embankment andreleased to Lower Lime Creek.

Fresh water will be transported by pumping from Clary Lake to the mill in a6” HDPE pipeline.

Fresh water will be transported by gravity from Roundy Creek to the camp in

a 3” HDPE pipeline.

1 9 . 6 W A S T E R O C K A N D T A I L I N G G E O C H E M I C A L C H A R A C T E R I Z A T I O N

19.6.1 W A S T E R OC K M E T A L L E A C H I N G A N D AC I D R O C K D R A I N A G E P O T E N T I A L

A C I D RO C K DR A I N A G E PO T E N T I A L

Initial characterization of the potential for metal leaching and acid rock drainage

(ML/ARD) has occurred as a component of Avanti’s exploration drilling program todefine the resource for the expansion of the existing pit. Every interval was analyzed

for sulphur concentrations as part of a multi-element ICP scan. Every fourth intervalwas also analyzed for carbonate content.

Sulphur concentration determined by the ICP method was used as a surrogate for

total sulphur concentration determined by direct methods. Typical sulphur concentrations were between 1 and 2%. Geological information and subsequent

analysis of a subset of samples showed that calcium sulphate minerals occur in





some parts of the proposed open pit resource. This indicates that total sulphur concentrations will tend to over-estimate acid potential (AP) due to sulphide mineral

oxidation where sulphate minerals are present. As a result, Avanti used a regressionequation to estimate sulphide content from sulphur content. The current assessment

is that correction of total sulphur for sulphate content does not significantly affect the

overall prediction of ARD potential for the waste rock but this is being investigatedfurther as the ML/ARD characterization of the project proceeds.

Carbonate content was used as a surrogate for acid neutralization potential (NP). In

situations where the main carbonate minerals are composed of calcium andmagnesium carbonates, the surrogate provides a reliable indication of NP becausethese minerals are effective acid consumers. If iron and other heavy metal

carbonates are present, carbonate content over-estimates NP because theseminerals provide limited acid neutralization. Ongoing mineralogical studies are

quantifying carbonate mineralogy to evaluate the suitability of carbonate as a

surrogate for NP.

Avanti used the sulphur and carbonate data to construct a waste block model for AP

and NP. Values for individual blocks were then used to calculate the ratio NP/AP.

The distribution of ratios was initially assessed with respect to conventional genericclassification for ARD potential. If NP/AP was less than 1, the rock was consideredto be potentially ARD generating (PAG), whereas if NP/AP was greater than 2, the

rock was classified as NPAG. Ratios between these values were considered to be

uncertain with respect to ARD potential. Site specific classification criteria can bedeveloped using testing and these tend to be between 1 and 2. The distribution of

NP/AP values (Figure 19.35) showed that 5% of samples were classified as NPAGand 43% were classified as PAG leaving 52% as having uncertain classification. For

the purpose of the PFS, the following was assumed:

95% of rock was nominally classified as PAG (NP/AP <2) because thedetailed data to develop a site specific criterion are being obtained and,based on experience, the BC MEMPR will likely consider rock with NP/AP

below this value as PAG for the purpose of assessment and permitting.

Limited previous testing for the site indicated that a site specific criterionnear 1 may be appropriate. For the purpose of defining rock with relatively

lower risk of ML/ARD potential, a placeholder NP/AP criterion of 1.1representing the 50th percentile of the distribution was used for wastemanagement planning.

Management of ARD potential should be considered in project designs.

This could include segregation of rock based on ARD potential to maximizeuse of rock with the lowest ARD potential for construction, provision for subaqueous disposal of rock with highest ARD potential, and inclusion of a

future water collection and treatment due to overall PAG nature of the rock.





Figure 19.35 Distribution of NP/AP Indicated by Block Modelling

The existing waste rock dumps at the site, some of which originate from mining in the1960s, are currently not showing ARD though acid generating rock is observed on

surface, and drainage contains elevated sulphate concentrations. The lack of ARD

may reflect a number of factors including composition of the rock and delay of acidicconditions due to slow depletion of carbonate minerals. This evidence implies that

ARD is not likely to start in the mine life and that waste rock and pit water

management plans can allow for a lag time of decades when consideringmanagement of PAG rock and ARD.

ME T A L LE A C H I N G PO T E N T I A L

Metal leaching refers to the release of regulated potential contaminants at levels that

may not be acceptable for direct discharge to the environment regardless of drainagepH. Depending on the site specific application of water quality guidelines, a number

of elements may be a concern in non-acidic drainage.

For Kitsault, parameters potentially of concern based on whole rock composition andwaste rock dump seepage monitoring are molybdenum (from oxidation of

molybdenite), cadmium and zinc (oxidation of sphalerite), and sulphate (oxidation of

pyrite and leaching of gypsum). Locally elevated concentrations of lead (associatedwith galena), and antimony and arsenic (associated with sulfosalts) are present in therock but are not currently observed at significant levels in the waste rock dump





seepage. Selenium is commonly associated with sulphide deposits and is regulatedat low levels. Seepage chemistry shows low levels of selenium. Data on the

occurrence of selenium in the bedrock is currently being obtained.

19.6.2 T A I L I N G

Tailing samples have been produced from metallurgical testing of three rock-type

based ore composites prepared by Avanti. The three composites produced similar tailing implying that rock type was not an important variable. Total sulphur

concentrations (dominantly as sulphide) in whole tailing varied from 1.2 to 1.8%.

NP/AP varied from 0.7 to 0.9 leading to the conclusion that the tailing have potentialfor ARD and need to be appropriately managed. The current closure strategyincludes a permanently f looded tailing impoundment.

Analysis of metals in the tailing and testwork supernatants showed that leaching of molybdenum, zinc, cadmium, sulphate, and fluoride may need to be considered for

disposal of the tailing.

1 9 . 7 R E C L A M A T I O N A N D C L O S U R E

19.7.1 G E N E R A L

The Kitsault Project will be developed, operated, and closed with the objective of

leaving the property in a condition that will mitigate potential environmental impactsand restore the land to an agreed to land use and capability. Closure andreclamation activities (Figure 19.36) will be carried out concurrent with mine

operation wherever possible, and final closure and reclamation measures will beimplemented at the time of mine closure.





Figure 19.36 Closure Activities





The proposed mine development is located on a very rugged and steep terrain with asignificant portion of the development area already disturbed from the former Kitsault

Mine activities. It is anticipated that there will be a very limited source of good qualitysoil/growth medium available for salvage and for use in reclamation. Therefore, it is

recommended to minimize the amount of new disturbance and utilize the already

disturbed areas for the proposed developments, as much as possible.

The following sections provide an overview of the proposed reclamation and closure

plan for the Kitsault Mine. A number of studies and investigations are already

underway that will be used to develop the final reclamation and closure plan. A brief overview of this work is provided in Section 19.7.4. Section 19.7.5 provides adiscussion of the proposed reclamation activities for each of the main components of

the mine. Section 19.7.6 presents the basis of costs to undertake the reclamationactivities.

19.7.2 R E C L A M A T I O N U N I T S

For the purposes of reclamation planning, the Kitsault Project has been broken down

into the following key reclamation units:

open pit

waste rock dump and LGO stockpiles

TMF including the tailing embankment

mine site facilities (process building, truck shop, conveyors, crusher, andcrushed ore stockpiles)

infrastructure (camp and wharf facility in Alice Arm)

access and haul roads

surface water diversion structures

borrow areas.

Under published guidance, in support of the Application Requirements for a Permit

Approving the Mine Plan and Reclamation Program Pursuant to the Mines Act

R.S.B.C. 1996, C.293, the BC MEMPR has established key information that is to beprovided in the reclamation program component of the Environmental Assessment

application. These requirements are summarized as follows:

proposed end land use objectives land capability or productivity and proposed post-mine capability or

productivity objectives for all significant land uses (this information is

required to create the property reclamation program and is used as ameasure of reclamation success)

plans for characterizing the soils and overburden resource for reclamation

purposes





plans for salvaging, stockpiling, and replacing soils and other suitable growth

media

consideration of future erosion and mass wasting for long-term stability

treatment of structures and equipment

reclamation of water courses

pit lake

tailing impoundment reclamation

road reclamation

pre- and post-mine trace element concentrations in soils and vegetation

the general composition, size, shape, and location of all consolidated and

unconsolidated geological units disturbed by the project

prediction of the geochemical performance of the various geological units in

the form which they will be exposed, and a determination of the potential for deleterious effects

determination of disposal and remediation methods, their effectiveness, and

quantities by area requirements;

determination of monitoring requirements for extraction, waste handling, anddisposal operations

determination of the time to onset of ML/ARD in materials where there is a

delay in the application of remedial measures

programs for prevention, treatment, and control of ML/ARD

toxic chemical disposal environmental monitoring

preliminary characterization of surficial and bedrock materials for

geotechnical assessments.

Additional requirements include the preliminary design of:

ore processing facilities

TMF

waste rock dump and the LGO stockpile

open pit

access roads

other significant transportation or utilities infrastructure.





19.7.3 R E C L A M A T I O N OB J E C T I V E S

Under the BC Mines Act and the Health, Safety, and Reclamation Code for BC, the

primary objective of the reclamation plan will be to return, where practical, all areasdisturbed by mining operations to acceptable land use and capability. The following

goals are implicit in achieving this primary objective:

long-term preservation of water quality downstream of decommissionedoperations at the compliance point

long-term stability of engineered structures, including the waste rock dumps,

open pit, and TMF

removal and proper disposal of all access roads, structures, and equipmentthat will not be required after the end of the mine life

long-term stabilization of all exposed erodible materials

natural integration of disturbed areas into the surrounding landscape, andthe restoration of a natural appearance to the disturbed areas after mining

ceases, to the best practical extent

establishment of a self-sustaining cover of vegetation that is consistent with

existing forestry and wildlife needs.

19.7.4 B A S E LI N E ST U D I E S

ST U D Y O B J E C T I V E S

Information gathered from the soils and terrain mapping program will be used in

developing the reclamation and closure plan for the mine site and associatedfacilities. Establishing an inventory of the soils that exist prior to mine development

creates a baseline that is used as a goal in land reclamation objectives at closure.The soil survey provides information on soil type, extent, depth, and suitability for use

in reclamation. The final reclamation plan will be based on the pre-mining land

capability, baseline soil metal levels, and the end land use objectives and will includestrategies such as soil salvage, stockpiling, and replacement strategies, progressivereclamation, and interim and final erosion control plans.

The objectives of the reclamation planning and closure work are to:

identify sand soil units and their distribution in the areas that will bedisturbed as part of the mine development

quantify topsoil and subsoil volumes that may be available for use as a

growth medium for reclamation at various mine facilities

determine the location, depth, and volume of the soil types for salvage, andidentify potential soil stockpile locations and develop replacement thickness

criteria and volume objectives





develop soil handling plans for salvaged and stockpiled soils and for final

reclamation use

estimate post-mine capability and/or productivity of the reclaimed land units

develop erosion control and revegetation plans for disturbed lands and

stockpiled soil

develop mitigation and management plans to address the potential effects

develop monitoring and feedback maintenance plans as a means to manage

the site through to a self sustaining ecosystem.

ST U D Y A R E A A N D METHODOLOGY

The study area for the soil reclamation planning and closure program is focused on

the areas that will be disturbed during the mine site and marine site facilitiesdevelopment. This includes the pit area, plant site, shops, waste rock dumps, TMF,

access roads, port, and camp.

The soil and terrain mapping that is being carried out as part of the terrestrialecosystem mapping will include information on the texture, surficial material,

qualifying descriptors, surface expression, and the geomorphological processes.

This information along with a further detailed field assessment of the soils on thedirect footprint of any project component will be used to determine the suitability of the soils for salvage and for use in reclamation. This soils information will be

integrated into the mine plan, spatially and temporally, to develop a soils handlingplan for the project.

Post-mine landforms will be provided by the mine planning consultants (primarily

based on geotechnical stability issues). Reclamation objectives for the various post-mine landforms will be developed on an inter-disciplinary basis with dueconsideration for current land use and other values (especially terrestrial and aquatic

resources).

W O RK C OMPLETED

A detailed field assessment of the soils for reclamation and closure planning wascompleted in September 2009. Relevant information was collected during the fieldassessment including soil depth, texture, coarse fragment content, root depth and

quantity, vegetation type, surface coverage, as well as signs of compaction and

erosion. The assessment was also conducted on the areas reclaimed at closure of the former Kitsault Mine.





W O RK R E M A I N I N G F O R PE R M I T A P P L I C A T I O N

Data interpretation is underway. The preparation of salvage maps is to follow the

receipt of a completed pre-feasibility mine plan. Preliminary closure costing will beprepared based on the pre-feasibility mine plan.

Integration of the pre-feasibility mine plan with soil resource information is to be

completed. A completed reclamation and closure planning report, including theestimated closure costing, will be required for the permit.

19.7.5 R E C L A M A T I O N A N D C L O S U R E AC T I V I T I E S

O P E N P I T

At the end of mine life, the open pit will be allowed to f ill with water. The Patsy Creekdiversion will be decommissioned in Year 16 and will be redirected into the pit so as

to speed up pit filling. If it is found during decommissioning that the water from thePatsy Creek diversion is required to dilute the surplus flow from the tailing pond to

meet receiving water targets in Lime Creek, an altered strategy to fill the pit may be

required, including allowing the pit to fill naturally while the Patsy Creek diversionremains in place until the pit spills naturally.

Currently, the pit is expected to fill with water to the low point of the pit rim at 515 m

elevation where it would discharge towards Lime Creek through a spillwayconstructed at the low point of the pit. Flow would then discharge into the permanent

TMF lake.

More detailed hydrological and hydrogeological studies will be undertaken at thefeasibility-stage to refine the flooding model for the pit. A safety berm will be

constructed around the pit.

W A S T E RO C K A N D LO W-G R A D E OR E

Over the LOM, it estimated that 162 Mt of waste rock will be generated. Of this,

96 Mt will be used to build the tailing dam and the remaining 66 Mt will be placed inthe southern end of the TMF. At closure, 25 Mt of this waste rock will be relocated to

the Kitsault pit for submergence under the pit lake, with the remaining waste rock left

in place, partially submerged by the TMF lake. An engineered cover and growthmedium will be placed over the resloped waste rock in the TMF. The cover would

then be vegetated. During mining operations, every attempt will be made to salvagethe growth medium that was placed in 2006 during the reclamation of the previous

mine.

Once active mining from the pit has ceased, the LGO stockpile will be processed for

the remainder of Year 14 and a small portion of Year 15. The full tailing stream will

be discharged to the pit. Reclaim water will continue to be sourced from the TMF, as





the settling of tailing solids in the pit will likely not be sufficient to source reclaimwater from the pit supernatant.

TA I L I N G

The closure measures for the TMF will include maintaining a water cover or lake over the tailing in perpetuity. During operations, a permanent riprap-lined spillway will be

constructed at the south abutment of the tailing dam. This spillway has beendesigned to accommodate the PMF and, after closure, will have the capacity to

discharge the entire flow from both Patsy and Lime creeks.

The flood spillway elevation is set at 470 masl.

An engineered cover and growth medium will be applied to the downstream face of the tailing dam as early as Year 8, when the tailing dam reaches its ultimate

embankment height of 475 m elevation. The downstream face would then be

vegetated with local native species to develop a stable cover of vegetation. Theaccess road on the face will be maintained for drainage control and to provide

access to the seepage collection pond at the toe of the dam. The seepage collectionpond at the toe of the tailing dam will remain active post closure.

D I V E R S I O N A N D SP I L L W A Y S

During the LOM, water will be pumped from the southern end of the tailing pond intothe Lime Creek Diversion Tunnel and into penstock pipes leading to a hydroelectric

plant. Once the mine ceases operation, the water will be routed via the permanent

spillway at the tailing dam and through the lower portion of the penstock to the

hydroelectric plant. A new 300 m-long tunnel connecting the TMF to the end of theexisting tunnel will allow surplus flow from the TMF to pass into the hydroelectricscheme, generating power in perpetuity.

Upon cessation of tailing deposition in Year 15, the Lime Creek Diversion Tunnel will

be decommissioned by sealing off the east portal. The Lime Creek/Patsy CreekDiversion Dam (West) will be breached allowing flow from the pit lake and from Limecreek to discharge into the TMF Lake.

The Patsy Creek Diversion Dam (East), which will divert Patsy Creek through achannel located in a pit bench during operations, will be breached at closure allowing

the pit to flood to 515 m elevation. The channel in the pit bench will be left in place

but sealed at one end with compacted fill. The south diversion ditch above the pit willbe breached.

B U I L D I N G S A N D IN F R A S T R U C T U R E

Buildings and structures that comprise the mine site facilities (mill, camp, power station, administration, maintenance shop, laboratory, site roads, fuel storage, and





explosives storage) will be removed at closure. These facilities will be demolished.Salvageable items within the buildings will be removed from site and sold.

Hazardous wastes will be removed from site and disposed of in an approved facility.

The majority of non-hazardous, inert building materials will be disposed of in a site

landfill or placed in the bottom of the open pit and covered with a layer of NPAGwaste rock. Concrete footings will be broken up and disposed of in the open pit. Anymetal-contaminated soils will be removed and disposed of in the tailing pond.

Hydrocarbon-contaminated soils will be excavated and treated on-site in a land farm.

Once successfully treated, these soils will be used a growth medium for reclamation.

Following removal of the facilities and any associated contamination, the disturbed

areas will be re-graded, capped with topsoil where needed, and fertilized and seeded

with native species. Non-essential mine site roads will be scarified and seeded, withall stream crossings returned to their pre-mining condition. The site landfill will beclosed using best practice methods.

The camp and storage areas, including administrative buildings, will be dismantledand removed. The area will then be re-graded, growth medium placed over the area,and revegetated.

Similarly, the fuel storage area will be removed and decontaminated. The fuelstorage is located just east of the wharf facility. Contaminated soil will be relocatedand growth medium spread over the area in order to be vegetated.

Over in the area north of the pit, the conveyor system will be dismantled and, asmuch as possible, will be salvaged and removed. The truck shop and truck wash willalso be dismantled and salvaged and non-hazardous material will be put into the pit.

Contaminated soil will be removed to a different location and then the area will beconsolidated, covered, and revegetated.

Both the process building and the primary crusher truck dump will also be dismantled

or demolished and any non-hazardous material will be disposed of in the pit. The

area will be re-graded and receive growth medium before being revegetated.

The wharf facility which serves the camp and the mine site will likely be dismantledand removed.

R O A D S

Some of the roads on site will be maintained post closure for ongoing sitemaintenance and monitoring. This will include the main road to the Kitsault town site,the roads accessing the area near the penstock and dam face, and the roadway on

the dam face itself. Roads that can be reclaimed will have the culverts removed,stream crossings re-graded, and their surfaces scarified to encourage vegetationgrowth.





P I P E L I N E S

Tailing and recycle pipelines will be removed and disposed of in Kitsault pit. The

pipeline routes will be revegetated. The 75 mm-diameter fresh water supply line willbe left buried.

W A T E R TR E A T M E N T

As previously discussed, ARD is not likely to start in the LOM and it is expected thatthere will be a lag time of decades before management of PAG rock and ARD will

need to be considered. The seepage volume differential, compared to the annual

surplus from the TMF (greater than 18 Mm3 annually), is expected to be sufficient tonegate the need for immediate water treatment of the seepage water from the toe of the tailing dam after closure. However, an allowance to construct and operate a

future water treatment plant has been included in the post closure cost.

PO S T- C L O S U R E MO N I T O R I N G

The level of post-closure monitoring will be a function of the environmental

performance of the mine site. Monitoring requirements are expected to reduce over time as the potential impacts to the receiving environment decrease. However,

under the current scenario, there may be a need for effluent treatment sometime inthe future. This will necessitate a post-closure monitoring program.

Post-closure monitoring will likely consist of the following:

environmental effects monitoring including studies on water quality,sediment quality, benthos, and fish to assess effects on the aquatic

receiving environment

engineering inspections by qualified persons of the retention pond, open pit,

and all engineered structures including the effluent treatment plant andlandfill.

Water quality monitoring will be done on a regular basis by the on-site staff.

19.7.6 C L O S U R E C OS T E S T I M A T E S

SC O P E O F ES T I M A T E A N D D E F I N I T I O N S

Completion of Production is the point in time when the mine has ceased to use

and/or is unlikely to use the plant and Avanti has decided that demolition should

start. Based on the current LOM plan, the date for Completion of Production isDecember 31, 2025.





Closure refers to the two-year period from January 1, 2026 to December 31, 2028,during which all demolition, cover construction, and contaminated soil remediation

would be completed. However early remediation activities, such as reclaiming theTMF dam borrow areas and the dam face, are current scheduled for Year 8 or 2019.

Post-closure refers to the period after January 1, 2029. Activities in this period willinclude inspection and maintenance of dams, surface and ground water monitoring,and water management.

B A S I S O F ES T I M A T E – C LOSURE

Format of Est imate

The estimate of closure costs is compiled in a Microsoft Excel workbook. The

workbook shows estimates of direct and indirect costs for each of the followingcomponents:

demolition and miscellaneous

diversions and spillways

tailing area and waste areas

contaminated areas and borrow areas.

Direct Costs by I tem – Closure

Demolition

Demolition direct costs were estimated from the footprint of the buildings and

estimates of equipment and crew time by area.

Earthworks

Direct costs for regrading were estimated by SRK using a regrading productivityspreadsheet and all-in equipment unit rates. Direct costs for placement of growth

medium were estimated by SRK from:

volumes calculated from average depths, and surface areas fromtopographic plans

equipment fleets and productivities estimated from excavate-load-haul-

dump-spread-compact calculations

all-in equipment unit rates provided by a local contractor

seeding and fertilizing costs from SRK experience.





The excavate-load-haul-dump-spread-compact calculations follow standard methods,as used by earthworks contractors. The calculations make use of equipment

specifications obtained from manufacturer’s data (in this case, the Caterpillar Handbook).

Equipment all-in unit rates were obtained from the Blue Book. The all-in ratesinclude equipment, operator, maintenance, parts, insurance, home office overhead,and contractor profit.

Contaminated Soils

Direct costs for removal of contaminated soils were estimated by SRK based on:

removal of an average 0.5 m depth of contaminated soil over an assumed

area

soil removal equipment and productivities estimated from excavate-load-

haul-dump calculations, with hauling of the removed material to the TMF

all-in equipment unit rates from the Blue Book

seeding and fertilizing costs from SRK experience.

Indi rects and Closure

Project Management

Project management, administration, and environmental management costs for the

closure period include:

a project manager at site for 24 months including 3 months on either side of the 24-month closure project

two project engineers assisting the project manager during active

construction periods

final design and engineering is included in the indirect costs

project administration by one contract administrator and two clerks

1 environmental manager and 1 environmental technician for 24 months

analytical costs for the post-closure monitoring

vehicles and fuel

office maintenance and utilities.

Project QA/QC costs, including materials testing and surveys carried out byindependent professionals, are included in the engineering cost.





Contractor Profit and Site Office

An allowance is provided for contractor profit, office space, and work staff at siteincluding office utilities and maintenance.

Engineering and QA

Engineering and field QA costs were estimated as a percentage of total direct costs.

Insurance

General insurance is covered in the all-in equipment unit rates. Additional insurance

of 0.5% of total direct cost was added to cover the additional requirementsassociated with demolition and contaminated soils remediation.

Bonding

Bonding of 0.5% of direct cost is provided. This item is not to be confused with a

Reclamation Bond, which will likely be required at the outset of the project.

Living-out Allowances

Living-out allowances for specialists is been included in the Engineering and QApercentages.

Taxes

Provincial Sales Tax (PST) is estimated at 7% of the taxable direct costs. Equipmentrentals, fuel, labour and supplies for demolition activities are PST exempt. For

construction activities, only fuel is PST exempt.

Contingency

Contingency was added to the direct and/or indirect costs and is an integral part of the estimate. The contingency used is 20% and is intended to offset the risk of unforeseen or under-predicted circumstances or conditions, which experience shows

will likely result, in aggregate, in additional costs

B A S I S O F ES T I M A T E – PO S T- C L O S U R E

Format of Est imate

The post-closure cost estimates, created in an Excel workbook, are formatted as acash flow table, showing annual expenditures from 2026 to 2119. Post-closure costs

are provided in Appendix L.





Earthworks Inspect ion and Maintenance

Average annual costs for inspection and maintenance of covers, ponds, and damswere estimated at 1.5% of direct earthworks cost (less contaminated soils) for thefirst 5 years and 0.5% in every year thereafter. Actual maintenance expenditures are

expected to occur only periodically, on an as-required basis.

Water Treatment System Construct ion

Construction costs for a future water treatment plant were estimated by SRK from

experience on other projects.

Water Treatment Operat ing Cost

Water treatment operating and maintenance costs were estimated by SRK. Noindirect costs or contingency are included in the operating cost estimate.

The plant will be capable of handling higher flows and more contaminated water

without significant cost increases.

Post-Closure Envi ronmental Management

Environmental management costs during the post-closure period include:

one environmental technician and one vehicle for six months of each year

analytical costs.

N E T PR E S E N T VA L U E A N A L Y S I S

Appendix L shows the closure and post-closure costs as annual cash flows for theperiod 2026 to 2057. The closure costs are incurred in 2026, 2027, and 2028. Then

there is a long period from 2029 to 2054 where there are only operating andmaintenance costs. In 2055, the capital costs for constructing the water treatment

plant are incurred, and the water treatment operating and maintenance costs start in

2056. Although subsequent years are not shown in Appendix L, operating andmaintenance costs are assumed to continue indefinitely.

In preparing cost estimates for activities that can take place many years in the future,

it is important to take into account the effects of interest. The conventional way to dothat is to use a method known as “discounting”. The most common form of discounting is the calculation of NPV. In simple terms, the NPV calculation shows

how much money would be needed to set aside today in order to have enoughmoney to carry out future activities.

To complete discounting calculations for the cash flows shown in Appendix L, a

discount rate of 3% was assumed. That rate is commonly used in discounting





closure cost estimates for the purposes of setting closure securities. The costs werediscounted back to 2026, the last year of operations.

All estimating was done in 2009 Canadian dollars, and the discounted values shown

in Appendix L are therefore also in 2009 Canadian dollars.

1 9 . 8 O N - S I T E I N F R A S T R U C T U R E

19.8.1 S I TE L A Y O U T

The overall mine and plant layout is shown in Figure 19.1. The location of theindividual facilities will take advantage of the challenging natural topography and, to

the extent possible, minimize negative impacts on the environment.

The Kitsault Project infrastructure will consist of:

a port facility

a camp/accommodation complex

an access road from the port to the mine site

a processing plant

maintenance and administration facilities

the TMF

various ancillary buildings

diversion ditches water diversion tunnel

pipelines for water and tailing

a hydro electric scheme

a revised transmission line alignment.

The plant site is located approximately 2 km northeast of the Ultimate Pit. Ore from

the pit will be trucked up a haul road to the primary crusher, which is located north-west of the Ultimate Pit. A conveyor material handling system and services will be

constructed to deliver the mill feed for processing from the primary crusher. The

conveyor system will extend from the west side of the primary crusher and runapproximately parallel to the haul and access roads. Tailing pipelines containing the

tailing and return water pipelines will be constructed to deliver the tailing to the TMF.

The mine areas will be accessed via the existing road connection from Cranberry

Junction and the town of Kitsault. This road will be utilized to transport equipment to

the mine site during the initial stages of the project. Avalanche sheds are notexpected to be constructed along access routes local to the site.





The site can also be accessed by float plane, barge, and helicopter.

The Patsy Dump rock landforms are located adjacent to the northeast side of the

Ultimate Pit. A low grade stockpile will be located northwest of the process plant

facility.

The process plant will consist of:

an ore crushing/grinding and material handling facility

the main process facility containing secondary grinding, flotation, regrinding,leaching, and concentrate dewatering

the TMF, located in Lime Creek Valley, approximately 4 km east of the

process plant facility; the TMF catchment will collect a sufficient annual

surplus of water during operation for plant water supply operations and willstore a one month plant water supply.

The overall plant layout (Drawing No. A00-10-003, Appendix M) is compact and, tothe degree possible, takes advantage of the natural ground contours. A significantvolume of cut-to-fill material will be required to ensure a solid foundation for plant

structures and related equipment. An MSE wall will be required along the haul roadfrom the pit to the primary crusher.

It is assumed that heavy equipment will require conventional foundations; heavy

equipment will be located on bedrock with spread footings and raft foundations over the remainder of the area. Concrete slabs will either rest on piles installed to

bedrock. Existing in-situ material will be removed to bedrock and replaced with non-

frost susceptible engineered backfill. The capital cost estimate includes a provision

for excavation and fill to an average depth of 1.5 m below grade.

There are no observed major avalanche run-out hazards in the plant site plateau

area. Plant domestic and process water supply will be provided from water diversions constructed around the perimeter of the tailing dam.

The plant site will be terraced with plant site roads to establish the construction grade

that has been assumed based on the Ultimate Pit and primary crusher elevation. Allterracing was based on nominal geotechnical information provided for the plant area.

Construction laydown areas have also been established; these areas will also be

cleared and levelled.

19.8.2 L I ME C R E E K D I V E R S I O N T U N N E L

The Lime Creek Diversion Tunnel is approximately 5 m wide by 5 m high by 3 km

long, and runs through a rock landform between the Lime Creek and Roundy Creekvalleys. The diversion tunnel will link the Lime Creek and Patsy Creek valleys and

divert flows from Patsy Creek and the surrounding catchments upstream of Lime





Creek into the lower Lime Creek Valley. The tunnel will be constructed using drilland blast techniques and driven from two headings.

The tunnel will slope downhill and emerge at the south wall of Lime Creek Valley.

Flows will then be channelled by a weir into the penstock pipe to a micro-hydro plant,

with an adjacent spillway to route flood flows around the hydro plant that exceed thecapacity of the penstock.

19.8.3 P L A N T C O N V E Y O R S

All conveyors were designed in accordance with latest version of Conveyor

Equipment Manufacturer’s Association (CEMA) manual. Mineralized materialcrushing will be carried out using a gyratory crusher located northeast of the pit ramp

exit. The crusher will discharge unto an apron feeder and then unto the inclined

crushed ore stockpile feed transfer conveyor. This conveyor discharges unto thecrushed ore stockpile feed conveyor with a capacity of up to 40,000 t/d to

accommodate surges. The crushed coarse materials will be reclaimed from the40,000-t live capacity stockpile by reclaim apron feeders onto the SAG mill feed

conveyor. Conveyors are also employed in the pebble circuit for the SAG mill

19.8.4 S I TE R O A D S

The main ring road around the plant is estimated is estimated to be 675 m long and10 m wide. Site roads will be supported on crushed rock, excavation, and fill to an

average depth of 1.5 m below grade, which has been used for estimating purposes.

19.8.5 D I V E R S I O N T U N N E L AC C E S S R OADS

Access roads will be constructed during the initial phases of construction from the

plant side to the diversion tunnel head end and the northern tailing diversion tunnelhead end in order to facilitate the driving of a north and south head of the Lime Creek

Diversion Tunnel.

19.8.6 P R O C E S S P L A N T

The process plant facilities will consist of a primary crushing station located near thepit. There will be an open 40,000 t live storage stockpile facility, a pebble crushing

section, and a SAG and regrind mill section. The process facility will also include

secondary grinding, flotation section, reagent preparation area, filtration area, aconcentrate handling area, and a tailing handling area. The grinding and flotation

building will be a stick-built structural steel building complete with and overheadcrane, electrical rooms, heating, ventilating and air conditioning (HVAC), and offices.





19.8.7 AN C I L L A R Y BU I L D I N G S

Ancillary buildings will be pre-engineered structures or stick-built structures as

required. Drawings of these facilities are included in Appendix M. The plant sitedesign includes the following buildings:

fuel storage facility

fuel station

administration building

assay and metallurgical laboratory

warehouse and maintenance building

400-person modular construction camp (set up at the port site)

truck shop

fuel station diesel fuel storage and dispensing

sewage treatment plant.

19.8.8 AS S A Y O F F I C E

An assay laboratory will be located in a separate building at the south side of the mill

building. The laboratory will be a 755 m2 single-storey pre-engineered structure, andwill be equipped to perform daily analyses of mine and process samples.

19.8.9 C O N C E N T R A T E ST O R A G E

The on-site concentrate storage facility will accommodate a 5-day storage capacityequal to 70 t/d of concentrate.

Concentrate will be loaded into 2-t super-sacks at the project site, loaded intocontainers, and the hauled and loaded onto ocean-going transport at the new portfacility.

1 9. 8. 10 W A R E HO US E /T R U C K SHO P /M I N E D R Y

The warehouse/truck shop/mine dry building will be a pre-engineered building,approximately 100 m long by 22 m wide. The building will be designed to provide

facilities for maintenance and repair, warehouse storage, minor office space, cleanand dry areas, and general storage. It will be located north of the Ultimate Pit,

adjacent to the existing haul road.

The truck shop/mine dry will house two maintenance bays, one light vehicle repair bays, a welding and machine shop, an electrical and instrument shop, a storage





warehouse with an upper level mezzanine area, and a dry area including lockers,offices, restrooms, first aid, and emergency vehicle storage. Waste oil will be

disposed of in the refuse incinerator with any remaining oil removed and discarded atan approved facility.

1 9. 8. 11 T R U C K W A S H /T I R E C H A N G E BU I L D I N G

The truck wash and tire change building is approximately 28 m long by 24 m wide.The building contains a wash bay, maintenance bay, tool crib, compressor room, hot

water pressure system, tire change, and an oil separator. Waste oil will be disposed

of in the refuse incinerator with any remaining oil removed and discarded at anapproved facility.

1 9. 8. 12 SE W A G E

The treatment plant will be a rotating biological contactor (RBC) system. The solid

and liquid material will be separated in the treatment plant with the liquid streamdischarging to the tailing pond and the solid material pumped out and trucked away

twice per year by a specialized licensed contractor. The treatment plant will beconstructed in modules with all modules used for the construction camp. Modules

will be removed after construction so that the remaining system is optimized toservice the operations facilities.

1 9. 8. 13 C O M M U N I C A T I O N S SY S T E M

A fibre optic cable has been included to the mill and mine site to providecommunications. Telecommunications to the plant site has been allowed for by an

allowance in the capital cost estimate.

Radio transceivers will be used for remote monitoring and control. A fibre opticbackbone will be installed throughout the plant site to facilitate the control systems

communication. A UHF radio system will be used for mobile communication.

1 9. 8. 14 P O T A B L E W A T E R SU P P L Y

Water storage reservoirs for the process plant can be combined with diversion intakestructures. These will be designed as project water requirements become better

established. During the winter months, well water from a field of wells near the plant

site may be needed to supply fresh water for process make up and domestic use atthe plant and camp facility.

Because of the port and plant sites, two potable water sources are required.





1 9. 8. 15 E X P LO S I V E S ST O R A G E A N D H AN D LI NG

For information on explosives storage and handling, refer to Section 19.1.7.

1 9. 8. 16 G E O T E C H N I C A L C O N D I T I O N S

The allowable ground pressure was assumed to be a minimum of 400 kPa and600 kPa for the crushing plant and coarse ore tunnel. The equipment foundation

may require a higher value, to be confirmed by selected vendors and a geotechnicalengineer.

Additional surface geotechnical investigation will be required to confirm the

foundation ground conditions for the plant and building foundations for the nextphase of the project.

1 9 . 9 O F F - S I T E I N F R A S T R U C T U R E

The off-site infrastructure relates to the new port facility. The port facility will be

comprised of the following:

an access causeway

barge loading ramp with counterweighted ramp lifting towers and stopfenders

a floating berthing lead

floating mooring buoys

fuel line to offload fuel barges

warehouse facility.

The largest components to be transported across the loading ramp at the port

facilities are expected to be delivered during construction and start-up of the mineoperations. Despite this short delivery period, the loading ramp lifting towers have

been permanently positioned to accommodate these large components.

Alternatively, the foundations for the lifting towers could be positioned closer to thesides of the ramp to reduce the size and cost of the ramp lifting beam. In thisscenario, the ramp would initially be raised and lowered with a mobile crane and the

tower superstructures would be installed after the largest components have been

moved across the ramp.

The onshore facilities will require site preparation for the fuel storage, propane

storage, warehouse storage structure and the accommodation complex. Access

roads will also be constructed, and the existing road and bridge over Lime Creek willbe upgraded. A layout of the proposed port facility is shown in Figure 19.37.





Figure 19.37 Port Facility Layout





1 9 . 1 0 P O W E R S U P P L Y A N D D I S T R I B U T I O N

Electrical power will be supplied to the mill via an existing 138 kV overhead

transmission line from the BC Hydro Aiyansh substation to the mine site, which is a

distance of approximately 42 km. A BC Hydro first-stage investigation confirmed thatthe system and the Aiyansh substation have adequate capacity for the Kitsault

Project.

The existing transmission line from Aiyansh substation to the mine site was originally

built to accommodate a 138 kV level. In recent years, the transmission line has been

operated by the utility at 25 kV level to match the limited electrical requirements of the Kitsault town site. For the purposes of this study, allowances have been made inthe capital cost estimate to extend the 138 kV power line to the new mine substation

location. A System Impact Study by BC Hydro will be required to determine the cost

to refurbish the transmission line from the Aiyansh substation to the mine site tap

location; no allowance has been included for this eventuality.

At the mine site, a new project substation will reduce the voltage to the mine

distribution voltage of 13.8 kV.

A small separate substation has been included to provide continued electrical power

at 25 kV to the existing Kitsault town site via a 25 kV pole line.

1 9. 10 .1 M A I N S I TE SU B S T A T I O N

The plant electrical design will include a step-down substation consisting of an

outdoor structure, incoming isolation switches and circuit breakers, and two identical

30/40 MVA three-phase outdoor power transformers complete with on-load tapchanging capabilities. Utility metering will be installed at this location. Each of the

main transformers will have sufficient capacity to carry the plant load at its forced

cooling rating. Secondary cable connections will transfer power to the main 13.8 kVswitchgear line-up, located indoors in an adjacent electrical room within the

substation area. This switchgear will also distribute power to the various plant loads.The single line diagram is presented in Appendix M.

1 9. 10 .2 P O W E R D I S T R I B U T I O N

The electrical distribution design is based on the site layout; the items entered in the

project mechanical equipment list, which includes process and mechanical designdata including required equipment motor sizes (kW), is based on information taken

from proposed equipment proposals and estimates of electrical power requirementsfor non-process based loads. Where available, information provided by others that

affects the overall power requirements for the project have been included.





The following factors have been applied to the basic kilowatt requirements in aconsolidated equipment database format:

demand factors

use factors typical reactive power ratings (relating load kW to KVA)

typical motor efficiency factors.

The project equipment list, in load-list format, was used to assist with single-line

design and planning; the equipment list is included in Appendix D. It should be notedthat the energy amount included in operating cost calculations includes anapproximate 5 % additional allowance as compared to the load-list summary (i.e.

36 MW vs. 34 MW demand).

Generally, motors greater than 200 hp will be connected at 4.16 kV. The ball mill

and SAG mill system will be connected at 13.8 KV (or at a level to best suit theequipment). Three-phase motors (up to 200 hp) will be connected for operation at

575 V.

Motors up to 200 hp are designated for full-voltage starting. Large motors at the4.16 kV level, unless otherwise indicated, are currently intended to be full-voltage

start. There are no provisions for soft starts unless otherwise noted on the

equipment list and/or single line diagram. Similarly, variable-speed drives will beprovided only for equipment so noted on the equipment list and/or the single linediagram.

Power from the main substation location will be delivered at 13.8 kV via a

combination of overhead lines and underground power cables. The main feeders willbe:

overhead power line to pit mining area

overhead power line to tailing pumping and barge areas

overhead power line and cable connection to primary crushing area

cable feeder to crushed ore stockpile area

cable feeders to serve main process area loads

cable feeders to the SAG mill (12 MW connected) and associated harmonic

filter

cable feeders to the ball mill drive systems (11 MW connected) andassociated harmonic filter.

For ease of maintenance, most electrical equipment will be located in indoor

ventilated electrical rooms wherever practical. Where the electrical rooms are notdesigned into the footprint of the main building, allowances for pre-manufactured

buildings (equipment installed and pre-wired) have been included.





The HPGR facility will have a separate electrical room housing its electricaldistribution as part of the building. The final size of this room will be determined by

the final equipment selection.

EM E R G E N C Y PO W E R

Allowances have been made for:

a 500 kVA diesel generator at the process plant location in order to provide

back-up power for critical load equipment; this process plant dieselgenerator will be interlocked with the plant 600 V distribution system to

prevent inadvertent paralleling of sources of energy

a 3,200 kVA diesel generator at the mining area in order to provide back-uppower for critical load equipment.

EQ U I P M E N T T Y P E AN D RA T I N G S

Wiring Mater ials and Instal lat ion

Power and Control Cables

Other than the 13.8 kV transmission lines, which are aluminum conductor steel

reinforced (ACSR) type, electrical power will be distributed at the various voltage

levels via armoured cables in cable tray or, in some instances, via buried ductsystems.

Generally, all power cabling will use copper conductor Teck-type cables inunderground, non-metallic ducts, or above ground in ladder-type cable trays. Theywill have interlocking aluminum armour with a PVC FT4-rated, UV resistant jacket,suitable for -40°C. Voltage, insulation class colour, and spacing will be as needed for

the application.

Underground Cable Ducts

Where underground cabling is required, cables will be installed in minimum

100 mm ø Schedule 40 rigid PVC ducts. The ducts will be buried typically to1,200 mm below final grade. Concrete pull boxes will be provided where required for

number of bend limitations and sized to suit existing cable fill plus minimum 25%spare capacity. A “Danger – Buried Cable” marker tape will be installed at both sides

of trench during backfilling. A 50 mm-thick concrete layer may be added above the

ducts in selected areas where heavy vehicle traffic is expected.





L I G H T I N G

Illumination levels within the plant facility shall meet the requirements of all relevant

codes and standards. The minimum illumination levels will be as follows:

electrical room: 300 lux

control room: 500 lux

offices: 600 lux

loading: 30 lux

yard: 10 lux.

Lighting for process area and exterior will be via 347 V, high efficiency HID typeindustrial fixtures.

Battery-powered emergency lights and exit signs allowances are included to

illuminate stairways and other interior egress routes within major buildings.

1 9 . 1 1 M A I N A C C E S S R O A D S

There are several road access routes to the mine and plant site, including an accessroad from the new camp and port facility to the mine site.

The site can be accessed from Smithers by travelling 110 km northwest on

Highway 16 to the Cassiar Highway Junction, and then north 76 km on Highway 37towards the Cranberry Junction. The westbound Nass Forest Service Road (FSR)

runs 32 km to the Kinsqyish FSR. This road winds 14 km past the Nass Bridge, theKinkush-Alice Arm Pass Road junction, several concrete bridges, and thedecommissioned transformer station next to the mine access road.

The site can also be accessed starting at Terrace and travelling 100 km north to

Nass Camp and then a further 95 km on an upgraded gravel road to site. Analternative access road route is from Terrace via Highway 37 to Cranberry Junction(175 km) then via a gravel connector road from Cranberry Junction to site

(approximately 106 km).

The re-routed access road from the existing town site of Kitsault heading easttowards the Clary Lake existing road is approximately 7 km in length. The re-routing

of the existing Kitsault town site access road is necessary as the TMF will inundate amajor section of the existing road within the first years of operation.

1 9. 11 .1 P L A N T R O A D D E S I G N R E Q U I R E M E N T S

The Kitsault project access roads are classified as resource development roads. The

design criteria proposed for the permanent access would be an 8-m wide gravel





surface, radio-controlled road capable of bearing the legal axle loading for trucks onBC highways on a year-round basis. The road will be used to provide vehicle access

for development of the mine site as well as year-round access for supplies,equipment, and crew transport.

Alignment controls, such as maximum 10% sustained adverse grades and 50 mminimum radius horizontal curves, are recommended to allow transport of the largestpieces of equipment to the mine site.

1 9. 11 .2 P O R T S I TE R O A D A N D E A R T H WO RK S

The proposed location for the port and accommodation complex/town site facilities isapproximately 3 km south of the Kitsault town site, in a flat delta-like area.

Earthmoving in this area is expected to pose no difficulties. The existing bridge on

this route will require upgrading. A geotechnical consultant has identified parts of this area as a possible borrow source.

The re-routed access road from the existing town site of Kitsault heading east

towards the Clary Lake existing road is approximately 7 km in length. The re-routingof the existing Kitsault town site access road is necessary as the tailing managementfacility will inundate a major section of the existing road within the first couple of

years of operation. The re-routed route starts at the edge of the existing town site at

and will climb before descending to the existing access road to Clary Lake and theNass valley.

Parts of the route pass through steep, mountainous side slopes that will require high

initial construction costs and regular maintenance costs. These sections will be kept

to a minimum by using switchbacks where possible. When negotiating swamp

areas, the route follows rock outcrops so as to minimize the import of futureconstruction materials. A gravel source is not readily available and crushing of road

construction surfacing/base materials will be required.

1 9. 11 .3 E X P LO S I V E S ST O R A G E F A C I LI TY E A R T H WO RK S /AC C E S S R O A D

The proposed explosives storage facility (ESF) access road will leave the plant site

access road approximately 1.5 km from the plant site. The road will beapproximately 0.6 km long and with a grade that will not exceed -2%.

1 9. 11 .4 O T H E R C I V I L W O R K S

Other civil works will include:

a haul road (all haul roads will be 30 m wide) from pit exit elevation of 570 mto the proposed primary crusher location

earthworks for the primary crusher and crusher product conveyor including

excavation/fill and MSE retaining walls





conveyor cut through from crusher conveyor transfer point to the plant site

improvement of existing haul road from pit to proposed plant site area.

1 9. 11 .5 E A R T H W OR K S I D E SL O P E S

The following criteria where used to layout and calculate cut and fill embankments,plant access roads, and haul roads:

Rock Cut: 1 Horizontal; 4 Vertical, with 6.5 m wide benches every 10 m of

vertical excavation

Earth and other materials:

cut: 1.5 Horizontal; 1 Vertical

fill: 1.5 Horizontal; 1 Vertical.

1 9 . 1 2 L O G I S T I C S

A preliminary logistics study was performed to determine the preferred method of

transporting for the PFS and is included in Appendix N.

The proposed location for the port and accommodation complex/town site facilities isapproximately 1.6 km south of the existing Kitsault town site (from an existing bridge

crossing) and is in a gently sloping area. Earthmoving in this area should berelatively easy and no problems are anticipated. The existing bridge on this route will

require upgrading.

The port facility allows for the barging of equipment and supplies to the mine site, as

well as the ocean transport of concentrate from the mine site. The port is expectedto handle concentrate throughput of up to two hundred 20 ft containers every two

weeks.

The existing highways/roads leading to the project area will require some bridge and

crossing upgrades. These upgrades will be identified and evaluated in more detail

during the next phase of the project study.

Equipment and supplies can be shipped via barge (probably from Prince Rupert,

140 km from the facility) to a new port facility near Kitsault, and then by the upgraded

and re-routed access road to the plant and mine site.

For the purposes of this study, Wardrop assumed that larger construction equipment,

supplies, and concentrates will be transported by truck to and from the new port

facility. A 60-ha Crown-owned site about 3 km from the site is available for lease.The existing road from the mine site to the port will require upgrading. The portfacilities and the construction/permanent accommodation complex facility will be

located in this area.





The mine port facility will ship mine concentrates and receive mine material andsupplies, including fuel. The largest components that will be transported across the

loading ramp at the port facilities are expected to be delivered during constructionand start-up of the mine operations. Despite this small delivery window, the lifting

towers have been permanently positioned to accommodate these large components.

Alternatively, the foundations for the lifting towers could be positioned closer to thesides of the ramp to reduce the size and cost of the ramp lifting beam. Under thisscenario, the ramp would initially be raised and lowered with a mobile crane and the

tower superstructures would be installed after the largest components have beenmoved across the ramp.

The currently-proposed location of the container storage yard is not directly adjacent

to the loading ramp, because of challenging terrain at the site. There are severalalternatives for unloading empty containers from a barge and loading full containers

onto the same barge, including:

one reach stacker moving containers between the storage yard to the barge

two reach stackers moving containers between the storage yard to the barge

one reach stacker located at the storage yard and one reach stacker located

at the loading ramp with one or more tractor units or bomb carts shuttling

containers between the storage yard and the loading ramp.

There are three primary alternatives for transporting fuel to the site for mobileequipment and liquid chemical reagents to support mine process facilities, including:

1. Both fuel and reagents will be transported in B-train liquids tanker trucks on

a flat deck barge.

2. Fuel will be transported within a fuel barge with reagents carried by liquidstanker trucks either on top of the fuel barge or the barge carrying containers.

3. Both fuel and reagents will be carried in separate compartments of the samefuel barge.

While Options 2 or 3 would reduce the number barge trips to the site from Prince

Rupert, the largest fuel barge operator in the area indicated that some of thereagents could not be carried within fuel barges. Therefore, Option 2 is

recommended.

The fuel will be pumped from the fuel barge through a dedicated 75 mm diameter

fuel line located on the access trestle and loading ramp to an upland storage tank.The costs associated with the fuel line and storage tank are not included in thisstudy.





1 9 . 1 3 C A P I T A L C O S T E S T I M A T E

An initial capital of US$641 M (Q4 2009) will be required for the project, based on

capital cost estimates developed by the following consultants:

Wardrop: mine capital costs, mine rock dump costs, process plant and

associated infrastructure costs, power supply costs, and access road costs

KP: tailing management, haul road, construction access roads for TMF and

tunnel diversion, fresh water supply, and water management costs


Rescan: associated environmental, bonding and permitting costs

WPW: marine port facilities costs.

All currencies in this section are expressed in US dollars. Costs in this report have

been converted using a fixed currency exchange rate of Cdn$1.00 to US$0.92. Theexchange rate is based on Q4 2009 Bank of Montreal rates, with the consent of

Avanti. The expected accuracy range of the capital cost estimate is ±30%.

“Initial capital” has been designated as all capital expenditures required to produce

concentrate. Some capital costs have been relocated to operations and sustaining

capital as directed by Avanti.

A summary of the major capital costs is shown in Table 19.28. The breakdown of the

capital cost estimate is included in Appendix O.

1 9. 13 .1 E X C LU S I O N S

The following items are excluded from the capital cost estimate:

force majeure

Federal and Provincial taxes

overtime

cost outside battery limits

interest during construction

sunk costs

provision for escalation beyond Q4 2009

risk analysis.





Table 19.28 Capital Cost Summary

Description US$000

Direct Works

Overall Site 37,300

Mining 51,000

Crushing 24,000

Crushed Ore Storage and Reclaim 12,900

Process 106,200

Tailing Managem ent Facility 122,200

Water Management 33,200

Site Services and Site Utilities (Mine Site) 4,500

Ancillary Buildings (Mine Site) 16,700

Plant Mobile Fleet 2,400

Temporary Services (Port Site) 13,300

Port Site Facilities 8,700

Subtotal 432,500

Indirects

Project Indirects 104,000

Owner's Costs 26,000

Contingencies 78,000

Subtotal 208,100

Total Capital Cost 641,000

1 9. 13 .2 M I N E C A P I TA L C O S T S

Mine capital costs are derived from a combination of supplier quotes and historical

data collected by Wardrop. This includes the labour, maintenance, major componentrepairs, fuel, and consumables costs.

The mine equipment capital costs include basic equipment capital, assembly, and

commissioning. Costs for delivery to site (excluding Federal and Provincial taxes or duties) are included under project indirects. Mine capital costs are shown in Table19.29.





Table 19.29 Mine Capital Costs

Cdn$ (000) US$(000)

Mobile and Support Equipment 52,023 47,862

Explosives Storage 136 125

Fuel Storage and Delivery - -

Dewatering (Pumping only) 775 713

Electrical 1,522 1400

Communication 598 550

Safety 80 74

Engineering Equipment 271 247

Other Mining Costs -

TOTAL MINE CAPITAL 55,405 51,000

Note: pre-production costs of US$13,829 (Cdn$15,031) have been included in the operating costs.

M I N I N G B A S I S O F ES T I M A T E

Unit costs for consumable and labour rates are estimated from sources listed below

while the magnitude of consumables and labour required are determined for each

specific activity from experience and first principles.

Currencies are expressed in US dollars. All costs in this section were calculated in

Q4 2009 Canadian dollars. A conversion to US dollars was implemented at an

exchange rate of 0.92. No allowance is included for escalation beyond this quarter.

The unit costs are based on the following data:

Salaries for the supervisory and administrative job category are based onWardrop’s experience of similar functions in BC mines. An average burdenrate of 39% has been applied to base salaries to include all statutory

Canadian and BC, social insurance, medical and insurance costs, pension,

and vacation costs.

For hourly employees, general labour rates expected in BC mines and

proposed projects were used. An average burden rate of 46% has beenapplied to base wages to include all statutory Canadian and BC, socialinsurance, medical and insurance costs, pension, and vacation costs.

Mine designs to determine the size and makeup of the mine fleet as well as

fuel requirements which is affected by distance from the pit to the variousdestinations over the existing and future topography.

Budgetary quotations, including freight for all consumables, tires, and fuel as

well as assembly and commissioning. The long term fuel price is estimated

at a delivered cost to site of Cdn$1.05/L.

Power costs were estimated as the sum of energy charges, demand

charges and estimated at an overall Cdn$0.0399/kWh





Mining equipment consumables, major equipment replacements, sustaining

capital, labour loading factors, equipment life, and costs are based on a

combination of vendor information, InfoMine USA’s 2008 Mine and MillEquipment Costs and Wardrop’s data base from similar mining operations.

Wardrop’s estimate of start-up capital costs includes the following:

major mine equipment

services and infrastructure

pre-production tasks

support and auxiliary equipment.

Mine capital costs are shown to mining Year 1 in Table 19.30. This schedule shows

capital by year that the respective equipment is required to be working. Actual

expenditure will be required sufficiently before that to allow for manufacturing,

delivery, and erection.

Table 19.30 Mine Mobile and Support Equipment Capital

Fleet Capital Cost Task Description

PP*

US$(000)

Year 1

US$(000)

Drilling

Diesel Rotary Drill – 269 mm Primary Drill 2,087 2,087

Secondary Drill – 127 mm Secondary Drill 536 -

Blasting

Hole Stemmer – 3 t Blast Hole Stemmer - -

Loading

Major

Electric Hydraulic Shovel – 18 m3

Loading Mineralized Material & Waste 12,988 -

Front End Loader – 18 m3 Loading Mineralized Material & Waste 4,540 -

Support

Track Dozer – 432 kW Shovel/Dump Support 2,760 1,380

Rubber Tired Dozer – 250 kW Pit Clean Up - 1,058

Fuel/Lube Truck Shovel Fuelling & Lube 175 175

Hauling

Major

Haul Truck – 177 t Hauling Ore/Waste 17,256 17,256

Support

Water Truck – 20,000 gal Haul Roads Water Truck 138 -

Track Dozer – 306 kW Dump/Road/Stockpile Maintenance 897 897

Motor Grader – 221 kW Road Grading 874 874

Vibratory Compactor Road Construction/Maintenance 129 -

Snow Plow/Sanding Truck Road Maintenance 92 -

Tire Manipulator Tire Handling 436 -

table continues…





Fleet Capital Cost Task Description

PP*US$(000)

Year 1US$(000)

Pit Maintenance

Hydraulic Backhoe – 4 m Utility Excavator 805

Hydraulic Excavator – 2 m

3

Utility Excavator 598 -Tractor & Lowboy – 136 t Utility Tractor Transport In-pit 349 -

Hydraulic Crane – 68 t Utility Crane 549 -

Hydraulic Crane – 36 t Utility Crane 265 -

Flat bed Truck – 20 t Field Mechanical Maintenance 117 -

Mechanics / Welder Truck Field Mechanical Maintenance 101 101

Integrated Tool Carrier Field Mechanical Maintenance 253 -

Bus Crew Transportation - -

Crew Cab Pickup – 4x4 Supervision and Crew Transportation 46 -

Pickup Trucks – 4x4 / ¾ t Supervision and Crew Transportation 373 207

RT Forklift – Sellick SD-100 Field/Shop Maintenance 85 -

Shop Forklift – Hyster H100XM Shop Maintenance 49 -Light Plant/Towers In-pit/Dump Lighting 110 -

Mine Pumps In-pit Sump Dewatering - -

Assembly and Commissioning 1,255 -

TOTAL CAPITAL COST 47,862 24,035

* PP = pre-production.

EL E C T R I C A L POWER D I S T R I B U T I O N C OSTS

Table 19.31 o capital costs for the open pit electrical power distribution system.

Table 19.31 Power Supply Construction Cost Summary

Description US$000

Overhead 13.8 kV to Pit Distribution Switchgear 405

13.8 kV Switchgear SG3 (Top of Mine) 40

5 MVA, 13.8/78.2 kV Transformer 199

7.2 kV Distribution Panel 74

13.8/4.16 kV Transformer (Initial 1 MVA) 53

4.16 kV MCC4 MCC5 65

13.8/4.16 kV Transformer (Initial 1 MVA) 53

600 V MCC6 MCC7 with Electrical Room 84

LV & MV Motor Wiring Allowance (trailing cables, etc.) 89

Area Lighting, Panels and Cable Tray Allowance 53

Grounding Allowance 35

8 kV, 3C-4/0 Shd GC Mining Cable to Shovel #1 105

8 kV, 3C-4/0 Shd GC Mining Cable to Shovel #2 105

Mining Cable Couplers 40

Total Construction Cost 1,400





For this PFS, it has been assumed that the power distribution to the open pit minewill be provided by BC Hydro under their current tariffs from the nearest transmission

interconnection point. The transmission line and associated facilities from that pointon to the mine will be the responsibility of Avanti as per BC Hydro policy.

The single line diagram for the study is provided in Appendix M.

1 9. 13 .3 P R O C E S S C A P I T A L C OS T

B A S I S O F ES T I M A T E

The estimating guidelines for the preparation of the capital cost estimate are

described in detail in the Basis of Estimate (Appendix O). The estimate has been

produced in Microsoft Excel 2007. The estimate breakdown structure is user-definedby area and commodity code. The contingency has been determined by historical

data from similar projects.

The project site condition specification is provided in Appendix P.

Currencies are expressed in US dollars. All costs in this section were calculated in

Q4 2009 Canadian dollars. A conversion to US dollars was implemented at an

exchange rate of 0.92. No allowance is included for escalation beyond this quarter.

The capital cost estimate for the process plant has been completed by Wardrop andis based on the information shown in Table 19.32.

Table 19.32 Process Capital Cost Basis of Estimate

Commodity Estimate Basis

Plant and Equipment

Major Equipment (>$1,000,000) Single budget price quotations based onspecifications & data sheets

Major Equipment (>$500,000) Telephone and e-mail budget price quotations

based on duty specifications

Minor Equipment (<$500,000) In-house database and/or factored equipment costs

from similar projects

Bulk Materials & Site Works

Site Preparation & Roads Estimated on a cost/unit area based on a

preliminary earthworks volumes calculated from a3D model (LAN desktop)

Concrete – Building Foundations Estimated on a cost/unit area based on historical

data for similar buildingsConcrete – Equipment Foundations

Structural – Equipment Supports

Structural – Building Steel

Architectural (incl. Ancillary Buildings)

Building Services

table continues…





Commodity Estimate Basis

Service Piping & Valves Percentages of direct equipment costs, by area,

based on the study equipment list and historical datafrom similar projects

Process Piping & Valves

Electrical

Instrumentation & ControlsInstallation

Installation Labour Hours calculated or based on historical data and in-

house experience

Productivity 1.15 productivity factor has been assumed for the

estimate

Vendor Representatives/ Supervision An allowance based on complexity

Contractor Distributables/ Preambles Included in the unit labour rate

Freight

Main Bulks & Major Equipment An allowance based on specific equipment and

complexity. Freight costs to site have been includedin the material section. Unless specifically quoted,

freight has been factored as 8% on equipment andmaterials, and 6% on mobile equipment.

Air Freight (for equipment and personnel) Minimum allowance included plus helicopter supportfor initial ore slurry tunnel construction and mining

pioneering work

Commissioning

Commissioning Start up Assessments based on in-house data

Construction & Commissioning Spares Based on 1.5% of process equipment costs

Mining Spares Based on 7.5% of major mine equipment stock

The following backup documentation and information is available in Appendix D:

process design criteria

preliminary flowsheets

general arrangement drawings

plant mobile equipment list

equipment load list.

1 9. 13 .4 P E R M A N E N T AC C O M M O D A T I O N A N D C O N S T R U C T I O N C A MP S

One construction camp, located at the port facility, is allowed for in the estimate. Thecamp is estimated on a modular basis and will be expanded to accommodateincreasing labour force during construction.

The construction camp will be refurbished to provide a permanent camp for 250

people.





1 9. 13 .5 L A B O UR R A T E S

Different labour rates were applied to various areas of the project. In general, a

labour rate of US$70.43/h (Cdn$76.55/h) has been used.

1 9. 13 .6 T A X E S

Federal and Provincial taxes have been excluded.

1 9. 13 .7 L O G I S T I C S

No logistics study has been performed for this project.

1 9. 13 .8 O W N E R S ’ C O S T S ( I N C L U D I N G O W N E R S C O M M I S S I O N I N G AL L O W A N C E )

An allowance has been included for the Owners’ costs. This cost has been providedby the Owner. The cost of the land required for the port facility is included in these

costs.

1 9. 13 .9 E X C LU S I O N S

The following are not included in the capital cost estimate:

force majeure

schedule delays such as those caused by:

major scope changes

unidentified ground conditions

labour disputes

environmental permitting activities

abnormally adverse weather conditions

receipt of information beyond the control of the EPCM contractors

cost of financing (including interests incurred during construction)

PST and GST

royalties or permitting costs

schedule acceleration costs working capital

cost of this study

sunk costs.





19.13.10 AS S U M P T I O N S

The following assumptions have been made in the preparation of this estimate:

All material and installation subcontracts are competitively tendered on an

open shop, lump sum basis.

Site work is continuous and is not constrained by the Owner.

There is a 70-hour work week with a rotation of 3-weeks in/1 week out for

the construction phase of the project.

Skilled tradespersons, supervisors, and contractors are readily available.

The geotechnical nature of the destination site is expected to be sound,

uniform, and able to support the intended structures and activities. Adverse

or unusual geotechnical conditions requiring piles or soil densification havenot been allowed for in this estimate.

19.13.11 C O N T I N G E N C Y

A contingency allowance was built up to cover additional costs, which will be incurredas a result of more detailed design and investigations. It is considered that this

estimate will adequately cover minor changes to the current scope to be expected

during the next phase of the project.

Several major costs (and allowances) are assumed to contain a certain amount of

contingency; therefore, a lower contingency was applied across the board. The

average contingency for the project is calculated to be approximately 14%.

19.13.12 TMF A N D R E L A T E D W A TER M A N AG E ME N T SYSTEMS

G E N E R A L

This section summarizes the PFS capital (initial and sustaining) and operating costestimate for the TMF and related water management systems. This estimate was

prepared using a combination of quoted, estimated, and factored costs and has an

expected level of accuracy of ±30%.

The initial and sustaining capital costs, as well as operating costs, were calculated in

third quarter 2009 Canadian dollars. A conversion to US dollars was implemented at

an exchange rate of 0.92. All costs listed are in US dollars.

SU M M A R Y O F TMF A N D W ATER MA N A G E M E N T ES T I M A T E

The capital cost estimate for construction of the TMF and water managementsystems is approximately US$212 M for initial construction (inclusive of an overall17% contingency), with yearly sustaining capital costs ranging from approximately





US$11 M to US$50 M incurred annually over the first 8 years of mine operations.The annual operating expenditure was estimated to be approximately US$4 M over

the life of the mine.

The cost estimate, broken down into major subdivisions according to the Work

Breakdown Structure (WBS), is summarized in Appendix O. In addition to the WBSdivision, the estimated costs were allocated as either capital (initial or sustaining) or operating costs depending on the activity. Initial capital includes all costs prior to the

start of Year 1 in the mine plan, including construction of the TMF starter

embankment, tailing distribution system, reclaim water system, surplus water system,seepage collection system, fresh water systems and diversion infrastructure.Sustaining costs generally included those costs necessary to raise the TMF

embankment. The annual operating cost was estimated from power andmaintenance requirements for the tailing and water distribution systems.

Escalation and camp costs have also been excluded from this estimate. The cost

estimate for the TMF and water management system has been included in Appendix O.

ME T H O D O L O G Y

The cost estimate of the TMF and water management systems was broken down into

the following subdivisions:

tailing disposal and reclaim

TMF earthworks and foundation preparation

tailing/borrow roads

seepage collection and sediment control

surplus water pipeline

diversion systems

fresh water supply

indirects.

The cost estimate for major components was generally prepared from an estimate of quantities derived from the PFS design drawings multiplied by unit costs developed

using analogous or bottom-up estimating techniques.

The earthworks component of the TMF and water management system, including

roads and diversion systems, were prepared by estimating the size and production

rate of an appropriate equipment fleet. Assumptions regarding the location of thevarious construction materials such as borrows, quarries, and waste rock from theopen pit were incorporated in the earthworks estimates. In addition, similar

techniques were used to develop unit rates for construction of site roads required for the TMF and water management systems. Lump sum allowances, based on





historical experience for similar work, were used where there was insufficient detail inthe design to develop a more detailed estimate. All TMF earthworks and foundation

preparation, tailing/borrow roads, and diversion system costs were included as either initial or sustaining capital costs in the estimate. Sustaining capital generally

consisted of construction activities necessary to raise the TMF embankments.

The capital (initial and sustaining) cost estimates for the tailing disposal and reclaim,seepage and sediment control, surplus water pipeline, and fresh water supply

(collectively referred to as ‘pipeworks’) were generally estimated based on a mixture

of vendor quotes and historical experience for similar work. Percentage-based mark-ups for labour and equipment were applied to the material costs to cover installation.Operating costs for pipeworks included power and maintenance costs. Power was

estimated based on pump sizes and a unit rate of US$39.9/MWh. Annualmaintenance costs were estimated as a percentage of the capital cost for the various

components of the pipeworks.

Indirects for the cost estimate included construction indirects, engineering andprocurement, and construction management. These costs were estimated from past

experience with similar work using fixed percentages of the subtotal of the major

TMF and water management components (8%, 3%, and 2%, respectively).

19.13.13 P O R T F A C I LI TI E S

The basis of estimate for the marine facility capital costs have been developed based

on in-house data and experience with similar projects. Costs included are generallybased on the current cost of construction in northern BC and do not allow for

escalation beyond the base date of the estimate. The capital cost estimate is

summarized in Table 19.28. The basis of estimate is included in Appendix O.

1 9 . 1 4 O P E R A T I N G C O S T E S T I M A T E

The operating costs are based on cost estimates developed by the following

consultants:

Wardrop: mine and process plant operating costs

KP: tailing management facility and water management operating costs


Rescan: associated environmental, bonding and permitting costs

WPW: marine port facilities operating costs.

The operating cost for the Kitsault Project was estimated at US$1.74/t mined(Cdn$1.89/t mined) or US$3.05/t milled (Cdn$3.32/t milled). The estimate was

based on an average daily process rate of 40,000 t milled.





Currencies are expressed in both Canadian and US dollars. The costs in this sectionwere estimated in Q3 2009 dollars.

When it was required, certain costs in this report have been converted using a fixed

currency exchange rate of Cdn$1.00 to US$0.92. The expected accuracy range of

the operating cost estimate is +30%.

The operating costs are defined as the direct operating costs including mining,

processing, tailing storage, G&A and other costs. Sustaining capital includes allcapital expenditures after the process plant has been put into production.

Power will be supplied by grid lines at an average cost of Cdn$0.0399/kWh. The

power cost for the electric shovels is included in the mining operating cost.

1 9. 14 .1 M I N I N G O P E R A T I N G C O S T S

All mining operating costs are shown in Canadian dollars, unless otherwise specified.Mine operating costs are derived from a combination of sources. The operating costincludes the labour, maintenance, major component repairs, fuel, and consumables

costs.

Consumables such as tires, explosives, and drill accessories were solicited fromsuppliers. Maintenance, major component, and wear item repairs were sourced from

InfoMine’s Equipment Costs and from Wardrop’s recent project data. The current

fleet hourly operating costs are used as a constant basis over the schedule periodsand estimates are input for sustaining and replacement capital.

Blasting costs are based on studies from similar projects and historical blasting

costs. In this cost estimate, a total blasting scenario is assigned to the blastingcontractor. Geotechnical costs for high wall control blasting, horizontal drains, etc.are based on other study data collected by Wardrop.

Staff and hourly operating rates are based on current salary and wage levels in

similar mines operating in BC. A total benefit package was applied to the base rateconsisting of vacation, statutory holidays, medical and health insurance, employmentinsurance, long term disability insurance, Canada Pension Plan, and Worker’s

Compensation Board.

From the basic operating capacities of the equipment, the travel speedcharacteristics of the trucks and cycle times for the various haul profiles are

calculated from Caterpillar’s FPC Version 3.05 program. The haul road profiles werelaid out in MineSight® and then entered into the FPC program. The haul truck speedis determined by the rimpull curves of the 177 t truck but is constrained by maximum

operating speed criteria. The equipment productivity and operating hours arecalculated separately in a spreadsheet scheduling program based on the totalequipment cycle times derived from FPC. The calculated operating hours are





multiplied by the hourly labour, maintenance, major component repairs, fuel, andconsumable costs to arrive at the total operating costs.

Labour requirements have been determined for each labour category. In the case of

operators, labour requirements are estimated based on the amount of equipment on

duty. Maintenance labour is estimated based on historical ratio between equipmentoperators and maintenance mechanics and electricians. All other labour and staff are estimated from experience with existing mines and anticipated operating

conditions for the project.

Labour costs are calculated by multiplying the annual wage by the total employeerequirement for each labour category. The hourly labour and mine salaried labour

requirements are summarized in Table 19.33 and Table 19.34, respectively.

Detailed information on labour requirements is provided in Appendix Q.

Table 19.33 Mine Hourly Labour Requirements

Hourly Labour Summary Year 5

Mine Operations

Shovel Operator 10

Loader Operator 5

Haul Truck Operator 62

Drill Driver 10

Dozer Operator 24

Grader Operator 10

Water/Sand Truck Operator -

Blaster -

Blaster Helper -

Equipment Trainee 4

Mine Labourer 8

Mine Maintenance

Mechanic JM - HD 8

Mechanic - HD 28

Mechanic JM - LD 4

Mechanic - LD 2

Electrician 10

Serviceman 8

Welder 8

TOTAL HOURLY 201





Table 19.34 Mine Salaried Labour Requirements

Salaried Labour Summary Year 5

Mine Operations

Mine Superintendent 1

General Mine Foreman 1

Drill & Blast Foreman 1

Mine Foreman 4

Training Coordinator 2

Mine Clerk 1

Maintenance Superintendent 1

Maintenance General Foreman 1

Maintenance Foreman 4

Electrical Foreman 2

Maintenance Planner 1

Maintenance Clerk 1

Technical Services

Chief Mine Engineer 1

Senior Mine Engineer 2

Dril ling / Blasting Engineer 1

Senior Surveyor 1

Survey Technician 1

Senior Geologist 1

Senior Geotechnical Engineer 1

Geologist – Grade Control 2

Total Salaried 30

Labour rates are based on hourly rates for mine operations and maintenancepersonnel in the area (Table 19.35), and current salaries for G&A employees (Table

19.36).

Table 19.35 Mine Operating and Maintenance Hourly Labour Rates

PositionBase Rate

(US$/h)Payroll Burden

(US$/h)Total Pay

(US$/h)

Mine Operations

Shovel Operator 29 13 42

Loader Operator 27 13 40

Haul Truck Operator 26 12 38

Drill Operator 27 13 40

Dozer Operator 27 12 39

Grader Operator 27 12 39

table continues…





Position

Base Rate(US$/h)

Payroll Burden(US$/h)

Total Pay(US$/h)

Water/Sand Truck Operator 26 12 38

Blaster 28 13 41

Blaster Helper 24 11 35Equipment Trainee 25 12 37

Mine Labourer 21 10 31

Mine Maintenance

Mechanic (Junior level) – Heavy Duty 29 14 43

Mechanic – Heavy Duty 27 13 40

Mechanic (Junior level) – Light Duty 29 14 43

Machinist – Light Duty 27 13 40

Electrician 29 14 43

Serviceman 27 13 40

Welder 29 14 43

Table 19.36 Mine Staff Salaries

Position

Base

Salary(US$)

Payroll

Burden(US$)

Salary With

Burden(US$)

Mine Operations

Mine Superintendent 129,000 46,000 175,000

General Mine Foreman 86,000 33,000 119,000

Drill & Blast Foreman 74,000 29,000 103,000

Mine Foreman 57,000 23,000 80,000

Training Coordinator 74,000 29,000 103,000

Mine Clerk 41,000 18,000 59,000

Maintenance Superintendent 118,000 43,000 161,000

Maintenance General Foreman 91,000 34,000 125,000

Maintenance Forem an 76,000 30,000 106,000

Electrical Foreman 76,000 30,000 106,000

Maintenance Planner 75,000 29,000 104,000

Maintenance Clerk 41,000 18,000 59,000

Technical Services

Chief Mine Engineer 109,000 40,000 149,000

Senior Mine Engineer 86,000 33,000 119,000

Dril ling/Blast ing Engineer 70,000 27,000 97,000

Senior Surveyor 59,000 24,000 83,000

Survey Technician 56,000 23,000 79,000

Senior Geologist 87,000 33,000 120,000

Senior Geotechnical Engineer 86,000 33,000 119,000

Geologist – Grade Control 75,000 29,000 104,000





LOM unit operating costs are listed in Table 19.37. The mine cost tables, includingmine capital and operating cost schedules, are provided in Appendix Q.

Table 19.37 Mining Costs per Tonne

LOM CostUS$/t Mill Feed LOM Cost US$/tMaterial Mined

Drilling 0.10 0.06

Blasting 0.27 0.15

Loading 0.18 0.11

Hauling 0.82 0.47

Labour 1.25 0.71

Auxiliary Equipment 0.30 0.17

Dewatering 0.02 0.01

Load/Haul Contract 0.11 0.07

Total Mining Cost 3.05 1.74

M I N E FU E L C O N S U M P T I O N

Fuel consumption rates are estimated in the mine schedule for each equipment type.

These consumption rates are applied to the operating hours of the equipment toestimate the total fuel consumption. Fuel costs have been included in the unitoperating costs estimated above.

1 9. 14 .2 P R O C E S S O P E R A T I N G C O S T S

SU M M A R Y

All process operating costs are shown in US dollars, unless otherwise specified. The

process operating costs are based on a processing rate of 40,000 t/d milled and 92%plant availability.

The average annual process operating cost has been estimated to be US$52.3 M, or

US$3.58 /t milled, excluding G&A costs.

The estimated process operating costs are summarized in Table 19.38 and include

the following:

personnel requirements include supervision, operations, and maintenancestaff with salary/wage levels based on current labour rates at comparableoperations in BC

liner and grinding media consumption has been estimated from the Bond

Ball Mill Work Index and Abrasion Index calculations and quoted budgetprices and/or Wardrop’s database





maintenance supplies have been estimated based on Wardrop’s database

for similar projects

reagents based on test results and quoted budget prices and/or Wardrop’sdatabase

other operating consumables include laboratory chemicals and labwarerequirements, filter cloths, and service vehicles consumables

the unit power cost value at US$0.037/kWh.

Table 19.38 Process Operating Costs and G&A Costs Summary


Annual

Cost(US$)

Unit Cost

(US$/tMilled)

Human Power

Operations Staff 12 1,438,342 0.10

Operations Labour 44 3,079,819 0.21Maintenance Labour 30 2,672,848 0.18

Met Lab & Quality Control 12 813,659 0.06

Sickness And Vacation Contingency Operations 7 548,864 0.04

Sickness And Vacation Contingency Maintenance 3 267,285 0.02

Sub-Total Staff 108 8,820,816 0.60

Supplies (Process Plant)


Maintenance Supplies 795,800 0.05

Power Supply 10,650,225 0.73

Sub-Total Supplies (Process Plant) 43,485,161 2.98

Total Process Operating Costs 108 52,305,977 3.58G&A Staff 27 2,238,209 0.15

G&A Expenses 5,819,000 0.40


Total G & A Costs 27 $8,057,209 0.55

PE R S O N N E L

The projected personnel requirements are 108 persons including:

12 staff for plant management and professional services

44 personnel for plant operations

30 personnel for plant maintenance.

12 personnel for the metallurgical laboratory, process optimization andassaying.





The mill labour requirements are given in Table 19.39, and the mill maintenancelabour requirements are presented in Table 19.40.

Salary/wage rates are based on current rates in northern BC and including base

salary, holiday and vacation pay, pension plan, various benefits, and tool allowance

costs.

The operating and maintenance personnel numbers also include a vacation,

sickness, and absenteeism (VS&A) contingency.

The total estimated annual personnel cost is US$8.8 M, or US$0.60/t milled. The

detailed personnel description and costs are shown in Appendix Q for each

processing plant area.

Table 19.39 Mill Labour Requirements


Base

Rate(US$/h)

Base

Salary(US$)

Loaded

Salary(US$)(33%

Burden)

Annual

Cost(US$)

UnitCost

(US$/tMilled)

Staff

Mill Superintendent 1 151,800 201,894 201,894 0.014

Chief Metallurgist 1 138,000 183,540 183,540 0.013

General Foreman 2 86,020 114,407 228,813 0.016

Chief Assayer 1 92,000 122,360 122,360 0.008

Senior Chemist 1 86,020 114,407 114,407 0.008

Metallurgists 4 86,020 114,407 457,626 0.031

Clerk 2 48,760 64,851 129,702 0.009

Sub-Total Staff 12 1,438,342 0.099

Operations

Mill Foreman (Shift Supervisor) 4 80,371 106,894 427,575 0 .029

Crusher Operators 4 23.98 52,506 69,833 279,330 0.019

Control Room Operators 4 26.08 57,120 75,969 303,876 0.021

Grinding Operators 4 23.98 52,506 69,833 279,330 0.019

Flotation Operators 4 26.08 57,120 75,969 303,876 0.021

Leach Plant Operators 4 23.98 52,506 69,833 279,330 0.019

Dewatering, Drying & Packaging 4 23.12 50,632 67,340 269,362 0.018

Tailing Management 4 21.89 47,932 63,750 254,999 0.017

Reagent Operator 2 23.12 50,632 67,340 134,681 0.009

Labourer/Trainee 10 18.80 41,162 54,746 547,459 0.037

Sub-Total Operations 44 3,079,819 0.211

table continues…






Base

Rate(US$/h)

Base

Salary(US$)

Loaded

Salary

(US$)

(33%Burden)

Annual

Cost(US$)

Unit

Cost

(US$/tMilled)

Assay Laboratory and Quality Control

Metallurgical Technicians 4 23.98 52,506 69,833 279,330 0.019

Assayer 4 23.98 52,506 69,833 279,330 0.019

Sample Bucker 4 21.89 47,932 63,750 254,999 0.017

Sub-Total assay Lab and QC 12 813,659 0.056

VS&A Contingency 7 78,409 548,864 0.038

Total Mill Labour 75 5,880,684 0.403

Table 19.40 Mill Maintenance Labour Requirements


Base

Rate(US$/h)

Base

Salary(US$)

LoadedSalary(US$)

(33%Burden)

Annual

Cost(US$)

Unit

Cost

(US$/tMilled)

Mill Maintenance

Mill Maintenance Superintendent 1 138,000 183,540 183,540 0.013

Mill Maintenance General Foremen 1 86,020 114,407 114,407 0.008

Mill Maintenance Foremen 2 80,371 106,894 213,787 0.015

Electrical Foreman 2 80,371 106,894 213,787 0.015

Mill Maintenance Planner 2 30.77 67,395 89,635 179,271 0.012

Millwrights-Journeyman 6 29.36 64,292 85,509 513,052 0.035

Welders-Journeyman 2 29.36 64,292 85,509 171,017 0.012

Apprentices (level 2) millwright 2 23.98 52,506 69,833 139,665 0.010

Apprentices (level 2) welder 2 23.98 52,506 69,833 139,665 0.010

Electrician 6 29.36 64,292 85,509 513,052 0.035

Instrument Technicians 2 26.08 57,120 75,969 151,938 0.010

Apprentices (level 2)(electrical) 2 23.98 52,506 69,833 139,665 0.010

Total Mill Maintenance 30 2,672,848 0.183

VS&A Contingency 3 89,095 267,285 0.018

Total Mill Labour 33 2,940,133 0.201

PL A N T O P E R A T I N G CO S T S

The total plant operating supplies costs have been calculated to be US$32.0 M/a, or US$2.19 /t milled. These costs include the crushing, grinding, and flotation areas, as

well as reagents, product dewatering, and assaying. The details are provided inTable 19.41. The major consumables include the liner and grinding media, and thereagents.





Table 19.41 Plant Operating Supplies

Supplies

Consumption

( kg/t Ore) Source

Unit Cost

(US$/kg) Source

Unit Cost

FOB Point

Total Cost

(US$/a)

Unit Cost

(US$/t Milled)

Crushing

Gyratory Crusher Liners 0.014 Industry 2.788 Supplier Mine Site 569,785 0.039

Pebble Crusher Liners 0.020 Industry 3.964 Supplier Mine Site 1,157,570 0.079

Screen Panels 2 sets/a Estimate $2,000/set Supplier Mine Site 4,000 0.000

Sub-total Crushing 1,731,355 0.119

Grinding

SAG Mill Balls, 125mm 0.564 Calc 0.856 Supplier Mine Site 7,045,353 0.483

Ball Mill Balls, 75mm 0.564 Calc 0.892 Vancouver 7,348,379 0.503

SAG Mill Liners 0.054 Calc 2.631 Supplier Mine Site 2,074,438 0.142

Ball Mill Rubber Liners 0.038 Calc 2.854 Supplier Mine Site 1,583,310 0.108

Regrind Media 5.87E-04 Calc 8.593 Vancouver 73,579 0.005

Isamill Wear Parts 3 sets/a Estimate $100,000/set Mine Site 300,000 0.021

Sub-total Grinding 18,425,059 1.262

Reagents (Mill)

Frother (DF250) 0.033 2.944 Supplier Vancouver 1,418,419 0.097

Lime 0.218 0.230 Supplier Langley 732,044 0.050

Collector (Fuel Oil) 0.211 0.966 Client 2,975,860 0.204

Flocculant 0.005 3.220 Supplier Vancouver 235,060 0.016

Dryer Fuel Oil 0.080 0.966 Client 1,128,288 0.077

Nokes Reagent 0.119 1.288 Supplier Vancouver 2,237,771 0.153

Hydrochloric Acid 0.050 0.920 Industry Vancouver 671,600 0.046

Sub-total Reagents (Mill) 9,399,042 0.644

table continues…





Supplies

Consumption(kg/t Ore) Source

Unit Cost(US$/kg) Source

Unit CostFOB Point

Total Cost(US$/a)

Unit Cost(US$/t Milled)

Dewater

Filter Cloth 0.002 Industry 1,471,680 0.101

Concentrate Bags Allowance Industry 276,000 0.019

Sub-total Dewater 1,747,680 0.120

Assay and Quality Control

Laboratory Supplies Allowance Industry 184,000 0.013

Environmental Sample Assaying Allowance Industry 46,000 0.003

Sub-total Assay and Quality Control 230,000 0.016

Others

Mill Light Vehicle Operation Allowance Industry 184,000 0.013

Conveyor Belt Replacements Allowance Industry 276,000 0.019

Miscellaneous Allowance Industry 46,000 0.003

Sub-total Others 506,000 0.035

TOTAL OPERATING SUPPLIES 32,039,136 2.194





The liner and grinding media costs will be US$18.4 M/a, or US$1.26/t milled. Theliner and grinding media consumption values were estimated using the Bond

Abrasion Index calculation, and the latest prices have been obtained from thesupplier as a budget price.

The annual reagent costs only amount to US$9.4 M, or US$0.64/t milled. Thereagent consumption values have been taken from the laboratory test resultsadjusted with a typical laboratory-to-plant scale-up factor. The reagent costs are the

current budget prices as provided by potential suppliers of chemical reagents.

The annual maintenance supplies costs have been calculated to be US$0.8 M, or US$0.05/t milled, and is given in Table 19.42. Maintenance supplies estimates have

been estimated based on comparable projects.

Table 19.42 Maintenance Supplies Allowances

Area Total Cost

(US$/a)

Unit Cost

(US$/t Milled)

Crushing 110,400 0.0076

Grinding 230,000 0.0158

Flotation 138,000 0.0095

Tailing Handling 69,000 0.0047

Concentrate Dewatering 82,800 0.0057

Reagents 55,200 0.0038

Assaying 36,800 0.0025

Miscellaneous Mill Supplies 36,800 0.0025

Misc. Building Complex Supplies 36,800 0.0025

Total Maintenance Supplies 795,800 0.0545

The crushing, primary grinding, concentrate regrinding, and other processes require

a 36 MW power source. The process plant power supply is shown in Table 19.43.The details of the estimate are presented in the Appendix Q.

Table 19.43 Total Mill Power Supply

Supplies kWhUnit Cost(US$/kwh)

Total Cost(US$/a)

Unit Cost(US$/t Milled)

Total Plant Power Supply 290,133,618 0.0367 $10,650,225 $0.73

Note: plant power running load = 36,000 kW, and includes a contingency allowance.

1 9. 14 .3 G E N E R A L A N D AD M I N I S T R A T I V E

The G&A costs, estimated based on similar projects undertaken by Wardrop, have

been summarized in Table 19.39 and are presented in detail in Table 19.44. The

total G&A annual costs amount to US$5.8 M, or US$0.40/t milled. The G&A labour





costs only account for US$2.2 M, or US$0.15/t milled. As indicated, no G&Acontingency allowance has been included in the G&A costs.

The G&A costs incorporate the cost items that do not relate directly to the mining or

the processing operating costs, and include the following:

personnel, including the general manager and staff in the accounting,

payroll, human resources, purchasing, safety, and the environmentaldepartments

expenses related to G&A services including insurance, administrativesupplies, medical services, legal services, human resources related

expenses, travelling, accommodation costs, crew transportation, roadmaintenance, and powerline maintenance.

Table 19.44 General & Administrative Expenses


Base

Salary(US$)

Loaded

Salary

(US$)33%

Burden

Annual

Cost(US$/a)

Unit

Cost(US$/

Milled)

G&A Labour

General Manager 1 174,800 232,484 232,484 0.016

Secretary 2 48,760 64,851 129,702 0.009

Human Resources Manager 1 80,371 106,894 106,894 0.007

Human Resources Assistant 1 57,960 77,087 77,087 0.005

Human Resources Clerk 2 48,760 64,851 129,702 0.009

Janitorial 2 48,760 64,851 129,702 0.009

Environmental Manager 1 80,371 106,894 106,894 0.007

Environmental Technician 1 57,408 76,353 76,353 0.005

Purchaser/Buyer 2 80,371 106,894 213,787 0.015

Warehouseman 2 48,760 64,851 129,702 0.009

Safety and Training Officer 2 67,160 89,323 178,646 0.012

First Aid 2 57,960 77,087 154,174 0.011

Public Relations Officer 1 57,960 77,087 77,087 0.005

Chief Accountant 1 80,371 106,894 106,894 0.007

Payroll Clerk 2 48,760 64,851 129,702 0.009

Accounting Clerk 2 48,760 64,851 129,702 0.009

IT Services 2 48,760 64,851 129,702 0.009

Sub-Total G&A Labour 27 2,238,209 0.153

table continues…






Base

Salary(US$)

Loaded

Salary

(US$)

33%Burden

Annual

Cost(US$/a)

Unit

Cost

(US$/Milled)

G&A Expenses

Purchasing and Logistics 46,000 0.003

Office Supplies 46,000 0.003

Communications (Ph/Fax/Internet) 46,000 0.003

Janitorial 69,000 0.005

External Assays/Testing 46,000 0.003

Safety & Training Supplies 46,000 0.003

Medical Service/First Aid 46,000 0.003

Security Supplies 46,000 0.003

Insurances 276,000 0.019

Legal Services 46,000 0.003

Regulatory Compliance 46,000 0.003

Payroll Administration 46,000 0.003

Benefit Administration 46,000 0.003

Audit 46,000 0.003

Consultants 92,000 0.006

Head Office Expenses 46,000 0.003

Travel & Expenses 138,000 0.009

Public Relations & Donations 46,000 0.003

Recruitment 69,000 0.005

Apprentice Training 69,000 0.005

Powerline Maintenance 184,000 0.013Road Maintenance 920,000 0.063

Miscellaneous 230,000 0.016

Crew Transportation 644,000 0.044

Crew Accommodation 2,484,000 0.170

Sub-total G&A Expenses 5,819,000 0.399

Contingency 0% 0 0.000

TOTAL G&A 27 8,057,209 0.552

1 9 .1 4 .4 T M F O P E R A T I N G A N D W A T E R M A NA G E MEN T C OSTS

Table 19.45 provides unit costs for operating and water management.





Table 19.45 TMF Operating and Water Management Costs

Description

Average

AnnualCost (US$)

Unit

Cost(US$/t)

Human Power 331,521 0.02Materials 1,511,094 0.10

Equipment 104,631 0.01

Power 1,904,338 0.13

Contingency 384,253 0.03

Total Tailing Disposal Costs 4,235,838 0.29

1 9. 14 .5 U N I T C A S H OP E R A T I O N C OSTS

The unit cash operating costs over the LOM are provided in Table 19.46.

Table 19.46 Unit Cash Operation Costs (LOM Average)

Description US$

(000s)US$/tMined

US$/tMilled

US$/RecoveredPounds

Mining 657,107 1.63 3.05 1.79

Processing 768,406 1.91 3.57 2.09

G&A 118,825 0.30 0.55 0.32

Power Generation 3,421 0.01 0.02 0.01

Tailings Management 61,085 0.15 0.28 0.17

Totals 1,608,844 4.00 7.47 4.38

1 9 . 1 5 F I N A N C I A L A N A L Y S I S

1 9. 15 .1 I N T R O D U C T I O N

An economic evaluation of the Kitsault Project was prepared by W ardrop based on a

pre-tax financial model and post-tax financial model. For the 15-year mine life, thefollowing pre-tax financial parameters were calculated for the base case pre-taxscenario:

25.5% IRR

3.6 year payback on US$641M initial capital

US$919 M NPV at an 8.0% discount rate.





The post-tax scenario of the base case resulted in:

20.6% IRR

3.8 year payback on US$641 M initial capital

US$551 M NPV at an 8.0% discount rate.

Table 19.47 summarizes the key results of the financial model for the evaluatedscenarios. The base case economic evaluation highlights for the life of the mine are

shown in Appendix R. The note accompanying Table 19.47 is an integral part of the

post-tax financial analysis.

Table 19.47 Summary of the Post-tax Financial Analysis

Units

CPM Market Study Wardrop/

EMCFPricesBase Case Low Case High Case

Molybdenum US$/lb 15.76 14.07 16.90 11.99Exchange Rate to US$ 0.92 0.92 0.92 0.92

NPV (at 8%) US$M 551 372 696 155

IRR % 20.6 16.9 24.2 12.1

Cash Cost/accountable lb Mo US$/lb 4.43 4.43 4.43 4.43

Payback Period years 3.8 4.3 3.3 5.2

Note: The financial analysis includes revenue estimated by KP from the run-of-river hydro-electric

project. Wardrop has used KP’s revenue estimate in this financial analysis but has not

independently verified it.

A sensitivity analysis was carried out to evaluate the project economics for the base

case metal price scenario.

A post-tax model was prepared to evaluate the impact of the base case with thefollowing:

Federal and Provincial corporation taxes

Provincial mining taxes – Net Proceeds and Net Revenue taxes.

1 9. 15 .2 P R E -T A X M O D E L

F I N A N C I A L EV A L U A T I O N S – NPV A N D IRR

The financial analysis has been prepared based in part on the mine and millproduction schedules, recoveries, grades, capital and operating cost estimates,metal deductions, smelter terms, and treatment terms as set out in this report.

Therefore, it should be assumed that the exclusions and assumptions that relate to

the cost estimates and other listed parameters also relate to the economic analysis





(i.e. the occurrence of any of the risk factors identified in related sections) might havea material impact on the accuracy of this economic analysis.

The production schedule has been incorporated into the 100% equity pre-tax

financial model to develop annual recovered metal production from the relationships

of tonnages milled, head grades, and recoveries. Market prices for molybdenumhave been adjusted to realized price levels by applying smelting, refining, andconcentrate transportation charges from the mine site to the smelter in order to

determine the NSR contributions for each metal.

Unit operating costs for mining, milling, and G&A areas were applied to annual milledtonnages to determine the overall mine site operating cost which has been deducted

from the net smelter return (NSR) to derive annual Net Revenues.

Initial and sustaining capital costs have been incorporated on a year-by-year basisover the mine life and deducted from the Net Revenue to determine the Net Cash

Flow before taxes. Initial capital expenditures include costs accumulated prior to the

first production of concentrate; sustaining capital includes expenditures for miningand milling additions, replacement of equipment, and waste management.

Working capital has been calculated on the basis of three months of mine site

operating costs and applied to the first year of expenditures. It will be recovered atthe end of the mine life and applied towards reclamation during closure.

The financial analysis model was developed based on the following factors.

All taxes included in the financial model have been reviewed and confirmedby PWC.

Capital and operating costs were provided as outlined in Sections 19.13 and19.14.

No metal penalties have been applied in the analysis.

Metal revenues projected in the cash flow models were based on averageproduction values as shown in Table 19.48.

Table 19.48 Metal Production from Kitsault Project

Units Life of Mine

Total Tonnes to Mill Mt 215.3

Annual Tonnes to Mill Mt 14.6

Average Grades

Molybdenum Mo % 0.085

Total Production

Molybdenum Mlb 367.91

Average Annual Production

Molybdenum Mlb 24.5





The undiscounted annual cash flows are illustrated in Figure 19.38.

Figure 19.38 Undiscounted Annual and Cumulative Cash Flows

ME T A L PR I C E A N D EX C H A N G E RA T E SC E N A R I O S

The project economics were completed for the four pricing scenarios based on theCPM Molybdenum Industry Analysis and Wardrop Energy & Metals ConsensusForecasts (EMCF) (Table 19.49) and have been used for comparison purposes.

Wardrop’s new policy utilizes the EMCF quarterly reports (The ConsensusEconomics Inc.) in calculating the Wardrop/EMCF prices. This new approach is toavoid large fluctuations in metal prices from study to study and to use the long term

price averaged from three EMCF quarterly reports. For this study, if executed

between August 31, 2009 and January 1, 2010, the long term metal prices would bederived by averaging the long term prices for previous January, April, and July

quarterly reports to drive the Wardrop/EMCF prices. The average long term price of molybdenum is calculated using CPM projections until 2018 and then maintained flat

for the rest of the mine life.





Table 19.49 Summary of Metal Price and Exchange Rate Scenarios

Period

CPM Market Study (US$)Wardrop/

EMCFPrices (US$)

Base

Case

Low

Case

High

Case

Exchange Rate to US$ 0.92 0.92 0.92 0.92

Year 1 (Q4 2013) 14.38 12.88 16.88 11.99

Year 2 13.56 12.75 15.25 11.99

Year 3 14.31 13.38 15.13 11.99

Year 4 14.88 13.88 15.56 11.99

Year 5 16.13 14.38 17.06 11.99

Year 6 to End of Mine Life 16.50 14.50 17.50 11.99

Average Mo Market Price 15.76 14.07 16.90 11.99

An FXR of 0.92 (US$:C$) has been used in development of the capital cost estimate

and in all financial scenarios and cases. The high case, low case, and theWardrop/EMCF prices were applied to the same base case financial model. Theresults of all four scenarios as described are presented in Table 19.50.

Table 19.50 Pre- and Post-tax NPV, IRR, and Payback by Metal Price Scenario

US$ Post-tax Report Units

BaseCase

LowCase

HighCase

Wardrop/EMCF

Discount Rate % 8.0 8.0 8.0 8.0

Weighted Average Metal Prices (LOM) US$/lb 15.76 14.07 16.90 11.99

Post Income Tax NPV US$ M 551 372 696 155

Pre-tax NPV US$ M 919 644 1,136 314

Post Income Tax IRR % 20.57 16.90 24.17 12.07

Pre-tax IRR % 25.5 21.1 29.9 15.2

Initial Capital US$ M 641 641 641 641

Total Sustaining US$ M 400 400 400 400

Total Operating Cost US$/t milled 7.47 7.47 7.47 7.47

Mine Life years 14.7 14.7 14.7 14.7

Average Cash Cost per Accountable lb Mo US$/acc lb 4.43 4.43 4.43 4.43

Pre-tax Payback Period years 3.6 4.1 3.1 4.9

Post-tax Payback Period years 3.8 4.3 3.3 5.2

1 9. 15 .3 P O S T - T A X M O D E L

PWC performed a due diligence review on a tax model presented by Wardrop. PWC

has confirmed that the tax model accurately predicts the after-tax cash flow.

Economic evaluation indicated a post-tax IRR of 20.6% and a post-tax NPV of US$551 M at a discount rate of 8.0%. The post-tax base case financial model used





the same inputs as the pre-tax economic evaluation with the following key taxationrates:

Federal corporate tax at 15%

Provincial corporate tax at 11% Provincial net proceeds tax (Stage I) at 2%

Net provincial revenue tax (Stage II) at 13%.

Capital Cost Allowance (CCA) pools have been categorized according to federal tax

provisions.

1 9. 15 .4 SE N S I T I V I T Y AN A L Y S I S

A sensitivity analysis was carried out on the following parameters:

molybdenum prices

exchange rate

initial capital expenditure

mine site operating costs.

The analysis is tabulated as financial outcomes in terms of pre-tax NPV (US$) and

pre-tax IRR. The project NPV (8% discount) is most sensitive to the exchange rateand molybdenum prices (Figure 19.39).

Figure 19.39 Pre-tax NPV Sensitivity Analysis

Similarly, the project IRR is most sensitive to molybdenum price and exchange rate

followed by operating cost and capital cost (Figure 19.40).





Figure 19.40 IRR Sensitivity Analysis

In addition, a sensitivity analysis has been performed on the post-tax NPV, analyzed

by variations of metal prices and discount rates; the results are shown in Table19.51.

Table 19.51 Sensitivity Analysis to Post Tax NPV (US$ M)

Mo Price(US$/lb)

Discount Rate

0% 5% 8% 10%

8.00 -204 -346 -392 -412

12.00 772 324 156 71

16.00 1,620 893 618 477

20.00 2,476 1,460 1,073 874

24.00 3,335 2,027 1,526 1,269

1 9. 15 .5 R O Y A L T I E S

The project is subject to a 9.22% net cash flow interest, which is owned by AZN. Netcash flow is only payable after the recovery of all capital costs associated with

construction or sustaining capital. In addition, there is a 1% NSR royalty to AlcoaInc. (Alcoa) that has been allowed for. Alcoa has an election to take a $10 M cash

payment at commercial production, 90 days after Avanti presents them with a

Feasibility Study. For the purpose of the economic model, the royalties were

modelled per the information provided by Avanti.





1 9. 15 .6 SMELTER T E R M S

Contracts will generally include payment terms for molybdenum as follows:

There will be 1% deduction from the recovered molybdenum by the smelter;

therefore, the mine will receive 99% of the recovered molybdenum.

The gross value of molybdenum is calculated by applying molybdenumprices to the accountable molybdenum. Accountable molybdenum is

recovered molybdenum less 1% deductions.

There is a roasting charge of US$1 per accountable pound of molybdenum.

1 9. 15 .7 C O N C E N T R A T E T R A N S P O R T L O G I S T I C S

Concentrate from the mine site will be transported via truck to the Kitsault barge.The costs for concentrate transportation are as follows:

truck transport to Kitsault Barge – Cdn$2.63/wmt

ocean transport – US$46.61/wmt.

C O N C E N T R A T E TR A N S P O R T IN S U R A N C E

An insurance rate of 0.15% will be applied to the provisional invoice value of theconcentrate to cover land-based transport as well as ocean transport from the mine

site to the smelter.

O W N E R ’S R E P R E S E N T A T I O N

There will not be any owner’s representation on the concentrate.

C O N C E N T R A T E LO S S E S

Concentrate losses are not applied to this evaluation as the concentrate is shipped

by 2-ton super sacks.





2 0 . 0 P R O J E C T E X E C U T I O N P L A N

The project will be designed and constructed to industry and regulatory standards,

with emphasis on addressing all environmental and safety issues.

Upon Mines Act approval and authorization to proceed in Q4 2010, the facility will

take approximately 35 months to construct and be completed in Q3 2013.

Construction will be concurrent with acquiring the key agency permits.

The project schedule is shown in Figure 20.1. The environmental and permitting

timeline is critical; other work, such as bulk sampling and detailed engineering, can

proceed and have some flexibility in their completion times.

The environmental assessment process has already been partially completed. The

longest lead items include the procurement time length and include the SAG mill, ball

mills, and large mining equipment.

The construction of the tailing starter wall, project financing, mining procurement of

equipment, mobilization of equipment, and subsequent pre-stripping are critical

activities for the delivery of the project.

The water diversion tunnel is a major construction activity but is not on the current

critical path. However, timely construction of the haul roads from the pit to the TMF

is vital to ensure the dam starter wall can be built by the end of Year -1, allowing the

tailing to be deposited in the TMF. The TMF has been designed as a flow-throughfacility.

Based on scheduling information from KP, the water diversion tunnel will be driven

from two headings by drill and blast techniques and will take approximately eight

months to complete.

The subsequent project implementation phase is based on an engineering,

procurement, and construction management (EPCM) approach. The engineering

work is assumed to be conducted by experienced engineering contractors. This

approach has worked successfully with similar past projects, allowing for tight control

of the project costs and schedule. The execution plan assumes approval for full

project release is received upon receipt of the project permits and prior to start of theEPCM phase.

The success of meeting the planned project completion date is subject to numerous

risks and uncertainties at this time. It is anticipated that any further study phase will

include a number of optimization studies, the timely completion of which is key to

meeting the project schedule. In addition, many events or circumstances beyond the





control of Avanti or its contractors can affect construction progress including, without

limitation: inclement weather, fire, floods, labour disruptions, interruptions, or late

receipt of supplies and equipment, interruptions of construction caused by

government restrictions on construction, failure of negotiations or reaching

agreement with third-party interests.

The project schedule assumes that no internal or external constraints, such as a

delay in funding approval or receipt of the project permits, will prevent the planned

start-up. Figure 20.1 indicates the key project milestones.





Figure 20.1 Project Schedule





2 1 . 0 E N V I R O N M E N T A L C O N S I D E R A T I O N S

2 1 . 1 R E G U L A T O R Y A P P R O V A L P R O C E S S

21.1.1 O V E R V I E W

Major mining projects in BC are subject to environmental assessment and reviewprior to certification and issuance of permits to authorize construction and operations.

Environmental assessment is a means of ensuring the potential for adverse

environmental, social, economic, health, and heritage effects or the potential adverseeffects on Aboriginal interests or rights are addressed prior to project approval.

There are generally two stages in the environmental assessment:

the pre-application phase when studies and consultations are undertaken

the application review phase during which the project details and effects onenvironment and communities are reviewed along with further consultations.

Generally, the scope, procedures, and methods of each assessment are flexible andtailored specifically to the project circumstances. These are defined in an approvedTerms of Reference. Depending on the scope of the project, assessment and

permitting of major mines in BC may proceed through either:

the BC Environmental Assessment (BC EA) process pursuant to theEnvironmental Assessment Act (BC EAA); or

the Major Mine Review Process pursuant to the Mines Act .

Federal government review under the Canadian Environmental Assessment Act

(CEAA) is independent of provincial review as the federal agencies, deemedresponsible authorities (RA), will ascertain if there are federal “triggers” invoked by

the project, thereby resulting in an applicable level of federal review (e.g. none,

screening level assessment, or comprehensive review). The BC EA and CEAA aresubject to harmonization agreements to expedite reviews, whereas the Mines Act

and CEAA are not yet harmonized.

In general each environmental assessment contains four common main elements(McLaren, 2008):

opportunities for all interested parties, including First Nations andneighbouring jurisdictions, to identify issues and provide input





technical studies of the relevant environmental, social, economic, heritage

and health effects of the proposed project

identification of ways to prevent or minimize undesirable effects andenhance desirable effects

consideration of the input of all interested parties in compiling theassessment findings and making recommendations about project

acceptability.

An EA certificate or Mines Act Permit, issued by the Minister of Environment or

Minister of Energy, Mines, and Petroleum Resources, respectively, at the conclusion

of the environmental assessment, represents government approval in principle andallows the proponent to seek any other statutory authorization needed to constructand operate the project.

The Kitsault Mine is not an abandoned mine because there is a Mines Act Permit inplace and there are outstanding reclamation obligations for which the MEMPR holdssecurity.

The Kitsault Mine is a permitted brownfield site with considerable past mining activityand basic infrastructure in place. The project meets the criteria that constitute a non-reviewable project under Part 3 of the Reviewable Project Regulation. Re-opening

the mine can well be considered a modification of an existing project.

Avanti has received confirmation that the project does not automatically trigger theprovincial Environmental Assessment Act review, in lieu of the grandfathering

provision of Section 51 of the Environmental Assessment Act which applies to Permit

M-10. Accordingly, Avanti has an option to request an opt-in to the BC EA review

process, or pursue assessment as a major mine under the Mines Act review processadministered by the Northwest Regional Mine Development Review Committee. Avanti is considering a number of factors in making this decision, including how eachreview process is able to integrate federal and Nisga’a treaties into the review

process.

Federal government agencies will conduct separate screening of the project toascertain whether triggers exist for additional federal review of the project, primarily

in the areas of fisheries, explosives use and manufacture, and navigable waters.

These determinations are still forthcoming and will be informed by the outcome of thepre-feasibility level of mine planning detail and environmental baseline studies.

21.1.2 BR I T I S H C O L U M B I A AU T H O R I Z A T I O N S, L I C E N S E S, A N D P E R M I T S

The Kitsault Project will require various provincial and federal authorizations,licenses, and permits to operate the project.

Table 21.1 shows a preliminary list of the BC authorizations, licenses, and permitsthat Avanti will be required to obtain. The completed technical studies and





environmental assessment will form the basis of the applications. The permitrequirements will be reviewed and updated as the project advances through the

environmental assessment and permitting process.

Table 21.1 BC Authorizations, Licenses, and Permits Required for the Kitsault

Project

BC Government Permits and Licenses Enabling Legislation

Permit Approving Work System & Reclamation Program (Mine Site –

Initial Development)

Mines Act

Amendment to Permit Approving Work System and ReclamationProgram (Pre-production)

Mines Act

Amendment to Permit Approving Work System and Reclamation

Program (Bonding)

Mines Act

Amendment to Permit Approving Work System and Reclamation

Program (Mine Plan – Production)

Mines Act

Permit Approving Work System and Reclamation Program (Gravel

Pit/Wash Plant/Rock Borrow Pit)

Mines Act

Mining Lease amendment (if required) Mineral Tenure Act

Water License – Notice of Intention (Application) Water Act

Water License – Storage and Diversion Water Act

Water License – Use Water Act

Water License – Construction of fences, screens and fish or game

guards across streams to conserve fish or wildlife

Water Act

Water License – Alteration of Stream or Channel Water Act

Authority to Make a Change In and About a Stream – Notification Water Act/

Water Regulation

Authority to Make a Change In and About a Stream – Approval to Makea Change

Water Act/

Water Regulation

Authority to Make a Change In and About a Stream – Terms andConditions of Habitat Officer

Water Act/

Water Regulation

Occupant License to Cut – Access Road modifications involving tree

removal

Forest Act

Occupant License to Cut –Mine Site/Tailings Impoundment Forest Act

Occupant License to Cut – Gravel Pits, Borrow Areas Forest Act

Road use Permit (existing Forest Service Road); or Forest Act

Special Use Permit – Access Road Forest Practices

Code of BC Act

License of Occupation – Staging Areas on Crown Land offsite Land Act

License of Occupation – Borrow/Gravel Pits on Crown Land offsite Land Act

Waste Management Permit – Effluent (Sediment, Tailings and Sewage) Environmental Management Act

Waste Management Permit – Air (Crushers, Ventilation, Dust) Environmental

Management Act

Waste Management Permit – Refuse Environmental

Management Act

table continues…





BC Government Permits and Licenses Enabling Legislation

Special Waste Generator Permit (Waste Oil) Environmental

Management Act

(Special Waste

Regulations)

Sewage Registration Environmental

Management Act

Camp Operation Permits (Drinking Water, Sewage Disposal, Sanitation

and Food Handling)

Health Act/

Environmental

Management Act

Waterworks Permit Drinking Water

Protection Act

Fuel Storage Approval (volume-based) Fire Services Act

Food Service Permits Health Act

Highway Access Permit Highway Act

21.1.3 F E D E R A L AU T H O R I Z A T I O N S, L I C E N S E S A N D P E R M I T S

Table 21.2 shows a preliminary list of the required and potential federal

authorizations, licenses, and permits required by Avanti to operate the project.

Table 21.2 Federal Authorizations, Licenses, and Permits Required for the

Project

Federal Government Approvals and Licenses Enabling Legislation

Canadian Environm ental Assessm ent Agency Approval CEAA

Metal Mining Effluent Regulations (MMER) Fisheries Act/

Environment CanadaFish Habitat Compensation Agreement Fisheries Act

Section 35(2) Authorization for Harmful Alteration, Disruption or Destruction Of Fish Habitat (HADD) 1

Fisheries Act

Navigable Water: Stream Crossings Authorization 2

Navigable Waters Protection

Act

Explosives Factory License Explosives Act

Explosives Magazine License Explosives Act

Ammonium Nitrate Storage Facilities Canada Transportation Act

Radio Licenses Radio Communications Act

Radioisotope License (Nuclear Density Gauges/X-ray Analyzer) Atomic Energy Control Act

1

required if a Fisheries Act HADD occurs.2 required if Transport Canada determines navigable waters involved in project

Note: The project will be subject to the MMER enabled by the Fisheries Act . The regulations

require Avanti to achieve the specified effluent discharge standards, to implement a comprehensive

Environmental Effects Monitoring program, and to provide compensation for the harmful alterationof fish habitat should this occur.





21.1.4 R E C L A M A T I O N SE C U R I T Y BON D

Section 10 of the BC Mines Act stipulates that the Chief Inspector of Mines may, as a

condition of issuing a permit, require that the mine owner provide monetary securityfor mine reclamation and to provide for protection of, and mitigation of damage to,

watercourses and cultural heritage resources affected by the mine. Security willremain in effect until such time as the Chief Inspector of Mines determines that allreclamation obligations have been met and Avanti can be indemnified. Avanti hasposted a Cnd$100,000 bond, which is required under the current M-10 permit.

During the mine planning and environmental assessment phase, the MEMPRreclamation costing spreadsheet will be completed as the basis for initiatingreclamation security negotiations with the province. The amount of security required,

and the form in which the security is to be provided, will be agreed between Avantiand the Chief Inspector of Mines (with input from the Ministry of Finance), as part of

the permitting process. The predicted capital and long-term operating costs of the

mine site water collection and treatment system (if required) will likely be taken intoconsideration when deciding the amount of security required (BC Ministry of Energyand Mines (MEM) and Ministry of Environment, Lands, and Parks (MELP), 1998).

Performance bonds are an acceptable means of providing this security. In addition,enough “hard" security must be posted so that at any point in time, the amount willfully cover the next five-year period of expected post-closure costs related to water

treatment and site management and monitoring (BC MEMPR, 2006). Reclamation

securities are reviewed periodically during the mine operation and post-closure

periods to ensure required levels of security reflect operational circumstances andprevailing financial conditions.

2 1 . 2 E N V I R O N M E N T A L S T U D I E S

As a requirement of project review, environmental baseline studies are required to

inform assessments of the project’s impacts on the physical and socioeconomicenvironments, management, mitigation, and monitoring plans. Until such time as a

formal government-approved project Terms of Reference for conducting such studies

is finalized, Avanti has proactively initiated a comprehensive review of past studiesand commenced new studies to fulfil the requirements for environmentalassessment. Beginning in October 2008, over 20 discipline-specific and inter-related

studies were initiated. The individual study objectives, and work completed for each

discipline, are outlined in the following sections.





21.2.1 SO I L S A N D T E R R A I N M A P P I NG

O B J E C T I V E S

The soils and terrain mapping is being undertaken for the project area as part of theTerrestrial Ecosystem Mapping (TEM) program. TEM is required for the effects

assessment of the project. In addition to providing input for the TEM program, the

soils and terrain mapping is used as a tool to establish pre-project baseline soilconditions and to identify features that are locally or regionally rare and/or sensitiveto disturbance. This includes the identification of the effects the project on the

various soils types and soils that are sensitive to disturbances, such as soils that are

susceptible to erosion and degradation due to project activities. The identification of such soils is important in developing management and mitigation plans that will

eliminate or minimize these effects.

The objectives of the soils and terrain mapping program are to:

establish the baseline soil and terrain conditions for the project study area

provide data to feed into the TEM program

provide the basic soil resource data on which potential effects of the project

may be assessed

assist in establishing pre-mine capability and/or productivity in the study

area

characterize the topsoil and subsoil for suitability as a growth medium for reclamation.

S T U D Y A R E A A N D M E T H O D O L O G Y

The soils and terrain mapping study area, also the TEM study area, is approximately

36 km2 (3,600 ha). It includes all areas for the proposed mine devolvement such as

the mine site facilities (pit, plant site, shop, waste rock dumps, etc.), tailing storagefacilities, access roads, and marine facilities (wharf and camp).

The soils and terrain mapping being undertaken is based on a combination of

desktop-based document review and field surveys that are developed usingstandards and methods established in BC (BC MELP and Ministry of Forests, 1998;Resource Inventory Committee (RIC), 1998, 2000). The soils will be described

according to the Canadian System of Soil Classification Agriculture (Agri-FoodCanada, 1998) for mapping purposes. The resulting soils and terrain information canbe displayed spatially using GIS technology and can also be linked with other spatial

data, resulting in a useful resource management and planning tool.





W O R K C O M P L E T E D

A detailed review of available information has been conducted, including all relevant

data collected during the former Kitsault Mine operations. Terrain pre-typing wascarried out using black and white (1994 stereo) aerial photographs at a scale of

1:15,000 that were available for the project area. A field survey of the soils andterrain was conducted in July and September of 2009. A total of 97 sites wereinspected between the two field programs – 66 sites on the first trip and 31 on thesecond field survey – resulting in an inspection intensity of approximately one

inspection per 40 ha. Detailed soil and terrain data was collected at each site. Soilsamples were taken from representative locations for the determination of baseline

metal levels, soil pH, and organic carbon.

W O R K R E M A I N I N G F O R P E R M I T A P P L I C A T I O N

Laboratory and data analysis, refining of the pre-typed terrain mapping based on the

field survey results, and the preparation of surficial material mapping are currentlybeing undertaken by Rescan.

21.2.2 W A S T E R OC K M E T A L L E A C H I N G A N D AC I D R O C K D R A I N A G E P O T E N T I A L

Refer to Section 19.6.1 of this report for details on waste rock metal leaching andacid rock drainage potential.

21.2.3 M E T E O R O L O G Y A N D AT M O S P H E R E

O B J E C T I V E S

On-site meteorological data are required for a variety of purposes. Wind speed anddirection data are usually required to select and locate infrastructure. Wind, air

temperature, and barometric pressure data are required to estimate the dispersion of fugitive dust from the project activities and to determine the project’s potential air quality effects. Wind, solar radiation, precipitation, and snow survey data are

required for the design of tailing and water storage facilities as well as water balance

calculations. On-site ambient air quality data, in terms of dust deposition, arerequired for determining background deposition values that will be incorporated into

the air dispersion modelling, during the environmental assessment.

Noise associated with industrial operations is often discussed in the context of worker health and safety. However, mining operations also expose the environment

outside the project fence-line to elevated noise levels. These can be perceived as aconsiderable nuisance by neighbouring communities even though typically noise

levels beyond the project fence-line will not negatively affect health. In addition,

wildlife populations, especially birds, have been shown to avoid areas of elevatednoise. This might lead to the fragmentation of valuable habitat and abandonment of migration routes.





The objectives of the 2009 atmospheric baseline program were as follows:

characterize the air quality and meteorology settings to sufficient detail for

the environmental assessment and to supply engineers with informationnecessary for project design, including the meteorological variability

characterize baseline noise conditions in the project area.


The meteorology, air quality, and noise baseline study areas focused on theproposed mine site and port areas. A meteorology station was installed in theproposed mine site area during November 2008. The site selected for the

meteorological station and the position of the station’s sensors, complied withestablished guidelines by Environment Canada – Meteorological Services of Canada

(MSC) in the Guidelines for Co-operative Climatological Autostations (MSC, 2004).

Two snow courses were established at the Kitsault Project site and adhere to therecommendations and procedures found in the BC Ministry of Environment (MOE)Procedure Manual for Snow Surveys (Volume 6, Section 9; December 1982). One

snow course is located near the existing automated meteorological station. Thesecond snow course is located near the proposed TMF.

Five locations were selected for dustfall monitoring – two near the location of theopen pit, one near the plant site, one along the proposed road corridor (plant or minesite road), and one near the proposed port site. Dustfall methodology complied with

the guidelines outlined in ASTM D 1739 – 98 (Reapproved 2004) (ASTM, 2004)Standard Test Method for Collection and Measurement of Dustfall (Settleable

Particulate Matter).

Baseline noise monitoring complied with the ASTM Method E 1686 – 03 StandardGuide for Selection of Environmental Noise Measurements and Criteria. Six

monitoring locations were selected at the proposed mine and port sites.


The following work has been completed as part of the 2009 baseline meteorology, air quality, and noise studies:

the dataset from the on-site meteorology station is continuous and inclusive

to July 22, 2009 (the date of the most recent data download was as of September 25, 2009)

the snow courses were sampled monthly from February to May 2009 (four

months)

dustfall monitoring took place on a monthly basis from May to September

2009 (five months)





baseline noise was monitored for 24 hours during each of the four seasons

(February, May, July, and September, 2009).


A baseline report summarizing the 2009 meteorology, air quality, and noise baselinestudies will be prepared. The baseline report will also provide regional analyses for

meteorological parameters (including snow surveys) based on the nearestEnvironment Canada meteorological stations. Following the baseline report, aneffects assessment will be conducted with respect to the atmospheric environment.

This assessment will include air quality modelling, noise modelling, and agreenhouse gas inventory as well as discussions of the potential effects of projectgenerated air emissions and noise on the surrounding environment and nearby

communities.

21.2.4 SU R F A C E H YDROLOGY

O B J E C T I V E S

The Kitsault Project is located in the humid environment of the Coast MountainRange. Proximity to the coast, relatively high precipitation rates, and mountainousterrain result in high amounts of surface water runoff within the project area. An

understanding of the surface hydrology of the project site will be essential to a robusteffects assessment of the project on the environment as well as the design and

operation of an appropriate water balance and water management plan. The primary

objective of the surface hydrology monitoring program is to characterize the currenthydrologic regime of the project area.


The hydrologic monitoring program focused on the Lime Creek watershed as theprimary study area. Three Water Survey of Canada hydrometric stations historicallyoperated within the watershed: at the mouth of Lime Creek, Lime Creek below Patsy

Creek, and at the mouth of Patsy Creek. These stations provide good spatial and

temporal coverage of the project area. The three stations are located downstream of the project area and at the two main drainages within the project area, and have data

sets that range from 1 year (Lime Creek below Patsy Creek) to 20 years in length (atthe mouth of Lime Creek). It is unusual to have such extensive historical data sets to

build upon within the immediate area of a proposed development in northwesternBC; the 2009 monitoring program has taken full advantage of this.

Monitoring stations consist of a pressure transducer paired to an automated data

logger that collects water level (stage) data at regular intervals (15 min to 1 h).

Continuous stream level data will be converted to stream flow data by use of a stage-discharge curve developed for each monitoring station. Stage-discharge curves will

be developed using manually measured stream flows over a range of water levels.






Continuous stream level monitoring has been re-established at the location of the

three former Water Survey of Canada hydrometric stations. The stations operatingon Lime Creek below Patsy Creek and on Patsy Creek were re-established during

the spring of 2009. The station operating at the mouth of Lime Creek was re-established in the fall of 2008.

In 2009, 17 manual flow measurements were conducted at the 3 monitoring stations

during 6 site visits over the winter and open water seasons.


Manual flow measurements will be used along with concurrently observed water

levels to generate stage-discharge rating curves for each station. The continuous

water level records obtained from each site will be translated to continuous flow

records using the stage-discharge curves. The resulting hydrographs will becompared to the historical data sets to assess and determine the applicability of the

historic data sets to current project area hydrological conditions. If applicable, the

combination of the historical and 2009 stream flow data will be used to estimate a

number of important hydrological indices for the project area including annual andreturn period runoff, seasonal flow distribution, and expected occurrence of peak andlow flow events.

It is expected that the combination of the historic and 2009 data sets alreadycollected will be adequate to allow for a robust environmental effects assessment for the project. However, it is recommended that stream flow data continue to be

collected throughout all phases of the project as this will provide essential information

to refine and adapt the water balance and water management plan, assess actualeffects to the environment during operations, and to allow for appropriate reclamation

and closure planning and design.

21.2.5 H Y D R O G E O L O G Y

O B J E C T I V E S

The aim of the hydrogeology program is to define the baseline groundwater quality

as well as the potentiometric surface and hence the groundwater flow regime of theproject area. Sufficient data was required to be obtained to develop a reasonable

understanding of the subsurface hydraulic conditions.

The fieldwork has involved the installation of groundwater monitoring wellsthroughout the property. Groundwater levels in the wells will be monitored, samples

will be taken to determine groundwater quality, and hydraulic conductivity testing willbe performed to determine the properties of the screened lithologies of the wells.

Additional hydraulic conductivity testing was also performed as drill holes were





advanced. These data sets will be related to the geology of the site with the aim of developing a representative groundwater model to characterize the project site.


The hydrogeology investigation and monitoring program focused primarily on theLime Creek watershed. It is within this watershed that Avanti plans to place most of

their proposed infrastructure. A total of 19 monitoring wells were completed withinthis watershed by KP during the 2009 field program. Two monitoring wells had been

installed by SRK in the open pit area during 2008.

A wharf area is planned at the lower elevations of the Roundy Creek watershed.Three monitoring wells were installed in the area of the proposed wharf to assess

groundwater levels and groundwater quality.


A total of 24 groundwater monitoring wells have been installed throughout theproperty. Monitoring wells have been installed by KP (22) and SRK (2); it is

understood that SRK plan to install a vibrating wire piesometer string. As the KP drillholes were advanced, packer tests were performed to determine the hydraulic

properties of the tested lithologies; a total of 95 packer tests have been performed.

In September 2009, Rescan completed the first round of groundwater sampling;further groundwater samples will be taken at quarterly intervals. Groundwater quality

parameters to be assessed from laboratory results include general chemistry, total

metals, dissolved metals, nutrients, and total organic carbon; field parameters

include oxygen reducing potential, total dissolved solids, electrical conductivity, pH,and temperature.


Additional field work to be completed will include the collection of additional quarterly

samples to define the baseline groundwater quality. During sampling eventsgroundwater levels will also be measured to determine the extent of seasonal

variation in the levels.

Upon completion of the field work, a baseline report will be produced, which willdefine the groundwater quality, the groundwater flow regime (including areas of

groundwater recharge and discharge), and will include all the data collected duringthe field work.

The information contained within the baseline report will be used to develop an

effects assessment for the project. A representative groundwater model will be

developed to predict potential impacts to the groundwater system, both in terms of groundwater quantity and quality, of the proposed development. The effects





assessment, largely based on the groundwater model results, will then be produced.The effects assessment will quantify the magnitude as well as spatial and temporal

extents of predicted environmental impacts caused by the proposed development of the Kitsault Project.

21.2.6 O C E A N O G R A P H Y

O B J E C T I V E S

To accommodate berthing barges for the transport of mining concentrates to globalmarkets, a port site development has been proposed approximately 2 km southwest

of the town of Kitsault on the south shore of Alice Arm, BC. This is immediatelydownstream from the former submarine tailing outfall that was present in the 1980s.To provide defensible data for an accurate effects assessment of the project on the

surrounding marine habitat, a comprehensive marine sampling was constructed.The objectives of the marine baseline studies are to characterize the followingcomponents of the study area:

physical oceanography (water column structure, inlet circulation, and oxygen

content)

chemical oceanography (water and sediment quality)

biological oceanography (phytoplankton, zooplankton, and benthos).


Sampling for the marine program was undertaken in July 2009 and encompassed the

shoreline habitat 1.5 km upstream and 1.5 km downstream of the proposed port site,a cross-channel reference station, and a pelagic station of Alice Arm. Selected

stations were based on current available knowledge related to the proposed projectwith key stations located at the proposed port location, at the outflow of Lime and

Roundy creeks, as well as the cross-channel reference station and a mid-channelpelagic station. In all, 17 stations were sampled with 14 occurring along the

proposed port shoreline.

At each station, a suite of physical, chemical, and biological components werecollected. Water column structure was determined using salinity, temperature, and

dissolved oxygen measurements. Water and sediment quality each included nutrient

(nitrogen and phosphorus) and trace metal chemistry with polycyclic aromatichydrocarbon (PAH), polychlorinated biphenyl (PCB), and tributyltin (TBT) analyses

included as part of the sediment quality program. To determine any biologicaleffects, abundance, biomass, and taxonomy samples were collected for phytoplankton, zooplankton, and benthos.






All marine components mentioned above were successfully sampled in July 2009.

Early results show that the water column structure in Alice Arm is typical of BC fjordswith low salinity, high temperature surface water overlying high salinity, cool

temperature bottom water. Pycnocline depths were ~10 m. The surface waters werelargely saturated with dissolved oxygen but were progressively subsaturated withdepth. No anoxia was found to 50 m depth.

Chemistry and biology analyses are ongoing.


A baseline report summarizing the 2009 marine survey will be prepared. The ability

to move into a successful environmental effects assessment will be largely

determined by the integrity of the benthos analysis, which was collected from several

difficult sediment habitat types (e.g. coarse sand and high woody debris habitats). Itis possible that additional sediment samples may need to be collected in 2010 if 2009 samples are deemed unsuitable to properly assess future biological effects. It

is further recommended that a scaled-down water quality survey be implemented for

the winter of 2010 to gather ‘true’ baseline water quality conditions (i.e. those withlimited biological effects).

21.2.7 F R E S H W A T E R AQ U A T I C S

O B J E C T I V E S

The 2009 sampling program was designed to provide an overview and characterizethe freshwater aquatic resources within the Kitsault Project area, including water

quality, sediment quality, primary and secondary producers. Sampling locationswere chosen based on the project information available at the time. The samplingprogram focused on selected receiving environments (sites downstream of potential

project activities) and reference sites including stream, lake, and wetland habitats.

The 2009 freshwater aquatic resources baseline report will summarize these data.


The general study area included streams and wetlands in the vicinity of the project

area for the 2009 aquatic resource studies. In total, 7 stream sites, 4 wetlands, andPatsy Lake were visited in August 2009. Selected stream sites are located either

upstream or downstream of the pit and/or tailing dam. Lake and wetland sites willnot receive any direct discharge from the project but were selected because they are

in close proximity to project activities and will be monitored for potential impacts.

Reference stream sites are located in an adjacent watershed.





During this sampling period water quality, sediment quality, benthic invertebrates,zooplankton (Patsy Lake only), phytoplankton (lake and wetlands), and periphyton

(streams) samples were collected. In addition to this work, monthly water qualitysampling began in May 2009 at three stream sites downstream from project activities

and one reference stream.


The majority of the 2009 fieldwork has been completed with ongoing monthly water

quality sampling. The warm and dry weather that preceded the August field trip

caused water levels in Patsy Creek and its tributaries to be too low to fully completethe sampling work. Detailed results of this work will be presented in a baselinereport.

Generally, at least one year of monthly stream water quality data is required toassess seasonal variability at a site. This data is being collected at four of the six

stream sites. It is common to have gaps in this type of dataset since samples can bedifficult to collect during low flow periods or during winter when snow and ice make

access the stream difficult or unsafe. For this reason, monthly sample collectionsbeyond one year may be undertaken in order to characterize the seasonal patterns in

the project area as accurately as possible.

Sampling sediment quality is completed once at each stream, wetland, or lake siteover the course of a sampling season; these samples were collected in August. The

nature of the stream sites within the project area is composed of large boulder andcobble substrates often combined with relatively steep grades. This results in very

few depositional areas for fine sediments to collect along the stream.

Primary (phytoplankton and periphyton) and secondary (zooplankton and benthicinvertebrates) producers were sampled at each site, if the stream volume permitted.

Monitoring aquatic resources in Lime Creek downstream of the TMF will be an

important component of the aquatic effects monitoring program. For this reasonemphasis has been placed on characterizing the aquatic communities in lower Lime

Creek as thoroughly as possible.


The current water quality sampling program will continue on a monthly basis until the

spring of 2010 at most sites.

21.2.8 F R E S H W A T E R A N D M A R I N E F I S H E R I E S

O B J E C T I V E S

The proposed Kitsault Project has the potential to affect the fisheries resources in thevicinity of the marine environment of Alice Arm and the freshwater fisheries





resources of Lime and Roundy creeks. The watershed of Lime Creek is the site of the mine infrastructure and proposed open pit and the TMF. The fisheries resources

of Lime and Roundy creeks are not well known and were the focus of the 2009baseline fisheries work.

The proposed marine loadout facility lies 1 km south of the town of Kitsault. Amarine loadout facility is under design to enable barges and small vessels to bringsupplies and load mine products. The marine fisheries resources at the proposed

site are not well known and studies were conducted in the intertidal zone and shallow

sub-tidal zones for up to 1 km on either side of the facility location. Finfish, shellfish,crustaceans, algae, and invertebrate species may utilize the intertidal and shallowsub-tidal habitats for spawning, rearing, migration, and feeding.

The objectives of the 2009 fisheries baseline program were as follows:

Freshwater:

determine fish species presence and distribution within the project studyarea of Lime and Roundy creek watersheds

characterize and gather baseline data on fish population and fish habitat

quality of the project study area for the environmental impactassessment and future environmental effects monitoring

provide fish habitat assessment data on areas that will be potentially lost

due to mine infrastructure.

Marine:

determine presence and utilization of fish, algae, and marineinvertebrates within the small marine study area

characterize and gather baseline data on fish community composition

and fish habitat quality of the project study area for the environmentalimpact assessment

provide fish habitat assessment data on areas that will be potentially lost

due to infrastructure.


The 2009 freshwater fisheries baseline program focused on the Lime Creek

watershed, which will potentially be affected by mine development and operation,and one reference watershed located 2 km south (Roundy Creek). Patsy Lake,

located at the headwaters of Patsy Creek upstream of the mine site, was alsosampled. Fish habitat, fish community, and barrier surveys were undertaken.Stream habitat assessments followed protocols outlined in the Reconnaissance(1:20,000) Fish and Fish Habitat Inventory Protocol (Resource Inventory Standards

Committee [RISC], 2001), Sensitive Habitat Inventory Methods, and the BCWatershed Restoration assessment protocol (Johnston and Slaney, 1996). Fish

community surveys by electrofishing, minnow trapping, angling, and gillnetting were

conducted in July and September 2009. Other baseline information on the physical,





chemical, and biological environment in these watersheds was collected by theaquatic ecology team through sampling the water quality, sediment quality, and

primary and secondary producers.

The 2009 marine fisheries program focused mainly on sampling the marine fish

community in the vicinity of the proposed marine site. Sampling was conducted bygillnetting, long-lining, minnow trapping, crab trapping, and beach seining.


Freshwater F isher ies

In summary, Dolly Varden were the only fish species captured in the lower 2 km of Lime Creek, downstream of an 8 m waterfall barrier while juvenile Coho salmon and

Dolly Varden were captured in the lower 1 km of Roundy Creek. No fish werecaptured near the proposed mine infrastructure in the middle and upper reaches of

Lime and Patsy creeks. Crews identified impassable barriers on Lime and Roundycreeks which limit the upstream migration of anadromous fish (i.e. fish which migrate

from the marine to freshwater environment) and resident fish.

Lime Creek, the site of the proposed mine and tailing facility, has a large waterfall

(approximately 8 m high) located about 2 km upstream of the ocean within a large

canyon section. The moderate gradient canyon section also contained numerous2 to 4 m high cascades, which were also barriers to fish. Crews sampled

downstream of the barrier falls on Lime Creek and between 3 and 5 km upstream.Dolly Varden was the only fish species found downstream of the falls in Lime Creek

and were low in abundance.

Fish habitat suitable for spawning, rearing, and overwintering was of fair to goodquality in the fish bearing reaches of Lime and Roundy creeks. The habitat consisted

of pools, glides, cascades, and riffles with mainly cobble and boulder substrates, and

clear water. A mature forest of coniferous trees lined the stable, confined middle andupper reaches in both systems.

Despite extensive fish sampling in upper Lime Creek and its main tributary, Patsy

Creek, no resident fish populations were found. Patsy Lake was sampled bygillnetting and minnow trapping in July and September but no fish were captureddespite extensive effort; therefore, it was concluded to be a non-fish bearing lake.

Marine Fisher ies

In the last week of July 2009, fish sampling in the marine environment between the

town of Kitsault and Roundy Creek found a more diverse fish community than in the

freshwater environment. Species utilizing the nearshore and intertidal habitatbetween the mouth of Lime and Roundy Creek were mainly salmonids (e.g. Dolly

Varden adults and juvenile chinook salmon, Coho salmon, cutthroat trout, rainbow





trout). Adult Dolly Varden were the most abundant salmonid captured but no adultsalmon or trout were caught. Low numbers of other non-sport fish species were

caught including ronquils, gunnels, sculpins, herring, and surf perch. AdultDungeness crabs were common along the sub-tidal areas. Bivalves, other than the

abundant mussels and barnacles, were generally not present along the gravel and

rocky beaches.

Marine fish habitat and communities in Alice Arm were less diverse than marine

foreshore settings on the outer coast near Prince Rupert. The intertidal zones at

Alice Arm were either gently sloping gravel and cobble beaches or steep boulder andbedrock shores. No eelgrass, bull, or Giant kelp beds were found in the intertidal or shallow sub-tidal zones. From the highest tide level to lowest, the intertidal zones

consisted of bands of sedges and vascular plants at the high tide line followed bysparse to dense bands of rockweed (e.g. Fucus spp.), then green algae (e.g. Ulva

spp.) in the mid-intertidal and a sparse zone of short brown kelp on rock at the lowest

low tide zone.


Baseline reports summarizing the 2009 work will be prepared, followed by review of

the potential effects or impacts of the projects on fisheries resources or habitat.Based on results obtained in 2009, a second year of baseline data collection may be

carried out in 2010 to focus on fish habitat locations that may be affected by the mine

infrastructure or effluent discharge such as the fish-bearing section of lower LimeCreek. Underwater habitat assessment using video at the proposed marine

infrastructure location will also need to be completed.

21.2.9 W E T L A N D S

O B J E C T I V E S

The objectives of the wetland study are to map the locations and size of wetlands

within the study area and classifying according to provincial and federal standards.The wetland functions (ecological, hydrological, habitat, and biochemical) will also be

measured/sampled at a sample of wetlands.

S T U D Y A R E A A N D M E T H O D L O G Y

Wetlands in the ecosystem mapping study area were surveyed in September 2009.Wetland surveys concentrated on delineating wetlands on an ortho-image andcollecting wetland ecosystem data consisting of wetland vegetation community

information, wetland soil information, and wetland hydrologic feature data. Wetland

surveys were conducted following MacKenzie and Banner (2001) and MacKenzieand Moran (2004). Wetlands were classified in the field following Warner and Rubec(1997) and MacKenzie and Moran (2004).






A total of 82 wetland surveys were completed. The majority of wetlands surveyed

were isolated fens and bogs; however, the greatest area of wetlands surveyed werewetland complexes of fens and shallow open water. Wetland hydrological monitoring

was also conducted in two wetland complexes through a series of shallow static andcontinuous groundwater monitoring wells and the aquatic biology of four wetlandswas also characterized. Samples of Eriophorum angustifolium were also collected atsix sites for plant tissue metals analysis.


In order to complete the proposed study, wetland ecosystems will be mapped andincorporated into a geographic information system. The field data will be used to

identify the wetland ecosystem associations and their distribution. The hydrological,

biochemical, ecological, and habitat functions of the areas wetlands will be

determined through data analysis of hydrological, aquatic biological, plant tissuesamples, and ecosystem observations.

2 1. 2. 10 E C O S Y S T E M S A N D V E G E T A T I O N M A P P IN G

O B J E C T I V E S

Ecosystem and vegetation mapping is being conducted for the project area as part of

the TEM. TEM mapping is required for the effects assessment of the project under

the Mines Act . These products will identify areas with rare or endangeredecosystems, sensitive ecosystems, and ecosystems identified by First Nations and

local residents as important for social, cultural, or economic, reasons. This programalso identifies potential and existing rare plants and invasive plants during fieldsurveys and collects plants for metals analyses to identify baseline metals content

information. In addition, the ecosystem and vegetation mapping is used to evaluate

wildlife habitat through the production of a habitat suitability mapping product.

The objectives of the ecosystems and vegetation mapping program are to:

establish the baseline ecosystems and vegetation conditions for the project

study area

conduct field surveys to provide data to ground-truth the TEM mapping

identify rare and endangered ecosystems or plants in the study area

collect baseline information on the metals content of plants in the study area

identify invasive plants in the study area.






The ecosystems and vegetation study area (the TEM study area) covers 3,600 ha

and extends to the height of land or 1 km from the mine site. It includes all areas for the proposed mine devolvement such as the mine site facilities (pit, plant site, shop,

waste rock dumps, etc), tailing storages facilities, access roads, and the marinefacilities (wharf and camp).

TEM is currently underway for the project area. TEM is required under the Mines Act

and uses standardized methods developed for BC (BC MELP and MOF, 1998; RIC,

1998 and 2000). This mapping technique uses a combination of aerial photointerpretation and ground surveys to produce a fine-scale map of the study area withtwo mapped “layers”. The first layer identifies terrain polygons by dividing the study

area into relatively homogenous areas with common soil properties (e.g. rocky bluffs,upper, middle and lower slope positions, alluvial fans, etc.). This map is used by the

Soils and Terrain group to identify and map soil distribution. These terrain polygons

are then sub-divided based on existing vegetation communities to produce theecosystems and vegetation layer. The resulting layers can be displayed spatiallyusing GIS technology and can also be linked with other spatial data, resulting in a

useful resource management and planning tool.

Field surveys are conducted to ground-truth the TEM product using standard

methods described by BC MELP and MOF (1998). These surveys collect vegetationinformation in standardized plots, recording location, slope, aspect, and vegetation

characteristics. Surveys are conducted in conjunction with the soils and wildlife

teams.

Field surveys also record rare and endangered ecosystems and plants in the study

area. Ecosystems are listed by the BC Conservation Data Centre (BC CDC), whichrates all identified ecosystem types as to their relative abundance. Certainecosystems are rated as rare and can be put on the Red or Blue lists for BC. Plantspecies are listed provincially by the BC CDC and federally by the Committee on the

Status of Endangered Wildlife in Canada (COSEWIC). Invasive plants were also

recorded in the study area, as identified and listed by the Invasive Plant Council of BC (IPC BC).

Characterizing the metal levels in plant tissue is a requirement of the mine permit

application (BC MEMPR, 1998) and is used to guide reclamation planning and endland use objectives. Plant species targeted for collection are usually those

commonly found throughout the study area and likely to be a food source for wildlife

or people. Plant tissue metal concentrations are used for future monitoring duringmine operations, closure, and reclamation. Future metal levels in plant tissue will be

compared to baseline values in order to determine if changes in metal levels are

occurring.






A detailed review of available information including all relevant data collected during

the former Kitsault Mine operations has been conducted. Terrain and vegetationphoto classification has been carried out using black and white stereo aerial

photographs (1994) at a scale of 1:15,000. A field survey of the ecosystems andvegetation in the area was conducted in July 2009. A total of 66 sites were observedfor ecosystems and vegetation. No rare plants were found during field surveys. Theclassification of rare ecosystems is ongoing through the TEM mapping. No invasive

species were observed during field surveys, although several agronomic specieswere found in disturbed and reclaimed areas of the mine site. Fifty plant samples

were collected for heavy metals analysis, which is ongoing at ALS Laboratories in

Vancouver.


TEM mapping and its supporting fieldwork has been completed and is currently beingprocessed to ground-truth the mapping and produce a digital version capable of being used in a GIS program. The rare ecosystem and plant, and invasive plant

surveys are also completed for the TEM study area near the mine site, as is the

collection of the plant samples for metals analysis, which is ongoing at ALSLaboratories in Vancouver.

2 1. 2. 11 T E R R E S T R I A L A N D M A R IN E W I L D L I F E

O B J E C T I V E S

The 2009 baseline sampling program focused on wildlife and wildlife habitatexpected to occur within the project area. Inventories for terrestrial mammals,marine mammals, birds, and herptiles were conducted in 2009. The information and

data gathered during baseline studies will be used to highlight potential concerns of

the proposed project and will focus on species at risk, umbrella species, or species of high priority (i.e. Valued Ecosystem Components [VECs]). The specific objectives of the 2009 wildlife baseline studies were to:

review existing literature and provincial data sources to determine thewildlife information that is already known for the area

determine the presence, distribution, and suitable habitat for black-tailed

deer, moose, and mountain goat during key periods of the year;

determine the presence and suitable habitat for freshwater and marine

waterfowl and waterbirds on lakes, rivers, and along the shoreline of Alice Arm

document upland breeding bird populations





evaluate the breeding locations of amphibian species, particularly the listed

Western Toad

determine the presence and locations of raptor species in the study area,particularly the listed Northern goshawk

determine the presence and species composition of the bat population in thestudy area;

determine the species and distribution of marine mammals in Alice Arm

conduct wildlife habitat suitability assessments in conjunction with TEM toproduce Habitat Suitability Ratings for VEC species.

The results of the 2009 baseline studies will be presented in the wildlife baseline

report.


The terrestrial wildlife study area covered 180 km 2 around the Kitsault mine, an area

that is largely a brownfield site. The marine wildlife study area included the waters of Alice Arm around Kitsault, as well as Observatory Inlet and portions of Portland

Canal.

Peer-reviewed literature, government publications, and unpublished reports wereused to collect existing pertinent information regarding wildlife species present in the

study area. In addition, a number of field surveys were employed to conduct

baseline wildlife inventories. Methods were consistent with the provincial RISC. The2009 field inventory of terrestrial wildlife species included the following:

winter aerial surveys for mountain goat, moose, and Sitka black-tailed deer

summer aerial surveys for mountain goat

upland breeding bird surveys variable radius point counts (VRPC)

Northern Goshawk call playback studies (summer)

waterfowl summer (brood) surveys of freshwater wetlands

bat surveys (echolocation detection and mist net capture)

western toad breeding pond survey (summer)

ecosystem mapping and preliminary habitat ratings.

The 2009 field inventory of marine wildlife species (seabirds and marine mammals)

included the following:

marine mammal boat-based surveys in summer (June) and fall (September)

seabird ground and/or aerial surveys in winter (March), spring (May),

summer (June and August), and fall (September)





waterfowl aerial and/or ground surveys in late winter (March), early spring

staging (April), spring (May), summer (August), and fall staging (September).


The field work identified in the work plan for 2009 was completed. The following is asummary of the key findings from these baseline surveys. Detailed results of this

work will be presented in a baseline report.

An initial literature search was conducted to identify species at risk or species of concern that may be present on or near the project site. Twenty-four species at risk

were identified that may occur in the project area, including 15 bird species, threeherptiles, and six mammals.

Ungulates

Mountain goats were observed during aerial surveys conducted in both the winter and the summer. Seven mountain goats were observed in the winter, indicating that

high quality winter habitat exists within the project area. Twelve mountain goatswere observed during summer aerial surveys, with interesting observations at lower elevations than expected for this season.

A total of 27 moose were counted during winter aerial surveys, with animalsobserved at low elevations. The majority of moose observed were wintering near theKitsault River estuary.

Sitka black-tailed deer typically winter in areas under conifer forest cover and assuch pose challenges for winter aerial surveys; however, one deer was observed,and evidence (tracks) were also noted on the north side of Alice Arm during the

winter aerial surveys, indicating that Sitka black-tailed deer are present in the area.

Bats

Bats were surveyed using mist-netting (summer) and an Anabat recorder (spring and

summer). During the spring surveys, no bats were detected with the Anabat, likelydue to cold and windy weather conditions. During the summer bat surveys, bat calls

were identified with the Anabat; however, no bats were captured in the mist net.Myotis spp. (probably little brown myotis, M. lucifugus), was the most frequently

detected species based on the Anabat sonograms. Observers also potentially

recorded a long-eared myotis and a third bat species with the Anabat. Ongoinganalysis of the Anabat sonograms will confirm the presence of these species.

Marine Mammals

Marine mammal boat-based surveys were conducted in June and September. Fivespecies were observed, including harbour seal (including pups), Steller sea lion,





Dall’s porpoise, harbour porpoise, and humpback whale. Harbour porpoise andSteller sea lion are listed as Special Concern under the Species at Risk Act , and are

blue-listed in BC, and humpback whales are listed as Threatened under the Species

at Risk Act , and are blue-listed in BC. In the waters around the proposed port,

harbour seals and harbour porpoises were observed.

Birds

Water-dependent birds (i.e. waterfowl, shorebirds, and seabirds) were surveyed

during the winter, spring, summer, and fall with 19, 18, 21, and 14 species observedduring the surveys, respectively. The majority of waterfowl observations during thewinter survey, and the spring and fall staging surveys, were within the Kitsault River

estuary, suggesting this is an important staging area for some waterfowl species,

such as mallard, Canada geese, American wigeon, northern pintail, Barrowsgoldeneye, lesser scaup, and sandhill crane (blue-listed). Great blue heron (blue-

listed) were also observed feeding along the estuary. During summer brood surveys,

eight water-dependent species broods were identified associated with fresh water lakes and ponds. Species with broods included Pacific loon, Barrows goldeneye,

Canada goose, and red-throated loon. In the marine environment no waterfowl

broods were detected and few waterfowl observations were made during summer.

Seabird species were observed in the inlet during winter, spring, summer, and fall

surveys, including a variety of gulls, grebes, loons, bald eagle, and a few shorebirds.

The most significant observation was of the provincially red-listed and federallyThreatened marbled murrelet, observed during all four surveys. Over 150 marbled

murrelets were observed along Alice Arm and in Observatory Inlet in May and June,including breeding pairs. The consistent presence of marbled murrelets in Alice Arm

suggests that nesting may occur in suitable forest in the region.

Upland breeding birds were surveyed in June, using the VRPC methodology. Forty-seven species of breeding birds were identified; the olive-sided fly catcher was the

only species observed of federal or provincial concern. The only raptor species

incidentally observed during surveys were bald eagles. Call-playback surveys for Northern Goshawk were conducted, however, no goshawks were observed.

Herpt i les

The study area was surveyed for the presence of western toad breeding ponds inearly August. No ponds were identified with tadpoles or toadlets, though some sites

were potentially suitable. Ten adult western toads were observed near roads.Considering these observations, it is likely that breeding occurs within identifiedwetlands in the study area.





Inc idental Wi ldl i fe

Incidental observations of other mammals and/or mammal signs include wolf, river otter, red squirrel, red fox, porcupine, mink, black bear, American marten, grizzlybear, lynx, cougar, and fisher. In addition, barn swallows (provincially blue-listed)

were observed nesting at the Avanti camp.


The work remaining includes additional field work to accommodate data collection

from a larger regional study area, development of baseline reports, and preparationof an environmental impact assessment and wildlife monitoring program. Thecomprehensive scope of studies is viewed as essential as this area is within the

Nisga'a Final Agreement (NFA) Wildlife Area. Additional field work to be completedduring 2010 would need to include the following:

mountain goat surveys of additional survey units to be conducted in winter (March) and summer (July), which focus on identifying mountain goatpopulation and distribution over a regional study area

habitat suitability mapping/modelling and evaluation in the field (June) for species such as moose, mountain goat, grizzly bear, Sitka black-tailed deer,

marten, and marbled murrelet

marbled murrelet nesting survey in a regional study area (spring)

riverine bird survey, specifically harlequin ducks in a regional study area

(spring)

winter furbearer snow-track survey to characterize habitat, specifically for

wolverine and fisher (March).

In addition, incidental wildlife observations will continue to be recorded, along with

evaluation of the literature, collection of traditional knowledge (TK), consultation withknowledgeable individuals, etc.

2 1. 2. 12 V I S U A L Q U A L I T Y

O B J E C T I V E S

A visual quality assessment is being undertaken for the project as part of the effects

assessment. In perspective view, visual impact assessments estimate the potentialvisual effect of proposed operations on the scenic landscape. The objectives of thevisual quality study are as follows:

1. conduct a visual landscape inventory; delineate and record areas that are

“visually sensitive” (i.e. communities, special features, viewpoints, parks,recreation sites)





2. conduct fieldwork; gain an on-the-ground familiarity with the area from avisual perspective to locate the pre-selected viewpoints and to gather data

3. develop and prepare visual simulations; perform a GIS analysis usingtopography and available forest cover data

4. assess the visual simulations and GIS analysis and prepare visual impactassessment report.

The first two steps are completed during baseline studies.


The visual quality study area was determined based on the desk-top visuallandscape inventory conducted and included Alice Arm, the town of Kitsault, and themain access road to Kitsault.

The steps taken follow procedures found in the Visual Impact Assessment Guidebook (Second Edition) (MoF, 2001).


A visual landscape inventory was completed using GIS, which delineated andrecorded potentially visually sensitive areas. Other sources of information used in

planning and preparation included recreational features inventory data from theIntegrated Land and Management Bureau as well as the North Coast Land and

Resource Management Plan (LRMP).

Wildlife and TEM field crews collected visual quality baseline information while doingtheir discipline work. The baseline information consisted of taking photographs in one

or more directions looking towards the proposed mine site, recording the GPS point

for the photograph location and taking a compass bearing to document the directionof each photograph. To date, photographs have been obtained from variouslocations in Alice Arm, from locations on shore near the proposed wharf site, and

from the road leading toward the town site.

W O R K R E M A I N I N G F O R T H E P E R M I T A P P L I C A T I O N

Photographs from various locations in the town site should be obtained to complete

visual quality baseline studies. Once all baseline photographs have been collected,

the baseline report will be written.





2 1. 2. 13 AR C H A E O L O G Y

O B J E C T I V E S

In BC, the Heritage Conservation Act (HCA) automatically protects all archaeologicalsites predating 1846 AD on Crown and private land. Sites such as burials and

Aboriginal rock art sites are protected regardless of age. The proposed Kitsault

Project has the potential to impact archaeological sites that may be present within or adjacent to the proposed project footprint.

An Archaeological Impact Assessment (AIA) was conducted for the Kitsault Project

under HCA Heritage Inspection Permit 2009-0085, issued by the BC ArchaeologyBranch (Ministry of Tourism Culture and the Arts). The objectives of the AIA were to:

identify and evaluate any archaeological sites located within and adjacent to

the proposed development’s footprint

identify and assess possible impacts of the proposed developments on anyidentified archaeological sites

provide recommendations regarding the need for and appropriate scope of further archaeological studies prior to development

recommend viable alternatives for managing adverse impacts, if any areidentified.


The AIA was conducted in accordance with the Archaeological Impact Assessment

Guidelines and with the methodology outlined in the application for Permit 2000-0085. The general stages of the AIA investigation are described below.

Firs t Nat ions Communicat ion

In March 2009, a copy of the permit application was forwarded by the Archaeology

Branch to the Nisga’a Lisims Government (NLG) for review and comment. Prior tofieldwork the NLG was invited to select individuals to participate in the field

assessment. When complete, a copy of the final AIA permit report will be sent to theNLG.

Background Research

A review of published information for the project area and surrounding region wasconducted prior to the fieldwork. This included a review of ethnographic, historic,

archaeological, and environmental literature, and a search of the BC ArchaeologicalSite Register using the Remote Access to Archaeological Data (RAAD) application.





Field Assessment

The field methods employed during the AIA were consistent with those outlined in theapplication for Permit 2009-0085. Fieldwork took place on May 12 and September 14 to 25, 2009. Nisga’a members participated on all days of fieldwork.


The AIA was conducted for the project footprint. No archaeological sites protected

by the HCA were identified within the project footprint. Evidence of 20th century

mineral exploration (e.g. blazed trees, cut stumps) and metal wire was identifiedduring the AIA; however, these are not protected by the HCA.


A TK study will be conducted as part of the EA baseline studies. The results of the

TK study will need to be reviewed and incorporated into the Archaeology baselinestudy where appropriate. Depending on the type of TK information that is provided,

additional archaeological fieldwork may be required. Additionally, in the event that

there are changes to the project footprint, further archaeological assessments maybe recommended.

Further work to develop an Archaeological Chance Find Procedure for the project will

be required. This procedure will address the unlikely possibility of chancearchaeological finds being discovered during construction activities.

2 1. 2. 14 H U M A N H EALTH

O B J E C T I V E S

The proposed Kitsault Project has the potential to affect the health of residents andland users in the vicinity of the project. Local residents and land users include

Nigsa’a First Nations, guide outfitters, trappers, and recreational land users such as

campers, fishers, and hunters. Activities related to the Kitsault Project have thepotential to affect water and soil quality, which in turn affect terrestrial animals,freshwater and marine fish, and plants in the vicinity of the proposed Project. Land

users who harvest these plants and animals as country foods may subsequently be

affected if there are any effects to the soil and water quality. However, previous or

currently existing human activities in the area may also have had an effect on thesefactors. The human health baseline study establishes the current quality of countryfoods in the project area.

The objectives of the human health baseline program are to determine the current

quality of country foods, defined by the total metal concentration in the tissues of meat, fish, and plants that are harvested from the area for food.






The study area for human health effects was based on the study boundaries used in

the ecosystem mapping studies that establish potential areas which may be affectedfrom the proposed project.

The methodology for assessing baseline country food quality is based on Health

Canada (2004).


The country foods baseline is a multi-year study which requires water, soil and tissue

(i.e. vegetation, fish, and other wildlife) samples in order to assess the current qualityof country foods. The field studies that collect these samples are only in their first

year and there is insufficient data to establish the baseline effects to human health.


The baseline report for country foods is primarily a desk-based study relying onsampling data collected from various field studies. A baseline report summarizing

the 2009 field relating to human health will be prepared. A second year of fieldstudies may be carried out in 2010 that will continue to collect sample data for soil,

water, and tissues.

For country foods, the 2009 baseline studies for vegetation, wildlife and fisheries willbe evaluated to select relevant country food species. The tissues of these species

will be modelled or collected for laboratory analysis and incorporated into the country

foods baseline for 2010.

2 1 . 3 S O C I O E C O N O M I C S T U D I E S

21.3.1 SO C I O E C O N O M I C

O B J E C T I V E S

The proposed Kitsault Project has the potential to affect the social, economic,

cultural, and health environment of communities within proximity to the project.

Effects may be direct or indirect, through related project components such astransport and shipping. The objectives of the human health baseline program

include:

identifying the socioeconomic regional perimeters that have the potential tobe affected by the project





profiling of the past and current socioeconomic environment and context

within which the project is proposed.

Study Area and Methodology

The socioeconomic study area of the project commences at a broad provincial level,with a focus on northwest BC, followed by the Kitimat-Stikine Regional District

(KSRD) and Nisga’a Nation communities. Terrace and Prince Rupert are included inthe study area due to their role as regional service centres that have the potential to

experience indirect socioeconomic effects from the project.

The study is informed by an initial scoping exercise to inform and frame socio-economic study scope and main areas of focus. Desk-based research is then

conducted on pre-existing literature, reports, and publicly available information

databases concerning:

local and regional population and demographics employment and unemployment patterns

education and training achievements and opportunities

business and economic development activities and plans

social and health issues

social and health services

cultural revitalization initiatives

current community and regional infrastructure and services.

Subsequent to desk-based research, a gap analysis is conducted to identify areas

necessitating further research. The results of the gap analysis inform and frame thefield-study program which aims to supplement and confirm desk-based research

findings. This entails conducting face to face interviews with individuals or groupsand/or telephone interviews.

Economic modelling is conducted to help determine the project’s direct, indirect, and

induced economic effects through the implementation of the BC Stats Input-OutputModel.


The socioeconomic baseline study is typically a two-year study, which requires thatdesk-based research be built on by field studies by way of local community-based

research. The results of the baseline study assist to identify valued socioeconomic

components (VSECs). A socioeconomic effects assessment is then conducted topredict the effects of the project on the social or human environmental surroundingthe project.






Issues scoping and preliminary socioeconomic desk-based research has been

conducted whereby general information has been collected to create an overallprofile of the socioeconomic study area and its communities.

Additional desk-based research is required. Upon completion of this additional work,

a baseline report summarizing the 2009 key findings relating to socioeconomics willbe prepared.

Field studies will be required in 2010. This will require the cooperation and

collaboration of local representatives, especially in the Nisga’a Nation communities.

The results of field studies will be integrated into the desk-based research, compiled,and finalized before the effects assessment process may take place.

21.3.2 N O N - T R A D I T I O N A L L A ND A ND R E S O U R C E U SE

O B J E C T I V E S

The proposed Kitsault Project has the potential to affect the existing land uses,tenures, asserted rights, and other interests. This includes the Nisga’a Nation and its

citizens. The objectives of the land and resource use baseline program include:

identifying the existing land uses, tenures, and asserted rights intersectingthe project study area

identifying other interests such parks, protected areas, and recreational

interests

describing LRMPs and objectives encompassing the project study area.


The land use study area is modelled after the terrestrial study area and takes into

consideration the species with the largest habitat range, unique ecosystems, andnatural landform barriers. Many of the land uses are related to wildlife (e.g. hunting

and trapping) and, as such, adopting the same study area ensures consistency and

the ability to compare across the baseline studies of different disciplines.

An initial step to inform and frame the land and resource baseline study andassessment will be to undertake an issues scoping exercise. A review of other

developments in the area, other relevant literature, the media, findings from the initialconsultation programme, existing baseline information, reports, and proposed projectcomponents will allow for an initial issues scoping exercise to frame the development

of the baseline studies and selection of VECs. This exercise will also identify key





stakeholders and users. Relevant information will be shared with the consultationcomponent.

An integrative approach including desk-based research and field studies, in addition

to close integration with the GIS database, will be central to outlining project scope

and approach to land and resource research. Land and resource studies will alsowork closely with other disciplines, including wildlife, terrestrial ecology, andtraditional use.

Information on land uses and tenures will be obtained from appropriate local,

regional, provincial government and NLG agencies. Relevant management plans for the study area will also be consulted, including the North Coast LRMP and the South

Nass Sustainable Resource Management Plan (SRMP).


The land and resource use baseline study is typically a two-year study, whichrequires that desk-based research be build on by field studies. The results of the

land and resource study assist to identify VECs. A land and resource use effectsassessment is then conducted to predict the effects of the project on the land usesand existing tenures intersecting and in proximity to the proposed project.


Issues scoping and preliminary land and resource desk-based research has been

conducted to identify key land users, tenure holders, and regional land use

management plans. Additional desk-based research is required. Upon completion

of this additional work, a baseline report summarizing the 2009 key findings relatingto land and resource use will be prepared.

Field studies will be required in 2010. This will involve the cooperation andcollaboration of local representatives, especially in the Nisga’a Nation communities.

The results of field studies will be integrated into the desk-based research, compiled,and finalized before the effects assessment process may take place.

21.3.3 N I S G A ’ A U S E A N D T R A D I T I O N A L K N O W L E D G E

O B J E C T I V E

The first objective of this study is to fulfil Chapter 10, Paragraph 8 of the NFA, inparticular the effects of the proposed project on Nisga'a rights, interests, and title.

Recognizing that citizens of the Nisga'a Nation who potentially use the land andwaters in and around the proposed project area have a wealth of knowledge andunderstanding that spans many generations. The second objective of the Nisga'a

Use and Knowledge study is to elucidate Nisga'a technical and academic informationabout the proposed project area and surrounding areas.





The study recognizes the need for specific information from Nisga'a citizens aboutwildlife, fish, and vegetation presence, abundance, and trends as well as the uses of

country foods in the proposed project study area. As a result, information from thisstudy from trappers, hunters, and other land users will supplement wildlife,

archaeology, vegetation, fisheries, human health, and other studies. The information

collected will be used to provide insight regarding local ecology and help identifywhich species, foods, and habitats are present and important in the area and howthey might be affected by the proposed project.


The proposed Kitsault project location falls within the Nass Area and Nass Wildlife Area, as defined by the NFA. These areas entitle citizens of the Nisga'a Nation to

non-exclusive rights to harvest fish, aquatic plants, wildlife, and migratory birds. The

Nisga'a Use and Knowledge study area will be determined in collaboration with NLG.

Rescan can provide varying degrees of technical support for a Nisga'a-led study. Ata minimum, Rescan will coordinate closely with the NLG with respect to reporting

and incorporation of Nisga'a Use and Knowledge into the environmental assessmentapplication. The following draft methodology framework is proposed to acquire

thorough and comprehensive information regarding Nisga'a Use and Knowledge, butwill be modified following direction from the NLG:

Background and ethnographic research: A literature review of pre-

existing studies or information which could help inform this study (e.g. theNFA, the Nisga'a Land Use Plan, and previous Nisga'a academic and other research).

Request for information from previous Nisga'a studies: As part of theNFA, information about hunting and fishing is collected by the NLG.Discussions with the NLG are hoped to provide access to relevant

information and previously conducted Nisga'a studies related to the studyarea.

Identification of participants: In association with the NLG and Nisga'a

villages, identification of, and agreement about, who holds relevant

knowledge and information on the study area would be sought. Participantsmay consist of individuals and families who have knowledge of the areathrough previous or current use by themselves or their family (e.g. Elders

and other users).

Information meeting: It essential that Nisga'a communities and knowledgeholders be informed about the study, why it is being conducted and how the

information will be used. An informal community meeting focusing on Eldersand other knowledge holders can serve this purpose. This meeting wouldalso allow knowledge holders to self-identify for participation in the study.

Interview consent and set up: Contact would be made with each identified

knowledge holder to seek consent to interview and record their knowledge.





Conducting interviews: The interview process could be guided using a

number of tools, including:

an introduction sheet to explain the process

a series of interview questions divided into categories such as personal

history, landscape and resources, climate, wildlife, fish, plants, traditionaland current use

maps of the study area to record sites and geographical information

photo books of common plants and fish.

Interviews would focus on the study area, particularly the areas that aremost relevant to the knowledge holder. Interviews would be tape-recorded

where permitted by the knowledge holder.

Site Visits/Ground-truthing: There may be an opportunity to conduct part

of the interviews and transfer of knowledge during a site visit of the area.This would allow for ground-truthing of information and can be a more useful

and realistic experience for knowledge holders. Transcription and verification of interviews: Interviews and mapping

information would be transcribed then returned for verification by eachparticipant. A copy would be made available to the knowledge holder and

the NLG. Study findings would be presented and verification sought duringworkshops with participants.

EA Integration: Work would be done in collaboration with Rescan

biophysical and social science team leads to incorporate Nisga'a informationthroughout the application, where possible and provided, and subject toconfidentiality provisions specified by the NLG. Nisga'a information would

be disseminated in a seminar-style format, allowing for an in-depth

discussion of findings and their relevance to discipline reporting.


Discussions are currently underway between Avanti and the NLG with respect to

NLG approval for and launch of this study. Background and ethnographic researchhas been conducted regarding the Nisga'a Nation, incorporating publicly-available

information on Nisga'a rights and title, the NFA, the Nisga'a cultural context, Nisga'ahistorical land and resource use (not site-specific), regional place names, and

Nisga'a economic resource use.


Launch and advancement of this study is pending approval by the NLG to undertakethe study, appointment by the NLG of a working contact, and direction from the NLG

with respect to study design and methodology. A NLG review of the backgroundresearch would also be requested.





21.3.4 C O M M U N I T Y E N G A G E M E N T A N D C O N S U L T A T I O N R E Q U I R E M E N T S

The Kitsault Project environmental assessment will be required to develop a

consultation program guided by the Northwest Regional Mine Development ReviewCommittee, which will be consistent with the guidelines derived from the

Environmental Assessment Act , the Act’s Public Consultation Policy Regulation(2002); the Provincial Policy for Consultation with First Nations (2002), and theSupplementary Guide to Proponents: BC Environmental Assessment Process.Consultations will be required with the NLG, communities, and public during the pre-

application and application review phases.

Consultation with the Nisga'a Nation is required as a provision of the NFA, one of Canada’s first modern-day treaties. Nisga'a interests will be reflected in agreements

to address and/or accommodate Nisga'a Nation issues, values, and concerns. Theprovisions of the NFA and, specifically, as defined in Chapter 2 (General Provisions)

and Chapter 10, Section 8 (Environmental Assessment and Protection) along with

the Provincial Policy for Consultation with First Nations (2002), will be followed tofulfill this requirement. The Public Consultation Policy Regulation (2002) is used withrespect to public consultation. This sets out guidelines that would apply to Avanti’s

consultation program, public notice, and public comment periods.

21.3.5 C O N S U L T A T I O N W I T H N ISGA ' A N A TI ON A ND F I R S T N A TI O N S

The Kitsault mine area falls outside of the lands owned by the NLG under the terms

of the 1998 NFA, which became effective on May 11, 2000. However, it is within theNass Wildlife Area (as defined in and governed by the Treaty), and as such it is

subject to the traditional hunting and fishing rights of the Nisga'a which, among other interests, have been recognized and protected under the terms of the NFA.

The NFA states that if the proposed mining activities at the Kitsault Mine area may

reasonably be expected to have adverse environmental impacts on the Nisga'apeople, Nisga'a lands, or their NFA interests, then a specific process for consultationby the Federal or Provincial government, as the case may be, is set out in the NFA.

In addition, when an environmental assessment is carried out under Provincial or

Federal law, the NFA grants specific rights to the Nisga'a Nation to participate in theprocess. The NFA also enumerates various requirements that are additional to the

requirements under environmental assessment legislation.

Other than the Nisga'a Nation, Avanti has not been informed by the BC government

agencies of additional consultation requirements pertaining to other First Nations thatmay have aboriginal rights, interests, or claims relevant to the proposed miningactivities at the Kitsault Mine area. Final confirmation as the project moves forward

will be obtained from BC Ministry of Aboriginal Relations and Reconciliations’ office

responsible for treaty negotiations.





2 2 . 0 M A R K E T I N G A N D C O N T R A C T S

The following industry summary is from CPM’s “Molybdenum Market Outlook”, April

2009, and CPM’s “Molybdenum Market Update”, September 2009.

2 2 . 1 I N T R O D U C T I O N

Structural shifts in the supply and demand of molybdenum brought life back into themarket in 2002 after many years of lacklustre price performance. With expanded

end uses for molybdenum and strong demand from emerging markets, the growth insupplies was unable keep pace with demand. The market moved into a deficit for six

years beginning in 2002 and shortfalls in supply drove prices sharply higher. By theend of 2004 prices were up nearly 11-fold from a low of US$2.33 per pound in March2002. Monthly prices then averaged roughly $30 from 2005 through the third quarter of 2008, with a high of $38.32 being touched in June 2005.

Molybdenum prices quickly changed course in late September 2008 as steel demandcollapsed and the world economic environment turned decidedly negative. Theglobal financial crisis and liquidity crunch led to distressed selling by metals traders in

need of cash, inventory liquidation by consumers, and a fire sale at some Chineseproducers. By early 2009, molybdenum prices had retraced practically all of their

gains made since 2004, touching a trough of $7.70 in April. Molybdenum prices thenmade a strong recovery in August 2009, after edging higher throughout the second

quarter of 2009. The abrupt and large slashes in global steel production kept steelinventories contained. As a result, molybdenum prices bounced back above $18 assteel fabricators tried to source supplies, which had been curtailed since the severe

price correction in the fourth quarter of 2008. Molybdenum supplies from China,

largely a marginal producer, were immensely affected by the price correction and inJanuary 2009 China became a net importer of molybdenum. A surge in Chinesemolybdenum imports to meet rapid inventory builds by ferromolybdenum producers

and steel fabricators is mostly complete, molybdenum prices declined back to

roughly $11 in November. The build-up of inventories in China may keepmolybdenum prices relatively subdued until 2010.





Figure 22.1 Molybdenum Prices – Monthly, through November 2009

0

5

10

15

20

25

30

35

40

45

0

5

10

15

20

25

30

35

40

45

1992 1994 1996 1998 2000 2002 2004 2006 2008

US$/Lb.US$/Lb.

Source: American Metals Market Prices (1992-2005) and Metals Week Mean Monthly Oxide Price

(2006-November 2009)

While molybdenum prices seem likely to decline in the near term as this rapid

bounce back in prices may be overdone, CPM sees the molybdenum price poised tostrengthen over the next several years as a period of even tighter supply and

demand conditions is forecast to unfold. This follows a long history of lacklustre price

performance, with prices averaging $3.20 between 1950 through 2002.Underdeveloped end-uses of molybdenum-bearing products coupled with reasonablyconsistent production from by-product producers, whose operations are largely

insensitive to the fundamentals of the molybdenum market, can be attributed to these

historically lower prices. These trends are changing, however, ushering in a newperiod in which demand is strong and growth in supply is limited. CPM concludes

that molybdenum prices have the capacity to rise and stay high for an extendedperiod due to the following factors:

The total contribution of molybdenum supplies from existing by-product

producers is declining. Even when incorporating additional supplies fromnew molybdenum recovery circuits at existing copper mines withmolybdenum mineralization as well as from fresh by-product output from

new copper/molybdenum mines, CPM forecasts a supply shortage re-

emerging in 2011 and 2012. The global credit crisis has created an even more competitive, expensive,

and challenging environment to bring on new supply.

Demand has been not only growing in its principal end uses, but demand for molybdenum has been increasing as the energy, transportation, and

construction industries have been seeking to utilize molybdenum’s robust

alloying properties.





2 2 . 2 M O L Y B D E N U M M A R K E T O V E R V I E W

Molybdenum is not yet traded on any commodities exchanges. Typically prices are

determined by negotiation between producers, trading houses, and end users withsupply and demand fundamentals in the background. Molybdenum is traded invarious forms, including raw molybdenum concentrates, molybdenum oxide,

ferromolybdenum, and molybdenum powders. The London Metals Exchange (LME)

is scheduled to introduce physically deliverable molybdenum and cobalt futurescontracts in the first quarter of 2010. Institutional and individual investors, along with

proprietary traders, investment banks, and brokerage companies have showninterest in molybdenum fundamentals in recent years. With the launch of the LMEmolybdenum contract in 2010, market fundamentals may become more transparent.

In 2008, over three-fourths of global molybdenum production was sourced fromChina (37%), the United States (25%), and Chile (16%). Output from Peru (8%),

Canada (4%), Russia (2%), and other countries (8%) accounted for the remainder.

As for global reserves, roughly 72% of the molybdenum’s global reserve base is inChina and the United States, according to United States Geological Survey (USGS)estimates.

Figure 22.2 World Molybdenum Reserves, 2008

US,31%

Canada,5%

Chile,

13%

China,

38%

F.S.U.,

8%

Other,

4%





Figure 22.3 World Molybdenum Reserve Base, 2008

US,28%

Canada,

5%

Chile,

13%China,

44%

F.S.U.,

7%

Other,

3%

Table 22.1 World Molybdenum Reserves and Reserve Base, 2008 (Billion

Pounds)

Reserves1

% of

World Total Rese rve Base2

% of

World Total

China 7.3 38.4% 18.3 96.5%

United S tates 6.0 31.4% 11.9 62.8%

Chile 2.4 12.8% 5.5 29.1%

Canada 1.0 5.2% 2.0 10.6%

Armenia 0.4 2.3% 0.9 4.7%

Russiae 0.5 2.8% 0.8 4.2%

Peru 0.3 1.6% 0.5 2.7%

Mexico 0.3 1.6% 0.5 2.7%

Kazakhstan 0.3 1.5% 0.4 2.3%

Kyrgyzstan 0.2 1.2% 0.4 2.1%

Uzbekistane 0.1 0.7% 0.3 1.7%

Iran 0.1 0.6% 0.3 1.6%

Mongolia 0.1 0.3% 0.1 0.6% ____ ____

World Total 19.0 41.9

- T he part of t he reserve base which could be economically extracted or

pr oduced at the tim e of determ inat ion .

- The measured and indicated resource from which reserves are estimated.

Note: Data may not sum to totals as figures are rounded.

e- Estimated

Source: USGS





2 2 . 3 M I N E P R O D U C T I O N

The molybdenum market is characterized by two distinct sources of mine supply,

which in the long-run adhere to different market fundamentals. The majority of theworld’s molybdenum supply comes from copper/molybdenum deposits located in theUnited States, Chile, and Peru, where molybdenum is produced as a by-product of

copper production. Primary molybdenum deposits – typically found in China, the

United States, and Canada – account for the largest portion of the remaining totalglobal production.

In the short-run, copper producers can extract high grade molybdenum in copper ore

bodies or add molybdenum recovery circuits, if molybdenum prices are high enoughto justify the expense. This type of activity was reflected in the ratio of by-product

production to total global output, which quickly increased to over 65% in 2005, upfrom roughly 54% in 2003. Over this three-year period, by-product output grew at a

CAGR of 21.7%. However, high grading is not sustainable throughout the mine life

of a copper mine. Molybdenum output then fell at copper/molybdenum producers,declining by an annual rate of 5.1% between 2006 and 2008. Similarly, by-productproducers also can alter their mining sequence to focus on other by-products such as

gold. Following a decline in the molybdenum prices, in second quarter of 2009Kennecott Utah Copper shifted production at its Bingham Canyon mine toward a bias

for copper and gold. In the long run, however, a molybdenum by-product producer’s

output is mainly influenced by copper prices. Beginning in 2009 by-product supply isexpected to record positive growth rates until 2018, but at a more sustainable CAGR

of 5.4%. This growth rate does not make provisions for supply disruptions, however.

Primary molybdenum producers, on the other hand, have historically acted as swingproducers because operating costs per pound of molybdenum have been higher for

these companies compared to by-product producers. Depending on the price of molybdenum, primary producers have brought their operations on and off line. Now

with comparably higher operating costs at the majority of small- and medium-scale

Chinese operations, many of the marginal cost producers are now located in China.

Two very significant structural shifts altering the origination of copper supplies also

are limiting the amount of molybdenum available from by-product production. Over

the next decade or so, primary producers (on average) are expected to account for alarger portion of total molybdenum production than in the past twenty years. Thecontinued use of the leach-solvent extraction-electrowinning (SX/EW) process limits

the potential molybdenum recovery from copper projects, as the SX/EW process

does not produce any molybdenum as a by-product. In addition, over the next fewyears, some of the largest new copper mines are expected to come on stream in the

Democratic Republic of the Congo and in Zambia. These are copper/cobalt depositsthat do not have recoverable mineralization of molybdenum.





Figure 22.4 Annual Mine Production of Molybdenum – Projected through 2018p

0

100

200

300

400

500

600

700

800

900

0

100

200

300

400

500

600

700

800

900

1 98 5 1 988 19 91 1 99 4 1 997 20 00 2 00 3 2 006 20 09p 2 01 2p 2 015 p 20 18p

Primary

By-Product

MillionLbs. Million Lbs.

Actual Productio n Projected Prod uction

Other

Source: CPM’s “Molybdenum Market Update”, September 2009.

2 2 . 4 O P E R A T I N G C O S T S

Western World average pro-rata operating costs, weighted for production (excluding

co-products such as copper, and by-products such as gold and silver), for both by-product and primary molybdenum producers increased nearly four-fold from

approximately $2.15 in 2000 to roughly $8.28 per pound in 2008. In China, 2008

operating costs for the high cost producers are estimated to have ranged from $10 to$13.50 per pound. Going forward, the growth in demand is expected to continue to

justify these higher costs.

2 2 . 5 S E C O N D A R Y S U P P L I E S

Molybdenum can be recycled from both steel and spent catalysts. Usingmolybdenum scrap versus ‘new’ molybdenum is not always cost advantageous.

Historically, scrap has not been consistently cheap enough or abundant enough

relative to mined molybdenum to spur a significant shift away from primary material.

Moreover, the quantity of molybdenum employed in steel is increasing and therecycling rates in developing countries are substantially lower than in the OECD. Assuch, secondary supplies are not expected to threaten mine supplies over the next

10 years although recycling may increase with the trends in environmental

sustainability.





2 2 . 6 C P M ’ S S U P P L Y O U T L O O K

The price correction in late 2008 led to a remarkably swift supply reaction as

molybdenum producers calibrated supply with poor demand to prevent significantoversupply conditions. In the past, primary molybdenum producers continuedoperations for extended periods of time even when molybdenum prices met or fell

below their costs of production. Primary, and to some extent by-product producers,

have been more attuned to the fundamentals in the molybdenum market during thiscycle. Supplies from China, a major swing producer, are expected decline as much

as 18% year-on-year in 2009. Delays in bringing new projects online, resulting fromthe current economic downturn and credit crisis, is likely to continue to have apronounced effect on the market over the next 10 years. The partial recovery in

molybdenum prices in 2009 is expected to spur some capacity restarts in 2010.However, growth in supply may be limited in 2011 and 2012. With broad recoveryforecast in global economic activity, the molybdenum market is forecast to return to a

deficit over that two year period.

The growth in output from primary producers is projected to increase in 2013 and2014, closing the supply gap. Given CPM’s demand projections, mine production is

forecast to narrowly meet demand through 2017. Additional supplies from recycledcatalysts should move the market into a slightly higher surplus over this period.

As of 2009, there was sufficient molybdenum roasting capacity globally. With China

Molybdenum’s recent addition to global roasting capacity, currently there areadequate supplies of global roasting capacity through 2017 to roast CPM’s forecast

molybdenum mine supply.

2 2 . 7 M O L Y B D E N U M D E M A N D

Molybdenum is a resilient metal used in a wide range of industrial applications. Itsproperties are instrumental additions to many metallurgical and non-metallurgical

products. Currently, roughly 70% of molybdenum is used by the steel industry withthe combined share of metallurgical uses representing nearly 87% of global

consumption. Molybdenum is commonly added to specialty and stainless steels

because of its effectiveness as a hardening agent, as well as for its strength,toughness, and corrosion resistance. In inhospitable environments such as extremetemperatures, deepwater, and other locations exposed to corrosive elements,

molybdenum-bearing products are employed to ensure the necessary corrosion

resistance, strength, hardness, and versatility. Molybdenum’s applications rangefrom oil and gas pipelines and offshore infrastructure to industrial plants and

automotive, ship, and aircraft components. The remainder of demand, roughly 13%,is attributed to specialty chemicals, such as specialty lubricants and pigments as well

as catalysts for petroleum refining and chemical processing. In line withmolybdenum’s industrial base, demand has increased above the historical trend over

the past decade. In periods of economic weakness, demand for molybdenum also

has been relatively resilient. Molybdenum consumption has been robust due to the





rise in global infrastructure spending, the urbanization of emerging economies, aswell as an expanding list of newer applications and the limited room for substitution.

Figure 22.5 Molybdenum End Uses, 2008

Stainless

26%

Full alloy

16%Tool

10%

HSLA

10%

Carbon

8%

Cast iron3%

Mo Metal &

Alloys

7%

Superalloys

6%

Catalysts

9%

Lubricants& Pigments

4%

Other

chemical

1%


Table 22.2 Growth in Molybdenum Demand by End Use – Compound Annual

Growth Rates for Selected Years

Steel

(%)

Other

Metallurgical

(%)

Non

Metallurgical

(%)

Total

(%)

2001 – 2007 10.0 7.9 6.9 7.0

2007 – 2009p -5.3 -3.7 0.02 -4.3

2009p – 2018p 6.1 4.6 5.2 5.7






Table 22.3 Molybdenum End-use Profiles

Applications

Steel

Full Alloy Construction/au tomotive industries, shipbuilding, heavy machinery, offshore pipelines

Stainless Biofuel tanks, flue gas , des ulphurization in oil and coal units, desa lination plants

Carbon Construction and transpo rtation, tunne ls, food storag e, communication

Tool Manufactu re of tools and the cutting or sh aping of power machinery

HSLA Oil and gas pipelines, cons truction and automotive industries, bridges

Other Metallurgical

Superalloys Superchargers, aircraft turbine engines, gas turbines, chemical and petroleum plants

Cast Iron Diesel engine motor blocks and cylinder heads , mining, milling, and crushing

Mo Metal & All oys Auto parts , lamp filaments , glass manufacturing, heat shields, optical coatings

Non-Metallurgical

Catalysts Petroleumh ydroproces sing and hydrod esulfurization

Lubricants High performance base oils, greases , syn thetic fluids, bond coatings, friction product s

Pigments Paints, inks, plastic and rubber products , and ceramicsOther Chemical Smoke sup press ants , PVC cabling, metal-based smoke sup press ants

Category






22.7.1 M O L Y B D E N U M C O N S U M P T I O N B Y T H E ST E E L I N D U S T R Y

The growth in global crude steel production has increased considerably since 2000.

Figures from the World Steel Association show that output rose at an annual average

rate of 6.0% between 2000 and 2008 after registering growth of under 1% during the1970s and 1980s. The relatively recent increase in production rates has been due to

urbanization and industrialization trends in emerging markets, significant growth insteel production capacity in China, and consecutive years of increased spending onboth private and public infrastructure projects worldwide. Along with these increases

in steel volumes, steel-based molybdenum demand grew at an average rate of 6.4%

per annum from 2000 through 2008. Growth in molybdenum demand has surpassedthat of steel production due to the gradual shift toward higher quality, specialty steels

in addition to increases in its intensity of use in major developing economies. InChina, the percentage of alloyed steel in total crude steel production is likely to reachabove 10% during the forecast period from roughly 5% in 2008 as a result of

investments in modern equipment, market demands, and industry consolidation in

the country.

Figure 22.6 Share of Alloy Steel as a Percentage of Crude Steel Production,

2008

0%

5%

10%

15%

20%

25%

Japan European Union South Korea China C.I.S.


As a rule of thumb, molybdenum is typically found in long products (i.e. bar, wire,

rods, and billets) as opposed to flat products. If molybdenum is found in flatproducts, it is usually of the hot-rolled variety. Both long and hot-rolled flat steels areused predominantly in industrial applications that require exacting specifications.

Molybdenum has a rather unique effect on steel whereby relatively small additionscan enhance the steel’s performance with a minimal increase in cost . Adding





molybdenum to iron allows the steel to respond more uniformly to heat treatment,increases the tensile, fatigue, and impact strength, in addition to imparting wear

resistance and toughness. These properties are often needed in non-idealenvironments such as seawater, chloride solutions, and elevated temperatures.

CPM’s Molybdenum Market Outlook segments special steels into the following

categories: stainless, full alloy, tool, high-strength low-alloy (HSLA), and carbon.

S T A I N L E S S

The stainless steel sector accounted for roughly 26% of global demand in 2008.

Molybdenum is employed in certain corrosion-resistant austenitic (300 series), highstrength super ferritic (400 series), and duplex grades. By weight, grade 316contains 2-3% molybdenum whereas grade 444 contains 1-2% molybdenum. In

duplex steels, the molybdenum content can reach 6%. Production of duplex steel

currently stands at about 600,000 t/a, and could reach above 1 Mt by 2014.Molybdenum is found in over 96% of all duplex steels. Led by emerging markets as

well as shifting consumer preferences, molybdenum-based stainless steel demand isforecast to be 205 million pounds by 2018, up from 123 million pounds in 2008. Thisrepresents a compound annual growth rate (CAGR) of 7.1% between 2009 and

2018.

The following is a list of key uses for molybdenum-bearing stainless steels:

tubular products for the oil and gas industry

desalination plants

offshore infrastructure

coal and oil burning power plants pulp and paper facilities

storage tanks

industrial food processing equipment

chemical processing

petrochemical facilities.

F U L L A L L O Y

Full alloy steels accounted for about 16% of global molybdenum demand in 2008.This steel category includes many different types of steels in the engineering,construction, and electrical steel categories. As such, not all types of full alloy steels

contain molybdenum. In the grades where molybdenum is employed, the contentcan range from 0.15% to 9.0% to optimize the material’s hardness and temperatureresistance. Molybdenum can be substituted in certain applications, and consumers

have shown a preference for micro-alloyed versus full alloy steel over the past

decade or so. CPM forecasts that molybdenum demand in full alloy steels could





register a CAGR of 5.2% in the 2009–2018 period, rising from a projected 68 millionpounds in 2009 to 107 million pounds in 2018.

The following is a list of key uses for molybdenum-bearing full alloy steels:

general construction

shipbuilding

heavy machinery

transport systems

automobiles

offshore pipelines.

T O O L S T E E L

Tool steels are used in the fabrication of industrial tools for cutting, punching, milling,or forming metals and plastics. Although tool steels account for less than a half-percent of total crude steel output on a tonnage basis, they are high-value, niche

products. They are manufactured with superior hardness, abrasion resistance, and

ability to withstand extremely high temperatures. Molybdenum can account for 0.1%to 9.5% of the total weight of a tool steel. Tool steel performance enhancements can

be obtained with higher molybdenum contents, which improve cracking and fatigueresistance. The quality of tool steel varies greatly by region. Although the largest

tool steel producer is China, higher quality products are generally found in parts of

the European Union, in Japan, and in the United States. Changes in manufacturingspending and fluctuations in industrial production in OECD as well as non-OECD

economies can cause variations in tool steel demand. Demand for tool steel couldgrow at compound annual rate of 5.1% per year between 2009 and 2018.

Key uses for molybdenum-bearing tool steels include machinery for automotive,heavy equipment, and electronics production.

H L S A S T E E L

HSLA steels are typically used in the oil and gas industry for pipeline infrastructureas well as in the automotive and construction industries. HSLA steels employ

different alloying elements to reduce the total weight of the steel while optimizing

performance for specific environments. Lower overall weight, reduced welding, and

a more resilient outer surface can be achieved when alloys such as molybdenum areadded. Molybdenum-bearing HSLA steels are commonly used throughout theenergy supply chain in pipelines, offshore platforms, storage tanks, and LNG

carriers. For example, China’s second East-West Gas pipeline, which is planned to

connect resources in Central Asia to 15 Chinese regions, reportedly will containroughly 4,100 pounds of molybdenum per mile of pipe. In total, the 5,408 milepipeline will employ an estimated 22 million pounds of molybdenum. Considering the





expected growth in energy infrastructure, notably in non-OECD economies,molybdenum demand in HSLA steel is forecast to grow at a compound annual rate of

6.5% between 2009 and 2018, rising from a projected 42 million pounds in 2009 tomore than 73 million pounds by the end of the forecast period.

The following is a list of key uses for molybdenum-bearing HSLA steels:

energy supply chain (pipelines, offshore platforms, storage tanks, and LNGcarriers)

automotive parts

general construction.

Figure 22.7 World Energy Consumption

0

50

100

150

200

250

300

350

400

450

1995 2000 2005 2010 2015 2020 2025 2030

OECD Non-OECD

Actual Projected

Quad. Btu


C A R B O N ST E E L

Carbon, or generic, steel accounts for approximately 90% of global steel production.In these steels, carbon is the primary alloying element, which adds strength and

hardness to iron, and molybdenum is rarely used, but if it is, it is used in small

amounts. Even so, carbon steel accounts for roughly 8% of global molybdenumdemand because of its large market share. In 2008, carbon steel consumed slightly

less than 39 million pounds of molybdenum. By 2018, this figure is projected to beabove 52 million pounds.

The following is a list of key uses for molybdenum-bearing carbon steels:

construction and transportation industries





tunnels

food storage

communication infrastructure.

22.7.2 O T H E R AP P L I C A T I O N S

C A T A L Y S T S

Molybdenum-bearing catalysts are instrumental in petroleum refining and

petrochemical processing. Molybdenum demand in catalysts recorded a CAGR of

7.7% between 2001 and 2008 as crude oil streams became heavier and sourer, oilproduct demand rose swiftly in emerging economies, new refineries were built in Asia

and the Middle East, and sulphur content regulations became more stringent. Thesestructural trends are expected to intensify over the coming decade. High complexity

refineries are still needed to cope with the shift in crude oil quality and the growth in

middle distillate demand. In addition, large-scale petrochemical projects continue tobe planned for the major producing centers in the Middle East and the largeconsuming countries in Asia. As a result, demand in catalysts could grow at a

compounded annual rate of 5.9% from 2009 through 2018.

Figure 22.8 Average Annual Growth in Crude Oil Production – 1996 to 2007, by

Crude Quality

0.00%

0.50%

1.00%

1.50%

2.00%

2.50%

3.00%

3.50%

4.00%

4.50%

Light Medium Heavy Sweet Medium Sour

Sulfur Co ntent API Gravity






S U P E R A L L O Y S

A superalloy is a metallurgical product designed for use in extremely high

temperatures where materials cannot afford to expand under extreme temperatures.Molybdenum can be added to superalloys to assist in creep strength and resistance

to oxidization; properties necessary in sulphuric-acid plants, land-based turbinesfound in gas-fired power plants, aircraft parts, and defence applications. Growth insuperalloy demand will trend broadly with fixed-asset investment as well asdevelopments in the aerospace industry. Molybdenum demand for superalloys may

grow from nearly 28 million pounds in 2008 to more than 43 million pounds in 2018.CPM’s Molybdenum Market Outlook forecasts a strong pickup after 2010 in line with

aerospace production trends and infrastructure investment timelines.

M O L Y B D E N U M M E T A L A N D A L L O Y S

Molybdenum as a pure metal or a molybdenum-based alloy is used primarily in the

coatings, lighting, and glass industries. Along with nickel and chromium,molybdenum-bearing coatings improve the wear and friction properties of automotiveparts such as gears, synchronizers, and piston rings. Molybdenum wire is employed

in the lighting manufacturing process, in sealing ribbons for automotive halogen

lamps, and in other manufacturing applications. Moreover, pure molybdenumpowder coating is used in emerging applications such as CIGS photovoltaic solar

cells, where 50–60 t of molybdenum metal powder is needed per 100 MW. Some of the uses in this category are susceptible to substitution where other elements cancompete on price and/or performance. Molybdenum demand in this sector is

expected to grow at an average rate of slightly more than 3% over the next decade.

C A S T I R O N

Cast iron refers to a family of multi-component ferrous alloys containing primarily

iron, carbon, silicon, as well as other major and minor alloying elements. Theproducts that contain molybdenum are employed in diesel engine motor blocks andcylinder heads, and in mining operations. Although molybdenum is not always used

as an alloy, it is generally chosen to increase the cast iron’s strength, heatresistance, and creep resistance. The molybdenum content of cast iron can be up to

3% by weight. Molybdenum demand growth in this sector is expected to generally

trend with global economic growth although years of higher growth are expected dueto the forecast expansion of the diesel engine market and higher levels of industrialproduction. Molybdenum-based cast iron growth is forecasted at a 4.8% CAGR

between 2009 and 2018.

L U B R I C A N T S

Molybdenum disulfide is used in lubricants due to a lower coefficient of friction than

other lubricants, durability to withstand high temperatures, high yield strength, and astrong affinity for metallic surfaces. Global lubricant demand for molybdenum is





projected to increase at a CAGR of 3.8% from 2009 through 2018, led by growth inthe manufacturing and automotive sectors of emerging markets.

P I G M E N T S

Molybdenum is used in molybdate-based pigments because of its stable color formation and corrosion inhibition properties. These light and heat-stable pigments

have a wide range of colors from bright red-orange to red-yellow, and they are usedin paints, inks, plastics, rubber products, and ceramics. The shift away from

inorganic pigments, which contain molybdenum and sometimes toxic metals such as

lead, toward organic-based pigments continues to threaten this market segment.Molybdenum demand in the pigment industry is expected to grow at a 3.4%compound annual rate between 2009 and 2018.

O T H E R

Molybdenum’s use in chemicals and other applications is dominated by its use as asmoke suppressant in PVC cabling. Molybdenum stabilizes char in combustible

situations and thus prevents the formation of smoke particles. PVC is made throughthe polymerization of a vinyl chloride monomer. PVC is used in construction,

housing, automobiles, airplanes, and medical devices. Molybdenum demand growthis forecast at a 4.4% CAGR in the 2009 – 2018 period, trending with GDP growthand influenced by the housing and construction industries.

2 2 . 8 S U B S T I T U T E S

Molybdenum demand, for the most part, is price inelastic. The key reason for molybdenum’s price inelasticity is that the majority of molybdenum is used in smallquantities as a microalloy. Buyers, therefore, tend to absorb higher prices as the

benefits of molybdenum outweigh its cost. In addition, steel producers, for example,were able the transfer the cost increases when prices were higher through alloy

surcharges without a reactionary decrease in demand.

Possible substitutes for molybdenum include chromium, vanadium, niobium, andboron in alloy steels; tungsten in tool steels; graphite, tungsten, and tantalum for

refractory materials in high-temperature electric furnaces. Substitution remains a

threat for some molybdenum-bearing products, but a significant changeover to other metals is unlikely as these are not perfect substitutes for the majority of

molybdenum’s existing applications. In addition, many of these substitute marketscannot support a large transition from the molybdenum market, given their relativesize. Over the past five years, conditions in the molybdenum market and its

substitutes’ markets have trended together. Prices for all these metals increased,thereby negating a cost advantage. Moreover, many these substitutes arecomplementary to molybdenum use. Although a large scale shift to other markets is





not feasible, some pockets of substitution are factored into the CPM’s high, low, andbase case scenarios.

2 2 . 9 C P M ’ S S U P P L Y A N D D E M A N D B A L A N C E

Increased demand for molybdenum, notably between 2003 and 2007, ushered in asubstantial restructuring of the molybdenum industry. To meet demand, the pace of

new supplies also has had to accelerate. As miners sought to develop deposits thatwere once uneconomic, operating costs for both primary producers and by-product

producers rose. All molybdenum producers are now subject to operating costs (pro-

rata) that are greater than the long-term average price nominal price.

CPM’s analysis reveals that demand for molybdenum may remain fairly priceinelastic and the growth in demand justifies the rising costs due to the following

developments:

Increases in mine supply are in the pipeline but physical, bureaucratic, and

financial constraints are slowing down this progress.

There are few substitutes for molybdenum in its major applications; thusfuture demand, for the vast majority of end uses, is not likely to be swayedby strong prices.

A period of tight supply and demand conditions is forecast to unfold in 2011 and

2012, pushing real molybdenum prices up to an average of $23.00 per pound over this two-year period. The ramp-up of larger molybdenum projects is estimated tomove the market into a surplus from 2013 through 2017. However, the running

balance of inventories suggests that stock levels will remain fairly low over this five-year period. While growth in secondary supplies from recycled catalysts may helpmitigate the pressure from market fundamentals, fresh supplies from either by-

production or primary producers seem necessary to fill the supply gap in 2018, thetail-end of CPM’s 10-year projections. In 2018, real molybdenum prices are forecast

to average $16.50 per pound.





Figure 22.9 Base Case: Real Molybdenum Prices and World Supply and

Demand Balance – Annual, Projected through 2018p

0

6

12

18

24

30

36

42

-36

-24

-12

0

12

24

36

48

1995 1998 2001 2004 2007 2010p 2013p 2016p

US$/Lb.Million Pounds

Molybdenum Prices (RHS)

Deficit (LHS)

Surplus (LHS)

Actual Projections

2018p

Source: CPM.

2 2 . 1 0 P R I C E O U T L O O K – C P M ’ S B A S E C A S E S C E N A R I O

During the late 1970s and early 1980s strong demand for oil-country-tubular-goodsalloyed with molybdenum led to a spike in prices. In 1980 prices averaged roughly

$20 (more than $60 in real terms). Even though molybdenum prices rose nearly 16-fold from the start of 2002 to the monthly average high of $38.32 in June 2005, on a

historical basis, real molybdenum prices only reached roughly 60% of the peak seenduring the rally in 1980.

With molybdenum supplies unable to recover from the global recession and credit

crisis as quickly as demand, a deficit combined with stronger molybdenum prices is

expected in 2011 and 2012. However, higher prices are likely to lead to an uptick inChinese mine supplies. This increased supply combined with the ramping up of several large primary producers may result in oversupply conditions in 2014 and

2015. CPM’s mine supply forecast includes the restart of Freeport McMoRan’sClimax project and the commencement of new Greenfield projects such as General

Moly’s Mount Hope project. It also includes additional supplies from new

molybdenum recovery units at existing copper producers, such as BHP Billiton’s andRio Tinto’s Escondida mine, as well as new copper development projects. Pricesmay fall back temporarily in 2013 and 2014 before edging higher.

By 2016 and 2017 prices may rise to average $14.75, even as supply exceedsdemand by roughly 10 million pounds per year.





The molybdenum market balance is forecast to revert to a deficit in 2018 in the basecase scenario. Increased output from either by-product or primary producers seems

necessary to fill the gap in supply during the tail-end of CPM’s projection period. In2018 molybdenum prices are forecast to average $16.50.

2 2 . 1 1 P R I C E O U T L O O K – C P M ’ S A L T E R N A T I V E S C E N A R I O S

The supply forecast could vary from the base case scenario, as multiple factors couldinfluence future supply including demand, credit markets, molybdenum prices, and

unforeseen disruptions in mine production. For CPM’s alternative scenarios, the

majority of the underlying assumptions made for the base case scenario are heldfixed including, macroeconomic variables such as exchange rates, inflation, andenergy prices. The high and low case scenarios are demand-driven in that changes

in the market’s demand fundamentals will spur calibrations on the supply-side as well

as oscillations in price levels.

2 2 .1 1 .1 C P M’ S H I G H C A S E SC E N A R I O

Demand for molybdenum is projected to increase by a compound average growthrate of 4.5% between 2008 and 2018 in the high case scenario. This is greater thanthe CAGR of 4.1% projected in the main scenario and is reflective of:

In the first few years of the forecast period, the market remains tighter than

in the base case. Deficit conditions emerge between 2010 and 2013, whichreflects a strong post-recession restocking cycle in the steel industry as well

as a return to strong economic growth.

Above-trend global energy demand as well as the continued need for energysupply infrastructure in emerging markets keeps growth rates high in thesecond half of the forecast period. This is particularly evident in the steel

category in addition to superalloys and catalysts.

2 2 .1 1 .2 C P M’ S L O W C A S E SC E N A R I O

Meanwhile, in the low case scenario, demand for molybdenum is projected toincrease by a CAGR of 3.9% between 2008 and 2018. This compares to a 4.1% inthe main scenario and is based on:

A less smooth transition into an economic expansionary phase after the2008–2009 global downturn. The increases in steel capacity utilizationrates, which began in the second and third quarter of 2009, begin to slow inline with the sluggish recovery in final demand. In addition, excess capacity

in key end-use industries takes longer to work off before robust growth

resumes. As a result, the rebound in molybdenum demand in 2010 is lower than anticipated in the base case.





On a longer term basis, global molybdenum registers a 5.5% CAGR

between 2010 and 2018 in the low case compared to 5.7% in the base case.

This discrepancy is the result of slightly lower economic growth during theforecast period as well as lower energy demand levels.

Table 22.4 Real Annual Molybdenum Prices per Pound

Base

Case

High

Case

Low

Case

2008 $28.73 $28.73 $28.73

2009p $11.11 $11.11 $11.11

2010p $17.50 $18.00 $15.00

2011p $22.00 $24.00 $20.00

2012p $24.00 $28.00 $18.00

2013p $18.50 $24.00 $15.50

2014p $13.00 $14.50 $12.00

2015p $13.75 $15.50 $13.002016p $14.50 $15.00 $13.50

2017p $15.00 $15.75 $14.00

2018p $16.50 $17.50 $14.50

Source: CPM.





2 3 . 0 R I S K A N D M I T I G A T I O N

Risk identification allows mitigating strategies to be devised and resources to be

allowed for their implementation, thus enhancing the project’s security. While

unforeseeable risks by their nature cannot be predicted, the effort to identify risks has

been comprehensive.

Table 23.1 outlines the risks identified for the project.

Table 23.1 Risks and Mitigations

Risk Mitigation

Process

Lead removal from

concentrate

Testwork to confirm lead removal by leaching to ensure that the

concentrate grade meets the market specification

Lead leach filtrate

treatment

Testwork to neutralize leach filtrate to ensure that it is

environmentally acceptable for disposal

Final concentrate

molybdenum grade too

low

Testwork to confirm number of cleaner and regrind stages

Rock lithologies

responding differently to

selected process

Locked-cycle variability test evaluation required

Quantify smelter penalty

elements

Concentrate to be analysed for characterization of impurity elements

Thickening and filtration Rheological tests including settling and filtration using concentrates

required

Project Execution

Project Financing Ensure appropriate resources and efforts are directed early to

secure financing

Delays in project schedule

due to delivery delays,

engineering delays

Execution plan includes early procurement of long-lead items;

selected portions of engineering have been advanced to near basic

engineering level during feasibility phase

Capital cost overruns Costs to be monitored and trended through project execution phase

Contractor non-

performance

Close monitoring and managing of contractor critical items

scheduling and costs to minimize unavoidable cost over-runs

Non-availability of key

personnel (management,engineering, supervisory,

and tradesmen)

Ensure early placement of contracts, prompt and effective recruiting

at start of project, and the expanded use of contractors andconsultants

table continues…





Risk Mitigation

Transportation and Logistics

Foul weather Backup and storage of essential supplies, medical, food, fuel, etc.

for foul weather conditions

Emergency transportation Allocate designated air evacuation plan for emergencies andessential material transport

Mining Operations

Initial production Productivity may be reduced and production targets not met due to:

training of new personnel

development schedule delays, etc.

mining and milling equipment start up and trouble shooting

Monitor and advance detail engineering work to ensure timely pre-

production mine development, assess mining extraction sequences,

and milling operational delays during start-up

Metal prices and mining

grade

Monitor exchange rate and metal price trends and implement

forward selling strategies

Mining ore grade Monitor and enhance quality control during mining of ore and

develop flexible mine plans with multiple ore sources

Skilled labour availability By retaining services from mining contractors and consultant to

cover temporary shortfall in certain key areas until appropriate staff

are employed; early recruitment and training during pre-production

South Wall failure causes

failure of diversion

channel

Ongoing monitoring and clean-up of the diversion channel

Water diversion drainage

channel within the pit

An open drainage channel will transfer Patsy Creek during the

construction and early operational phases of the mine; audit design

and location options; review design and construction options in the

FS

High demand & short

supply of miningequipment

Contracts are required from tire equipment providers in 2011 to

ensure that mining equipment will be available in time for projectstart up

Tire supply Supply contracts are required from tire manufactures in 2011 to

ensure that tires will be available in time for project start up.

Tailing Management Facility

Deep seated slope failure Facility is designed to required codes and standards

Surface slope failure Facility is designed to required codes and standards

Tailing construction water

management

Construction diversion channels, pipes, and intake structures will

need to be validated by geotechnical investigations along potential

diversion alignments during the FS.

Starter Embankment The topographical relief in the area of the embankment is

significant, with near vertical valley walls in some areas. A greater

level of detail in the construction sequencing, including siteinvestigations, is required to refine the construction duration.

table continues…





Risk Mitigation

Geotechnical investigation A PFS site investigation program has been conducted to date.

Further investigations, including drillholes in the embankment

alignment, borrow areas, rock quarries and foundations of the intake

structures will be required to advance designs during the feasibility

study and detailed engineering stages.

Embankment Seepage A synthetic geomembrane liner on the upstream slope of the

embankment is used on the Stage 1 embankment. The Stage 2

and 3 expansions utilize a compacted glacial till core within the

zoned embankment.

Dam overtopping All stages of the TMF will incorporate an emergency spillway. Dam

designed to meet CDA guidelines.

Sediment Control Sediment control structures (ponds and ditches) will be constructed

within borrow areas, and downstream of the main embankment to

collect sediment and decrease suspended solids prior to water

release.

Seepage through

Foundation

Keying the liner system into the bedrock will prevent seepage

through the embankment. Localized grouting may be required toreduce seepage potential through any fracture zones in the

foundations.

Embankment Failure Design embankment with required Factors of Safety, as per

Canadian Dam Association, Dam Safety Guidelines. Furthermore,

as part of the Feasibility Design, a dam breach and failure runout

analysis will need to be conducted, as per CDA requirements.

Finally, a rockfill dam is inherently stable under seismic events.

Water Management

Water diversion tunnel Historical geotechnical investigations and one 2009 drillhole provide

insight into the construction of the water diversion tunnel. Further

investigations, including drillholes intersecting the tunnel alignment,

will be required to advance designs to feasibility level.

Infrastructure

Main access roads Inadequate definition of upgraded road system carrying capacity

could lead to incorrect project implementation scheduling or higher

capital costs; conduct an extensive logistics study to assess roads

and give recommendations

The port foreshore is

susceptible to liquefaction

under the design seismic

event and will require

marine densification

Conduct an offshore geotechnical investigation including a drilling

program.

Port and mine worker

camp development

Inadequate definition of marine docking and environmental

conditions, as well as upland terrain conditions, could result in

permitting delays of the developing site, or the need for higher

costs. Conduct terrestrial and marine foreshore investigations at a

sufficient level of detail to ensure all geotechnical, marine

navigation/docking and environmental (fisheries) values are

identified and available to site design.

table continues…





Risk Mitigation

Environmental

Environmental process

review delay

Confirm appropriate EA process as soon as possible. Develop

proactive agency, working group, and consultation strategies and

plans to ensure the shortest feasible pre-application period and best

issue resolution during the application review period. Proactively

address provincial, federal and Nisga’a interests and requirements.

IFC imposed constraints

from EA process

Delay in construction due to delays in the EA schedule and

contingent on First Nations approval; coordination reviews will be

required between Avanti and Rescan





2 4 . 0 O P P O R T U N I T I E S A N D

R E C O M M E N D A T I O N S

The following sections outline areas to investigate for project improvements. To

complete the recommendations, a high-level budgetary estimate is provided inCanadian dollars.

2 4 . 1 G E O L O G Y

It is recommended to advance the project to feasibility-level studies. The resource

model as it stands is sufficient but the secondary metals, if recoverable, maycontribute to project revenues. Therefore, it is recommended to perform an audit on

the current model for lead and silver grades to establish potential by-product

recovery ($10,000).

2 4 . 2 M I N I N G

Recommendations related to mining are as follows:

Investigation of a higher production rate (greater than the current base caseof 40,000 t/d). Considering the available mineral resources, there is

potential to increase production rate and potentially improve the overallproject economics

Further study of the Patsy Creek Diversion, in the south pit wall, isnecessary in order to confirm geotechnical parameters and pit slope

stability. The study should look at minimizing additional waste stripping

caused by the water diversion.

Alternative primary crusher location should be investigated. There may bepotential to move the crusher to the northwest corner of the pit that will

reduce initial capital cost for construction of the access to the crusher; andmine operating and capital costs.

Further study of the transportation system(s) for moving waste from the pit tothe tailing dam for use in embankment construction.

Detailed Hydro-Geology evaluation of the area is needed in order to improvethe accuracy of pit dewatering design. Vertical and horizontal dewatering

wells have not been included as part of the required PFS activities. It may

be necessary to lower the water table within the pit prior to mining and toprevent water inflow from the nearby creeks and surface. It is





recommended that a water management plan for the location of the pit water discharge be included in future studies.

Considering hydro-geology evaluation, a hydrology assessment is requiredso that the diversion and water management plan will be developed for the

mining area, for both surface and groundwater quantities. Foundation testing of the waste dump and low grade stock pile is required.

Detailed drill and blasting studies are recommended for more future studies.These will help to determine the penetration rate expected for the selecteddrills and the specific rock types in the pit and the most applicable powder

factor for the rock types that are present in the Kitsault pit. It is

recommended that a detailed blasting study will include a cost benefitassessment between owner run blasting versus full contractor blasting

responsibility. The most effective system will result in better blastingefficiency and lower costs.

It is recommended that the optimization of the shovel fleet be completed

during the feasibility phase.

The use of contractor to move the material volumes in the schedule bereassessed in the next study in line with shovel optimization.

It is recommended that a study to address the optimization of the haulage

fleet be carried out during the feasibility phase. There is the potential to use

a larger-sized truck fleet for the current and possibly higher production ratescenarios.

2 4 . 3 T A I L I N G A N D W A T E R M A N A G E M E N T

Recommendations for the TMF and site water management systems include thefollowing:

additional and complimentary geotechnical site investigations in the footprint

of the TMF main embankment (inclusive of the overflow spillways requiredfor each stage of embankment)

geotechnical site investigations at the intake structures and alignment for the

Lime Creek Diversion Tunnel, at the intake structures of the Patsy Creek

Diversion, at access roads and at the construction borrow areas

develop updated sediment control plans for site development and borrowarea excavations

trade-off study for main haul road to TMF to consider incorporatingengineered retained wall along steep sidehill sections to reduce rock

excavation costs

detailed laboratory testing program on samples of the tailing and waste rock

hydrological and water quality analyses of Clary Lake and Roundy Creek





continued hydrological assessment of the Lime Creek and Patsy Creek

drainages

trade-off study for placing the balance of the waste in the TMF versus intothe pit at mine closure.

The following is a list of recommendations for future studies:

All input data used in the PFS represents best estimates/assumptions

available at the time of the study and need to be confirmed.

The exact locations and elevations of various system components need to

be confirmed.

The size distribution and rheology for bulk tailings need to be assessed.

Water quality in the TMF surface water pond and suitability for releasewithout treatment needs to be assessed and confirmed.

Seepage discharge that combines seepage through the embankment andrunoff from the surrounding terrain need to be confirmed.

Clary Lake water quality and intake location need to be assessed.

Roundy Creek water quality and flow rate need to be assessed.

As part of the feasibility design, a dam breach and failure runout analysis will

need to be conducted, as per CDA requirements.

The costs for the tailing and water management recommendations fall into twocategories (site investigations and engineering), as follows:

Site investigation and laboratory testing (including drilling contractors,helicopter time, laboratory costs, etc.) are estimated at $600,000.

All engineering work related to elevating the TMF and water management toa feasibility-level is estimated at $600,000.

2 4 . 4 P R O C E S S

The 2009 SGS test program must be augmented with additional flotation testwork to

confirm that the target concentrate grade of 52% Mo, and the overall molybdenumrecovery value of 90.6% can be achieved ($150,000).

The removal of lead from the molybdenum concentrate (in order to meet market

specifications and to ensure that there will be no smelter penalties for above-specification lead content) will require additional testwork and mineralogical

evaluation to determine the nature of the contaminating lead particles ($100,000).

Continued evaluation of bi-product revenue possibilities from the removal of lead andsilver from the ores should be conducted ($100,000).





2 4 . 5 I N F R A S T R U C T U R E

Wardrop has made the following recommendations for infrastructure:

Geotechnical site investigations should be conducted at the primary crusher,truck shop, plant site, camp and port to confirm the foundation conditions for

these major site structures.

Marine foreshore investigations should be completed to determine optimal

camp and docking facility location ($170,000).

Further investigation of port access to ensure that a handymax vessel couldsafely navigate into Alice Arm. The results of such a study would have to be

confirmed by the BC Coast Pilots. We recommend that Avanti negotiate a

contract with a shipping company for delivery of supplies and material andpossible export of concentrate ($20,000).

Proceed with additional engineering work to provide information required to

request initiation of a system impact study by BC Hydro. This study willproduce utility related costs and requirements ($125,000).

Consideration and active design of feasible energy conservation measures

(process design, equipment selection, energy recovery schemes) inconjunction with BC Hydro Power Smart initiatives to take advantage of energy cost reductions. Prudent planning with sufficient lead time and

involvement by BC Hydro experts can create opportunities to reduce the

relative magnitude of Tier II energy rates with respect to the overall project($75,000).

Further investigate the viability of a roasting facility in Prince Rupert, and the

possibility of transporting the molybdenum concentrate to this facility($150,000).

A risk analysis would need to be carried out at the front end of a futureFeasibility Study ($75,000).

2 4 . 6 E N V I R O M E N T A L

Rescan recommends that Avanti initiate environmental studies to support the

investigation of a roasting facility in Prince Rupert ($550,000.00).





2 5 . 0 I N T E R P R E T A T I O N A N D C O N C L U S I O N S

Based on the work carried out in this PFS for the Kitsault Project and the resultant

economic evaluation, this study should be followed by the recommended trade-off

studies as well as a Feasibility Study in order to further assess the economic viability

of the project.





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Berry, T.F. (May 1964), A Flotation Investigation for British Columbia Molybdenum

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Criteria Section IV, 1978, Volume 1.

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Drummond, A.D. et al, 1969, “Neyite a new Sulphosalt from Alice Arm, B.C.”; Can.

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the Kitsault Project 12,000 ton per Day Mill and By-products Plant, Climax

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Laboratory Work Relating to Expansion of Kitsault Milling Facilities to 12,000Tons per Day, Climax Western Operations Mine Evaluation Group.

Excell, J.D. (September 1971), Report on Hot Leach Process.

Ford, D.,(August 1971) ,Experimental Sandmill , British Columbia Molybdenum.

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A P P E N D I X A

C E R T I F I C A T E S O F Q U A L I F I E D P E R S O N S



C E R T I F I C A T E O F Q U A L I F I E D P E R S O N

I, Frank Grills, of Vancouver, BC, do hereby certify that as a co-author of this NI 43-101 PRE-

FEASIBILITY STUDY – AVANTI MINING INC., KITSAULT MOLYBDENUM PROPERTY,

BRITISH COLUMBIA, CANADA, dated December 15, 2009, I hereby make the following

statements:

I am a Project Manager with Wardrop Engineering Inc. with a business address at #800

– 555 W est Hastings St., Vancouver, BC, V6B 1M1.

I am a graduate of University of Surrey, (B.Sc. (Hons), 1979) and the University of

Witwatersrand (M.Sc. (Eng), 1985).

I am a member in good standing of the Association of Professional Engineers and

Geoscientists of British Columbia (Registration No. 29591).

I have practiced my profession continuously since graduation.

I have read the definition of “qualified person” set out in National Instrument 43-101

(NI 43-101) and certify that, by reason of my education, affiliation with a professional

association (as defined in NI 43-101) and past relevant work experience, I fulfill the

requirements to be a “qualified person” for the purpose of NI 43-101.

My relevant experience with respect to studies include 22 years experience in

engineering, estimating, construction management, and project management of major

process plants and studies. Recent studies I have worked on include Kerr Sulphurets

Mitchell and the Xietongmen Project.

I am responsible for the preparation of Sections 1.0, 2.0, 3.0, 5.0, 19.8, 19.9, 19.10,

19.11, 19.12, 20.0, 23.0, 24.0, 25.0, and those portions of 19.13 pertaining to the

process plant and associated infrastructure costs, of this technical report titled “NI 43-

101 Pre-feasibility Study – Avanti Mining Inc., Kitsault Molybdenum Property, British

Columbia, Canada”, dated December 15, 2009. In addition, I visited the property during

the period July 20 to 21, 2009.

I have no prior involvement with the Property that is the subject of the Technical Report.

As of the date of this Certificate, to my knowledge, information, and belief, this Technical

Report contains all scientific and technical information that is required to be disclosed to

make the technical report not misleading.

I am independent of the Issuer as defined by Section 1.4 of the Instrument.

I have read National Instrument 43-101 and the Technical Report has been prepared in

compliance with National Instrument 43-101 and Form 43-101F1.

Signed and dated this 15th day of December, 2009 at Vancouver, BC

“Original document signed and sealed by Frank Grills, P.Eng., M.Sc.(Eng)., PMP”

Frank Grills, P.Eng., M.Sc.(Eng), PMPProject Manager Wardrop Engineering Inc.




I, Marinus Andre de Ruijter, of Delta, British Columbia, do hereby certify that as a co-author of

this NI 43-101 PRE-FEASIBILITY STUDY – AVANTI MINING INC., KITSAULT

MOLYBDENUM PROPERTY, BRITISH COLUMBIA, CANADA, dated December 15, 2009, I

hereby make the following statements:

I am a Senior Metallurgical Engineer at Wardrop Engineering Inc. with a business

address at #800 – 555 West Hastings St., Vancouver, BC, V6B 1M1.

I am a graduate of the University of Witwatersrand, Johannesburg, South Africa (B.Sc. –

Physics, Mathematics, 1970; B.Eng., 1973; M.Eng., 1979).


Geoscientists of British Columbia (Registration #31031).

I have practiced my profession continuously since graduation, except during the years

2000 to 2004.





My relevant experience with respect to this project includes sulphide mineral flotation

and flotation research, and project development work on molybdenum ores.

I am responsible for the preparation of Sections 17.1 to 17.4, 17.6 to 17.7, 18.0, and

those portions of 19.14 pertaining to process operating costs, of this technical report

titled “NI 43-101 Pre-feasibility Study – Avanti Mining Inc., Kitsault Molybdenum

Property, British Columbia, Canada”, dated December 15, 2009. In addition, I visited

the property during the period July 20 to 21, 2009.









“Original document signed and sealed by Marinus Andre de Ruijter, P.Eng.”

Marinus Andre de Ruijter, P.Eng.Senior Metallurgical Engineer Wardrop Engineering Inc.




I, Miloje Vicentijevic, of Vancouver, British Columbia, do hereby certify that as a co-author of

this NI 43-101 PRE-FEASIBILITY STUDY – AVANTI MINING INC., KITSAULT

MOLYBDENUM PROPERTY, BRITISH COLUMBIA, CANADA, dated December 15, 2009, I

hereby make the following statements:

I am the Manager of Mining (Vancouver) for Wardrop Engineering Inc. with a business

address at #800 – 555 West Hastings St., Vancouver, BC, V6B 1M1.

I am a graduate of the University of Belgrade (B.Sc. in Mining Engineering, 1990) and

University of Alberta (M.Eng. in Engineering Management, 2003).

I am a member in good standing of the Association of Professional Engineers,

Geologists, and Geophysicists of Alberta (#75678).






My relevant experience includes 17 years of mining operations and consulting

experience.

I am responsible for the preparation of Sections 19.1, 19.15, and those portions of 19.13

and 19.14 pertaining to mining capital and operating costs, of this technical report titled

“NI 43-101 Pre-feasibility Study – Avanti Mining Inc., Kitsault Molybdenum Property,

British Columbia, Canada”, dated December 15, 2009. In addition, I visited the property

during the period June 28 to 29, 2009.

I have no prior involvement with the Property that is the subject of the Technical Report. As of the date of this Certificate, to my knowledge, information, and belief, this Technical







“Original document signed and sealed by Miloje Vicentijevic, P.Eng., M.Eng.”

Miloje Vicentijevic, P.Eng., M.Eng.Manager of Mining (Vancouver)Wardrop Engineering Inc.



SRK Consulting (U.S.), Inc.

7175 West Jefferson Avenue, Suite 3000

Lakewood, CO

USA 80235

[email protected]

www.srk.com

Tel: 303.985.1333

Fax: 303.985.9947

Page 1 of 2


I, Jeffrey Volk , of Lakewood, Colorado, do hereby certify that as a co-author of this NI 43-101

PRE-FEASIBILITY STUDY – AVANTI MINING INC., KITSAULT MOLYBDENUM PROPERTY,


statements:

I am a Principal Resource Geologist with SRK Consulting (U.S.), Inc. with a business

address at 7175 West Jefferson Avenue, Suite 3000, Lakewood, CO, USA, 80235.

I graduated with a Master of Science degree in Structural Geology from the Washington

State University in 1986. In addition, I have obtained a Bachelor of Arts degree in

geology from the University of Vermont in 1983.

I am a fellow of the Society of Economic Geologists and a Certified Professional

Geologist (CPG#10835) and member of the American Institute of Professional

Geologists (AIPG). I am also a fellow and member of the Australian Institute of Mining

and Metallurgy (FAusIMM).






I am responsible for the preparation of Sections 6.0 through 15.0, and 16.1.1 through

16.1.14 of this technical report titled “NI 43-101 Pre-feasibility Study – Avanti Mining

Inc., Kitsault Molybdenum Property, British Columbia, Canada”, dated December 15,

2009. In addition, I visited the property on July 15, 2008.

I have had prior involvement with the Property that is the subject of the Technical

Report, and was a “qualified person” responsible for the previously filed report entitled

“Amended NI 43-101 Technical Report on Resources, Avanti Mining, Inc. Kitsault

Molybdenum Property British Columbia, Canada” dated January 23, 2009. I was also a

“qualified person” for the previously filed report entitled NI 43-101 Preliminary Economic

Assessment Avanti Mining, Inc. Kitsault Molybdenum Property, British Columbia,Canada” dated March 3, 2009.







SRK Consulting (U.S.), Inc.

7175 West Jefferson Avenue, Suite 3000

Lakewood, CO

USA 80235

[email protected]

www.srk.com

Tel: 303.985.1333

Fax: 303.985.9947

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Signed and dated this 15th day of December, 2009 at Denver, Colorado

“Original document signed and sealed by Jeffrey Volk, CPG, FAusIMM, MSc”

Jeffrey Volk, CPG, FAusIMM, MScPrincipal Resource GeologistSRK Consulting (U.S.), Inc.




I, Michael Levy, of Denver, Colorado, do hereby certify that as a co-author of this NI 43-101

PRE-FEASIBILITY STUDY – AVANTI MINING INC., KITSAULT MOLYBDENUM PROPERTY,


statements:

I am a Senior Geotechnical Engineer with SRK Consulting (U.S.), Inc. with a business

address at 7175 West Jefferson Avenue, Suite 3000, Lakewood, Colorado 80235.

I have a B.Sc. in Geology from the University of Iowa in 1998 and a M.Sc. in Civil-

Geotechnical Engineering from the University of Colorado in 2004.

I am a registered Professional Engineer in the states of Colorado (#40268) and

California (#70578) and a registered Professional Geologist in the state of Wyoming

(#3550).






My relevant experience with respect to the pit slope evaluation includes data collection,

analysis, and m odelling for pit slope design projects at various mines throughout North

and South America.

I am responsible for the preparation of Section 19.2 of this technical report titled “NI 43-

101 Pre-feasibility Study – Avanti Mining Inc., Kitsault Molybdenum Property, British

Columbia, Canada”, dated December 15, 2009. In addition, I visited the property during

the period between September 8 and 11, 2008.




I have had prior involvement with the Property that is the subject of the Technical

Report, and was a “qualified person” responsible for the previously filed report entitled

“NI 43-101 Preliminary Economic Assessment” Avanti Mining, Inc. Kitsault Molybdenum

Property, British Columbia, Canada” dated March 3, 2009.




Signed and dated this 15th day of December, 2009 at Denver, Colorado.

“Original document signed and sealed by Michael Levy, P.E., P.G.”

Michael Levy, P.E., P.G.Senior Geotechnical Engineer SRK Consulting (U.S.) Inc.




I, Peter Healey, of Vancouver B.C., do hereby certify that as a co-author of this NI 43-101 PRE-



statements:

I am a Prinicipal with SRK Consulting (Canada) Inc with a business address at 2200 –

1066 West Hastings, Vancouver, BC.

I am a graduate of Auckland University (BE, 1976).


Geoscientists of British Columbia (License 13202).






My relevant experience with respect to reclamation and closure includes preparation of

closure plans for a number of mines in BC and the Yukon.

I am responsible for the preparation of Sections 1.21 and 19.7 of this technical report



the property during the period June 28 to 30, 2009.

I have been involvement with the reclamation and closure of the Kitsault Mine, the

subject of the Technical Report, since 1996. As of the date of this Certificate, to my knowledge, information, and belief, this Technical







“Original document signed and

sealed by Peter Healey, P.Eng” Peter Healey P.Eng,PrincipalSRK Consulting (Canada) Inc.




I, Stephen Day, of Vancouver B.C., do hereby certify that as a co-author of this NI 43-101 PRE-



statements:

I am a Prinicipal with SRK Consulting (Canada) Inc with a business address at 2200

1066 West Hastings Vancouver BC.

I graduated from the Department of Geological Sciences, University of British Columbia

with a Bachelor of Science Degree in 1985 and Master of Science Degree in 1988.


Geoscientists of British Columbia (License 18467).

I have worked as a geochemist for a total of 18 years since graduation from university.





I am responsible for geochemical characterization of mine and process wastes for the

proposed mine.

I am responsible for the preparation of Sections 1.20.2 and 19.6 of this technical report



the property during the period September 21 to 22, 2009.

I have no prior involvement with the Property that is the subject of the Technical Report. As of the date of this Certificate, to my knowledge, information, and belief, this Technical







“Original document signed and sealed

by Stephen Day, P.Geo” Stephen Day P.Geo.PrincipalSRK Consulting (Canada) Inc




I, Michael Royle, of Vancouver, BC, do hereby certify that as a co-author of this NI 43-101 PRE-



statements:

I am a Principal Hydrogeologist with SRK Consulting (Canada) Inc. with a business

address at 2200 – 1066 West Hastings Street, Vancouver, BC, V6E-3X2.

I am a graduate of the University of British Columbia, (Bachelor of Science, 1987).


Geoscientists of British Columbia (License # 27830).

I have practiced my profession for more than 20 years since graduation.





My relevant experience with respect to pit hydrogeology includes program review.

I am responsible for the review of Section 19.3 of this technical report titled “NI 43-101

Pre-feasibility Study – Avanti Mining Inc., Kitsault Molybdenum Property, British

Columbia, Canada”, dated December 15, 2009.


As of the date of this Certificate, to my knowledge, information, and belief, this TechnicalReport contains all scientific and technical information that is required to be disclosed to





Signed and dated this 15th day of December, 2009 at Vancouver, B.C.

“Original document signed and sealed

by Michael Royle, M.App.Sci., P.Geo. (BC)” Michael Royle, M.App.Sci., P.Geo. (BC)Principal HydrogeologistSRK Consulting (Canada) Inc.




I, Ken Brouwer, of Vancouver, B.C., do hereby certify that as a co-author of this NI 43-101 PRE-



statements:

I am Managing Director with Knight Piésold Ltd., Vancouver with a business address at

1400 – 750 West Pender Street.

I am a graduate of University of British Columbia, (BASc., 1982; M.Eng., 1985).


Geoscientists of British Columbia (15117).






My relevant experience with respect to geotechnical engineering, tailings management,

and water management for mining projects includes over 25 years of practice.

I am responsible for the preparation of Sections 19.4 and 19.5 with associated inputs to

Sections 19.13, 19.14, 23.0, and 24.0 of this technical report titled “NI 43-101 Pre-

feasibility Study – Avanti Mining Inc., Kitsault Molybdenum Property, British Columbia,

Canada”, dated December 15, 2009. In addition, I visited the property during the period

from July 20 to 21, 2009.

I was involved as a senior reviewer of the tailings management concepts that weredeveloped by Knight Piésold Ltd. during the previous Preliminary Economic Assessment

completed by SRK for the Kitsault project.







Signed and dated this 15th day of December, 2009 at Vancouver, B.C.

“Original document signed and sealed by Ken J. Brouwer, P.Eng.”

Ken J. Brouwer, P.Eng.Managing Director Knight Piésold Ltd.



Page 1 of 2


I, Harold Rolf Schmitt, of Victoria, British Columbia, do hereby certify that as a co-author of this

NI 43-101 PRE-FEASIBILITY STUDY – AVANTI MINING INC., KITSAULT MOLYBDENUMPROPERTY, BRITISH COLUMBIA, CANADA, dated December 15, 2009, I hereby make the

following statements:

I am a Senior Project Manager with Rescan Environmental Services Ltd. with a

business address at 6th

Floor, 1111 West Hastings Street, Vancouver, BC V6E 2J3.

I am a graduate of University of BC (B.Sc.,1973, M.Sc., 1985) graduating in Geology

and Regional Planning, and University of Ottawa (M.Sc., 1993) in Applied

Geochemistry.


Geoscientists of British Columbia (License # 19824).

I have practiced my profession continuously since graduation. I have read the definition of “qualified person” set out in National Instrument 43-101




My relevant experience with respect to Mine Environmental Assessment and Permitting

includes 36 years of employment in my profession, much as a senior government

professional involved in mine and energy resource regulation, including 10 years in

progressively more responsible positions as a Project Manager, with the past 4 years

involved in major mine environmental assessment and permitting in British Columbia.

I am responsible for the preparation of Sections 4.0, 21.0, and 24.6, as well as the

‘Environmental’ portion of Table 23.1, of this technical report titled “NI 43-101 Pre-feasibility Study – Avanti Mining Inc., Kitsault Molybdenum Property, British Columbia,

Canada”, dated December 15, 2009.

I have been involved with the Kitsault Project since March, 2008. Prior work has

included the following:

the preparation of a report reviewing the status of environmental liabilities

associated with the former Kitsault Mine submarine tailings disposal facility

(September 2008),

QP for the Preliminary Economic Assessment report (March, 2009) on matters

concerning environmental assessment and permitting,

Rescan’s Senior Project Manager for development of the environmental assessment

workplan, including discussions with the project owner and provincial and federal

government regulators on environmental assessment review process and

permitting,

Senior Project Manager for the environmental scoping report on a Roaster facility

(September, 2009), and e) as the QP for the PFS (December, 2009).







I have read National Instrument 43-101 and the Technical Report has been prepared incompliance with National Instrument 43-101 and Form 43-101F1.

Signed and dated this 15th day of December, 2009 at Vancouver, British Columbia

“Original document signed and sealed by Harold Rolf Schmitt, M.Sc., P.Geo.”

Harold Rolf Schmitt, M.Sc., P.Geo.Senior Project Manager Rescan Environmental Services Ltd.

ni 43-101 pre-feasibility study – kitsault molybdenum

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