Download - Iron and Steel Making

Smarajit SarkarDepartment 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◦ Al2O3 – 20.45%◦ CaO – 32.23%◦ SiO2 – 33.02%◦ MgO – 9.95%◦ S – 0.89%◦ MnO – 0.54%◦ TiO2 – 1.01%◦ FeO – 0.41%◦ K2O+Na20 – 1%◦ Trace Oxides – 0.5%

(Curtsey TATA STEEL)


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.

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 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/m3. The

quantity of gas produced depends upon the amount of fuel burnt.

For one tonne of coke burnt nearly 4000 m3 of effluent gas may

be produced. Hence a blast furnace requiring 1000 t of coke per

day would generate nearly 4 x 106 m3 of gas with a total energy

content of 3600 x 106 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/m3. 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/m3. 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/m3. 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/m3 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 m3/ton of pig iron

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.

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)= 5.55V0.24

Useful height (H0) =0.88×H

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 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.

dbelly =0.59 ×(V)0.38

HbelIy = 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× V0.45

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 Hstack = 0.63 H- 3.2 m

Stack angle

The stack angle usually ranges from 850 to 870

(i) 850 for weak and powdery ores; (ii) 860 for mixture of strong and weak, lumpy or fine ores; (iii) 870 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 formulaStack angle (α)= Cot-1(D-d1/2xStack Height) Where, D= Bosh parallel Diameterd1= Throat Diameter

Bosh angle can be calculated from the formulaBosh angle (β)= Cot-1(D-d/2xBosh Height) Where, D= Bosh parallel Diameterd= 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 V0.35

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 tuyeresPavlov: n = 2d +1Rice: n = 2.6d-0.3Tikhomirov et al : n = 3d-8

Where n= Number of tuyeres, d=hearth diameter

Capacity → (THM/Day)Parameter↓

2000 3000 5000

Useful Volume (m3) 1700 2550 4250Total Height (m) 33.08 36.46 41.22Useful Height (m) 29.11 32.08 36.27Bosh Parallel Dia (m) 9.96 11.62 14.11Bosh Parallel Height (m) 2.32 2.55 2.89Bosh Height (m) 4.37 4.81 5.44Hearth Dia (m) 9.1 10.92 13.74Hearth Area (m2) 65.04 93.66 148.27Hearth Height (m) 3.308 3.646 4.122Stack/Shaft Height (m) 17.64 19.77 22.77Throat Dia (m) 6.87 7.85 9.29Bosh Angle (0) 84.32 85.84 88.05Stack Angle (0) 85 84.55 83.96Nos. 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 theirfractional oxygen removal rates in gaseous reducing atmosphere. The percent degree of reduction orpercent fractional oxygen removal is given by

Wheren0 = 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 H2 starts between 200-

5000C, depending upon the physical characteristics and mineralogical

composition. However, the rate below 5000C 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 CO2 would

decrease if aFeO 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.Si02, FeO.AI203, FeO.Cr203, FeO.TiO2 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 SiO2, Al2O3 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 + O2 = 2CO: ▲H0 = - 2300 kcal/kg.C

On complete combustion to CQ2 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 O2, CO2 or steam

(H2O).

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 ft3/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. CO2 gas at 5 kg/cm2 pressure is passed through the coke bed for two hours. CO formed (by reaction C + CO2 = 2CO) is burnt in a burner and is exhausted out. Carbon of coke reacts with CO2 (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 CO2 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

- 13000C 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 m3 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. :

o Nucleii formation region

o Transition region

o Ball 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/m3 .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.

o The installation cost of a pelletising plant will be 30-40% more than

that of sintering plant of an equal size.

o The operating cost of sintering is slightly less than that of pelletising.

o Difficulty of producing fluxed pellets.

o Swelling and loss of strength inside the furnace

o Fluxed pellets break down under reducing conditions much more

than acid and basic sinters and acid pellets.

o Strong highly fluxed sinters, especially containing MgO, are being

increasingly preferred to pellets.


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+ CO2(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 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

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.

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-AI203 system 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 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-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.

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 Fe203 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 H2

9500C to softening temperature:

Direct reduction; gasification of carbon (solution loss

reactions) by CO2 and H2 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:

Ts 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 Ts and T m (Tm - Ts), 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 ds/ d1 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.


NIT Rourkela

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 bosh by 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..

Removal of Si from metal by slag-metal reaction at the hearth by:

o Lowering the hearth temperature o Producing 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 (SiO2)

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 asSiS(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

0.8-0.9t0.5-0.6t1.7-1.8t

2500 m3

0.6t 1t

•Fuel•Reducing agent supply•Permeable bed (spacer)

3200m3

+ 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 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.

CO2 emission

Industry Contribution %Power 51Transport 16Steel 10other 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 (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;

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

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.

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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