PREPARATORY PROCESSES

These are processes that alter the chemical or physical state of ores. Minerals recovered from ores

are not always in the optimum chemical or physical state for conversion to metals. Oxides are

more conveniently reduced to metals than sulphides, or the metal might be more readily leached

from the ore if it were present as a sulphate, a chloride or an oxide. Chemical conversion to the

desired species often is an integral segment of the extractive process. Sulphide ores or

concentrates, for example, usually are heated in an oxidized atmosphere (roasted) to convert them

to an oxide or sulphate.

The physical state of an ore may be too fine for charging to a process. Fine ores often are

agglomerated by sintering prior to charging to a blast furnace, the principal smelting unit for lead

and iron. In the case of iron ore, pelletizing is another very important agglomeration process that

has achieved commercial adaptation in the iron and steel. In the sections below, the following

pre-treatment processes will be explained.

1. drying

2. calcinations

3. roasting

4. agglomeration

Agglomeration

Methods of agglomerating ores have been under consideration since the last century. They may

be classified as follows:

briquetting

nodulizing (rotary kiln sintering)

vacuum extrusion

sintering (grate sintering)

pelletizing

Pelletizing

It is a relatively new process developed for the agglomeration of iron ore fines but which has

been adopted widely in the non-ferrous metallurgical industry. Originally the pelletizing process

was developed in the States to treat ultra-fine mineral dressing products obtained from the

upgrading of iron ores. This upgrading was carried out because of concern felt at the time that

the available resources of high grade ore appeared to be inadequate for the future development of

the steel industry. Pelletizing consists of three distinct operations: preparation of ore feed,

forming the pellets at atmospheric temperature (balling) and then firing them at a temperature in

the region of 1300oC (hardening).

The pelletizing process is based on the formation of green balls by rolling a finely ground ore or

concentrate to which bentonite is usually added together with critical amount of water. These

balls are then dried, pre-heated and fired, all under oxidizing conditions, to temperature of 1250-

1350oC. As a result oxide bridging, grain growth, and some slag bonding occurs, and pellet

strength is developed.

Preparation of Ore Feed

Feed to a pelletizing plant is wet concentrate. Frequently concentrate are reground to about 80% -

50μm before they are pelletized. Concentrate slurry is thickened and filtered to provide material

with desired moisture content, normally 10%. At this stage a small quantity of binder (bentonite)

is often mixed with the moist concentrate. Partial drying is sometimes needed; this can be

achieved either by heating or by mixing a proportion of dry fines with the wet material. The

moisture content for balling for the production of good green balls is quite critical.

Balling

The moist material is then balled, normally by passing it through a drum which rotates at about

10-15rev/min depending on its diameter, and which is inclined at about 5-10 o to the horizontal.

The output from the drum is screened and the oversize, usually +9mm, goes forward, the

undersize being returned. The drums that are used are normally 9m long and 2.7m in diameter,

and produces about 40-50tons/hr of green balls. More rarely, inclined discs are used for balling.

These discs are normally about 3.7-5.5 in diameter and are inclined at about 45o to the horizontal.

Hardening

The green balls are next hardened. Though this operation is often carried out in a single piece of

equipment, it consists of three operations-drying, firing and cooling. Three different types of

firing equipment are in general commercial use at the present time:

1. the vertical shaft furnace

2. the traveling grate, and

3. the grate kiln

Vertical Shaft Furnace

It was the earliest device used for pellet hardening. See the diagram provided. The shafts have

effective height of about 14m and are rectangular in section typically 4.2m by 1.8m. Fuel is

burned in the combustion chambers and the waste gases are introduced into the furnace through

multiple ports. Cooling air is introduced near the bottom of the furnace. At the base of the

rectangular section, shafts carrying toothed wheels pass through the furnace. These toothed

wheels, known as chunk-breakers, break up any clusters of pellets that may have formed and

regulate the flow of pellets.

The shaft furnaces operate on the counter-current floe principle, the heat extracted from the

pellets during cooling being used to heat the pellets in the high temperature zone. Shaft furnaces

are particularly adapted to the production of relatively small tonnages of pellets. For larger

outputs a multiplicity of shafts is required, as individual furnace capacity is not normally in

excess of 1000-1200tons/day.

Traveling Grate

The green pellets are laid on a traveling grate, similar to a sinter strand and subjected to drying,

firing and cooling as they travel along the strand. Strands are typically 3m wide and 60-90m in

length; the depth of the pellet bed, including the hearth layer is, is not more than 400mm. It is

usual to protect the grate bars by covering them with layer of fired pellets, and to protect the

sidewalls of the pallets and minimize heat loss by using side layers of fired pellets. Production is

in the region of 150-200tons/hr.

The heat for the process is supplied by oil or gas burned in the hood covering the firing zone of

the strand. Hot air drawn from under the firing section of the grate and from above the cooling

section is used both to dry the green pellets and as combustion air for the firing zone. On a typical

strand, 25% of the area is used for drying, 40% for firing, including pre-heating to firing

temperature, an 35% for cooling.

Grate Kiln

In this process the green pellets are dried and pre-heated to about 1000oC on a traveling grate,

then fired in a rotary kiln, and finally cooled in a separate cooler. The strand is shorter, 30-36m,

and simpler than that for the traveling grate; hearth and side layers of fired pellets are not used.

Hot waste gases from the kiln are first used to pre-heat the dried pellets and then passed through

the strand a second time to dry the green pellets.

The preheated pellets are charged into a rotary kiln inclined at a few degrees to the horizontal. An

oil or gas fired burner is set at the discharge end of the kiln, and the pellets travel down the kiln

counter-current to the combustion gases. The kiln which is refractory-lined, is typically 36m long

and 4.5m in diameter. From the rotary kiln, the pellets discharge into an annular cooler, 12m

average diameter with an annulus 1.8m wide. The air from the hotter section of the cooler is used

as secondary air for the rotary kiln burner. A single grate and kiln can produce more than

200tons/hr of fired pellets.

Desirable Properties of Pellets

Physical Properties

As pellets are normally transported over considerable distances from the mine to the blast

furnace, great attention should be paid to the physical strength of the pellets as this is a measure

of their ability to withstand the rigours of the handling involved.

Resistance to tumbling and abrasion

Compression strength

Size of pellets- in order to obtain high production rates in the blast furnace, a closely

sized burden, i.e. within a narrow size range, free from fine material less than 5mm in

size is essential

Porosity-an important fundamental property of green balls and fired pellets, and plays an

important role in each of the stages in the palletizing process. In blast furnaces reducing

gases, CO and H2, enter the pellets via the pores. A porosity of 22-30% for fired pellets

is associated with good reducibility

Chemical Properties of Pellets

Of importance is the:

Chemical analysis

Reducibility:- the rate of removal of oxygen under reducing conditions

ROASTING

The process is usually applied where sulphides are concerned. Many important metals such as

copper, lead, zicn, nickel etc occur in nature as sulphides. However sulphides are not readily

converted into metal. The roasting stage allows the conversion of sulphides or other chemical

states into oxides or sulphates which are more readily reduced to metal by carbon or leached into

solution.

Roasting of sulphides is a process (gas-solid reaction) where air in large amounts, sometimes

enriched with oxygen, is brought into contact with the sulphide mineral concentrates. This is done

at elevated temperature where oxygen will combine with sulphide sulphur to form gaseous SO2

and with the metals to form metallic oxides.

This oxidation must be done without melting the charge, which would destroy the required

maximum particle surface-oxidizing gas contact area. Stirring of the charge in some manner also

ensures exposure of all particle surfaces to the oxidizing gas. Only exception to this general

procedure is sintering (blast roasting).

The degree of sulphur elimination is controlled by regulating the air supply to the roaster and by

the degree of affinity the mineral elements have for sulphur and oxygen. Consequently minerals

such as iron sulphide, which have a higher affinity for oxygen than for sulphur, aamy all be

oxidized, while a copper mineral in the same roaster feed, with a greater affinity for sulphur than

for oxygen, will emerge in the calcine still as sulphides.

Roasting is essentially a surface reaction, with oxide layer first formed remaining as a porous

layer through which oxygen can pass into still unreacted inner sulphide portion of the particle and

through which the SO2 gas then formed can pass out. This passage becomes more difficult as the

porous oxide layer thickens, and there will be some reversing reactions in the particle interior as

the concentrations of SO2 build up:

MS + 3/2O2 MO + SO2

MO + SO2 MS + 3/2O2

This makes it difficult to remove the last interior amounts of sulphur. Particle size also is

important, and large particles will take much longer to react to their centres.

The oxidizing roast of a sulphide is an exothermic reaction and this heat of reaction helps to keep

the roaster at the required roasting temperature so that the process can continue with extra heat

supplied by burning fuel. Autogenous roasting can be achieved when a high sulphide roaster feed

material has sufficient exothermic heat generated from its oxidation reaction to be self-

propagating and to require no extraneous fuel.

Types of Roasting

There are several types of roasting.

a) Oxidizing Roast

This type of roasting converts sulphides to an oxide as follows:

PbS + 3/2O2 PbO + SO2

FeS + 3/2O2 FeO + SO2

b) Volatizing Roast

Normally done when there is need to eliminate unwanted volatile materials, e.g. if we have a

sulphide having arsenic and antimony, then this type of roating can be used to remove arsenic

as its oxide, As2O3 and also antimony as Sb2O3. Other impurity elements and oxides that can

also be removed include Cd and ZnO. These may be recovered from the process fume using

bag filters.

c) Sulphating Roast

To achieve a particular end product e.g. a sulphate, if the next stage of processing is leaching.

A sulphating roast is used when the metal sulphate is subsequently leached with a dilute

sulphuric acid solution. To achieve this method, availability of oxygen and temperature

should be optimum. Low temperatures are required to achieve this method. Metal sulphates

decompose at low temperature therefore sulphating roasting is normally conducted at about

600-800oC, i.e. below the corresponding decomposition temperature, with a restricted amount

of air.

d) Chloridizing Roast

Applied for converting a material to a chloride for easy subsequent processing e.g. are

titanium oxides which require very high temperatures to be reduced by carbon. But by using a

chloridizing roast we can use lower temperature to reduce it. It is generally used for the

conversion of a reactive metal such as titanium and zirconium which, forms extremely stable

oxides, to a less stable chloride or other halide. The halide is relatively easy to reduce with

another element which forms a more stable halide. In the production of titanium from TiO 2

containing concentrate, the TiO2 is subjected to a chloridizing roast in the presence of carbon

at 500oC.

TiO2 + C + 2Cl2 TiCl4 + CO2 ΔG= -295KJ

Since TiCl4 is thermodynamically less stable than TiO2, carbon must be added to make the

ΔGo for the reaction more negative. This is the case in most chloridising roast operations i.e.

TiO2 + 2Cl2 TiCl4 + O2 ΔGo500 = +97KJ

TiCl4, gas is subsequently reduced with Mg (Kroll Process) at about 850oC as MgCl 2 is more

stable (more negative free energy of formation) than TiCl4

TiCl4 (gas) + 2Mg (L) Ti(solid) + 2MgCl2(L)

e) Magnetizing Roast

If magnetic separation is the next stage of processing then magnetic roasting is applied to an

iron ore (hematite, Fe2O3 to magnetite Fe3O4).

3Fe2O4 2Fe3O4 + ½ O2

This process uses controlled reduction of hematite to magnetite which can be subsequently

magnetically separated from the gangue.

f) Reducing/Reduction Roast

MS + O2 M + SO2

A less common type of roasting reaction may result at high temperature and low oxygen

potential where oxide and sulphide interact to produce the metal.

2MO + MS 3M + SO2

g) Sinter Roasting

The process combines two important stages.

i. achieve an oxidizing roast

ii. to alter the physical state of the material

Most roasting is carried out to completion, a so called dead or sweet roasting for e.g. PbS (galena)

may be roasted to remove practically all S as follows.

MS + 3/2O2 MO + SO2

Similar reaction may be written for sulphides of Cu, Zn and Fe. Partial roasting to reduce the

level of S may also be carried out. At lower temperatures sulphates are the likely products of

roasting and at higher temperature oxides may be reduced by sulphides.

Industrial Roasters

Processes that have been developed to carry out roasting include mechanical or multi-hearth

roasters, Dwight-Llyod sintering machines, and fluid bed roasters. The choice of the roasting

process depends on the kind of smelting process to which the calcines are to be subjected after

roasting. Roasting in multiple-hearth and fluid-bed roasters requires fine feed material and

furnishes fine calcines which are then treated in reveberatory furnaces, flash smelting, or electric

furnaces. The multiple-hearth roaster is the oldest type, while fluid bed roaster is the more recent

development. Sulphide concentrates that must be both desulphurised and agglomerated are

usually roasted on blast roasters (sintering machines). This gives a coarse, porous, oxidized sinter

cake as a product, and is then a smeltable shaft furnace feed material from the one single roasting

operation.

Multiple-Hearth Furnace

This unit consists of a number of horizontal, circular, refractory hearths enclosed in a steel shell,

with the feed material placed on the top hearth and worked downward to be discharged as roasted

calcines from the bottom hearth. A central slowly rotating shaft turns air cooled or water-cooled

rabble arms on each hearth, which serves two purposes.

the rotating rabble blades turn over the roaster charge to stir fresh material to the surface

for the roasting, gas-solid, oxidizing reaction, and

push the charge across the hearth to drop holes to be passed on down to the hearth below

The drop holes are located to be not directly below one another, but at the outer periphery of one

hearth and at the centre of the hearth below. The charge then follows a lengthy, zigzag path down

through the roaster, allowing time for the oxidizing reactions to take place.-Tulani

As the feed material progresses downward in the roaster, it is heated by the rising hot gases from

the exothermic roasting reaction taking place on the lower hearths, until finally this feed material

is also heated to the reaction temperature, begins to burn, and oxidizes at rapid rate. This reaction

will continue until the roasted calcines are discharged from the bottom hearth of the roaster and

cool in air to below the roasting reaction temperature.

Gas burners are provided on the lower hearths to ensure that reaction temperature is reached if

roasting is not autogenous. The flow of air into the roaster is regulated by opening doors on lower

hearths, and natural draught provided sucks in air to supply oxygen for oxidizing.

Roaster capacity will average about 2.8 ton of pyrite per square metre of hearth area per day and

roasters range in size from 4-12 hearths of diameter 3 to 7m. SO2 concentration of the roaster gas

will be 4.5-6.5%.

Flash Roaster

In a flash roaster several intermediate hearths are removed to facilitate rapid oxidation. The top

hearths are used for drying the concentrate and the open space is used for combustion before the

material settles on the bottom hearth. Associated with this type of roaster there has to a dust

recovery system.

Fluid Bed Roaster

The roaster consists of a cylindrical brick-lined steel shell which is closed at the bottom by a

grate. A wind box below the grate blows in air at a sufficient volume and is distributed evenly by

the grate to hold the solid feed particles in suspension and give excellent gas-solid contact on all

surfaces.

Slurry of material to be roasted, with maximum particle size kept at about 6.3mm, is continuously

fed through a downpipe to the turbulent layer in the roaster. This turbulent layer with its solid

particles in suspension has the flow characteristics of a fluid. If the feed material has mixed size

and densities, the smallest and lightest particles will migrate to the top of the turbulent layer,

while the larger and heavier particles collect at the bottom. Part of the roasted calcines leave by a

side discharge overflow pipe, and part is carried off in the effluent gases to be recovered as flue

dust in a gas-cleaning system. Cooling coils remove the excess reaction heat from the turbulent

layer.

The oxidizing reaction is autogenous, and the high turbulence of the suspension and the resulting

excellent gas-solid contact and heat exchange account for the very high reaction rate of the

process and accompanying high capacity. This capacity will be of the order of 22 ton of pyrite

feed material treated per day per square meter of grate area. SO 2 content of the roaster gas is 9-

12%.

Blast Roaster (Sintering)

A sintering machine consists of a number of linked grate sections forming an endless belt which

moves on rollers. A suction box is located under the linked grates, and the speed of the belt

movement is adjustable.

A charge of fine feed material, generally 12.5mm in diameter or less, or pre-formed pellets, are

moistened, mixed, and fed in a layer several cm deep onto the moving pallets ahead of the suction

box. As the charged pallet comes over the suction wind box, the sulphides in the charge are

ignited from above by a fuel fired burner. The process requires no additional fuel, and the

reaction temperature is maintained by the exothermic heat given off as the sulphide is oxidized by

the air drawn through the charge.

The roasting zone travels downward through the pallet charge as the grates move forward over

the sectionalized windbox, and the combustion zone gradually passes through the entire layer

thickness from top to bottom before the roasted material is discharged from the sintering

machine.

The high temperature of roasting heats the charge components high enough to make them sticky,

adhering to one another and forming a strong porous cake. However the thinness of the charge

layer and the cooling effect of the air drawn into the wind box prevent any extensive melting and

it is only the particle surface layers that become soft and sticky. Any molten material formed

would shut off air penetration and terminate roasting, so excessive temperatures must be avoided.

At the end of the horizontal travel of the moving grates, and when roasting is completed to the

bottom of the charge layer on the pallet, the grates are dumped inside a dust hood. The sinter cake

is sized, with the coarse portion going on to become furnace or retort feed and fines returned as

revert feed to the sintering machine.

The sintering machine described above is a downdraught type, which has the suction windbox

below the pallet grates and air is sucked down through the bed from top to bottom. There is a

second type, the up-draught type, which also has a wide industrial acceptance. With the up-

draught type the wind-box is above the grate, drawing air up through the charge on the pallet.

Ignition is made initially on a thin layer of feed placed on the grate. Then after ignition has started

a thicker layer of feed material is added on the top of the burning portion and burns upward from

bottom to top as the pallet moves under the wind-box toward the discharge end.

Grate speed for both types varies over a wide range, from 25-120cm/min and will depend on the

degree of roasting and/ agglomeration desired, the bed depth, and the length of the machine.

SMELTING

Smelting is a concentration process where some of the impurities in the charge are gathered into a

light waste product called slag, which can be separated by gravitational flow from the heavier

portion containing practically all of the desirable metal components. The charge to a smelting

furnace is made up principally of solids, though some molten material may also be charged in

certain operations. The heat supplied to melt this solid charge can be fossil fuel, electricity, or

exothermic heat from oxidation of the charge if its sulphide. The charge is smelted to liquid state

to allow the gravitational separation of the slag and metal/matte layers and also to facilitate the

mobility and contact of the reacting compounds within the charge.

The slag component is made up mostly of oxides in the charge (both those found naturally in the

ore, and converted oxides). These oxides are high melting points materials and are usually fluxed

with SiO2 or CaO to form low melting slag making phases. The slag should be quite fluid at the

furnace reaction temperature so that metallics can easily settle through it to collect in their lower,

heavier layer.

Types of Smelting

There are two main types of smelting:

Reduction smelting

Produce an impure molten metal and a molten slag from the reaction of metallic oxide with a

reducing agent. The metallic values in the charge, as well as the slag-forming compounds, will be

present as oxides. A reducing condition is arranged in the furnace whereby these metallic values,

which have a greater reduction rate from oxides to metals than do oxides in the gangue portion,

will reduce to an impure metal and leaves the gangue still as oxides to make up the slag. Any type

of furnace can be adapted to reduction smelting. Most commonly used are the blast furnace and

electric furnace

Matte smelting

It produces a molten mixture of metal sulphides and slag. Matte is formed by the combination of

the liquid sulphides of copper, nickel, iron, and cobalt into a homogenous solution. The PGMs

present and small amounts of other base metals are absorbed in the matte. The remaining portion

of the charge combines to form an oxide slag. The process can be done in reverberatory furnace,

blast furnace, electric furnace, flash smelting furnace and continuous smelting process consisting

of three furnaces in series.

REVERBERATORY SMELTING PROCESS

The reverberatory furnace is a fossil-fuel hearth furnace in which a solid charge of fine

concentrate and flux (usually silica) is melted to 1200-1230oC by hot combustion gases sweeping

over the entire hearth to produce two immiscible liquids. The combustion gases are provided by

burning fossil fuels in end-wall or roof burners. The products of the process are:

molten matte in which most of the valuable metals are found (Ni, Cu, Co and PGMs)

molten slag in which most of the gangue minerals are found (waste oxides) and the

values are low (<0.8% valuables)

waste gases (< 2 volume %SO2)

The reverberatory furnace is primarily a melting process i.e. it makes little use of the energy from

S and Fe oxidation for heating and melting.

Advantages of the Process

high degree of versatility, all types of material, lumpy or fine, wet or dry can be readily

smelted

Disadvantages of the Process

it uses large quantities of hydrocarbon fuels, making little use of the energy which is

potentially available from oxidizing the sulphide charge

its off-gas is dilute in SO2, (<2 volume%), making its efficient removal difficult (only a

small amount of S is oxidized to SO2 and is greatly diluted in CO2, H2O and N2 from

fossil fuel combustion)

Description of the Process

It is continuous process, i.e. the furnace is continuously fired and matte and slag are continuously

produced from the solid charge. The solid feed is periodically charged along the sidewalls near

the burners where it forms banks which serve as reservoir for continuous melting. The heat for

the smelting is supplied by fossil fuel burners firing through an end-wall or the roof. A variety of

fuels can be used (pulverized coal, natural gas, oil) giving a long flame which extends half the

length of the furnace. Part of the heat from this flame radiates directly on the charge lying on the

furnace hearth below, and part radiates to the roof and sidewalls from where it is also reflected

down on the charge.

The temperature of the furnace will be about 1600oC at the firing end and 1200oC at the flue end.

Molten slag produced is periodically tapped through an end-wall away from the charging zone,

and matte is periodically tapped from the sides of the furnace into ladles and transported to the

converters for further processing. The off-gas is continuously withdrawn from the slag tap end of

the furnace and is passed through a waste heat boiler and electrostatic precipitators for heat and

dust recovery respectively before being discharged through a high chimney into the atmosphere.

Recycle converter slag is also charged into the furnace from the burner end to recover the

valuables (Cu, Ni, Co). Both roasted calcines and unroasted concentrate can be used for feed, and

the burner flames can be oxygen enriched to increase smelting capacity and decrease the fuel

consumption per ton of charged material.

Construction Details

It is rectangular in plan, 8-10m wide, 30-36m long and 3-4m high, hearth to roof. The furnace is a

refractory chamber supported on solid foundation and held together by a steel superstructure. The

walls and roofs are erosion and high temperature deterioration resistant magnesite or chrome-

magnesite (Cr2O3.MgO) bricks. Roof bricks are usually suspended from a steel superstructure

above the furnace.

Temperature Profile within a Reverberatory Furnace

The temperature can reach a maximum 7-8m from the burner end. The gases then cool as they

transfer their heat to the charge and the slag. Gases leave the furnace 50-100 oC hotter than the

slag and carry a considerable amount of sensible heat. With all types of fuel, the burner system is

designed to delay part of the combustion until the fuel is well along the furnace

Reverberatory Heat Balances

Heat balance across a reverberatory furnace can be broken into three categories;

1. Heat Input and Production

includes the sensible heat (above 25oC) of the charge, the converter slag and the

air used in sulphide oxidation

also includes heat produced by the sulphide oxidation reactions ΔHR

2. Heat Output

includes sensible heat (above 25oC) of the slag, matte and the gases produced by

oxidizing the sulphide charge

also included are the heat used to vaporizes the water in moist charge, and convective

and radiative heat losses

The amount of heat which must be provided for smelting by the combustion fuel is given by :

Heat Output – (Heat Input and Production)

3. The 3rd item in the heat balance is:

efficiency at which the energy of the hydrocarbon fuel is applied to the smelting process

Thermal Efficiency =100 (Qi-Q0)/Q

where Qi= gross calorific value of the fuel (25oC) + sensible heat of the air used in its

combustion

Qo= sensible heat in the effluent gases originating from combustion of the fuel

Q= gross CV of the fuel (defined as the heat produced by unit weight of fuel (25oC) with air (at

25oC) to give products of combustion at 25oC)

Sensible heat is the heat content of air or gas with reference to a heat content of zero at 25oC

Thermal Efficiency represents the heat of combustion less the heat carried out of the furnace.

Production Rates

Smelting rates in reverberatory furnaces are in the order of 2-4tons of charge per square metre of

hearth per day. Lower rates are for wet concentrates and higher rates for hot calcines. Smelting in

reverberatory furnace is primarily a melting operation and hence the rate of smelting is:

directly proportional to the rate of heat transfer to the charge, and

inversely proportional to the quantity of heat required per unit of charge

smelting rate (tons/hr) q/C

where q= rate of heat transfer to the charge (KJ/hr); C= heat required per tonne of charge (kJ/ton)

Table: Simplified heat balances for the smelting of wet concentrate, dry concentrate and

calcine in Reverb Furnace

Item Wet concentrate (10.7%H2O) Dried concentrate Calcine

Heat Input and Production (x105 kcal)

Sensible heat in converter slag

(1200oC)

1.7 1.7 1.7

Sensible heat in air for

sulphide oxidation (220oC)

0.2 0.2 -

Sensible heat in solid charge 0.7 (500oC)

Heat of matte and SO2

formation, i.e.

CuFeS2+O2matte+SO2

1.8 1.8

Total heat input 3.7 3.7 2.4

Heat Output (x105 kcal)

Sensible heat in matte

(1150oC)

1.9 1.9 1.9

Sensible heat in slag (1200oC) 1.9 1.9 1.9

Sensible heat in SO2 (1250oC) 0.4 0.4

Sensible heat in N2 from the

air used in S oxidation

1.0 1.0

Heat of vaporizing H2O from

the charge

0.8

Sensible heat in water 0.9

Heat losses (convection and

radiation)

2.3 2.3 2.3

Total heat output 9.2 7.5 6.1

Net Deficit to be made up by

fuel

5.5 3.8 3.7

Fuel energy required at 47%

efficiency (220oC air

preheat)

12x105 kcal 8.1x105 kcal 7.9x105 kcal

ELECTRIC SMELTING FURNACES

The electric furnaces are used for both reduction and matte smelting. The common direct-arc,

non-conduction hearth, three-electrode electric furnace is mostly used for the reduction smelting

and smaller matte smelting furnaces, while for large tonnage matte smelting the submerged arc

type resistance furnace with rectangular shape and six electrodes in line, 3 pairs individually

connected, is most common used.

The direct-arc furnace charges are heated principally by radiation from the arc, as current flows

from the electrode to the charge, and especially where the arc strikes the charge. Some heat is

also generated by passage of current through the charge. Most common arc furnaces are 3-phase

types and use 3 electrodes, one connected to each phase. The charge then, in effect, completes the

circuit of each pair of electrodes in turn.

The matte smelting furnace is not an arc furnace but is a resistance furnace, with the electrodes

dipping into the slag layer. The electric matte smelting furnace is an electrically heated hearth

furnace. It performs the same functions as the reverberatory furnace:

melt dried or roasted concentrates to produce molten matte, molten slag and SO2 bearing

off-gas

treats molten recycle converter slag for Cu/Ni/Co recovery

Much of the heat for heating and melting the charge is provided by passing electric current

through molten slag between electrodes. It is used in several locations where electricity is

inexpensive and where SO2 must be tightly controlled.

Advantages of the Process

it is completely versatile and it can be used to smelt any material, lumpy, fine etc

it produces small volumes of effluent gas

the SO2 concentration of its effluent gas is readily controlled by adjusting the amount of

air which is infiltrated into the furnace

makes efficient use of the electrical energy (high electrical-heat conversion efficiency)

easy and accurate control of temperature

Disadvantages of the Process

makes little use of the energy which is potentially available from oxidizing the sulphides

in the charge

operating cost tend to be high due to high price of electrical energy

SO2 concentration is too low for efficient removal as H2SO4

Description of the Process

The electric matte smelting furnace is a submerged electrode process. Heat for melting is

generated by the passage of electric current through molten slag between submerged carbon

electrodes. The furnace is continuously heated, and matte and slag are continuously produced

from the charge. Electric furnaces can have a number of shapes, but for matte smelting the

predominant shape is rectangular in which there are six equi-spaced in-line self baking carbon

electrodes. (Soderberg type).

Dry charge (concentrate + flux + reverts) is added to the top of the slag through roof ports. Some

of the Fe and S in the charge is oxidized by in-leaked air when concentrate falls into the furnace

and while it rest on the surface of the blanket. Molten converter slag is recycled to the furnace for

base metal recovery and is charged at the opposite end to the slag tap hole end. Slag from the

furnace is tapped intermittently and discarded usually after granulation. Molten matte is also

tapped intermittently and sent forward to converting.

Construction Details

The furnace sits on steel base plates seated on concrete or brick piers to avoid leakage of

electrical current to the ground. The bottom of the furnace is cooled by natural convection of air

beneath the furnace or sometimes air can also be blown with fans if additional cooling is required.

The roof is the sprung arch type made of light inexpensive fireclay bricks. Magnesite and

chrome-magnesite bricks are used only where there is contact with matte or slag.

Diagram

Electrical Systems

The electrodes are the self-baking Soderberg type, formed by adding a paste of tar (20%) and

ground anthracite coal (80%) into cylindrical steel finned casings at the top of the electrodes.

Heat from the furnace and from electrode resistance melts the paste and vaporizes its volatile

components to form a new baked section of the electrode. Once in use the electrodes are slowly

eroded away as they are oxidized to CO by reaction with slag.

Fe3O4 + C 3 FeO + CO

To counter this, electrodes are slowly lowered through hydraulic slipping mechanism and by

adding new paste on top. Electrode consumption is usually 2-3kg carbon/ton of charge.

Power Input/Productivity Control

Power being applied to the electric furnace is made by altering the voltages between the

electrodes.

Power = (voltage between the electrodes)2/(resistance between the electrodes)

Power increased (to increase concentrate smelting rate) by increasing voltage and vice versa.

(voltage altered by changing transformer load taps)

Power rating of the furnace is calculated from:

Power (KW) = (tons of charge/dy /24)*(energy requirement, KWh/ton of charge)

Control

Power is automatically controlled at set point by raising and lowering the electrodes. The control

is based on:

a) low electrical resistivity of matte and high electrical resistivity of slag

b) existence of parallel current paths between electrodes

i. electrode-slag-electrode

ii. electrode-slag-matte-slag-electrode

Raising the electrodes favour path (i) which increases resistance between electrodes and

decreases applied power. Lowering the electrodes favour current flow through low resistance

matte, increasing applied power. Furnace temperature is also controlled by raising and lowering

the electrodes. Temperature below the set point are increased by lowering the electrodes (i.e.

increasing power) whilst temperature above the set point are decreased by raising the electrodes.

Matte and slag can also be controlled somewhat independently.

Example

You are given an electric furnace of power rating 14MW and energy consumption of 610kWh

per ton of charge. Given also that the furnace has an availability of 95%, calculate its monthly

throughput.

FLASH SMELTING

These are recent development post world war II for matte smelting of copper. Their application

has since been extended to matte smelting of other metals. The essential feature is that they are

autogenous i.e. they use oxygen to oxidize the sulphur and iron in the concentrate, and this

produces enough heat to self-sustain the process. The reactions are exothermic, for example in the

flash smelting of chalcopyrite the reaction may be represented by:

2CuFeS2 + 5/2O2 + SiO2 Cu2S.FeS + FeO.SiO2 + 2SO2 + heat

The reaction provides much or all of the heat required for heating and melting. Sulphur dioxide

concentration in the off-gas is very high (>10 volume %) and can be efficiently removed as

H2SO4, liquid SO2 or elemental sulphur.

There are two types of flash smelting:

the Outokumpu process (uses pre-heated O2 enriched air)

the INCO process (uses commercial oxygen as the oxidant)

Flash smelting consists of blowing dry concentrate together with oxygen, hot air or a mixture of

both into a hot hearth type furnace. Once in the furnace, the sulphide particles react rapidly with

their accompanying oxidizing gases to form two molten liquids.

Advantages of the Process

make use of the energy which is available from oxidizing the sulphide minerals (hence

fuel cost are low)

the waste gases are rich in SO2 which can be efficiently removed as H2SO4, liquid SO2

(INCO)

production rates are high due to the rapid rates at which the minerals particles are heated

and oxidized (8-12tons charge/day/m2 of hearth area, which is 2-4* that of reverberatory

or electric furnace smelting)

Disadvantages of the Process

metal content of their slags is high

OUTOKUMPU FLASH SMELTING PROCESS

It uses air or oxygen enriched air, preheated to 400-100oC as the oxidizing gas. The furnace has

got burners which are situated at the top of a combustion tower at one end of the furnace and the

concentrate, fluxes and gases are blown down the tower and onto the slag surface. Effluent

furnace gases leave via an off-take tower at the opposite end of the furnace. The main features of

the furnace are:

1. concentrate burners (1-4), which combine dry particulate feed with O2 bearing blast and

blow them downward into the furnace

2. a reaction shaft where most of the reaction between O2 and the sulphide feed particles

takes place

3. a settler where molten matte and slag droplets collect and form separate layers

4. water cooled copper blocks tap-holes for removing matte and slag

5. an off-take for removing hot SO2 bearing gases, 10-40 volume % SO2 (containing about

10% of feed as dust)

Oil burners are sometimes placed at the top of the combustion tower and in the hearth region to

supplement the heat requirements.

The Outokumpu Flash furnace is surrounded by considerable equipment, all essential to its

successful operation: These are:

Input Side

solids feed dryer and delivery system

O2 production plant

blowers

Output Side

waste heat boiler

dust recovery and recycle system

off-gas extraction fans

H2SO4 plant

slag treatment system

OPERATIONS CONTROL

Matte Composition and Smelting Rate Controls

The concentrate feed rate is usually pre-determined and is based on concentrate availability,

overall production capability and economic goals of the company. The grade of the matte is set

by adjusting the ratio:

(O2 in blast input rate)/(concentrate feed rate)

Slag Composition Control

Controlled by the ratio SiO2/Fe mass ratio (~3/4 usually chosen). The quantity of silica based

upon producing a slag which has:

1. small solubility for metals (Cu/Ni)

2. sufficiently fluid for a clean matte-slag separation and easy tapping.

The ratio is controlled by adjusting the rate at which flux is fed to the furnace.

Temperature Control

Flash furnace can be controlled to give matte and slag temperature independently. Matte

temperature controlled by adjusting the N2/O2 ratio of the blast (which affect the reaction shaft

temperature). Can also be adjusted by altering the rate of fossil fuel combustion at the top of the

reaction shaft tower.

Slag Temperature is also generated by shaft temperature (N2/O2). It can be controlled

independently of matte temperature by burning fossil fuel in auxiliary burners above the slag

around the hearth.

INCO FLASH SMELTING

The concentrate, flux and O2 are blown horizontally into the furnace from both ends and the

waste gases are drawn off via a large central off-take. This design results in high flame

temperature over the entire length of the hearth area. The matte is withdrawn from the centre of

one sidewall and the slag beneath the burners at one end of the furnace. The features that

distinguish the INCO from the Outokumpu process are:

a. INCO process uses commercial oxygen, 95-98 volume % O2

b. its blast and particulate feed are blown horizontally into the furnace rather than

downwards

c. has no waste heat boiler

d. it is used to settle matte from molten converter slags

All of the energy for smelting comes from oxidizing the Fe and S of the concentrate feed. SO 2

concentration in off-gas is 70-80 volume % SO2.

The basic components of the furnace are:

a. concentrate burners, 2 at each end, through which commercial O2, dry concentrate, dry

flux and ground recycle materials are blown horizontally into the furnace

b. a central off-take through which the off-gas is drawn into cooling, dust removal and

sulphur dioxide capture system.

c. tapholes for withdrawing matte and slag

Diagram

Auxiliary equipment for the furnace includes:

solids feed dryer

O2 plant

off-gas cooling equipment

dust recovery and recycle system

SO2 capture system

Cu from slag recovery system (optional)

Comparison of INCO and Outokumpu Processes

INCO process has several advantages over the Outokumpu process:

it has a much lower overall energy requirement

its volume of effluent gas (per ton of charge) is small due to the absence of N2 and

hydrocarbon combustion products

SO2 concentration in its effluent gas is very high (80%) compared to 10-40% for

Outokumpu

its dust losses are low due to its relatively small volume gas flow rate

productivity (tons of charge/day/m2 of hearth area) is about 30% higher than that of the

Outokumpu process

slag losses relatively low (~0.7%Cu and usually discarded) compared to the 1% Cu for

Outokumpu process

ISASMELT (SIROSMELT)

SIROSMELT is a submerged-combustion and smelting process invented by the CSIRO in the

early 1980’s and commercialized during the 1980’s, largely by Mount Isa Mines Ltd. The

motivation for the development was to address the environmental issues of lead smelting. Stricter

regulations concerning lead emissions and ambient lead in air levels, and the necessity to reduce

capital and operating costs encouraged the development of alternative lead smelting processes to

replace the sinter plant/blast furnace combination. This technology has however since been

extended to other metals such as Cu, Ni-Cu.

The unique feature of the process is the top-entry lance which is used for injecting air (or oxygen

enriched air) and fuel (gas, oil or coal) into molten slag or matte baths to achieve submerged-

combustion. The resultant high heat and mass transfer rates make the process very intense and

high smelting rates are achieved. SIROSMELT is a versatile process and has a number of

metallurgical applications; e.g.,

1. Matte-smelting: Sulphide ores or concentrates can be partially oxidized to remove iron

and volatile impurities to form matte for further processing. The 15t/h plant at Mt Isa for

treating copper concentrate is an example of this application.

2. Oxidation smelting of sulphides: Sulphide ores or concentrates can be oxidized to form

slags with high metal oxide contents which can be subsequently reduced to produce

metal. The first stage of the ISASMELT for treating lead concentrates is an example of

oxidation smelting.

3. Reduction smelting of oxides: Oxidized materials can be reduced with coal to produce

metals. Materials treated this way have included oxidized ores and concentrates and slags

from oxidation-smelting and reduction processes. Large scale applications include the

first stage smelting of cassiterite concentrates the second stage of the ISASMELT process

for producing lead bullion.

4. Sulphide fuming: Under appropriate oxygen and sulphur potentials, volatile metal

sulphides can be fumed for subsequent recovery as oxide. Tin fuming can be done and

there is potential for fuming of lead and zinc sulphides.

Other applications have included scrap melting and secondary metal reclamation, battery paste

processing and chemical modifications of slags for further processing.

Isasmelt Matte Smelting

Process Description

The process entails:

a. dropping moist palletized solid feed into a tall cylindrical furnace

b. blowing oxygen enriched air through a vertical lance into the furnace slag bath

The products of the process are a matte/slag emulsion and strong SO2 bearing off-gas (35 volume

% SO2). The emulsion is tapped periodicall into an electric or fuel fired settling furnace for

matte/slag separation. The matte usually sent for conventional converting, (at Empress Nickel

Refinery, Kadoma it is usually sent for leaching. The slag (0.6-0.7%Cu) is discarded. Most of the

energy for smelting comes from oxidation of concentrate charge. Additional energy is provided

for by combusting:

coal fines in the solid charge

oil or gas blown through the vertical lance.

Construction Details

It is a vertically aligned steel barrel ~13m high and 3.5m Ø. It is lined inside with water cooled

copper slabs and chrome-magnesite refractory. The roof is water cooled copper slabs, water-

cooled steel panels. The lance consists of a stainless steel outer pipe (~0.5m Ø) for oxygen

enriched air and a steel inner pipe for oil or natural gas. The outer pipe is always immersed

~35cm in the smelting furnace slag, while the inner pipe terminates about ~1m above the slag

surface.

The lance is cooled by swirling its oxygen enriched air in the annulus between the pipes. This has

the effect of extracting heat from the outside pipe and causing a protective slag layer to freeze on

the pipe surfaces. The immersed lance tip slowly erodes away and is lowered to compensate for

this erosion, only to be replaced when almost 1m has eroded. (usually after one week)

Diagram

Operation and Control

Matte/slag temperature is measured by thermocouples inserted in the furnace walls. It is

controlled by adjusting the rate at which fossil fuel is supplied to the lance. Matte and slag

composition determined by XFF analysis of the matte/slag taps.

Matte composition: controlled by adjusting

(O2 input rate)/(concentrate feed rate) ratio

Slag composition: controlled by adjusting

(flux input rate)/(concentrate feed rate) ratio

CONVERTING (CONVERTERS)

These are devoted to the conversion of metallic compounds to the final metal. For e.g. Cu 2SCu

and Ni3S2Ni. Converting is oxidation of molten matte to form molten blister copper or Ni-Cu

alloy. It entails oxidizing Fe and S from the matte with air, sometimes oxygen enriched. It is

mostly done in the rotatable horizontal Peirce-Smith converters, which blows air or O 2-enriched

air molten matte into its molten matte through submerged tuyeres. Several other processes are

also used.

The main raw materials for converting are the molten matte from smelting. The other raw

materials are silica flux, air and commercial oxygen. Several value metal bearing materials are

also recycled to the converter, mainly as solidified matte reverts. The heat for converting is

supplied entirely by Fe and S oxidation.

Converting usually takes place in two chemically and physically distinct stages:

a. Slag-forming stage

Iron and sulphur are oxidized to FeO, Fe3O4 and SO2 by the following reactions:

FeS + 3/2O2 FeO + SO2

3FeS + 5O2 Fe3O4 + 3SO2

The melting points of these oxides are very high in-excess of 1350oC so silica flux is

added to form a liquid slag with the FeO and Fe3O4.

2FeO + SiO2 2FeO.SiO2

This stage is finished when the iron in the matte has been lowered to about 1%. The

principal product of the slag forming stage is impure molten Cu2S or Ni3S2-Cu2S, white

metal. Metal Making Stage

The remaining S attached to Cu2S or Ni3S2 is oxidized to SO2. The metal values are not

appreciably oxidized until it is almost devoid of S. Thus the blister copper or leach alloy

is low in both S and O (0.001-0.03%S, 0.1-0.8%O).

There are two types of converters encountered:

horizontal type: Peirce-Smith converters

vertical type

Peirce-Smith Converters

Industrially, hot matte (1200-1250oC) is tapped furnace into ladles. The ladles of matte are

poured into the converter. The matte contains nickel, copper, cobalt, iron and sulphur. Air is

blown into the converter through the tuyeres

Peirce-Smith Converters

CONTINUOUS MATTE SMELTING AND CONVERTING

Many different methods are employed in the extraction of copper from sulphide ores. The

Mitsubishi Process allows the direct production of copper blister, and has been practiced in many

plants the world over.

Sulphide copper ores are normally treated in two stages: the matte making stage, in which matte

and slag are produced by smelting; and the converting stage, in which iron and sulphur in matte

are progressively oxidised for the production of blister copper. Although the chemical reactions

in both process steps are oxidations, there is a significant difference in the partial oxygen pressure

employed. The smelting stage is less oxidising to ensure higher copper recovery, while the

converting stage is more oxidising so that residual iron and sulphur is removed. These two quite

distinct reaction conditions are most efficient when each stage is performed in separate furnaces.

The Mitsubishi process is therefore a three- furnace system comprising continuous smelting, slag

cleaning, and converting. An anode casting stage is then used to produce copper anodes.

The potential benefits of continuous multi-furnace are:

i. lower energy consumption

ii. elimination of the Peirce-Smith converting with its SO2 collection and air infiltration

difficulties;

iii. continuous production of high strength off-gas, albeit from two sources

iv. lower capital investment

v. lower production cost

vi. allow possible automation and minimal materials handling.

The Mitsubishi process is distinct from conventional smelting methods in that the matte grade can

be quickly raised to 65-70% copper while maintaining a low level of copper loss in the discard

slag. At converting a limestone flux is introduced. This collects oxidised iron (magnetite)

effectively and results in a very small tonnage of converter slag. This is readily water-granulated

and recycled to the smelting furnace.

Since the three furnaces are stationary, good roof design and small furnace outlets ensure

complete off-gas capture. In this way the expensive and difficult to maintain hooding systems

used on most rotary furnaces are eliminated. The molten materials exit the furnaces either by

continuous overflow or by syphon, and are conveyed by gravity to the next furnace by sloping

launders. This entirely eliminates ladles, fugitive ladle gases, overhead cranes, and ladle skulls.

Each furnace in the three-furnace system can be regarded as a "steady state" reactor. The bathline

within each furnace is constant (no tapping), simplifying refractory design and brick cooling.

Process control holds the matte grade steady and ensures optimum operating conditions right

through to continuous blister delivery to the anode furnaces. A major advantage of the process is

its effectiveness in capturing SO2. It produces two continuous strong-SO2 streams which are

combined to make excellent sulphuric acid or liquid SO2 plant feed. Also, the near absence of

crane and ladle transport of molten material minimizes workplace emissions from that source.

The Mitsubishi Process

The Mitsubishi process employs three furnaces connected by continuous gravity flows of molten

material. They are:

smelting furnace

electric slag cleaning furnace

converting

Smelting Furnace

It blows oxygen-enriched air, concentrate and SiO2 flux into the furnace liquids via vertical

lances. It oxidizes the Fe and S of the concentrate to produce matte and iron silicate slag. Its

matte and slag flow together into the electric furnace.

Solid particulate feed and oxidizing gas are introduced through 9 or 10 vertical lances placed in 2

rows across the top of the furnace roof. Each lance consists of 2 concentric pipes inserted through

the furnace roof. The Ø inside pipe is 4-6 cm, the Ø of the outside pipe 8-11cm. Dried feed is air-

blown from bins through the annulus between the pipes. The outside pipes extend downward to

0.5 -0.75m above the molten bath, the inside pipes above or just through the roof. The outside

pipes burn back and are periodically slipped down to maintain their designated tip positions.

The concentrate/flux/recycle feed meets the oxidizing gas at the exit of the inside pipe. The

mixture jets onto the molten bath to form a matte-slag-gas foam/emulsion in which the liquids,

solids and gas react with each other to form matte and slag. These continuously overflow together

through a tap-hole and launder into the electric slag cleaning furnace where they separate into

layers. The off-gas (30 to 45 vol%SO2) is drawn up a large uptake.

Electric Slag Cleaning Furnace Details

It is elliptical with three or six graphite electrodes in one or two rows along the long axis. It

accepts the matte/slag emulsion from the smelting furnace and separates it into matte and slag

layers.

Matte continuously underflows from its layer out of the electric furnace, and into the converting

furnace. Slag continuously overflows through a tap-hole. It is granulated in water and discarded.

Liquid residence times in the furnace are 1-2 hours. The purpose of the electrodes and electrical

power in the slag cleaning furnace is to keep the slag hot and fluid. heat being obtained by

resistance to electric current flow in the slag between the electrode and the slag leave the furnace

at 1250oC. Only a tiny amount of off-gas is generated in the electric furnace.

The Mitsubishi process is the most widely used of all the new technologies. Other technologies

are the Noranda, Worcra and the KIVCET processes.

Converting Furnace Details

The converting furnace continuously receives matte from the electri furnace. It blows O2-

enriched air blast (30-35 vol% O2) and CaCO3 flux onto the surface of the matte. It produces

blister copper (~0.7%S), molten slag (12-18%Cu) and SO2 bearing gas (~25 volume% SO2)

The blister copper continuously underflows from the furnace through a siphon and launder into a

holding furnace.

The smelting of concentrate and flux with oxygen enriched air takes place in the first unit. A

mixture of matte and slag overflows into and electrically heated furnace. This furnace also acts as

a settling furnace for the separation of matte from slag. Pyrite and coke are used to clean the slag

which may then be tapped and discarded with a composition of 0.4-0.5%Cu. Converting takes

place in the last vessel. Air is oxygen enriched to about 25% and blown into the bath. Cu2S is

converted to blister copper to about 98-99% Cu. Slag from this stage contains 7-15%Cu and is

returned to the first vessel. In this Mitsubishi process lime is added as flux instead of silica so

their principal component of the slag is CaO-FeO-Fe2O3

The process is highly automated with computer controls. The comparative smelting rates with

other technologies are:

Mitsubishi 4.7 tons/day/m2

Flash 2.0 tons/day/m2

Reverberatory 0.8 tons/day/m2

These are on an unroasted concentrate.

NORANDA PROCESS

The Noranda continuous furnace is a cylindrical type furnace very similar to Peirce-Smith

converters. Smelting produces three layers of slag, matte and metal, since the process is

continuous control of feed, flux and air is essential.

This process has not developed to work as a stand alone process because the products are of

lower quality than conventional one. Copper metal produced contains up to 2%S instead of the

usual <1%S. The slag contains 8-12% Cu and therefore cannot be discarded. There is therefore

need for an accompanying conventional plant to further treat the slag.

The furnace is divided into three zones:

smelting zone

converting zone- oxidation of Cu2S to Cu

settling zone

A SO2 gas is of low quality and cannot be used for H2SO4 manufacture.

KIVCET SMELTING

The feed mixture (from Feed handling) for the smelter consists of silica and limestone for

fluxing; lead concentrate; zinc plant residues rich in iron, zinc, and lead; recycled battery scrap;

fine, dry coal for fual and moderately coarse coke (about 5 to 15 mm). The coke is an important

part of the process chemistry.

The dry feed is injected at the top of the reaction shaft along with oxygen. In the reaction shaft,

the sulphur in the lead sulphide concentrate and the fine coal ignite instantly to for a hot,

concentrated sulphur dioxide gas and the lead, zinc, iron, and other metals form metal oxides. The

fluxing agents and the oxides form a semi-fused slag which falls to the bottom of the first

compartment in the furnace along with the coarse coke. The coke collects as a surface layer,

called a "coke checker", floating on top of the molten slag. When the metal oxides percolate

through this layer of burning coke, they are reduced and the lead is converted to metal as bullion.

The bullion continues to settle through the molten slag layer beneath the coke checker. Together

with the zinc-bearing iron slag, the bullion passes under a partition wall into a compartment,

which is an electric furnace. This partition wall extends into the molten slag forcing the hot

sulphur dioxide gas to pass through the waste heat boiler and on to the electrostatic precipitator

rather than into the electric furnace compartment.

The larger second compartment serves primarily as a settling area where the heat from large

graphite electrodes keeps the bullion-slag bath in a molten state. The lighter slag continues to

float to the surface and the heavier bullion sinks to the bottom of the compartment. This

separation enables them to be tapped separately from the furnace.

Slag Fuming Furnace where fine coal and air are injected into it. This injection generates more

heat and causes the zinc to vapourize to form a mainly zinc oxide fume (also contains residual

lead and silver, cadmium, indium and germanium), which is collected and further treated in the

oxide leaching plant to recover the zinc, indium, germanium and cadmium. The molten slag is

held in the slag fuming furnace until the practical limit of 2% zinc in slag is reached. Then the

slag is poured into a stream of water to solidify it into a black sand-like barren slag which is

collected and sold to various cement manufacturers.

The bullion produced in the Kivcet furnace contains silver, gold, bismuth and copper which must

be removed before the lead can be sold to customers, primarily battery manufacturers. The copper

is removed in the drossing plant adjacent to the Kivcet furnace. There, the continuous drossing

furnace cools the lead bullion down from 900ºC to just over 400ºC. This cooling step forces

copper matte to form and float to the surface where it can be removed for processing in the

copper plant. The bullion, which still contains silver, gold, bismuth, arsenic and antimony, is next

put through a "softening" stage which uses oxygen to remove some of the arsenic and antimony

in the form of a slag. The bullion is then ready for electro-refining in the lead refinery

Smelting Stage

C(s) + O2(g) CO2(g)

C(s) + ½O2(g) CO(g)

S + O2(g) SO2(g)

Oxidation of Sulphides

PbS + 3/2 O2(g) PbO + SO2(g)

ZnS + 3/2 O2(g) ZnO + SO2(g)

Formation of Metallic Lead

PbS + 2 PbO 3 Pb + SO2(g)

Reduction Stage

PbO + C Pb + CO(g)

PbO + CO(g) Pb + CO2(g)

ZnO + C Zn + CO(g)

ZnO + CO(g) Zn + CO2(g)

C + CO2(g) 2 CO(g)