rsp 4-2003 s.sarkar
TRANSCRIPT
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Smarajit SarkarDepartment of Metallurgical and Materials Engineering
NIT Rourkela
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Blast furnace productivity depends upon an optimum gas
through flow as well as smooth and rapid burden
descent.
The character of the gas and stock movements is
intimately associated with the furnace lines.
The solid materials expand due to heating as they
descend and their volume contracts when they begin to
soften and ultimately melt at high temperatures in the
lower furnace.
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A further volume contraction occurs when the solid coke burns
before the tuyeres.
An enormous volume of the combustion gas has to bubble
through the coke grid irrigated with a mass of liquid metal and
slag.
An optimum furnace profile should cater to the physical and
chemical requirements of counter flow of the descending solid,
viscous pasty or liquid stock and the ascending gases at allplaces from the hearth to the top
cont
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Only then, an optimum utilization of the
chemical and thermal energies of thegases as well as a smooth, uniform and
maximum iron production with minimumcoke rate will be realized.
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o In an integrated steel works the capacity of the
Blast Furnace depends upon
The capacity of the works.
The process of steelmaking adopted.
The ratio of hot metal and steel scrap in thecharge.
Consumption of foundry iron in the works.
Losses of iron in the ladle and the casting
machine. The number of furnaces to be installed
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It is the volume of Blast Furnace occupied by the charge
materials and the products , i.e. the volume of furnace
from the stock line to the tap hole.
Useful volume = the furnace capacity C.U.U.V.
C.U.U.V = coefficient of utilization of useful volume.
The value of C.U.U.V. varies in a wide range from 0.48-
1.50 m3/ton of pig iron
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V =k D2H
V=Useful volume
H=Total height
D=Diameter at the bottom of the shaft
K=A coefficient usually lies with in the range of 0.47
to 0.53. High value is for slim profile.
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Total height = useful height +distance betweenstock line and the charging platform (it isgoverned by the construction of gas off-take and
charging platform, this dimensions varies from 3to 4m.)
Useful height= height from the tapping hole tothe stock line.
The height of the blast furnace is mainlygoverned by the strength of the raw materials,particularly that of coke.
cont
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The strength of the coke charged to the
furnace should be sufficient to withstand the
load of raw materials without gettingcrushed. Coke provides permeability(in thedry as well as wet zones )and also
mechanical support to the large chargecolumn, permitting the gases to ascendthrough the voids.
Total height (H)= 5.55V0.24
Useful height (H0) =0.88H
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Diameter:The belly /bosh parallel is the cylinder that
connects the tapers of the shaft and the bosh.Its diameter, dbll, and the ratio of this diameter to
the useful or inner height of the furnace as wellas to the diameter of the hearth play animportant role in the operation of the furnace.The correct descent of the stock, ascent of the
gas and efficient utilization of the chemical andthermal energies of the gas depend greatly uponthese ratios.
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The importance of an adequate belly diameter lies in the
fact that softening and melting of the gangue and
formation of the slag occurs in this region.
An increase in the diameter facilitates gas passage
through the sticky mass and also slows down stock
movement, thus increasing the residence time for indirect
reduction.
However, the belly diameter cannot be increased
arbitrarily as it is directly related to bosh angle, bosh
height, hearth and throat diameters and useful height.
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The belly height depends upon the softenability of the
ferrous burden and also on the shaft angle desired.
If the slag fusion occurs at higher temperatures and in a
narrow temperature range as in the case of pre-fluxed
burden, the hydraulic resistance decreases in the
vertical cross-section and the belly height can be
correspondingly reduced.
dbelly =0.59 (V)0.38
HbelIy = 0.07H
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The hearth is designed such that its volumebetween the iron notch and tuyeres is sufficientto hold the molten metal and the slag.
The dia of hearth depends upon:
The intensity of coke consumption.
The quality of burden.
The type of iron being produced.
D hearth =0.32 V0.45
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A very approximate relationship between
the coke burning rate and hearth diameteris given by the following equation:D = c Q 0.5D = hearth diameter, m
Q = coke throughput, tonnes/24hc = throughput coefficient which variesbetween 0.2-0.3 depending upon burdenpreparation.
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For highly prepared burden, the value of
c = 0.2 has been achieved in modern largefurnaces .
Therefore, for a furnace planned to produce10,000 THM per day with a coke rate of 500kg/THM, i.e., a coke throughput of 5,000tonnes per day, the hearth diameter shouldbe about 14.1 m.
The value will be 21.2 m if the value ofc=0.3.
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With increasing diameter of the hearth,the gas penetration must be ensured
by providing adequate bedpermeability with the use ofmechanically strong, rich, pre-fluxed
burden of uniform size and low slagbulk as well as strong lumpy coke.
The Hearth height should be 10% of thetotal height of the furnace
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The shaft height must be sufficient to allow theheating, preparation and reduction of ore beforethe burden reaches the bosh. In the upperregions of the shaft , volume changes due toincrease in temperature and carbon deposition.These demand an outward batter for smoothflow of materials. In the lower region of the shaft
, the material starts fusing and tends to stick tothe furnace wall. So to counteract the wall dragan outward butter is necessary.
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Stack height Hstack
= 0.63 H- 3.2 m
Stack angle
The stack angle usually ranges from 850to870(i) 850for weak and powdery ores;
(ii) 860for mixture of strong and weak, lumpy or
fine ores;(iii) 870for strong, lumpy ore and coke.
Th i i i h l
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The variations in the angles are necessaryfor obtaining an adequate peripheral flowwhich is an essential pre-requisite for
forcing of the blast furnace.
Since the ore hump is located in theintermediate zone and it moves almost
vertically downwards pushing the lightercoke towards the wall and the axis.
A smaller shaft angle in the case of weak
and powdery ore helps to loosen theperiphery.
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Stack angle can be calculated from the formula
Stack angle ()= Cot-1(D-d1/2xStack Height)
Where, D= Bosh parallel Diameter
d1= Throat Diameter
Bosh angle can be calculated from the formula
Bosh angle ()= Cot-1(D-d/2xBosh Height)
Where, D= Bosh parallel Diameter
d= Hearth Diameter
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When the raw materials are charged into the
blast furnace, little volume change takesplace for a few meters of their descent andhence the walls of the throat are generallyparallel
Throat diameter can not be too small as ithas to allow the enormous volume of the gasto pass through at a reasonably low velocity
to maintain adequate solid gas contact andto decrease the dust emission, throathanging and channeling.
Cont..
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Throat diameter can not be too wide as itmay compact the charge. A certainvelocity and lifting power of gas is
necessary for losening the charge at top.
Throat Diameter d throat =0.59 V0.35
Where, V= useful volume
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A considerable amount of slag and iron
descends to the hearth through the inter-tuyere
zones. If they do so without having beenadequately heated, the thermal state of the
hearth may be disturbed with attendant high
sulphur in iron, sluggish slag movement, erratic
metal analysis, frequent tuyere burning, etc.
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The distance between the adjacent tuyeres
around the hearth circumference should be such
as to obtain, as far as possible, a merging of theindividual combustion zones of each tuyere into
a continuous ring.
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The number of tuyeres mainly depend upon thediameter of the hearth. The diameter of thetuyeres depend upon the blast volume.
The following formulae can be used to determinethe number of tuyeres
Pavlov: n = 2d +1
Rice: n = 2.6d-0.3
Tikhomirov et al : n = 3d-8Where n= Number of tuyeres,
d=hearth diameter
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Capacity
(THM/Day)
Parameter
2000 3000 5000
Useful Volume (m3) 1700 2550 4250
Total Height (m) 33.08 36.46 41.22
Useful Height (m) 29.11 32.08 36.27
Bosh Parallel Dia (m) 9.96 11.62 14.11
Bosh Parallel Height (m) 2.32 2.55 2.89
Bosh Height (m) 4.37 4.81 5.44
Hearth Dia (m) 9.1 10.92 13.74
Hearth Area (m2) 65.04 93.66 148.27
Hearth Height (m) 3.308 3.646 4.122
Stack/Shaft Height (m) 17.64 19.77 22.77
Throat Dia (m) 6.87 7.85 9.29
Bosh Angle (0) 84.32 85.84 88.05
Stack Angle (0) 85 84.55 83.96
Nos. of Tuyeres 20 25 34
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Smarajit SarkarDepartment of Metallurgical and Materials Engineering
NIT Rourkela
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Burden distribution is one of the key operating
parameters influencing blast furnace
performance, particularly the productivity and
the coke rate.
The proper distribution of burden materials
improves bed permeability, wind acceptance,
and efficiency of gas utilisation.
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In a typical Indian blast furnace equipped with a bell-
less (Paul Wurth) distribution system, the decrease
in coke rate that is due exclusively to burden
distribution was found to be 1012 kg/thm.
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Design of the blast furnace
and its charging device
(effect of these factors is
constant).
Angle and size of the big bell.
Additional mechanical
device(s) used for obtaining
better distribution.
Speed of lowering of large
bell.
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Inconsistency inphysical properties ofcharge materials(deficiencies caused bythis should be
eliminated by improvingquality of the burden.
Size range of the various
charge materials
Angle of repose of raw
materials and other
physical characteristics of
the charge.
Density of charge
materials.
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Level, system andsequence ofcharging, programmeof revolving thedistributor (conditionsdetermining majormeans of blast
furnace processcontrol from top).
Distribution of chargeon the big bell
Height of the big bell
from the stock-line i.e.charge level in thefurnace throat.
Order and proportion
of charging of variousraw materials.
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The density of three important raw materials viz. the ore, the
coke and the limestone are quite different.
The heaviest is iron ore with around 5-6 glcc, the lightest is
coke with density of around 15 glccand the limestone is
intermediate with-a value of density around 30-35 glcc.
It means that the rolling tendency of coke particles is maxi-
mum and that of the ore is minimum. Since the density values
cannot be altered, the sizes may be so chosen that their
differential rolling tendencies are offset to some extent.
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The problem of very dense ores is serious
from the point of view of their sluggish
reduction rates rather than their tendency
towards segregation. Such ores are therefore
invariably crushed and sintered to obtain re
porous agglomerates before charging these
in the furnaces.
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When a multi-particle material is allowed to
gently fall on a horizontal plane it tends to form
a conical heap. The base angle of this cone is
known as angle of reposeof that material.
This angle depends upon the particle size, its
surface characteristics, moisture content, shape,
size distribution, etc.
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For an iron ore of 10-30 mm size, with an
average mean size of 18 mm, the angle of
repose is around 33-35. For coke of 27-75 mm
size, with an average size of 45 mm, the same is
around 35-38. Similarly the angle of repose for
sinter is in the range of 31-34 and for pellets it
is around 26-28.
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The higher is the angle of repose the more it has the tendency to
form ridges on charging in a blast furnace.
The more dried is the ore and the more it is free from fines the
less pronounced is the angle of repose and thus less is the
tendency towards segregation.
The clayey ores tend to form ridges because of their high angle
of repose. The effective way to reduce the angle of repose of any
iron ore is to eliminate the fines, dry the ore if wet and to wash
off clay, if any, adhering the ore.
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On dumping, as the materials fall on the stock
surface, they take a parabolic path and mainly
two different profiles of the accumulated mass
emerge depending upon whether the particles
hit the in-wall directly(V- shape) or the stock
surface (M-shape)
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The M-profile itself is generally obtained if the material
strikes the stock surface. This happens when the
bell/throat diameter ratio is small (larger bell-inwall
distance) or the charging distance is small . It is clear
that the peak of the M-contour approaches the inwall
(hence the peripheral permeability decreases) as the
charging distance increases and ultimately the M
changes to V profile.
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Smarajit SarkarDepartment of Metallurgical and Materials Engineering
NIT Rourkela
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Right at the top of the furnace is the granular zonethat contains
the coke and the iron bearing materials charged, sometimes
along with small quantities of limestone and other fluxes. The
iron-bearing oxides charged get reduced to wustite and metallic
iron towards the lower end of the granular zone.
As the burden descends further, and its temperature rises on
account of contact with the ascending hot gases, softening and
melting of the iron-bearing solids takes place in the so-calledcohesive zone(mushy zone).
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Further down the furnace, impure liquid iron and liquid slag are
formed. The absorption of carbon lowers the melting point of iron
drastically. For example, an iron alloy containing 4 wt. % carbon
melts at only 1185C..
In the cohesive zone and below it, coke is the source of carbon forcarburisation of liquid iron. However, carbon directly does not
dissolve in liquid iron at this stage. The possible mechanism of
carburisation of iron entails the formation of CO by gasification of
carbon, followed by the absorption of carbon by the reaction:
2CO(g) = [C]in Fe+ CO2(g)
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Coke is the only material of the blast furnace charge which descends to
the tuyere level in the solid state. It burns with air in front of the tuyeres
in a 1-2 m deep raceway around the hearth periphery.
Beyond the raceway there is a closely packed bed of coke, the central
coke column or dead man's zone.
The continuous consumption of coke and the consequent creation of an
empty space permit the downward flow of the charge materials.
The combustion zone is in the form of a pear shape, called 'raceway'in
which the hot gases rotate at high speeds carrying a small amount of
burning coke in suspension.
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The raceway is a vital part of the blast furnace since it is the heat
source in a gigantic reactor and at the same time a source of re-
ducing gas.
The salient features of Combustion zone are summarized below:
The force of the blast forms a cavity the roof of which is formed of
loosely packed or suspended coke lumps and the wall more closely
packed.
The CO2 concentration tends to increase gradually from the centre
and reaches a maximum value just before the raceway boundary
where most of the combustion of coke occurs according to:
C+O2 (air) =CO2+94450 cal
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The temperature of the gas rises as the coke
consumption proceeds and reaches a maximum just
before the raceway boundary. Thereafter, it falls sharply
as the endothermal reduction of CO2 by C proceeds;
CO2 +C =2CO-41000 cal
The concentration of CO2 fall; rapidly from the raceway
boundary and the gasification is completed within 200-
400 mm from the starting point of the reaction.
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The primary slag of relatively low melting point which forms in the lower
part of the stack or in the belly consists of FeO-containing silicate and
aluminates with varying amounts of lime which has become incorporated
depending upon the degree of calcination undergone .
As the slag descends, ferrous oxide is rapidly reduced by carbon as well
as by CO. As the lime is continually absorbed, the original
FeO-Si02-AI203system rapidly changes to the CaO-Si02-AI203system
with some minor impurities accompanying the burden. The dissolution
of lime and the approach to the CaO-Si02-Al203 system is morepronounced,
.
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As the liquid primary slag runs down the bosh and loses its fluxing
constituent FeO, the liquidus temperature also increases. If,
therefore, the slag has to remain liquid it must move down to hotter
parts of the furnace as rapidly as its melting point is raised. As the
reduction of FeO is almost complete above the tuyeres the resulting
bosh slag, composed mainly of CaO-Si02-AI203
The hearth slag is formed on dissolution of the lime which was not
incorporated in the bosh and on absorption of the coke ash released
during combustion. The formation is more or less complete in the
combustion zone.
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This slag runs along with the molten iron into thehearth and accumulates there and forms a poolwith the molten metal underneath. During thepassage of iron droplets through the slag layer,
the slag reacts with the metal and a transferenceof mainly Si, Mn and S occurs from or to themetal, tending to attain equilibrium betweenthemselves as far as possible.
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Below 600C :
Pre-heating and pre-reduction
600 -950C:
Indirect reduction of iron oxides by CO and H2
9500C to softening temperature:
Direct reduction; gasification of carbon (solution loss
reactions) by CO2 and H2 becomes prominent.
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The formation o{ cohesive layers or partiallyreduced and partially molten iron oxide takesplace.
The coke slits provide passage for gaseous flow.
Dripping or Dropping Zone Semi fluidized region in which liquids drip and
fragments of cohesive layers drop.
Zone through which liquids trickle down to thehearth. It is the final stage of iron oxide reduction
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Blast, injectants and coke are converted to hot reducing gas. This
gas reduces the ore as it moves counter currently towards the top of
the furnace.
Hearth
It is a container for liquids and coke where slag/metal! coke/gas
reactions take place. Metal droplets pass through the slag/coke
layer. Liquid metal/coke layer in which chemical reactions take
place only to a small extent.
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fluidization of small particles when the local gas
velocity is excessive;
diminution of void age due to swelling andsoftening-melting;
flooding of slag in the bosh zone when the slag
volume and gas velocity are excessive.
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The charge in the blast furnace descends under gravity against the
frictional forces of solids and buoyancy of gas. With increasing gas
velocity, the pressure drop increases approximately quadratically
until the upward thrust of the gas and downward thrust of the solids
are held in balance. When this critical velocity is exceeded (the point of incipient
fluidization), the packing in the bed becomes loose, the finer
particles begin to teeterand the pressure drop ceases to increase,
i.e., the resistance to gas flow drops (due to increase in void age at
places where the fines become suspended).
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The mechanism of the softening-melting phenomenais schematically illustrated in previous Figure. It isevident that with the onset of softening, the voidage inthe bed decreases and the bed becomes more
compact (origin of the terminology cohesive). As aconsequence, further indirect reduction of iron oxideby gases becomes increasingly difficult. Upon melting,dripping of molten FeO-containing slag through the
coke layers increases the flow resistance through thecoke slits and the active (i.e. dripping) coke zonebecause of loss of permeability.
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The cohesive zone has the lowest permeability.
Hence, for proper gas flow:
Tsshould be as high as possible
The thickness of the cohesive zone should be
as small as possible. This thickness depends on
the difference between Ts
and Tm
(Tm
- Ts
), and
therefore, the difference should be as low as
possible.
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Gas flow through Granular zone:
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For resistance to gas flow, more important than theparticle diameter is the relative size of the materials inthe bed.In a mixed bed of widely varying particle size, thesmall particles land in the interstices of the large onesand decrease the void age .Starting with large uniform spheres, the void agedecreases as the small ones are introduced and the bedbecomes more and more compact as the proportion ofthe latter increases.The bed is most dense, i.e., the voidage is minimumwhen 60-70 percent of the total volume of theparticles consists of the large ones for about all thecases.
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The m increases on either side of theminimum, i.e., with increasing or decreasingvolume fraction of the small particles(approaching more uniformity of the sizedistribution).The voidage decreases greatly as the ratio ds/d1 decreases.This shows that for a good and uniformpermeability and low resistance to gas flow in amixed bed, the size fractions should be asnarrow as possible.One can easily visualize the adverse effects ofmulti-granular bed of particles of varyingdiameter on the voidage.
A narrow size distribution has the following advantages:
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charge permeability increases and the gas distribution is
more uniform with better utilization of the chemical andthermal energies of the gases;
more even material distribution at the stock level and less
material segregation in the shaft during descent;
gas flow is not impeded if the size ratio is within limits but
at the same time gives rise to a tortuous flow of gases with
continuous changing of flow directions, providing a larger
gas/solid contact time.
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The fraction of iron bearing material below the limiting size is
therefore termed as 'fines'by the blast furnace technologists and
is invariably eliminated by screening at every possible stage.
From the point of view of reduction the maximum top size of
an iron bearing material should be as low as possible, since the
rate of reduction decreases, perhaps exponentially, with
increasing size.
The size range of materials charged in the blast furnace
represents a compromise to give both good stack permeability
and adequate bulk reducibility.
Gas flow in wet zone:
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Wet zones consist of the coke beds in the bosh andbelly regions, i.e. inactive coke zone, active coke zone,and the coke slits in the cohesive zone. Here molten iron and molten slag flow downwardsthrough the bed of coke. This reduces the free crosssection available for gas flow, thus offering greaterresistance, thereby increasing the pressure drop.An extreme situation arises when, at high gasvelocity, the gas prevents the downward flow of liquid.This is known as loading. With further increase in gasvelocity, the liquid gets carried upwards mechanically,causing flooding.
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Scientists have tried to estimate pressure
drop in blast furnace. However, they are
approximate. Moreover, they are only for the
granular zone and coke zones.
The situation in the cohesive zone is very
complex, and reliable theoretical estimates
are extremely difficult to come by.
Therefore for practical applications in blast
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Therefore, for practical applications in blast
furnaces, an empirical parameter, calledFlow
Resistance Coefficient(FRC) has become
popular. The FRC for a bed is given as
where the gas flow rate is for unit cross section
of the bed, i.e. either mass flow velocity or
volumetric flow velocity .
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FRC=1/ bed permeability
The FRC for a furnace can be empirically determined
from measurements of pressure drop and gas flow rate.
Since it is possible to measure pressures at various
heights within a furnace, the values of FRC for individual
zones can also be determined.
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These measurements have indicated that
FRCs for the granular, cohesive, coke +
tuyere zones are approximately 20%, 50%
and 30% of the overall furnace FRC.
This means that the cohesive zone is
responsible for the maximum flow resistance
and pressure drop, to a very large extent.
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Smarajit SarkarDepartment of Metallurgical and Materials Engineering
NIT Rourkela
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Decreasing the extent of SiO formation by:o Lowering ash in coke, and the coke rate
o Lowering RAFT
o Lowering the activity of Si02 in coke ash by lime
injection through the tuyeres.
Decreasing Si absorption by liquid iron in the boshby enhancing the absorption of Si02 by the bosh
slag. This can be achieved by:o Increasing the bosh slag basicity.
o Lowering the bosh slag viscosity..
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Removal of Si from metal by slag-metal reaction atthe hearth by:
o Lowering the hearth temperature
o Producing a slag of optimum basicity and fluidity.
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Desulphurisation of metal droplets through slag-
metal reaction in the furnace hearth :
(CaO) + [S] + [C ]= (CaS) + CO (g)
Desulphurisation through the coupled reaction:
(CaO) +[S] +[ Mn] = (CaS) + (MnO)
(CaO) + [S] + [ Si] = (CaS) + 1/2 (SiOz)
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Sulphur pick-up through the vapour-phase reaction:
CaS( in coke ash) + SiO (g) = SiS(g) +
CaOFeS( in coke ash) + SiO (g) = SiS(g) +CO(g) +[Fe]
In the bosh and belly regions, SiSdecomposes as
SiS(g) = [Si] + [S]
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Reducing slag i.e. FeO content should be low
High basicity
High temperature, since desulphurisation is an
endothermic reaction
Kinetic factor
Contact surface of metal and slag ( by agitation)
Fluidity of slag( by adding MgO , MnO)
Time of desulphurisation
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Smarajit SarkarDepartment of Metallurgical and Materials Engineering
NIT Rourkela
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P=Q/K
Where,
P= Productivity, THM per day
Q= Coke burned, tonnes per day
K= Coke consumed, tonnes per day
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0.8-0.9t0.5-0.6t1.7-1.8t
2500m3
0.6t1t
FuelReducing agent supplyPermeable bed (spacer)
3200m3
+80kg dust
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The efficiency of operation of a blast furnace may be
measured in terms of coke rate which should of course
be as low as possible. The achievement of a satisfactory
coke rate depends on optimising the extent to which the
carbon deposition reaction proceeds. If the top gas is
high in C02 sensible heat is carried from the furnace as a
result of the exothermic reaction.
2CO=CO2+C
If on the other hand the top gas is high in CO, chemical
heat leaves the furnace.
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I d C ib i %
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CO2emission
Industry Contribution %Power 51Transport 16Steel 10other 23
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The purpose of HTP is to introduce more
oxygen to burn more carbon by blowing moreair and at the same time maintaining thelinear gas velocity (and pressure drop)identical to that in the conventional practice
without any formation of channels,maldistribution of gas, increase in coke rateor flue dust emission
Advantages:
For the same volume flow rate, a greater mass of air(hence, oxygen) can be blown with HTP; higheroutput;
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A major benefit that is so obvious is increased
production rate because of increased time of contact of
gas and solid as a result of reduced velocity of gases
through the furnace. Increased pressure also increases
the reduction rate of oxide;
Suppression of Boudouard reaction (C02 + C= 2CO) and
hence savings in fuel;
More uniform distribution of gas velocity and reduction
across furnace cross-section; smoother furnace
operation due to increased permeability;
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less flue dust losses, less variation of coke input, better
maintenance of the thermal state of the hearth, more
uniform iron analysis;
More uniform operation with lower and more consistent
hot metal silicon content have been claimed to be the
benefit of high top pressure;
Bhilai Steel Plant (operative), RSP yet to implement
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SiO2 +C ={SiO} +{CO}
From above equation it can be seen that partial
pressure of SiO can be brought down by increasing
the partial pressure of CO; in other words the SiO2
reduction reaction can be discouraged by application
of top pressure which enables a higher blast pressure
and hence an increase in partial pressure of CO.
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'raceway adiabatic flame temperature
This is the highest temperature available inside thefurnace. There is temperature gradient in vertical
direction on either side of this zone. This temperature
is critically related to the hearth temperature known as
operating temperature of the furnace. It is equally
related to the top gas temperature such that the hot
raceway gasses have to impart their heat to the
descending burden to the extent expected and leave
the furnace as off-gases at the desired temperature.
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The primary purpose of using injectants with the
blast is profitability which depends upon the
relative price of coke and injectants and the
amount of coke that can be saved per unit of the
latter, i.e., upon the replacement ratio:
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H20 + C = CO + H2(1)
HO (1200C) = + 2700 kcal/kg C
Presence of moisture in the blast generates double the
volume of reducing gas per mole of carbon burnt. As per
Eq.1 for every carbon burnt one mole of CO and an
additional mole of hydrogen will be available as product
of burning of coke for reduction in bosh and stack.
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The more the moisture the more will be this additional
hydrogen available. Kinetically hydrogen reduction of iron oxide is faster than
that by CO because of its small size. Presence of
moisture helps to burn coke at a faster rate with its
attendant favorable effects.
Some of the endothermic heat of moisture disintegration
is compensated by way of exothermic reduction of iron
oxide by hydrogen.
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higher gasifying power which intensifies coke
consumption In the raceway;
smoothens the temperature gradient and facilitates stock
descent ;
enlarges the combustion zone and accelerates stock
descent; heats up the axial zone; maintains thermal
state of the hearth;
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even with incomplete temperature compensation, the coke
rate may not rise because of higher reducing power and
higher heat transfer coefficient of hydrogen;
decreases pressure loss due to lower density and viscosity ofhydrogen.
The blast pressure may drop even by0.1-0.2 atm. which
means the furnace can be blown at a higher blast rate.
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It has been estimated that for an increase of 20 g/Nm3
moisture in the blast the endothermicity can be
compensated by a rise of 200C in the blast preheat.
By increasing moisture and compensating it by
additional rise of preheat means that cheaper heat
energy can be used to feed the furnace and thereby
decrease the coke consumption and economise the
operation.
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Oxygen enrichment of the blast and moisture enrichment
have quite opposite thermal effects. The two can be saddled
together to obtain better inputs.
Hot blast temperature, extent of oxygen enrichment and
humidification of blast have to be adjusted as interrelated
parameters simultaneously to obtain optimum conditions of
operation for maximum benefits such as minimum coke rate,
higher productivity and so on.
Th f th i j ti f l h b
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The reasons for the injection of coal have beeneconomic as well as operational flexibility and include
the following: After the steep rise in oil prices following the oil
crisis, iron makers were compelled to abandonheavy oil injection and were looking for a less
expensive auxiliary fuel. PCI accommodate shortages of coking capacity, by
replacing coke by coal in the blast furnaces. After athorough investment analysis, it has been found that
a reliable coal injection system requires much lowercapital cost and involves operating cost than theextension of coking capacity.
.
C l l d ti i fl
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Coal causes a lower reduction in flametemperature per unit injection than oil or natural
gas. It, therefore, allows more scope for blasttemperature adjustment/oxygen enrichment forincreased rates of injection and consequently, lesscoke consumption.
The PCI system design is capable ofinjecting coal on a continuous and stable basis and
ensure accurate and uniform distribution
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The coke savings from fluxed burden emanate from the following
causes :
better reducibility and enhanced indirect reduction (6-7 kg.C saved
from every 1 percent increase in indirect reduction);
use of higher blast temperatures because the thermal load is
smaller and the slag is pre-made; the primary slag melts at higher
temperatures and does so within a vertically narrow softening zone;
avoidance of carbon dioxide generated from limestone in the stack
which adversely affects indirect reduction;
transference of heat of calcination from the furnace to the
agglomerating plant.
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This is a unique design in which
large bell is replaced by a distributor
chute with 2 hoppers
A rotating chute is provided inside
the furnace top cone
Advantages: Greater charge distribution
flexibility
more operational safety and
easy control over varyingcharging particles
Less wearing parts: easy
maintenance
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Smarajit SarkarDepartment of Metallurgical and Materials Engineering
NIT Rourkela
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Oxidation of carbon:Bottom blowing increases sharply the
intensity of bath stirring and increases the area of gas-metal
boundaries (10-20 times the values typical of top blowing) .
Since the hydrocarbons supplied into the bath together
with oxygen dissociate into H2, H2O and CO2gas
bubbles in the bath have a lower partial pressure of
carbon monoxide (Pco )
All these factors facilitate substantially the formation
and evolution of carbon monoxide, which leads to a
higher rate of decarburization in bottom blowing
The degree of oxidation of metal and slag
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The degree of oxidation of metal and slag
Removal of phosphorous: Since the slag ofthe bottom-blown converter process have alow degree of oxidation almost during thewhole operation, the conditions existing
during these periods are unfavorable forphosphorus removal
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A small amount of inert gas, about 3% of the volume of oxygen
blown from top, introduced from bottom, agitates the bath so
effectively that slopping is almost eliminated.
However for obtaining near equilibrium state of the system
inside the vessel a substantial amount of gas has to be
introduced from the bottom.
If 20-30% of the total oxygen, if blown from bottom, can cause
adequate stirring for the system to achieve near equilibrium
conditions. The increase beyond 30% therefore contributes
negligible addition of benefits.
However at 30% oxygen blowing from bottom leads to formation of
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However at 30% oxygen blowing from bottom leads to formation of
very dry slag and possibility of its ejection during refining unless it is
accompanied by lime also.
The more the oxygen fraction blown from bottom the less is the post
combustion of CO gas and consequently less is the scrap ons of
processing.Blowing of inert gas from bottom has a chilling effect on bath and
hence should be minimum. On the contrary the more is the gas blown
the more is the stirring effect and resultant better metallurgical results.
A optimum choice therefore has to be made judiciously.
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The more the oxygen fraction blown from bottom the
less is the post combustion of CO gas andconsequently less is the scrap consumption in the
charge under identical conditions of processing.
Blowing of inert gas from bottom has a chilling effect
on bath and hence should be minimum. On the
contrary the more is the gas blown the more is the
stirring effect and resultant better metallurgical results.
A optimum choice therefore has to be made
judiciously.
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Better mixing and homogeneity in the bath offer the following
advantages:
Less slopping, since non-homogeneity causes formation of
regions with high supersaturation and consequent violentreactions and ejections.
Better mixing and mass transfer in the metal bath with closer
approach to equilibrium for [C]-[O]-CO reaction, andconsequently, lower bath oxygen content at the same carbon
content
Better slag-metal mixing and mass transfer andconsequently closer approach to slag metal equilibrium
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consequently, closer approach to slag- metal equilibrium,leading to:
o lower FeO in slag and hence higher Fe yieldo transfer of more phosphorus from the metal to the slag
(i.e. better bath dephosphorisation)
o transfer of more Mn from the slag to the metal, and
thus better Mn recoveryo lower nitrogen and hydrogen contents of the bath.
More reliable temperature measurement and samplingof metal and slag, and thus better process control
Faster dissolution of the scrap added into the metal bath
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As compared to top blowing, the hybrid blowingeliminates the temperature and concentration
gradients and effects improved blowing control, less
slopping and higher blowing rates. It also reduces overoxidation and improves the yield. It leads the process
to near equilibrium with resultant effective
dephosphorisation and desulphurisation and ability tomake very low carbon steels.
What is blown from the bottom, inert gas or oxygen?
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How much inert gas is blown from the bottom?
At what stage of the blow the inert gas is blown,although the blow, at the end of the blow, after theblow ends and so on?
What inert gas is blown, argon, nitrogen or theircombination?
How the inert gas is blown, permeable plug, tuyere,etc.?
What oxidising media is blown from bottom, oxygen orair?
If oxygen is blown from bottom as well then how muchof the total oxygen is blown from bottom ?
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The processes have been developed to obtain the combined advantages of
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both LD and OBM to the extent possible. Therefore the metallurgical
performance of a hybrid process has to be evaluated in relation to these
two extremes, namely the LD and the OBM. The parameters on which this
can be done are :
Iron content of the slag as a function of carbon content of bath
Oxidation levels in slag and metal
Manganese content of the bath at the turndown
Desulphurisation efficiency in terms of partition coefficient
Dephosphorisation efficiency in terms of partition coefficient
Hydrogen and nitrogen contents of the bath at turndown
Yield of liquid steel
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The advantages of continuous casting (over ingotcasting) are:
It is directly possible to cast blooms, slabs andbillets, thus eliminating blooming, slabbing mills
completely, and billet mills to a large extent. Better quality of the cast product.
Higher crude-to-finished steel yield (about 10 to20% more than ingot casting).
Higher extent of automation and process control.
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Solidification must be completed before the withdrawal
rolls.
The liquid core should be bowl-shapedas shown in theFigure and not pointed at the bottom (as indicated by thedotted lines), since the latter increases the tendency forundesirable centerline (i.e. axial) macro-segregation andporosity
The solidified shell of metal should be strong enough at
the exit region of the mould so that it does not crack orbreakoutunder pressure of the liquid.
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The surface area-to-volume ratio per unit length ofcontinuously cast ingot is larger than that for ingotcasting. As a consequence, the linear rate ofsolidification (dx/dt) is an order of magnitude
higher than that in ingot casting.
The dendrite arm spacing in continuously castproducts is smaller compared with that in ingot
casting.
Macro segregation is less and is restricted to the
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Macro-segregation is less, and is restricted to the
centreline zone only. Endogenous inclusions are smaller in size, since they
get less time to grow. For the same reason, the blow
holes are, on an average, smaller in size. Inclusions get less time to float-up. Therefore, any
non-metallic particle coming into the melt at the later
stages tends to remain entrapped in the cast product.
In addition to more rapid freezing, continuous castingdiffers from ingot casting in several ways. These arenoted below.
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Mathematically speaking, continuously cast ingot isinfinitely long. Hence, the heat flow is essentially in the
transverse direction, and there is no end-effect as is the
case in ingot casting (e.g. bottom cone of negative
segregation, pipe at the top, etc.).
The depth of the liquid metal pool is several metres long.
Hence, the ferrostatic pressure of the liquid is high
during the latter stages of solidification, resulting in
significant difficulties of blow-hole formation.
Since the ingot is withdrawn continuously from the mould, the frozen
layer of steel is subjected to stresses. This is aggravated by the
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stresses arising out of thermal expansion/ contraction and phase
transformations.
Such stresses are the highest at the surface. Moreover, when the
ingot comes out of the mould, the thickness of the frozen steel shell
is not very appreciable. Furthermore, it is at around 1l00-1200C, andis therefore, weak. All these factors tend to cause cracks at the
surface of the ingot leading to rejections.
Use of a tundish between the ladle and the mould results in extratemperature loss. Therefore, better refractory lining in the ladles,
tundish, etc. are required in order to minimise corrosion and erosion
by molten metal.
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Smarajit SarkarDepartment of Metallurgical and Materials Engineering
NIT Rourkela
P i l ki i i d f l i
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Primary steelmaking is aimed at fast melting
and rapid refining. It is capable of refining ata macro level to arrive at broad steelspecifications, but is not designed to meetthe stringent demands on steel quality, and
consistency of composition and temperaturethat is required for very sophisticated gradesof steel. In order to achieve suchrequirements, liquid steel from primary
steelmaking units has to be further refined inthe ladle after tapping. This is known asSecondary Steelmaking.
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improvement in quality improvement in production rate
decrease in energy consumption
use of relatively cheaper grade oralternative raw materials
use of alternate sources of energy
higher recovery of alloying elements.
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Lower impurity contents . Better cleanliness. (i.e. lower inclusion
contents)
Stringent quality control. (i.e. less variationfrom heat-to-heat)
Microalloying to impart superior properties.
Better surface quality and homogeneity inthe cast product.
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The term clean steelshould mean a steel
free of inclusions. However, no steel canbe free from all inclusions.
Macro-inclusions are the primary harmfulones. Hence, a clean steel means acleanersteel, i.e., one containing a much
lower level of harmful macro-inclusions.)
I i i i di id
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In practice, it is customary to divide
inclusions by size into macro inclusionsandmicro inclusions. Macro inclusions ought tobe eliminated because of their harmfuleffects. However, the presence of microinclusions can be tolerated, since they donot necessarily have a harmful effect on theproperties of steel and can even bebeneficial. They can, for example, restrictgrain growth, increase yield strength and
hardness, and act as nuclei for theprecipitation of carbides, nitrides, etc.
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The critical inclusion size is not fixed butdepends on many factors, including servicerequirements.
Broadly speaking, it is in the range of 5 to 500
m (5 X 10-3 to 0.5 mm). It decreases with anincrease in yield stress. In high-strength steels,its size will be very small.
Scientists advocated the use of fracturemechanics concepts for theoretical estimation ofthe critical size for a specific situation.
Precipitation due to reaction from molten steel or during
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Precipitation due to reaction from molten steel or duringfreezing because of reaction between dissolved oxygen
and the deoxidisers, with consequent formation of oxides(also reaction with dissolved sulphur as well). These areknown as endogenous inclusions.
Mechanical and chemical erosion of the refractory lining
Entrapment of slag particles in steel Oxygen pick up from the atmosphere, especially during
teeming, and consequent oxide formation.
Inclusions originating from contact with external sourcesas listed in items 2 to 4 above, are called exogenousinclusions.
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With a lower wettability (higher value of Me inc),
an inclusion can be retained in contact with the
metal by lower forces, and therefore, can break
off more easily and float up in the metal. On the
contrary, inclusion which are wetted readily by the
metal, cannot break off from it as easily.
Carryover slag from the furnace into the ladle
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Carryover slag from the furnace into the ladleshould be minimised, since it contains high
percentage of FeO + MnO and makes efficientdeoxidation fairly difficult.
Deoxidation products should be chemically
stable. Otherwise, they would tend todecompose and transfer oxygen back into liquidsteel. Si02 and Al203 are preferred to MnO.Moreover the products should preferably beliquid for faster growth by agglomeration and
hence faster removal by floatation. Complexdeoxidation gives this advantage.
Stirring of the melt in the ladle by argon flowing throughbottom tuyeres is a must for mixing and homogenisation,
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faster growth, and floatation of the deoxidation products.
However, very high gas flow rates are not desirable fromthe cleanliness point of view, since it has the followingadverse effects:
o Too vigorous stirring of the metal can cause
disintegration of earlier formed inclusion conglomerates.o Re-entrainment of slag particles into molten steel.
o Increased erosion of refractories and consequent
generation of exogenous inclusions.
o More ejection of metal droplets into the atmosphere with
consequent oxide formation.
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The varieties of secondary steelmakingprocesses that have proved to be ofcommercial value can broadly be categorisedas under:
Stirring treatments Synthetic slag refining with stirring Vacuum treatments
Decarburisation techniques Injection metallurgy
Plunging techniques
Post-solidification treatments.
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Ladle degassing processes (VD, VOD, VAD)
Stream degassing processes
Circulation degassing processes (DH and RH).
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Why RH-OB Process?
To meet increasing demand for cold-rolled steel sheets with improved
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mechanical properties, and to cope with the change from batch-type to
continuous annealing, the production of ULC steel (C < 20 ppm) is
increasing.
A major problem in the conventional RH process is that the time
required to achieve such low carbon is so long that carbon content at
BOF tapping should be lowered. However, this is accompanied by
excessive oxidation of molten steel and loss of iron oxide in the slag.
It adversely affects surface the quality of sheet as well.
Hence, decarburization in RH degasser is to bespeeded up. This is achieved by some oxygenblowing (OB) during degassing.
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The RH-OBprocess, which uses an oxygenblowing facility during degassing, was originallydeveloped for decarburization of stainless steel byNippon Steel Corp., Japan, in 1972.
Subsequently, it was employed for the manufactureof ULC steels.
The present thrust is to decrease carbon contentfrom something like 300 ppm to 10 or 20 ppmwithin 10 min. Cont
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Simplified by Hiltey and Kaveney
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