Iron making

Smarajit Sarkar
Department of Metallurgical and Materials Engineering
NIT Rourkela

Ahindra Ghosh and Amit Chatterjee: Ironmaking and Steelmaking Theory and Practice, Prentice-
Hall of India Private Limited, 2008
Anil K. Biswas: Principles of Blast Furnace Ironmaking, SBA Publication,1999
R.H.Tupkary and V.R.Tupkary: An Introduction to Modern Iron Making, Khanna Publishers.
R.H.Tupkary and V.R.Tupkary: An Introduction to Modern Steel Making, Khanna Publishers.
David H. Wakelin (ed.): The Making, Shaping and Treating of Steel (Ironmaking Volume), The
AISE Steel Foundation, 2004.
Richard J.Fruehan (ed.): The Making, Shaping and Treating of Steel (Steeelmaking Volume),
The AISE Steel Foundation, 2004.
A.Ghosh, Secondary Steel Making – Principle & Applications, CRC Press – 2001.
 R.G.Ward: Physical Chemistry of iron & steel making, ELBS and Edward Arnold, 1962.
 F.P.Edneral: Electrometallurgy of Steel and Ferro-Alloys, Vol.1 Mir Publishers,1979
 B. Ozturk and R. J. Fruehan,: "Kinetics of the Reaction of SiO(g) with Carbon Saturated Iron":
Metall. Trans. B, Vol. 16B, 1985, p. 121.
B. Ozturk and R. J. Fruehan: "The Reaction of SiO(g) with Liquid Slags,” Metall. Trans.B,
Volume 17B, 1986, p. 397.
B. Ozturk and R. J. Fruehan:”.Transfer of Silicon in Blast Furnace": , Proceedings of the fifth
International Iron and Steel Congress, Washington D.C., 1986, p. 959.
P. F. Nogueira and R. J. Fruehan:” Blast Furnace Softening and Melting Phenomena - Melting
Onset in Acid and Basic Pellets", , ISS-AIME lronmaking Conference, 2002, pp. 585.

There are as many as two thousand odd
varieties of steels in use. These specifically
differ in their chemical composition. However, a
couple of hundred varieties are predominantly in
use. The chemical composition of steels broadly
divide them into two major groups, viz. (i) plain
carbon steels and (ii) alloy steels.

The plain carbon steels are essentially alloys of
iron and carbon only whereas, if one or more of
elements other than carbon are added to steel in
significant amounts to ensure specific better
properties such as better mechanical strength,
ductility, electrical and magnetic properties,
corrosion resistance and so on it is known as an
alloy steel. These specifically added elements
are known as alloying additions in steels.

Steels may contain many other elements such as AI, Si,
Mn, S, P, etc. which are not added specifically for any
specific purpose but are inevitably present because of
their association in the process of iron and steelmaking
and can not be totally eliminated during the known
process of iron and steelmaking. These are known as
impurities in steel.
 Every attempt is made to minimise them during the
process of steelmaking but such efforts are costly and
special tech-niques are required for decreasing their
contents below a certain level in the case of each
element.

For cheaper variety of steels therefore their contents at
high levels are tolerated. These high. levels are however
such that the properties of steels are not signifi-cantly
adversely affected. These tolerable limits of impurities
are considered as 'safe limits' and the impurity levels are
maintained below these safe limits.
 For example, for ordinary steels sulphur contents up to
0.05% are tolerable ,whereas for several special steels
the limit goes on decreasing to as low as 0.005% or
even lower. For most high quality steels now the total
impurity level acceptable is below 100 ppm and the aim
is 45 ppm.

Plain carbon steels are broadly sub-divided into four
major types based on their carbon contents. These are
not strict divisions based on carbon contents but are
generally broad divisions as a basis of classification.
This division is definitely useful. These are:
(i) Soft or low carbon steels up to 0·15% C
(ii) Mild steels in the range 0·15-0·35% C
(iii) Medium carbon steels in the range 0·35-0·65% C
(iv) High carbon steels in the range 0·65-1·75% C

The alloy steels are broadly sub-divided into three groups
on the basis of the total alloying elements present. This
division is also only a broad division and not a rigid one.
This is :
(i) Low alloy steels up to 5% total alloying contents
(ii) Medium alloy steels 5-10% total alloying
(iii) High alloy steels above 10% total alloying

B.F. process is the first step in Producing Steel From
Iron Oxide.
This Would remain so probably at least for the first
quarter of the century despite

◦ Speedy depletion of Coking coal reserves
◦Enhanced adoption of alternate routes for iron making for
ultimate conversion to steel.

The B.F. works on a counter current principle
Ascending hot gases meet Descending solid charge
The charge includes Iron bearing materials (ore, sinter,
pellets), coke & flux (Lime stone, Dolomite)
The ascending gases cause reduction of Iron oxide in
the Iron bearing materials while progressively heating
it.
The result is Production of
◦ Liquid slag
◦Liquid Metal
◦B.F. Gas of considerable calorific value

All the reduced elements join the metal. A typical
composition of the Metal (Iron) produced in Blast
Furnace is presented below.

The Slag is a low melting chemical compound formed by the
chemical reaction of the gangue and the flux in the charge.
All unreduced ones join the slag
The major constituents of the slag include the following
◦Al
2O
3 – 20.45%
◦CaO – 32.23%
◦SiO
2
– 33.02%
◦MgO – 9.95%
◦S – 0.89%
◦MnO – 0.54%
◦TiO
2 – 1.01%
◦FeO – 0.41%
◦K
2
O+Na
2
0 – 1%
◦Trace Oxides – 0.5%
(Curtsey TATA STEEL)

Smarajit Sarkar
Department of Metallurgical and Materials Engineering
NIT Rourkela

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.

cont…

 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 all
places from the hearth to the top
 cont…

 Only then, an optimum utilization of the
chemical and thermal energies of the
gases as well as a smooth, uniform and
maximum iron production with minimum
coke rate will be realized.

oIn 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 the
charge.
Consumption of foundry iron in the works.
Losses of iron in the ladle and the casting
machine.
The number of furnaces to be installed

Stock line: The distribution pattern at the top.
Charge or stock level in the furnace throat
The materials or the stock or the burden should
be properly distributed for uniform distribution of
the ascending gas.
Zero stock line: Horizontal plane formed by
bottom of big bell when closed. 6ft stock level for
instance located 6ft below zero stock line.

 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:Advantages:
 Greater charge distribution
flexibility
 more operational safety and
easy control over varying
charging particles
 Less wearing parts: easy
maintenance

The advantages accruing from improved distribution
control can be summarised as follows:
Increased productivity, decreased coke rate, improved
furnace life .
Reduced refractory erosion
Improved wind acceptance and reduced hanging as well
as slips
Improved efficiency of gas utilisation and its indirect
reduction
Lower silicon content in hot metal and consistency in the
hot metal quality
Reduced tuyere losses and minimisation of scaffold
formation
Lower dust emission owing to uniform distribution of fines.


As has been made clear that even the most efficient of the
modern blast furnace would produce an effluent gas containing a
significant proportion of CO which could not be used for iron
oxide reduction. The actual CO content may vary around 20-30%
by volume. This has a calorific value of nearly 900 kcal/m
3
. The
quantity of gas produced depends upon the amount of fuel burnt.
For one tonne of coke burnt nearly 4000 m
3
of effluent gas may
be produced. Hence a blast furnace requiring 1000 t of coke per
day would generate nearly 4 x 10
6
m
3
of gas with a total energy
content of 3600 x 10
6
kcal which is nearly equivalent to 500 t of
coke.

The effluent gas from the furnace cannot directly be
used as a fuel since a substantial quantity of dust from
the burden is also discharged along with. It may lead
to accumulation of dust and wear in the equipment
using the gas. The gas is, therefore, cleaned before its
use and in so doing the sensible heat of the gas is
invariably lost. The chemi-cal heat of the cleaned gas
is what is utilised.

The average dust content may vary in the range of 7-30 g/m
3
. In
general cleaning is carried out in three stages viz. coarse, semi-fine
and fine cleaning. The coarse cleaning is done in dust catchers and
cyclones in dry condition. The dust content of the coarse cleaned gas
is nearly 5-10 g/m
3
. The semi-fine cleaning is carried out in
scrubbers, ventury washers, cyclone separators, centrifugal
disintegrators, feld washers or even in electrostatic precipitators. The
dust content is thereby reduced to 0·5-1·5 g/m
3
. Fine cleaning is
carried out mainly by electrostatic precipitators or at times by high
speed rotary disintegrators, The dust content is thereby reduced
down to 0.01 g/m
3
The semi-fine and fine cleaning is carried out either
in wet or dry condition. Wet methods are generally preferred to dry
methods for their better efficiency and smooth working.



Two adjacent uptakes are joined together to form one single duct
and the two such ducts, thus formed, are connected to form only
one duct which carries the gas downwards into the dust catcher.
The downcoming pipe or duct is called downcomer.
A bleeder valve is a safety device, which opens automatically or is
opened, to release extra pressure developed inside the furnace and
thereby eliminate the danger of explosion.
The uptakes and the downcomers are steel pipes and are lined
from inside with firebricks. The sizes of the uptakes and
downcomers and the angle of their joints are such that gas flows out
of the furnace smoothly without any hindrance.

The uptakes should be located on the furnace-top
periphery at those points which are not directly vertically
above the iron-notch, slag notch, blast main entrance to
the bustle pipe, etc. These are active points of the
furnace and if the uptakes are located right above these
points it may cause uneven distribution of the gas
through the burden. The entire design should also
ensure that minimum of dust is carried form the furnace
with the gases.

It essentially consists of a tall cylindrical structure
comprising of a combustion chamber and heat
regenerator unit of checker bricks. The clean blast
furnace gas is burnt in the combustion chamber
and the hot products of combustion later heat up
the checker bricks. In this case the stove is said to
be on 'on-gas' and is maintained on gas until the
checker bricks are heated to a certain
temperature.

Firing is stopped and cold blast is passed through
checkers which impart the heat stored in them and
there by produce preheated blast. The stove is
said to be 'on blast'. It can continue heating the
blast till a certain minimum temperature of the
blast is obtainable. The stove is again put on gas
and the cycle is repeated.

The stove design and the number of stoves, employed
should ensure a steady supply of preheated blast to the
furnace. This duty demands that the amount of heat
generated by way of combustion of gas per unit time
should be adequate to heat up the required amount of
blast to the required temperature per unit time, taking
into account the usual efficiency of heat transfer via
checker system and the usual heat losses from the
system.

The thermal efficiency of the stove varies between
75-90%. The checker work cools more rapidly
whereas it takes longer time to heat it up. In practice
a stove may be on gas for 2-4 hours and on blast for
1-2 hours. For an uninterrupted steady supply of
blast at specified temperature therefore a battery of
at least three stoves is necessary. A two stove
system is quite unsatisfactory and hence three or
four stove system is preferred.

The checkerwork has to absorb maximum heat at faster rate while
heating and should desorb heat equally rapidly to the incoming cold
blast. The larger the weight of bricks the more will be its heat
storing capacity. The larger is the surface area exposed as flues
the faster is the heat exchange with gas. The bricks should have
maximum weight with maximum surface area of flues i.e. maximum
openings to allow free passage of gases. It has been found that a
ratio of weight of bricks in kilogram to heating surface in
square metres of about 5-6 in minimum. Below this struc-tural
difficulties may arise.

The checker bricks are supported on steel grids which in turn
are supported by cast iron or steel columns. Since the
maximum temperature during combustion is generated near
the dome and since the top portion of checker bricks have to
stand higher temperatures, with progressively decreasing
value downwards, the quality of checker bricks used also very
accordingly. Heavy duty fire-bricks are essential for dome
construction. The top 3-6 m height of the checkers is made up
of higher alumina bricks or semi-silica bricks while the
remainder as of good quality firebricks.

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 m
3
/ton of pig iron

V =k D
2
H
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.

Total height = useful height +distance between
stock line and the charging platform (it is
governed by the construction of gas off-take and
charging platform, this dimensions varies from 3
to 4m.)
Useful height= height from the tapping hole to
the stock line.
The height of the blast furnace is mainly
governed by the strength of the raw materials,
particularly that of coke.
cont… …

The strength of the coke charged to the
furnace should be sufficient to withstand the
load of raw materials without getting
crushed. Coke provides permeability(in the
dry as well as wet zones )and also
mechanical support to the large charge
column, permitting the gases to ascend
through the voids.
Total height (H)= 5.55V
0.24
Useful height (H
0
) =0.88×H

Diameter:
The belly /bosh parallel is the cylinder that
connects the tapers of the shaft and the bosh.
Its diameter, d
bll
, and the ratio of this diameter to
the useful or inner height of the furnace as well
as to the diameter of the hearth play an
important role in the operation of the furnace.
The correct descent of the stock, ascent of the
gas and efficient utilization of the chemical and
thermal energies of the gas depend greatly upon
these ratios.

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.

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.
d
belly
=0.59 ×(V)
0.38

H
belIy = 0.07×H

The hearth is designed such that its volume
between the iron notch and tuyeres is sufficient
to 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× V
0.45

A very approximate relationship between
the coke burning rate and hearth diameter
is given by the following equation:
D = c Q
0.5
D = hearth diameter, m
Q = coke throughput, tonnes/24h
c = throughput coefficient which varies
between 0.2-0.3 depending upon burden
preparation.

For highly prepared burden, the value of
c = 0.2 has been achieved in modern large
furnaces .
There-fore, for a furnace planned to
produce 10,000 THM per day with a coke
rate of 500 kg/THM, i.e., a coke throughput
of 5,000 tonnes per day, the hearth
diameter should be about 14.1 m.
The value will be 21.2 m if the value of
c=0.3.

With increasing diameter of the hearth,
the gas penetration must be ensured
by providing adequate bed
permeability with the use of
mechanically strong, rich, pre-fluxed
burden of uniform size and low slag
bulk as well as strong lumpy coke.
The Hearth height should be 10% of the
total height of the furnace

The shaft height must be sufficient to allow the
heating, preparation and reduction of ore before
the burden reaches the bosh. In the upper
regions of the shaft , volume changes due to
increase in temperature and carbon deposition.
These demand an outward batter for smooth
flow of materials. In the lower region of the
shaft , the material starts fusing and tends to
stick to the furnace wall. So to counteract the
wall drag an outward butter is necessary.

Stack height H
stack
= 0.63 H- 3.2 m
Stack angle
The stack angle usually ranges from 85
0
to

87
0
(i) 85
0
for weak and powdery ores;
(ii) 86
0
for mixture of strong and weak, lumpy or
fine ores;
(iii) 87
0
for strong, lumpy ore and coke.

The variations in the angles are necessary
for obtaining an adequate peripheral flow
which is an essential pre-requisite for
forcing of the blast furnace.
Since the ore hump is located in the
intermediate zone and it moves almost
vertically downwards pushing the lighter
coke towards the wall and the axis.
 A smaller shaft angle in the case of weak
and powdery ore helps to loosen the
periphery.

Stack angle can be calculated from the formula
Stack angle (α)= Cot
-1
(D-d
1
/2xStack Height)
Where, D= Bosh parallel Diameter
d
1= Throat Diameter
Bosh angle can be calculated from the formula
Bosh angle (β)= Cot
-1
(D-d/2xBosh Height)
Where, D= Bosh parallel Diameter
d= Hearth Diameter

When the raw materials are charged into the
blast furnace, little volume change takes
place for a few meters of their descent and
hence the walls of the throat are generally
parallel
Throat diameter can not be too small as it
has to allow the enormous volume of the gas
to pass through at a reasonably low velocity
to maintain adequate solid gas contact and
to decrease the dust emission, throat
hanging and channeling.
 Cont..

Throat diameter can not be too wide as it
may compact the charge. A certain
velocity and lifting power of gas is
necessary for losening the charge at top.
Throat Diameter d
throat =0.59 V
0.35
Where, V= useful volume

A considerable amount of slag and iron
descends to the hearth through the inter-tuyere
zones. If they do so without having been
adequately 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.


The distance between the adjacent tuyeres
around the hearth circumference should be such
as to obtain, as far as possible, a merging of the
individual combustion zones of each tuyere into
a continuous ring.

The number of tuyeres mainly depend upon the
diameter of the hearth. The diameter of the
tuyeres depend upon the blast volume.
The following formulae can be used to determine
the number of tuyeres
Pavlov: n = 2d +1
Rice:n = 2.6d-0.3
Tikhomirov et al : n = 3d-8
Where n= Number of tuyeres,
d=hearth diameter

Capacity →
(THM/Day)
Parameter↓
2000 3000 5000
Useful Volume (m
3
) 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 (m
2
) 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

Richness: Richness means the percentage of
metallic iron in the ore. e.g. In order to produce a
tonne of pig iron about1.5tonnes of ore is required
in Australia (68% Fe), about 2 tonnes are required
in India (55-60%) and nearly 3 tonnes are required
in U.K. (30-35%)
Composition of the gangue : The
composition of gangue associated with an ore
may reduce the value of an otherwise rich ore or
in some case may even enhance that of a lean
ore.

e.g. Value of an ore is drastically reduced by the
presence of alkali oxides , reduced to some extent
by the presence of alumina and is in fact
enhanced by the presence of lime and/or
magnesia.
Location: The location of an ore, both
geographical and geological, is very important
Treatment and preparation needed
before smelting

Cold strength
Porosity
Decrepitation
Low-temperature breakdown under reducing
conditions (LTB)
Hot compression strength
Softening temperature and range
Swelling and volume change
High-temperature bed permeability under
compressive load and reducing conditions.

Cold strength measurement comprises of tumbler or
drum test for abradibility, shatter test for impact and
compression test for load during storage.
Tumbler or drum test: It measures the susceptibility of
ferrous materials (coke as well) to breakage due to
abrasion during handling, trans-portation, charging on to
the blast furnace bells as well as inside the furnace itself.
In this test, a certain weight of the material within a
selected size range is rotated in a drum of given size for
a given time with certain number of revolutions.

The abrasion strength is given by the percentage
weight of + 6.3 mm surviving the test and dust
index by the percentage of - 0.6 mm. For good
pellets the respective percentages are 85-95 and
3-7, for sinters 60-80 and 5-10 and for ores they
vary greatly, 60-95 and 2-25.

In order to minimize the amount of fines delivered to the
furnace, a practice attracting an interest is to deliberately
subject the materials, especially coke and sinter, to
mechanical breakdown and stabilize the charge, e.g., by
means of vibrating screens. They break where the bonds
are weak and the undersize screened out.
 However, it cannot be helped if any fines are generated
between charging into the skip car and then into the
furnace.


•Shatter test: It measures the susceptibility to breakdown due to
impact during loading, unloading and charging into the furnace.
• In this test a certain weight of material is allowed to fall on a steel
plate from a certain height for a pre-determined number of times
and the amount of undersize measured. For strong sinters the
percentage +10mm surviving is above 80.
Compression test: It is used mainly for pellets. Pellets, unreduced
or reduced to various degrees, are subjected to compressive load at
ambient or high temperatures and the percentage of + 5 mm yield
measured and correlated with blast furnace performance.

Porosity: While ores and pellets possess mostly open pores, in
sinters there are macro- and micro-pores as well as open and
closed pores (cut off from outside and cannot be reached by
gas).
True porosity and hence closed porosity can be determined from
open porosity which can be measured from the true and bulk
densities.
Although reducibility increases with increasing open porosity, the
latter changes continuously during reduction on load. Generally,
a high initial porosity results in earlier softening of the material.

Decrepitation : When iron bearing materials are suddenly
exposed to the ex-haust gas temperature at the stock level on
charging, breakdown may occur due to thermal shock. This is
known as decrepitation.
•Experimentally it is measured by dropping a known weight of
material in a furnace previously heated to a temperature level
of 400-600°C, under normal atmosphere, inert atmosphere or
under mildly reducing conditions. After the charge attains the
temperature it is removed, cooled and sieved to measure the
breakdown.

•In a typical test 500 g of 20-40 mm size undried ore is
dropped in a furnace previously heated to a temperature
level of 400°C and retained there for 30 min under a flow
rate of 5000 litres of nitrogen per hour. The sample is
then removed, cooled and the percentage of 0·5 mm and
-5·6 + 0·5 mm material in the product is determined by
sieving.
•It is believed that ores with more than 10% porosity will
not decrepitate.

Low-Temperature Breakdown Test (L.T.B.T.)
It has been observed in the experimental blast furnace that the iron
bearing materials do disintegrate at low temperatures under mildly
reducing conditions, that is in the upper part of the stack, affecting
the furnace permeability and consequently the output adversely. It is
believed that deposition of carbon in this region of the stack is also a
contributory factor although with sinters the breakdown has been
associated with the presence of micro-cracks.
In essence the test consists of subjecting the charge to static bed
reduction at low temperatures in a rotating furnace for a fixed dura-
tion. The percentage of fines generated is quoted as the
L. T.B. T. index.

Reducibility is the ease with which the oxygen
combined with iron can be removed indirectly.
A higher reducibility means a greater extent of
indirect reduction that may be obtained in the
blast furnace resulting in a lowered coke rate
and higher productivity.

Reducibility of ferrous materials is characterized by their
fractional oxygen removal rates in gaseous reducing
atmosphere. The percent degree of reduction or
percent fractional oxygen removal is given by
Where
n
0
= number of moles of oxygen originally combined with iron only;
n = number of moles of oxygen left combined with iron after experi-
mental time, t.

A schematic representation of relationship between reduction at
40% degree of reduction and 60% degree of oxidation levels,

particle size
porosity
crystal structure
pore size
volume change
impurities

Reduction of natural hematite ores by CO or H
2
starts between 200-
500
0
C, depending upon the physical characteristics and
mineralogical composition. However, the rate below 500
0
C is
sluggish.
Hematite is more reducible than magnetite although the amount of
oxy-gen to be removed per unit weight of iron is about 12 percent
higher in the former.
The better reducibility of hematite may be due to:
 formation of porous wustite from hematite, easily accessible to
reducer gas whereas magnetite forms dense wustite during
reduction;

tendency of hematite to break down and expose larger
surface due to expansion in volume during reduction to
magnetite ;
pores in hematite are more elongated and the microporosity
larger; magnetite has larger grain size and is more closely
packed;
a higher value of overall rate constant for wustite reduction
since the wustite lattice formed during reduction of hematite
exhibits a higher degree of disorder than that formed from
magnetite.

Chemical Influence
It is well known that the reduction rate of wustite is critical in the
overall kinetics of iron oxide reduction.
The equilibrium partial pressure or concentration of CO
2
would
decrease if a
FeO
is lowered by solution and/or compound
formation. Hence, the reduction rate would also decrease.

Natural ores can contain iron oxides as compounds with gang materials, such as,
2FeO.Si0
2
, FeO.AI
2
0
3
, FeO.Cr
2
0
3
, FeO.TiO
2
etc where wustite exists in a state of low
activity. The activity of wustite can also decrease when it undergoes sintering with
the impurities present, such as SiO
2
, Al2O
3
etc.

The reduction rate of ore increases with increase in linear
velocity of the reducing gas due to the reduction of the
boundary layer thickness at the bulk-gas/particle interface.
After a critical gas velocity is reached, there is no further
increase in the rate with increasing gas velocity since the
overall rate becomes controlled or limited by other processes.
The figure shows that the limit is only 0.4 m/s. The figure
also shows that the critical velocity is independent of the
degree of oxidation. In blast furnace, the linear gas velocity
does not affect the reduction rate since it ranges between 1-
20 m/s and is often exceeded.


For the reduction of iron ores the reducing gas has to diffuse
into the interior of the body where transformations can occur.
In general, the reduction rate increases with temperature but
the degree depends upon the mechanism of the reaction .
The overall reduction rate depends upon the relative
contributions of chemical control and gaseous mass
transport and hence depends upon the particular reactions
occurring and the reaction temperature. Since chemical
reaction has higher activation energy than gaseous diffusion,
the former will increase at a much· greater rate with increase
in temperature than the latter.

Hence, a stage will arrive where diffusion will become
rate-controlling. Depending upon the degree of
reduction, at lower temperatures of about 500-600°C,
the chemical reaction rate controls the reduction
rate forming what is known as the kinetic region in
the blast furnace. At temperatures above 600°C,
gaseous diffusion becomes the dominant rate
controlling mechanism. The temperature regime in the
blast furnace shaft is such that it can be assumed a zone
of mixed-control exists.


In the blast furnace , the reducing gas is
predominantly CO with varying amounts of
hydrogen depending upon the moisture content of
the blast and other blast additives like fuel oil or
natural gas. Study shows that a mixture of CO and
hydrogen appears to be a more efficient reductant
than either of them.

The function of coke in the blast furnace is five-fold, namely,
(i) it acts as a fuel by providing for the thermal requirements in the furnace,
the reaction being,
2C + O
2
= 2CO: ▲H
0
= - 2300 kcal/kg.C
On complete combustion to CQ
2
the heat evolved is 8150 kcallkg.C. Thus
only about 28 percent of the obtainable heat is supplied by coke;
(ii) it provides CO for the reduction of iron oxides;
(iii) it reduces the oxides of metalloids, such as, Mn, Si, P and others if
present;
(iv) it carburizes the iron and lowers its melting point;
(v) it provides permeability (in the dry as well as the wet zones) and also
mechanical support to the large charge column, permitting the gases to
ascend through the voids.

Coke is the universal fuel used in the blast furnace. It
acts both as a reductant as well as a supplier of heat. It
also comprises the major portion of iron production cost.
Now-a-days other fuels are also being used as part
replacement of coke. These fuels cannot be charged
from the top and as such they are injected into the
furnace through the tuyeres along with the blast. In
some countries, especially in Brazil, charcoal is used as
a blast furnace fuel.

Coke size: Coke comprises about 50-60 percent of the volume of the
charge material. The coke size is important as it provides
permeability in the dry as well as in the wet bosh zone The coke size
is always 3-4 times larger than the ore size, since coke is partially
burnt as it descends. It also has a lower density, and hence a greater
tendency for fluidisation. Of course, in the lower bosh region of a
blast furnace, coke is the only solid that remains, and which helps to
support the burden. The optimum size range for lump ore is 10-30
mm and for coke is 40-80 mm. Since the coke size becomes
smaller as it descends through the blast furnace due to mechanical
breakdown, gasification, attrition, etc., the factor of prime importance
is the strength of coke.

Coke strength: Mechanically considered, it is the quality cohesion that
prevents the coke from collapsing and tends to avoid the formation of
small particles. High cohesion or strength is related to several coke
making properties. On the basis of breakage by impact, compression
or abrasion, the coke strength should be assessed both at ambient as
well as high temperatures. Studies of the structure of different coke
samples show that the best varieties have a regular distribution of pores:
with adequate thickness and hardness of the walls between the pores
and are free from cracks generated internally. Such a structure ensures
withstanding of high compressive forces and high temperatures in the
all-important lower furnace.

The strength of coke produced in the coke-ovens is
influenced by:
blending ratio of coals of varying caking components and
proportion of the fibrous portion;
particle size and distribution of charging coal;
 coke-oven temperature and combustion conditions;
moisture and addition of oil;
soaking time;
width, height and method of heating.

It is defined as the ability of coke to react with O
2
, CO
2
or steam
(H
2
O).
More reactive cokes have higher thermal values of their volatile matter.
Coke of high reactivity ignites easily and gives rapid pick up of fuel bed
temperature. However, low reactivity coke gives a higher fuel bed
temperature than a highly reactive coke
Reactivity is inversely proportional to the absolute density. It is affected
by the presence of easily reducible iron compounds in ash.
Coke of high reactivity is obtained from weakly caking coals or blends.
Strongly coking, high rank coals produce coke with low reactivity.

For blast furnace coke, size and hardness are more
important than reactivity. Satisfactory hearth temperature is
obtained with unreactive coke containing little breeze.
Reactivity of coke is measured by Critical Air Blast method and
is reported as Critical air blast (CAB) value of coke. The CAB
value of coke is the minimum rate of flow of air in ft
3
/minute
necessary to maintain combustion in a column of closely graded
material (14 to 25 B.S.) which is 25 mm deep and 40 mm in
diameter. The typical CAB value for oven coke is 0.065
ft3/minute. More reactive coke has got lower CAB value.

Another modern and current method of expressing the reactivity and
strength of coke is Coke Reactivity Index (CRI) and Coke Strength
After Reaction (CSR) which is being followed in Indian steel plants.
Coke Reactivity Index (CRI).
To determine CRI, 200 gm of coke sample (size + 20 - 25 mm) is taken in
a stainless steel tube and heated in electric furnace to 1100°C. CO
2
gas at
5 kg/cm
2
pressure is passed through the coke bed for two hours. CO
formed (by reaction C + CO
2
= 2CO) is burnt in a burner and is exhausted
out. Carbon of coke reacts with CO
2
(depending upon the reactivity level of
the coke) and there is a loss of weight of coke depending upon its
reactivity. More is the loss in weight of the coke, reactivity is more. % loss
in weight of coke is reported as coke reactivity index (CRl). Ideal CRI
value of a good blast furnace coke should be about 20%. Typically
CRI of Indian blast furnace coke is about 25%.

Coke Strength after Reaction (CSR). The left out coke
from the CRI determination test is rotated for 60 rotation
in a micum drum. And the % of coke retained on a 10
mm size screen is reported as coke strength after
reaction (CSR). Stronger the coke, more is its CSR
value. Ideal value of CSR for blast furnace coke is a
minimum of about 55%. Typically CSR of Indian blast
furnace coke is about 60-65.

Agglomeration of Iron Ore Fines
About 65 – 75 % of iron ore gets converted into fines
( - 5 mm ) during various operations from mining to conversion
into CLO. Majority of these fines are exported to other countries
at throwaway price resulting in greater financial loss to the
nation. Most widely used methods for the agglomeration of these
fines to render them useful for BF are Sintering and Pelletization.
 Sintering – Sintering is essentially a process of heating of mass of
fine particles to the stage of incipient fusion for the purpose of
agglomerating them into lumps.

To increase the size of ore fines to a level acceptable
to the BF
To form a strong and porous agglomerate
To remove volatiles like CO
2
from carbonates, S from
sulphide ores etc
To incorporate flux in the sinter
To increase the BF output and decrease the coke rate

Iron ore sintering is carried out by putting a mixture Iron
bearing fines mixed with solid fuels on a permeable bed. The
top layer of sinter bed is heated up to the temperature of 1200
- 1300
0
C by a gas or oil burner. The combustion zone initially
develops at the top layer and travels through the bed raising
its temperature layer by layer to the sintering label. The cold
blast drawn through the bed cools the already sintered layer
and gets itself heated.


In the combustion zone, bonding takes place
between the grains and a strong and porous
aggregate is formed. The process is over when
the combustion zone reaches the lowest layer of
the bed. The screened under size sinter is
recycled and over size is sent to B.F.

Two types of bonds may be formed during sintering.
Diffusion or Recrystallization or Solid State Bond : It is formed as a result of
recrystallization of the parent phase at the point of contact of two particles in solid
state and hence the name.
Slag or Glass Bond: It is formed as a result of formation of low melting slag or glass
at the point of contact of two ·particles, depending upon the mineral constitution, flux
addition, etc.
As a result the sinter can have three different types of constituents:
Original mineral which has not undergone any chemical or physical change during
sintering.
Original mineral constituents which have undergone changes in their physical
structure without any change in their chemistry. Recrystallization is the only change
at some of the particle surfaces.
Secondary constituents formed due to dissolution or reactions between two or more
of the original constituents

The proportion of each of the physical and chemical change during
sintering depends upon the time-temperature cycle of the process.
The higher is the temperature more will be the proportion of new
constituents by way of solutions and interactions whereas lower is
the temperature and longer is the duration more is the process of
recrystallization in solid state.
The more is the slag bonding, stronger is the sinter but with less
reducibility and, more is the diffusion bonding, more is the
reducibility but less is the strength. Since ores are fairly impure
slag bond predominates. On the other hand in rich sinters slag
bond is of minor importance.

The area under the time-temperature curves
essentially determines the nature and
strength of the bonds developed during
sintering of a given mix. For a given mix it is
most unlikely the bonds of sufficient
strength will be formed below a certain
temperature level within a reasonably short
time. Hence the area under the curve above a
certain temperature, which may be around
1000°C for iron ores, is the effective factor in
deciding the extent of sintering

rather than the whole area under the curve from
room temperature to the combustion temperature
level. The nature of the time-temperature graph will
depend upon the rate of heating and cooling of a
given mix. The nature of this graph is of paramount
importance in assessing the sintering response. The
factors that affect this curve are then the variables of
the process and which should be adjusted properly
for obtaining effective sintering.

Bed permeability
 Total volume of air blast drawn through the bed
Particle size of iron ore
Thickness of the bed
Rate of blast drawn through the bed
Amount and quality of solid fuel incorporated in the sinter
mixture
Chemical composition of ore fines
Moisture content in the charge

During sintering, heat exchange takes place between the solid
charge and air drawn. At any time, the air takes the heat from
combustion zone and then transfers to the lower layer of the bed.
For faster rate of heat exchange, the volume of air drawn
should be more. If suction rate of air is too high, transfer of heat
may become less efficient. On the other hand, the flame front will
not move down the bed properly if suction is less. Higher the bed
permeability, more will be the air drawn. But, higher permeability
leads to loss of strength in the resulting sinter due to reduction in
bond strength. Hence a compromise is made between these two
factors. It is usual practice to draw about 700 – 1100 m
3
of air/ton of
charge.


An increase in particle size increases bed permeability and the
volume of air drawn.
Strength of sinter gets reduced with an increase in particle size of
the ore due to reduction in contact area.
For effective sintering, the use of larger ore lumps is undesirable.
Iron ore size > 10mm is rarely preferred.
Higher proportion of –100 mesh size fines adversely affects the bed
permeability. Better is that – 100 mesh size fraction should be
screened off and used for pelletization. Ideal size of iron ore for
sintering is 0.07 – 10 mm.

During mining and ore dressing operations, especially
where very fine grinding is necessary for wet
concentration, a large amount of - 0.05 mm fines is
generated which are not amenable to sintering
because of very low permeability of the bed. They can,
however, be agglomerated by balling them up in the
presence of moisture and suitable additives like bentonite,
lime, etc. into 8-20 mm or larger size. These green pellets
are subsequently hardened for handling and transport by
firing or indurating at temperatures of 1200-1350°C.

Pelletisation essentially consists of formation of green
balls by rolling a fine iron bearing material with a critical
amount of water and to which an external binder or any
other additive may be added if required. These green
balls of nearly 8-20 mm size are then dried, preheated
and fired, all under oxidising conditions, to a temperature
of around 1250-1350°C. Bonds of good strength are
developed between the particles at such high
temperatures.

The pelletisation process consists of the following
steps:
Feed preparation.
Green ball production and sizing.
Green ball induration:
(a) Drying
(b) Pre-heating
(c) Firing
Cooling of hardened pellets.

The observations on ball formation that eventually led to the
development of the theory of balling are as follows:
Dry material does not pelletise and presence of moisture is
essential to roll the powder into balls. Excessive water is also
detrimental.
Surface tension of water in contact with the particles plays a
dominant role in binding the particles together.
Rolling of moist material leads to the formation of balls of very high
densities which otherwise is attainable by compacting powder only
under the application of a very high pressure:
The ease with which material can be rolled into balls is almost
directly proportional to the surface area of particles, i.e. its fineness.

The capillary action of water in the interstices of the grains causes a
contracting effect on them. The pressure of water in the pores of the
ball is sufficiently high so as to compact the constituent grains into a
dense mass. The compressive force is directly proportional to
fineness of the grains since the capillary action rises with the
decrease in pore radius and the latter decreases with increasing
fineness. An optimum moisture is important since too little of
water introduces air inclusions in the pores and too much of water
would cause flooding and destruction of capillary action. The
optimum moisture content usually lies between 5-10 percent or
more, the finer the grains the larger the requirement.

Besides the bonds formed due to surface tension mechanical
interlocking of particles also pays a significant role in developing the
ball strength.
Maximum strength of a green ball produced from a given material
will be obtained by compacting the material to the minimum porosity
and with just sufficient water to saturate the voids. The rolling action
during pelletisation is beneficial in reducing the internal pore space
by effecting compaction and mechanical interlocking of the
particles.

From fundamental studies it has been concluded that there are
three different water-particle systems:
The pendular state, when water is present just at the point of
contact of the particles and surface tension holds the particles
together.
The funnicular state, when some pores are fully occupied by water
in an aggregate system.
The capillary state, when all the pores are filled with water but
there is no coherent film covering the entire surface of the particles.


The ball formation is a two stage process, i.e. nucleation or seed
formation and their growth. The formation of balls on a pelletiser depends
primarily on the moisture content. Seeds are formed only if critical
moisture level is maintained and without which the process cannot
proceed properly. Growth takes place by either layering or assimilation. It
has been observed that the size of the balls produced in a pelletiser from a
charge containing right amount of moisture depends on the time and speed
of the pelletiser, i.e. number of revolution.. Three regions can be clearly
observed, during ball formation. :
oNucleii formation region
oTransition region
oBall growth region.


When a wet particle comes in contact with another wet
or dry particle a bond is immediately formed between the
two. Similarly several such particles initially join during
rolling to form a highly porous loosely held aggregate
and crumbs which undergo re-arrangement and partial
packing in short duration to form small spherical, stable
nucleii. This is the nucleation period, a pre-requisite for
ball formation since these very nucleii later grow into
balls.

After nucleii are formed they pass through a transition period
in which the plastic nucleii further re-arrange and get
compacted to eliminate the air voids present in them. The
system moves from a pendular state through funicular state
to the capillary state of bonding. Rolling action causes the
granules to densify further. The granules are still plastic with a
water film on the surface and capable of coalescing with other
granules. The size range of granules in this region is fairly
wide.

The plastic and relatively wet granules grow if they are
favorably oriented. In this process some granules may even
break because of impacts, abrasion, etc. Growth takes place
by two alternative modes.
growth by assimilation is possible when balling proceeds
without the addition of fresh feed material.
growth by layering is possible when balling proceeds with
the addition of fresh feed material.

Growth by Assimilation
If no fresh feed material is added for balling the rolling action may break
some of the granules, particularly the small ones, and the material
coalesces with those which grow. The bigger the ball the larger it will grow
under these conditions. Since smaller granules are weaker they are the first
victim and growth of the bigger balls takes place at their expense.
Growth by Layering
Growth of the seeds is said to be taking place by layering when the balls
pick up material while rolling on a layer of fresh feed, The amount of
material picked up by the balls is directly proportional to its exposed
surface, i.e. the increase in the size of the balls is independent of their
actual size.
Growth by layering is more predominant in the disc pelletisers and
growth by assimilation is more predominant in drum pelletisers, at
least beyond the feed zone.

In general natural lumpy ore or sinter or pellets or a suitable com-bination of
two or more of these form the burden.. The modern large capacity furnaces
necessarily need fully prepared burden to maintain their productivity since
the required blast furnace properties cannot just be met by natural lumpy
ore. The selection of the process of agglomeration, whether sintering or
pelletising, will depend upon the type of ore fines available, the location of
the plant and other related economic factors involved.
Sintering is preferred if the ore size is -10 mm to + 100 mesh and if it is
-100 mesh pelletising is generally adopted. Pelletising in fact requires
ultrafines of over 75% of -325 mesh. These processes are there-fore not
competitive.

Minimum closure of pores by fusion or slagging; open pore
system; very good reducibility due to high microporosity .
Porosity of sinter is 10-18% and that of pellets is 20-30%.
The shape of pellets is near spherical and hence bulk
permeability of the burden is much better than that obtained
from sinter which is non-uniform in shape.
The shape, size and low angle of repose give minimal
segregation and an even charge distribution in the furnace.

More accessible surface per unit weight and more iron per unit of furnace volume
because of high bulk density, 3-3.5 tonnes/m
3
.Larger surface and increased time
of residence per unit weight of iron give better and longer gas/solid contact and
improved heat exchange;
Degradation of sinter during its transit is much more than that of pellets. The
sinter therefore has to be produced nearby the blast furnace plant while pellets
can be carried over a long distance without appreciable degradation. Ease in
handling
It should also be noted that If high rates of productivity demand elimination of
fines and since sinter happens to contribute more to the generation of fines than
that of pelllets, the later will have to be chosen as the burden in preference to
sinter.

oThe installation cost of a pelletising plant will be 30-40% more than
that of sintering plant of an equal size.
oThe operating cost of sintering is slightly less than that of pelletising.
o Difficulty of producing fluxed pellets.
oSwelling and loss of strength inside the furnace
oFluxed pellets break down under reducing conditions much more
than acid and basic sinters and acid pellets.
oStrong highly fluxed sinters, especially containing MgO, are being
increasingly preferred to pellets.

Smarajit Sarkar
Department of Metallurgical and Materials Engineering
NIT Rourkela

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.


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 10–12 kg/thm.

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.

Inconsistency in
physical properties of
charge materials
(deficiencies caused by
this should be
eliminated by improving
quality of the bur-den.

Size range of the various
charge materials
Angle of repose of raw
materials and other
physical characteristics of
the charge.
 Density of charge
materials.

Level, system and
sequence of
charging, programme
of revolving the
distributor (conditions
determining major
means of blast
furnace process
control from top).

Distribution of charge
on the big bell
 Height of the big bell
from the stock-line i.e.
charge level in the
furnace throat.
Order and proportion
of charging of various
raw materials.

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 1·5 glcc and the limestone is
intermediate with-a value of density around 3·0-3·5 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.

When a multi-particle material is allowed to gently fall
on a hori-zontal plane it tends to form a conical heap.
The base angle of this cone is known as angle of repose
of that material.
 This angle depends upon the particle size, its surface
characteristics, moisture content, shape, size
distribution, etc.

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
more porous agglomerates before charging these in the
furnaces.

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°.

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.

On dumping, as the materials fall on the stock
surface, they take a para-bolic 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)

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.

Right at the top of the furnace is the granular zone that 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-called
cohesive zone (mushy zone).


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 1185°C..
In the cohesive zone and below it, coke is the source of carbon for
carburisation 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
+ CO
2
(g)

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.

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
reducing 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 CO
2
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+O
2
(air) =CO
2
+94450 cal

 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 CO
2
by C proceeds;
CO
2
+C =2CO-41000 cal
The concentration of CO
2
fall; rapidly from the raceway
boundary and the gasification is completed within 200-
400 mm from the starting point of the reaction.


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-Si0
2-AI
20
3 system rapidly
changes to the CaO-Si0
2-AI
20
3system with some minor impurities accompanying the
burden. The dissolution of lime and the approach to the CaO-Si0
2
-Al
2
0
3
system is more
pronounced,
.

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-Si0
2
-AI
2
0
3
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.

This slag runs along with the molten iron into the
hearth and accumulates there and forms a pool
with the molten metal underneath. During the
passage of iron droplets through the slag layer,
the slag reacts with the metal and a transference
of mainly Si, Mn and S occurs from or to the
metal, tending to attain equilibrium between
themselves as far as possible.

0.81 kg. C is required for indirect reduction of 1 kg. Fe
from Fe
20
3 and about 1790 kcal of heat is evolved in the
process.
 for direct reduction of 1 kg. Fe, only 0.23 kg. C is
consumed but results in an absorption of 656 kcal of
heat.

Below 600°C :
Pre-heating and pre-reduction
600 -950°C:
Indirect reduction of iron oxides by CO and H
2

950
0
C to softening temperature:
Direct reduction; gasification of carbon (solution loss
reactions) by CO
2
and H
2
becomes prominent.



The formation of cohesive layers or partially
reduced and partially molten iron oxide takes
place.
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 the
hearth. It is the final stage of iron oxide reduction

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.

fluidization of small particles when the local gas
velocity is excessive;
diminution of void age due to swelling and
softening-melting;
flooding of slag in the bosh zone when the slag
volume and gas velocity are excessive.

The charge in the blast furnace descends under gravity against the
fric-tional 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 teeter and 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).

The mechanism of the softening-melting phenomena
is schematically illustrated in previous Figure. It is
evident that with the onset of softening, the voidage in
the bed decreases and the bed becomes more
compact (origin of the terminology cohesive).
As a consequence, further indirect reduction of iron
oxide by gases becomes increasingly difficult. Upon
melting, dripping of molten FeO-containing slag
through the coke layers increases the flow resistance
through the coke slits and the active (i.e. dripping)
coke zone because of loss of permeability.

The cohesive zone has the lowest permeability.
Hence, for proper gas flow:
T
s
should be as high as possible
The thickness of the cohesive zone should be
as small as possible. This thickness depends on
the difference between T
s
and T
m
(T
m
- T
s
), and
therefore, the difference should be as low as
possible.

Gas flow through Granular zone:
For resistance to gas flow, more important than
the particle diameter is the relative size of the
materials in the bed.
In a mixed bed of widely varying particle size, the
small particles land in the interstices of the large
ones and decrease the void age .
Starting with large uniform spheres, the void age
decreases as the small ones are introduced and
the bed becomes more and more compact as the
proportion of the latter increases.
The bed is most dense, i.e., the voidage is
minimum when 60-70 percent of the total volume of
the particles consists of the large ones for about all
the cases.

The €
m
increases on either side of the
minimum, i.e., with increasing or decreasing
volume fraction of the small particles
(approaching more uniformity of the size
distribution).
The voidage decreases greatly as the
ratio d
s
/ d
1
decreases.
This shows that for a good and uniform
permeability and low resistance to gas flow
in a mixed bed, the size fractions should be
as narrow as possible.
One can easily visualize the adverse
effects of multi-granular bed of particles of
varying diameter on the voidage.

A narrow size distribution has the following advantages:
 charge permeability increases and the gas distribution is
more uniform with better utilization of the chemical and
thermal 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 chang-ing of flow directions, providing a larger
gas/solid contact time.

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 de-creases, 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:
Wet zones consist of the coke beds in the bosh and belly
regions, i.e. inactive coke zone, active coke zone, and the coke
slits in the cohesive zone.
 Here molten iron and molten slag flow downwards through
the bed of coke. This reduces the free cross section available for
gas flow, thus offering greater resistance, thereby increasing
the pressure drop.
An extreme situation arises when, at high gas velocity, the
gas prevents the downward flow of liquid. This is known as
loading. With further increase in gas velocity, the liquid gets
carried upwards mechanically, causing flooding.

Lump ores, sinter and pellets disintegrate into smaller pieces during their
downward travel through the blast furnace owing to the weight of the
overlying burden, as well as abrasion and impact between the burden
materials.
It has been found that this tendency gets aggravated when the oxides are in
a reduced state. Reduction of hematite into magnetite occurs in the upper
stack at 500-600°C, and this is accompanied by volume expansion even to
the extent of 25%.
This results in compressive stresses being developed and contributes
significantly to breakdown of the iron oxides.
Blast furnace operators prefer a low RDI (below 28 or so) since the adverse
effect of high RDI has been clearly demonstrated in practice.

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
furnaces, an empirical parameter, called Flow
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 .

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.

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.

Smarajit Sarkar
Department of Metallurgical and Materials Engineering
NIT Rourkela

Decreasing the extent of SiO formation by:
oLowering ash in coke, and the coke rate
oLowering RAFT
oLowering the activity of Si0
2
in coke ash by lime
injection through the tuyeres.
Decreasing Si absorption by liquid iron in the bosh
by enhancing the absorption of Si0
2
by the bosh
slag. This can be achieved by:
oIncreasing the bosh slag basicity.
oLowering the bosh slag viscosity..

Removal of Si from metal by slag-metal reaction
at the hearth by:
oLowering the hearth temperature
oProducing a slag of optimum basicity and fluidity.

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] + ½[ Si] = (CaS) + 1/2 (SiO
2
)

Sulphur pick-up through the vapour-
phase reaction:
CaS( in coke ash) + SiO (g) = SiS(g) +
CaO
FeS( in coke ash) + SiO (g) = SiS(g) +
CO(g) +[Fe]
In the bosh and belly regions, SiS
decomposes as
SiS(g) = [Si] + [S]

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

0.8-0.9t
0.5-0.6t
1.7-1.8t
2500

m3
0.6t
1t
•Fuel
•Reducing agent supply
•Permeable bed
(spacer)
3200m
3
+
80kg dust

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 C0
2
sensible heat is carried from the furnace as a
result of the exothermic reaction.
2CO=CO
2
+C
If on the other hand the top gas is high in CO, chemical
heat leaves the furnace.

CO
2
emission
Industry Contribution %
Power 51
Transport16
Steel 10
other 23

The purpose of HTP is to introduce more
oxygen to burn more carbon by blowing more
air and at the same time maintaining the
linear gas velocity (and pressure drop)
identical to that in the conventional practice
without any formation of channels,
maldistribution of gas, increase in coke rate
or flue dust emission
 Advantages:
◦For the same volume flow rate, a greater mass of air
(hence, oxygen) can be blown with HTP; higher output;

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 (C0
2
+ 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;

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

SiO
2 +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 SiO
2

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.

The blast volume and therefore the coke
throughput can be increased by 30
percent with the maintenance of identical
pressure drop and gas velocity conditions
in the blast furnace by increasing the top
pressure to 2.1 from 1.1 ata and bottom
pressure to 3.5 from 2.5 ata under the
given blowing conditions.

'raceway adiabatic flame temperature‘
This is the highest temperature available inside the
furnace. 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.

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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