Zinc smelting process using oxidation zone and reduction zone

A smelting process of superior energy efficiency, particularly suitable for the extraction of zinc, but also of lead and copper from raw materials comprising sulphide concentrates, and oxidized metalliferous materials, is carried out by establishing at least one oxidation zone and at least one reduction zone in a furnace, each zone comprising a slag bath and a gas space; the gas spaces of the oxidation and reduction zones being separated by a partition; feeding raw materials comprising metal sulphides, together with fluxes, to the or each oxidation zone, and feeding oxygen or oxygen-rich gas into the slag bath in the or each oxidation zone to oxidize metal compounds therein and produce an off-gas containing SO.sub.2 ; transferring metal-oxide-bearing slag from at least one oxidation zone to at least one reduction zone; introducing coal into the or each reduction zone to reduce metal oxides therein, and produce gases containing carbon monoxide and hydrogen; introducing a controlled stream of oxygen into the or each reduction zone, the amount of oxygen in the said stream being sufficient to partly post-combust carbon monoxide and hydrogen, thereby producing heat which is transferred back to the slag bath in the reduction zone, while leaving the gases sufficiently reduced to ensure that reduced metal is not re-oxidized; and recovering reduced metal from at least one reduction zone.

SUMMARY OF THE INVENTION 
This invention relates to a pyrometallurgical smelting process for the 
extraction primarily of zinc, but also of lead and copper from raw 
materials consisting of sulphide concentrates, particularly mixed sulphide 
concentrates and oxidised metalliferrous materials. 
The process uses oxygen to oxidise the sulphides in the charge, with the 
exothermic heat released melting the charge to form a high zinc oxide 
molten slag. Coal is used as the reductant together with additional coal 
and oxygen to provide supplementary heat to the reduction stage. Zinc and 
lead vapours from the reduction stage are condensed to produce liquid 
metals. The process is amenable to a high degree of control and to 
automation. 
The process eliminates the need for sintering or roasting operations and 
gives high metal recoveries. 
BACKGROUND TO THE INVENTION 
Historically, zinc has been produced in retorts, in electrolytic cells and 
in the Imperial Smelting Furnace (ISF). Today almost all zinc is produced 
by the electrolytic process or the ISF process. The ISF is used to treat 
dirty and low grade concentrates as well as some secondary materials. To 
some extent its economic viability is threatened because of its large 
usage of high cost coke, its low thermal efficiency and the need for 
separate sintering and refining stages. To an extent the economics of the 
electrolytic process are dependent on low electricity costs. The process 
also produces a residue that presents disposal problems. 
This invention constitutes an advantageous alternative to the ISF and has 
the advantages of better utilisation of exothermic heats of reaction, use 
of low cost coal as reductant instead of metallurgical coke, and 
production of a high calorific value by-product gas. 
The KIVCET process (Australian Pat. Nos. 421,261; 456,550) has been used 
commercially to treat dirty copper concentrates, and in a modified form 
has been piloted to treat lead concentrates including some containing 
amounts of zinc. It has not been applied commercially to materials in 
which zinc is the predominant metal. 
The copper variant of the KIVCET process involves cyclone smelting of dried 
concentrates with oxygen to produce a molten bath containing any lead and 
zinc present in oxidised form as a slag and the copper as a matte. The 
sulphur dioxide off-gas is of high tenor and can be used for acid 
production. Slag and matte pass under a partition wall into an 
electrothermic furnace zone where reducing agents are added, to the bath. 
Heat is provided by electricity. Lead metal may be recovered from the 
slag, zinc may be fumed off and copper matte remains unchanged. The latter 
is transferred to a converter to produce blister copper. 
The KIVCET process for lead is similar, but is not considered suitable for 
adaptation to predominantly zinc containing materials unless low cost 
electricity is available on account of the high energy requirements 
involved. 
The Outokumpu company in Finland has investigated the smelting of lead and 
is also studying the fuming of zinc from slag produced in lead smelting. 
According to Outokumpu (Australian Pat. No. 501,911), concentrates are 
flash smelted to produce molten lead metal and a liquid slag which are 
passed to an electrothermic furnace where zinc is fumed off by injecting 
fine coal entrained in nitrogen. When applying this process to materials 
containing high percentages of zinc, a large amount of heat is produced as 
a result of the oxidation of zinc sulphide. In flash smelting, this heat 
is immediately used up in the volatilisation or large amounts of lead 
which gives rise to a large unwanted circulating furnace load. 
In the QSL process (U.S. Pat. No. 4,266,971), developed by Lurgi in West 
Germany, a molten bath of slag is fed with pelletised lead concentrates. 
The pellets enter the bath into which oxygen is injected. A high lead 
oxide slag is produced which is subsequently reduced to metallic lead in a 
second part of the furnace by injecting coal. As far as is known, this 
process has not been developed for the fuming of zinc. 
The ISASMELT process (Australian patent application No. 90762/82), is 
another bath smelting process for the smelting of lead concentrates. It 
consists of adding lead sulphide to a molten slag, injecting air into the 
slag bath to agitate the bath and to oxidise the lead sulphide to lead 
oxide and subsequently reduce the oxide to lead metal. Any zinc in the 
lead concentrates and remaining in slag, may be recovered by the addition 
of a slag fuming stage to the process. 
A process which is aimed at predominantly zinc containing raw materials has 
been proposed by Davey and Yazawa. The process concept involves the 
following reaction at elevated temperatures 
EQU ZnS+O.sub.2 =Zn+SO.sub.2 
As this reaction is endothermic, addition of carbonaceous fuel is required 
resulting in a mix of carbon oxides and sulphur dioxide in the off-gases. 
This makes it difficult to recover the heat value in any unburnt carbon 
monoxide and may lead to thermal inefficiency. There is also a potential 
problem of recovering zinc vapour from the complex gas stream.

DESCRIPTION OF THE INVENTION 
The invention involves the oxidation of zinc sulphide concentrates in a 
molten slag bath with industrial oxygen to produce an oxide slag, followed 
by reduction with coal to fume off metallic zinc which is collected in a 
lead splash condenser. 
The oxidation reactions are exothermic and when using dry sulphide feed 
materials, oxidation is autogenous. The reactions are typified by 
EQU 2ZnS+3O.sub.2 =2ZnO+SO.sub.2 
Most importantly the excess heat of oxidation of sulphide materials is used 
to melt the oxidised zinciferous materials and fluxes and to cause them to 
enter the oxide slag. During oxidation, volatile impurities, such as 
arsenic, antimony and cadmium, are eliminated in the fume phase. 
The oxidation stage may be modified by including and maintaining a molten 
sulphide phase (matte). Such matte phase aids in the oxidation of slower 
reacting feed materials, e.g. coarser particles. If there is sufficient 
copper in the feed materials the production of a copper matte or copper 
metal may be warranted. 
Reduction of zinc oxide to zinc vapour is carried out under strongly 
reducing conditions. The reduction reactions are endothermic. Thus 
additional coal must be burnt in the slag bath with a limited oxygen input 
so as to ensure that the combustion product is principally CO. 
Additional heat is introduced into the slag bath by partially burning the 
off-gases above the slag bath. To this end heat is transferred to the bath 
while still maintaining sufficiently reducing conditions to prevent the 
reversion of zinc to zinc oxide. Such "post-combustion" can be achieved by 
injecting oxygen into the top layers of the bath or the top space above 
the slag bath so as to burn as much of the CO as possible while still 
maintaining a suitable reversion temperature. This would give a 
CO/CO.sub.2 ratio of about 2. 
Some of the heat released in the post combustion step that is not 
transferred back to the slag bath can be utilised to reduce recycle zinc 
and lead materials by contacting such materials with the hot reducing 
off-gases before they enter the condenser. The post-reduction step can be 
used to adjust the off-gas composition and temperature to approach the 
zinc reversion conditions. 
The volume of discard slag produced on reduction in a reduction zone or 
reactor can easily be lessened-and so lessen zinc losses and flux 
requirements-by adding a further stage to reduce iron from part of the 
discard slag stream. The reduced iron is then returned to the preceding 
zinc reduction stage where it will contribute to the reduction of zinc 
oxide. The slag from the iron reduction stage is returned to the oxidation 
stage to flux iron contained in fresh feed materials thereby lessening the 
required quantity of flux. 
Any iron formed is not molten at 1350.degree. C. unless the slag is reduced 
to such an extent that additional carbon dissolves in the iron to lower 
its melting point. It is preferable, however, to have a bath of molten 
copper in the iron reduction reactor to dissolve the iron produced and to 
return it to the zinc reduction stage in the form of a metallic solution. 
FIG. 1 shows a flowsheet for the process with stages for oxidation, 
reduction plus post combustion, condensation and impurity bleed from 
recycle fume. 
Essentially the slag comprises FeO, Fe.sub.2 O.sub.3, CaO, SiO.sub.2 
Al.sub.2 O.sub.3, PbO, ZnO components; a typical compositon would be 
EQU 20% FeO, 15% CaO, 20% SiO.sub.2, 5% Al.sub.2 O.sub.3, 10% PbO, 30% ZnO. 
The oxidation reaction is generally driven at a temperature of about 
1350.degree. C. The feed materials (concentrates, residues, reverts etc.) 
are injected below the surface of the molten slag bath or by feeding them 
onto the surface of slag baths. Oxygen is injected through tuyeres or 
lances. 
Primary oxidation will occur in the course of submerged combustion of 
particulates at or near the tuyere tip. Injected oxygen will also oxidise 
some of the iron oxide in the slag from ferrous to ferric. Part of the 
sulphides added will be oxidised by the ferric iron in the slag. The 
reaction sequence is believed to be as follows 
In the vicinity of the tuyere or tuyere plume 
EQU 2ZnS+3O.sub.2 =2ZnO+2SO.sub.2 (1) 
EQU 12FeO+3O.sub.2 =6Fe.sub.2 O.sub.3 (2) 
In the bulk of the slag 
EQU ZnS+3Fe.sub.2 O.sub.3 =ZnO+6FeO+SO.sub.2 (3) 
In one important aspect is it desirable to maintain conditions such that 
zinc oxide concentration in solution is at a maximum to lessen the slag 
volume and hence the loss of zinc in slag. To achieve this the temperature 
must be kept as high as possible, however, too high a temperature will 
result in excessive fuming, especially in the presence of lead compounds. 
A temperature of 1350.degree. C. has been found particularly suitable. 
In another aspect the oxidation potential of the slag is significant. If 
the slag is excessively oxidised in the presence of high zinc 
concentrations then a spinel phase is precipitated, making the zinc 
viscous and unworkable. It is necessary, in accordance with the invention 
to inject the concentrates and oxygen in such a way that reaction (2) is 
minimised. Alternatively concentrates in close proximity to the oxygen may 
be injected through the same tuyere. This is possible in the presence of 
liquid matte and copper phases and will be described later. 
The fume trom the oxidation stage is separated from the SO.sub.2 gases. The 
latter are usually greater than 50% tenor due to the use of oxygen and may 
readily be used for making sulphuric acid. The fume contains a high 
proportion of volatile impurities, such as As, Sb, Cd, Se. All the fume or 
a bleed stream of it can be treated hydrometallurgically to remove these 
impurities before returning the lead and zinc fume residues to the 
process. 
The liquid high zinc slag from the oxidation stage is passed to the 
reduction stage where coal and oxygen are injected below its surface. The 
reactions occurring are believed to be 
EQU ZnO+C=Zn+CO (4) 
EQU 2CO+O.sub.2 =2CO.sub.2 (5) 
The practice of injecting coal and air in batch slag fuming processes is 
well known and needs no description here. The continuous process of the 
invention uses oxygen or a rich oxygen/air mix instead of air to lessen 
the volume of off-gases and to yield a greater calorific value by-product 
gas which can be used to advantage as an energy source for zinc refining 
or for production of oxygen. 
As is generally known the reduced gases will leave the slag bath and be 
either, fully oxidised above the bath to form zinc oxide as in zinc slag 
fuming, or be condensed in a lead splash condenser as in the ISF process. 
It is a feature of the present process to introduce a controlled stream of 
oxygen into the upper part of the reduction stage or reactor so as to 
partly post-combust the contained CO and H.sub.2 to CO.sub.2 and H.sub.2 
O, but to still leave the gases sufficiently reduced to ensure that the 
zinc is not oxidised to zinc oxide. Maintenance of a residual CO/CO.sub.2 
ratio of 2 will ensure that the zinc is not oxidised as the gases pass to 
the lead splash condenser. Zinc is recovered from the condenser and 
refined by distillation as in the ISF process. The gases from the 
condenser have a higher fuel calorific value than the gases from an ISF as 
they are not diluted by a large volume of nitrogen and may be used as an 
energy source for an associated oxygen plant. 
It should be noted particularly that the aforesaid post-combustion is 
carried out in such a way that most of the heat of combustion is 
transferred back to the slag bath. This provides some of the heat required 
for the endothermic reduction reactions and for the volatilisation of 
zinc. It thereby reduces the amount of coal that would otherwise need to 
be burnt with additional oxygen to provide the heat. Such post combustion 
may be brought about by injecting oxygen into the top space above the slag 
bath and transferring heat back to the slag bath by radiation and 
convection. 
The temperature of the gases leaving the reduction reactor will be 
significantly higher than the slag bath temperature as they will contain 
portion of the heat liberated in the post-combustion step. A technique can 
be adopted whereby fine oxidic zinc and lead recycle materials, produced 
in the condensation, refining and fume treatment stages of the process of 
the invention, are injected into the hot gas stream. Such oxides are 
reduced to metal vapours, by the CO present in the gas together with 
sensible heat in the gases. The net effect of post-reduction is to improve 
the energy efficiency of the process by using some of the chemical energy 
and lowering the temperature of the gases to just above the reversion 
temperature; this also reduces the heat load on the condenser. 
As an alternative to post combustion, a similar heat gain or release may be 
achieved by injecting oxygen into the slag bath at different positions to 
those at which coal is injected, thus providing strong local reducing 
conditions at coal injection positions and producing heat at the more 
oxidising conditions near the oxygen injection points. 
The slag (tail slag) produced in the reduction zone or reactor has been 
found to contain only a few percent of zinc and can be discarded. 
FIG. 2 shows a conceptual diagram of one type of reactor for carrying out 
the process described above. The reactor 1 is a water-cooled 
refractory-lined vessel. Partition 2 in the gas space above the slag bath 
separates the SO.sub.2 gases leaving the oxidation zone 3 through off take 
4 from the zinc laden gases leaving the reduction zone 5. Optionally a 
partition 6 in the slag bath may be provided between the oxidation and 
reduction zones of the bath. Additional partitions in the slag bath could 
be provided to give multiple reduction zones. Tuyeres 7 serve the 
injection of oxygen and/or charge materials, tuyeres 8 the injection of 
oxygen and/or coal, tuyeres 9 are for effecting post-combustion. Waste 
slag is withdrawn through a continuous tapper 10. Zinc laden gas passes 
out through off-take 11. Other arrangements or even a series of separate 
reactors could achieve the reaction sequences of this method. 
The charge and oxygen are injected into each zone through tuyeres 
positioned below the slag level. Post combustion ports are positioned 
above the slag level in the reduction zone. Slag is withdrawn from the 
reduction zone through a continuous tapper. 
FIG. 3 shows a flowsheet for a variant of the process which includes an 
iron reduction stage and a low iron slag recycle. The oxidation and zinc 
reduction stages are as described for FIG. 1, with the following 
additions: 
The tail slag from the zinc reduction reactor is split into two streams. 
One stream is discarded as before. It provides a discard for iron oxide 
and other gangue materials introduced with the feed materials and coal. 
The other stream is passed to a further stage of reduction where coal and 
oxygen are injected to reduce the iron oxide in the slag to iron metal. In 
another version of this variant a molten copper phase is provided in the 
reduction zone to dissolve the metallic iron produced and convey it in 
molten form to a zinc reduction furnace or reactor. 
The iron dissolved in copper will act as a reductant for the zinc and be 
reoxidised to iron oxide. The iron acting as a reductant in this way will 
replace coal that would otherwise have been added as reductant. The 
overall effect is that the coal requirement remains the same with the 
addition of the iron reduction reactor, except for a small increase to 
account for heat losses from the additional reduction furnace. The iron 
depleted copper is recycled to an iron reduction reactor to collect more 
iron. The reaction sequence is believed to be 
In the iron reduction reactor 
EQU (FeO)slag+C=CO+(Fe)copper (6) 
In the zinc reduction reactor 
EQU (Fe)copper+(ZnO)slag=(FeO)slag+(Zn)gas (7) 
It is economically important that the off-gas from the iron reduction 
reactor be used effectively. By putting the coal into the iron reduction 
reactor instead of the zinc reduction reactor, the two beneficial effects 
described below are largely lost 
(1) The effect of the combustion gases to dilute the zinc concentration in 
the zinc reduction reactor off-gas and so provide a lower equilibrium zinc 
concentration in the waste slag 
(2) The effect of the CO gas to release heat back to the slag bath in the 
zinc reduction reactor by post-combustion above the bath. 
To regain corresponding benefits when an iron reduction stage is included 
it is necessary to pass the iron reduction off-gas back to the zinc 
reduction reactor. This can be done by cleaning the gases and injecting 
them into the zinc reduction reactor slag bath. 
One aspect mentioned earlier as an aid to control in the oxidation zone was 
the presence of a molten matte phase. A cuprous sulphide rich matte phase 
is believed to play two roles in controlling the chemistry in the 
oxidation zone or reactor. 
Firstly, the matte has the capacity to dissolve the sulphide concentrates, 
resulting in a lower sulphur content in the slag phase removed from the 
reaction zone. This is thought to be due to the matte phase trapping and 
dissolving any unreacted zinc sulphide concentrate particles. In the 
absence of the matte phase these particles surrounded by a gas envelope of 
SO.sub.2 would be dispersed through the slag phase as discrete entities as 
they react with the Fe.sub.2 O.sub.3 in the slag (refer Reaction (3)). 
Secondly, the equilibrium, between copper matte, copper metal and SO.sub.2 
gas, buffers the oxidation potential at a level at which fuming is not 
excessive and spinels associated with slag iron will not form. It buffers 
by reversibly producing copper metal if the oxygen addition is excessive, 
rather than oxidising the iron in the slag to ferric oxide and 
precipitating spinel. Spinel phase precipitation can cause increased 
viscosity and slag foaming. If there is an inadequate oxygen supply the 
copper metal will reform cuprous sulphide. Under such conditions the matte 
also has the above mentioned capacity to dissolve the sulphide 
concentrates added rather than their being dispersed through the slag and 
giving rise to a high residual sulphur content in the slag. 
The reaction sequence described above could be summarised as below 
With balanced oxygen addition 
EQU 2ZnS+3O.sub.2 =2(ZnO)slag+2SO.sub.2 (8) 
With excessive oxygen 
EQU (Cu.sub.2 S)matte+O.sub.2 =2Cu+SO2 (9) 
EQU 2(ZnS)matte+3O.sub.2 =2(ZnO)slag+2SO.sub.2 (10) 
With insufficient oxygen 
EQU 2Cu+2ZnS+O.sub.2 =(Cu2S)matte+2(ZnO)slag (11) 
EQU 2ZnS=2(ZnS)matte (12) 
It should be noted that, while reaction (8) represents the overall reaction 
for balanced oxygen supply, reactions (9-12) can also be occuring 
simultaneously. 
A possible disadvantage of using matte in the oxidation reactor is that 
there will be an equilibrium level of about 5-10% Cu in the slag. This 
will need to be reduced in the zinc reduction reactor to recover the 
copper. The ratio of copper to zinc in the slag is about 1:4. If the feed 
materials have a lower ratio than this then the copper reduced from slag 
has to be returned to the oxidation reactor to maintain the matte phase. 
The slag will build up to the same Cu:Zn ratio and so imposes an 
additional reduction load on the process. If the feed materials have a 
higher ratio of copper to zinc, then a matte or preferably a copper metal 
phase is produced which is returned to the oxidation reactor. 
If lead is present in the feed materials, it will mainly oxidise into the 
slag phase. Some will report in fume which is recycled to the oxidation 
zone. Lead in slag is reduced in the reduction zone; for low ratios of 
lead to zinc it will be volatilised along with the zinc; at higher lead to 
zinc ratios, a separate lead bullion phase will be produced which can be 
tapped from the reactor for further processing to refined lead. 
Precious metals report in the molten matte or metal phases tapped and can 
be recovered from them by traditional methods. In the absence of matte or 
metal co-products, there will be no collector for the precious metals. The 
last variant described here is therefore most suitable whenever 
significant quantities of precious metals are present in the process. 
EXAMPLES 
Example 1 
This example is of a small scale experiment representative of the oxidation 
stage of the process of the invention (i.e. no matte phase present). 
A starting bath of 1733 gram of slag was melted in an alumina crucible. In 
commercial operation a suitably refractory lined steel vessel would be 
used. Oxygen and zinc concentrates were injected through a lance into the 
slag bath at a rate of 11 gram per minute and 5 normal litres per minute 
respectively for 60 minutes. The reactor was held at 1350.degree. C. Fume 
from the reactor was caught in a gas filter. The final slag and fume were 
weighed and analysed. The mass balance was as follows: 
__________________________________________________________________________ 
Material 
Mass g 
% Zn 
% Pb 
% Cu 
% Fe 
% SiO.sub.2 
% CaO 
% Al.sub.2 O.sub.3 
% S 
__________________________________________________________________________ 
Zn--Pb conc 
686 
35 16 0.4 13 0.8 0.2 0.2 31 
Start slag 
1733 
22 0.6 0.44 
19.2 
26.8 11.4 6.2 0.7 
Final slag 
1880 
27.7 
3.6 0.41 
18.7 
18.6 8.2 4.8 2.4 
Fume 113 
17.7 
56 0.05 
0.3 -- -- -- 6.4 
__________________________________________________________________________ 
When oxygen alone was blown into a slag bath of similar composition to the 
final slag above the sulphur content of the slag decreased to 0.6%. 
The experiment indicates that sulphur is removed from the bath at a 
practical intensity equivalent to 0.35 tonne per hour of feed per tonne of 
bath capacity while maintaining a fluid slag of relatively high zinc 
concentration. 
Example 2 
This example is of a small scale experiment representative of the reduction 
stage of the process of the invention. 
As stated zinc oxide slags are reduced in batch slag fuming processes using 
coal and air to produce zinc oxides. The method of the invention produces 
elemental zinc vapour which is then condensed to liquid zinc. The 
invention uses coal and oxygen or rich oxygen/air mixes, the latter to 
reduce the quantity of gas, and to increase its calorific value. This 
raises the zinc content of the off-gas from the slag bath as demonstrated 
below 
A starting slag bath of 300 g of low zinc content (2% Zn) was prepared and 
held at 1300.degree. C. Oxygen plus coal and nitrogen were injected 
through a lance below the surface of the slag. 
Oxygen feed rate was 1.4 l min., nitrogen feed rate (carrier gas for the 
coal) 1.5 l min., and the coal feed rate 1.5 g/min. 
Continuous feeding of high zinc slag was simulated by adding 25 g of slag 
containing 18% Zn every 3 min. Under these conditions, the slag bath 
equilibrated at 3.4% Zn after 30 min. 
Of the 1.5 g/min of zinc added in feed slag, 0.3 g/min reported in the slag 
bath and 1.2 g/min reported in the off-gas. The resultant off-gas was 
calculated to contain 10% Zn at a CO/CO.sub.2 ratio of 1:4. 
Example 3 
The following table gives examples of a range of high zinc slag 
compositions that were found to be fluid at a temperature of 1350.degree. 
C., and at oxidation potentials relevant to the desulphurisation reactions 
as in reaction (1) page 6. Fluidity was measured by a Herty fluidity gauge 
and found to very suitable for operation of the process. They were also 
judged by experience to be suitably fluid. 
__________________________________________________________________________ 
No. 
Temp. C 
% ZnO 
% PbO 
% Fe tot 
Fe.sup.3+ /Fe.sup.2+ 
% SiO.sub.2 
% CaO 
% Al.sub.2 O.sub.3 
__________________________________________________________________________ 
1 1350 48.8 0.3 
7.4 0.25 29.9 6.6 4.8 
2 1350 49.1 1.0 
7.4 3.5 24.0 10.3 3.3 
3 1350 38.6 14.1 
19.4 0.49 11.4 2.7 4.3 
4 1280 38.3 12.3 
13.7 0.23 18.1 7.3 3.8 
5 1350 34.2 11.0 
20.9 0.7 11.4 5.2 7.1 
__________________________________________________________________________ 
Example 4 
This example is illustrated in the attached flow sheet, FIG. 4. 
52 kg of a charge, of dried mixed lead and zinc sulphide concentrates and 
recirculated fumed material and fluxes, containing PbS 14%, ZnS 41%, FeS 
9%, Volatile Impurities 0.5%, Remainder gangue materials was injected into 
a slag bath located in a small hearth furnace with the amount of oxygen 
injected totalling 16 kg. The exothermic heat of the oxidation reactions 
melted the incoming materials and resulted in a bath temperature of about 
1250.degree. C. 
The slag produced during the oxidation stage contained 28% Zn and 12% Pb, 
with about 30% of the lead reporting into fume. The slag of this 
composition was quite fluid at the temperature of 1250.degree. C. 
This hot slag was passed to a separate vessel for reduction. Oxygen was 
injected into the slag bath and 11 kg of coal added as a reductant and to 
provide heat to keep the bath at 1250.degree. C., and maintain strongly 
reducing conditions. The lead and zinc was reduced to metal and 
volatilized from the bath. Oxygen was introduced above the bath via a 
lance to burn some of the CO to CO.sub.2 and reverberate the heat released 
back into the slag bath. The CO/CO.sub.2 ratio in the off-gas was kept 
above 2, so that the lead and zinc in the off-gas could be recovered in 
metallic form in a lead splash condenser, and to provide a gas with a high 
calorific value for subsequent use as a heating fuel. The zinc content of 
the waste slag was 3%, accordingly 97% of the zinc was recovered. Recovery 
of lead was practically complete with only some 0.1% reporting in the 
waste slag.