CA1279198C - Zinc smelting process using oxidation zone and reduction zone - Google Patents
Zinc smelting process using oxidation zone and reduction zoneInfo
- Publication number
- CA1279198C CA1279198C CA000505638A CA505638A CA1279198C CA 1279198 C CA1279198 C CA 1279198C CA 000505638 A CA000505638 A CA 000505638A CA 505638 A CA505638 A CA 505638A CA 1279198 C CA1279198 C CA 1279198C
- Authority
- CA
- Canada
- Prior art keywords
- slag
- zinc
- zone
- reduction
- oxidation
- Prior art date
- Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
- Expired - Fee Related
Links
Classifications
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B13/00—Obtaining lead
- C22B13/02—Obtaining lead by dry processes
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B15/00—Obtaining copper
- C22B15/0026—Pyrometallurgy
- C22B15/0028—Smelting or converting
- C22B15/003—Bath smelting or converting
- C22B15/0041—Bath smelting or converting in converters
-
- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B19/00—Obtaining zinc or zinc oxide
- C22B19/04—Obtaining zinc by distilling
- C22B19/10—Obtaining zinc by distilling in reverberatory furnaces
Abstract
ABSTRACT
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 oxidised 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 oxidise metal compounds therein and produce an off-gas containing SO2; 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-oxidised;
and recovering reduced metal from at least one 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 oxidised 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 oxidise metal compounds therein and produce an off-gas containing SO2; 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-oxidised;
and recovering reduced metal from at least one reduction zone.
Description
- l 12~9198 Smelting Process SUMMARY OF THE INVEN~ION
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 (IS~). Today almost all zinc is produced by the electrolytic process or the ISF process. The IS~ 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 laege 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.
.,,, . ,,, ~ . . ...
1279~L9~3 This invention constitutes an advantageous alternative to the ISF and has the advantages of better utilisation of exothermic heats of reaetion, use ot low cost eoal as reduetant instead of metallurgiaal coke, and production of a high calorific value by-product gas.
The KIVCET process (Australian patents 421,261; 456,550) has been used commercially to treat dirty copper concentrates, and in a modified form has been piloted to treat lead eoneentrates ineluding some eontaining amounts of zine. It has not been applied eommereially to materials in whieh zine is the predominant metal.
The eopper variant of the KIVCET proeess involves cyclone smelting of dried concentrates with oxygen to produce a molten bath containing any lead and zine present in oxidised form as a slag and the eopper as a matte. The sulphur dioxide off-gas is of high tenor and ean be used for aeid produetion. Slag and matte pass under a partition wall into an eleetrothermie furnaee zone where reducing agents are adqed to the bath. Heat is provided by eleetrieity. Lead metal may be reeovered from the slag, zine may be fumed off and eopper matte remains unehanged. The latter is transferred to a eonverter to produee blister eopper.
The XIVCET process for lead is similar, but is not eonsidered suitable for adaptation to predominantly zine eontaining materials unless low cost eleetrieity is available on aceount of the high energy requirements involved.
The Outokumpu company in Finland has investigated the smelting of lead and is also studying the fuming of zine from slag produced in lead smelting. Aeeording to Outokumpu (Australian patent 501,911), concentrates are flash smelted to produce molten lead metal and a liquid slag which are passed to an electrothermic furnaee where zine is fumed off by injecting fine coal entrained in niteogen. When applying this process to materials cont~ining high percentages of zinc, a lar~e amount of heat is produce~ as a result of the ---` 1279~L98 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 cireulating furnace load.
In the QS~ process (US patent 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 whieh oxygen is injeeted. A high lead oxide slag is produeed which is subsequently redueed to metallie lead in a seeond part of the furnace by injeeting eoal. AS far as is known, this process has not been developed for the fuming of zine.
The ISASMELT process (Australian patent applieation 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 zine in the lead concentrates and remaining in slag, may be recovered by the addition of a slag fuming stage to the proeess.
A proeess whieh is aimed at predominantly zine eontaining raw materials has been proposed by Davey and Yazawa~-- The proeess eoncept involves the following reaction at elevated temperatures -Zn~ + 2 = Zn + SO
As this reaction is endothermic, addition of carbonaeeous fuel is re~uired resulting in a mix of earbon oxides and sulphur dioxide in the off-gases. This makes it diffieult to recover the heat value in any unburnt carbon monoxide and may lead to ther~al inefficiency. There is also a potential problem of recovering zinc vapour from the eomplex gas stream ~'~79198 .
~I
DE~C~IPTION OF T~E INVENTION
The invention involves the oxidation of zinc sulphide concentrates in a molten slag bath w-ith 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 -2ZnS + 32 = 2ZnO + SO
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 conditioos to prevent the reversion of zinc to zinc oxide. Such ~post-combustion~ can be achieved by injecting oxygen into the top layers of the bat~l 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~ 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 1350C 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.
In one embodiment the present invention is directed to a zinc smelting process which comprises ta) establishing at least one oxidation zone and at least on 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;
(b) maintaining a molten matte phase rich in cuprous sulphide in the or each oxidation zone, feeding raw material comprising zinc sulphides and iron sulphides, together with :, ' 12~9~9~3 - 5a -fluxes, to the or each oxidation zone, and feeding o~ygen or oxygen-rich gas into the molten bath in the or each oxidation zone to oxidise metal compounds therein to produce a slag bearing zinc oxide and iron oxide and to produce an o~f-gas containing fume and SO2, said fume containing zinc and other volatiles;
(b.1) removing from said oxidation zone the off-gas produced in step(b), removing fume from said off gas, separating volatiles other than zinc from the removed fume to produce a zinc-containing fume resi.due, and recycling zinc-containing fume residue to the oxidation zone;
(c) transferring zinc oxide-bearing slag from at least one oxidation zone to at least one reduction zone;
(d) introducing coal into the or each reduction zone to reduce zinc oxide present in said zinc oxide-bearing slag to produce reduced zinc, discard slag, and reducing gases containing carbon monoxide and hydrogen, said discard slag containing iron oxide;
(e) 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 said carbon monoxide and hydrogen, thereby producing heat which is transferred back to the slag bath in the reduction zone, while leaving said reducing gases sufficiently reduced to ensure that reduced zinc vapor is not re-oxidised;
(f) recovering reduced zinc from at least one reduction zone; and (g) removing discard slag from at least one reduction zone, dividing the removed discard slag into two streams, passing one of the streams to waste, passing the other stream to an iron reduction reactor to reduce iron oxide present in said discard slag to produce iron metal, off-gases and a low-.. iron slag, recycling iron metal to the zinc reduction zone, and recycling low-iron slag from the reduction reactor to the oxidation zone.
~279~9~3 - 5b -In drawings which illustrate preferred embodiments of the present invention Figure 1 shows a flowsheet for the process with stages for oxidation, reduction plus post co~bustion, condensation and impurity bleed from recycle fume.
Figure 2 shows a conceptual diagram of one type of reactor for carrying out the process of the present invention;
Figure 3 shsws a flowsheet for a variant of the process which includes an iron reduction stage and a low iron slag recycle; and Figure 4 shows a flowsheet for the process described in Example 4.
Essentially the slag comprises FeO, Fe203, CaO, sio2 A12O3, PbO, ZnO components; a typical composition would be 20% FeO, 15% CaO, 20% sio2, 5% A12O3, 10% PbO, 30% ZnO.
--- 12~79~98 The oxidation reaction is generally driven at a temperature of about 1350C. 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 2ZnS + 32 = 2ZnO + 2SO2 (1) 12FeO + 32 = 6Fe23 (2) In the bulk of the slag ZnS + 3Fe2O3 = ZnO + 6FeO + SO2 (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 1350C
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 ~L~79~i9~3 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 from the oxidation stage is separated from the SO2 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 - ¦
ZnO + C = Zn + CO (4) 2CO + 2 = 2CO2 (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 H2 to CO2 and H2O, but to still leave the gases sufficiently reduced to ensure that the zinc is not .~ ~
~2~98 oxidised to zinc oxide. Maintenance of a residual CO/CO2 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.
.. ~ . . .
-- ~279198 - 8a -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 pcints.
The slag (tail slag) produced in the reduction zone or reactor has been found to contain only a few percent o~
zinc and can be discarded.
~ r 12791g8 ' Reference is now made to Figure 2. The reactor 1 is a water-cooled refractory-lined vessel. Partition 2 in the gas space above the slag bath separates the SO2 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. Addit:ional partitions in the slag bath could be provided to give multiple reduction zone.
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. ~inc 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.
The variant of the process described in the flowsheet of Figure 3 includes an iron reduction stage and a low iron slag recycle. The oxidation and zinc reduction stages are as described for Figure 1, with the following additions: -2~ 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 slagto 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.
---` 1279198 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 (FeO)slag + C = CO + (Fe)copper (6) In the zinc reduction reactor (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 ceduction reactor slag bath.
12~9~g8 .
One aspect Inentioned 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 SO2 would be dispersed through the slag phase as discrete entities as they react with the Fe2O3 in the slag (refer Reaction ~3)).
Secondly, the equilibrium, between copper matte, copper metal and SO2 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 -inade~uate 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 2ZnS + 32 = 2(ZnO)slag + 2SO2 (8) - 1279~8 With excessive oxygen (Cu2S)matte + 2 = 2Cu + SO2 (9) 2(ZnS)matte + 32 = 2(ZnO)slag + 2SO2 (10) With insufficient oxygen 2Cu + 2ZnS + 2 = (Cu2S)matte + 2(ZnO)slag (11) 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 occurring simultaneously.
A possible disadvantage of using matt 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 lS 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 manly 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 l~ad to zinc ratios, a separate lead bullion phase will be produced which can be tapped from the reactor for further processing to refined lead.
``` ~2'79~9B
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 1350C. 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 -~L2~9198 ~n . .
_I o ~ ~o o ~ ~ ~ oo ¢
o C~ `. . .
o o~o X
.,l . . .
U~o ~CO
_I
~P
G~00 0 ~r ~~. o o o o o . .
~o o ~ ~o C 1~ .
~ .
d~
~ .
V~ oo ~ o~ , U
_~ o ~ _~
UV~
.., ~ ,o ~ _~
S ~ u~ ~ L .
.,, U~
`-` 1279198 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 1300C. Oxygen plus coal and nitrogen were injected through a lance below the surface ofthe slag.
Oxygen feed rate was 1.4 1/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. ~nder these conditions, the slag bath equilibrated at 3.4% Zn after 30 min.
~279~g8 :
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/CO2 ratio of 1:4.
Example 3 The following table gives examples of a range or high zinc slag compositions that were found to be fluid at a temperature of 1350C, 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.
. ~.. ; .
o~ ~L27~
oP
o . . . .
~D O r~
d~
Oa~ o ~,~
... . . . .
+
a +
o ~ o o o aJ
a)I_1-- c,~ ~ o ~ ~ , d~
-~ , o Q~) o ~1~ o 0~1 ~ ~ ~1 C ~ ~ ~D
C~ . .
a~ o ~ o o o Eu~u ) u~o:) In Z ~ ~ ~ ~ n Lt~
.
. ' . ' .
18 - gL279198 EXAMPLÆ 4 This example is illustrated in the attached flow sheet~ ~ig. 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 1250C.
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 1250C.
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 1250C, 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 CO2 and reverberate the heat released back into the slag bath. The CO/CO2 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.
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 (IS~). Today almost all zinc is produced by the electrolytic process or the ISF process. The IS~ 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 laege 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.
.,,, . ,,, ~ . . ...
1279~L9~3 This invention constitutes an advantageous alternative to the ISF and has the advantages of better utilisation of exothermic heats of reaetion, use ot low cost eoal as reduetant instead of metallurgiaal coke, and production of a high calorific value by-product gas.
The KIVCET process (Australian patents 421,261; 456,550) has been used commercially to treat dirty copper concentrates, and in a modified form has been piloted to treat lead eoneentrates ineluding some eontaining amounts of zine. It has not been applied eommereially to materials in whieh zine is the predominant metal.
The eopper variant of the KIVCET proeess involves cyclone smelting of dried concentrates with oxygen to produce a molten bath containing any lead and zine present in oxidised form as a slag and the eopper as a matte. The sulphur dioxide off-gas is of high tenor and ean be used for aeid produetion. Slag and matte pass under a partition wall into an eleetrothermie furnaee zone where reducing agents are adqed to the bath. Heat is provided by eleetrieity. Lead metal may be reeovered from the slag, zine may be fumed off and eopper matte remains unehanged. The latter is transferred to a eonverter to produee blister eopper.
The XIVCET process for lead is similar, but is not eonsidered suitable for adaptation to predominantly zine eontaining materials unless low cost eleetrieity is available on aceount of the high energy requirements involved.
The Outokumpu company in Finland has investigated the smelting of lead and is also studying the fuming of zine from slag produced in lead smelting. Aeeording to Outokumpu (Australian patent 501,911), concentrates are flash smelted to produce molten lead metal and a liquid slag which are passed to an electrothermic furnaee where zine is fumed off by injecting fine coal entrained in niteogen. When applying this process to materials cont~ining high percentages of zinc, a lar~e amount of heat is produce~ as a result of the ---` 1279~L98 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 cireulating furnace load.
In the QS~ process (US patent 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 whieh oxygen is injeeted. A high lead oxide slag is produeed which is subsequently redueed to metallie lead in a seeond part of the furnace by injeeting eoal. AS far as is known, this process has not been developed for the fuming of zine.
The ISASMELT process (Australian patent applieation 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 zine in the lead concentrates and remaining in slag, may be recovered by the addition of a slag fuming stage to the proeess.
A proeess whieh is aimed at predominantly zine eontaining raw materials has been proposed by Davey and Yazawa~-- The proeess eoncept involves the following reaction at elevated temperatures -Zn~ + 2 = Zn + SO
As this reaction is endothermic, addition of carbonaeeous fuel is re~uired resulting in a mix of earbon oxides and sulphur dioxide in the off-gases. This makes it diffieult to recover the heat value in any unburnt carbon monoxide and may lead to ther~al inefficiency. There is also a potential problem of recovering zinc vapour from the eomplex gas stream ~'~79198 .
~I
DE~C~IPTION OF T~E INVENTION
The invention involves the oxidation of zinc sulphide concentrates in a molten slag bath w-ith 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 -2ZnS + 32 = 2ZnO + SO
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 conditioos to prevent the reversion of zinc to zinc oxide. Such ~post-combustion~ can be achieved by injecting oxygen into the top layers of the bat~l 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~ 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 1350C 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.
In one embodiment the present invention is directed to a zinc smelting process which comprises ta) establishing at least one oxidation zone and at least on 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;
(b) maintaining a molten matte phase rich in cuprous sulphide in the or each oxidation zone, feeding raw material comprising zinc sulphides and iron sulphides, together with :, ' 12~9~9~3 - 5a -fluxes, to the or each oxidation zone, and feeding o~ygen or oxygen-rich gas into the molten bath in the or each oxidation zone to oxidise metal compounds therein to produce a slag bearing zinc oxide and iron oxide and to produce an o~f-gas containing fume and SO2, said fume containing zinc and other volatiles;
(b.1) removing from said oxidation zone the off-gas produced in step(b), removing fume from said off gas, separating volatiles other than zinc from the removed fume to produce a zinc-containing fume resi.due, and recycling zinc-containing fume residue to the oxidation zone;
(c) transferring zinc oxide-bearing slag from at least one oxidation zone to at least one reduction zone;
(d) introducing coal into the or each reduction zone to reduce zinc oxide present in said zinc oxide-bearing slag to produce reduced zinc, discard slag, and reducing gases containing carbon monoxide and hydrogen, said discard slag containing iron oxide;
(e) 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 said carbon monoxide and hydrogen, thereby producing heat which is transferred back to the slag bath in the reduction zone, while leaving said reducing gases sufficiently reduced to ensure that reduced zinc vapor is not re-oxidised;
(f) recovering reduced zinc from at least one reduction zone; and (g) removing discard slag from at least one reduction zone, dividing the removed discard slag into two streams, passing one of the streams to waste, passing the other stream to an iron reduction reactor to reduce iron oxide present in said discard slag to produce iron metal, off-gases and a low-.. iron slag, recycling iron metal to the zinc reduction zone, and recycling low-iron slag from the reduction reactor to the oxidation zone.
~279~9~3 - 5b -In drawings which illustrate preferred embodiments of the present invention Figure 1 shows a flowsheet for the process with stages for oxidation, reduction plus post co~bustion, condensation and impurity bleed from recycle fume.
Figure 2 shows a conceptual diagram of one type of reactor for carrying out the process of the present invention;
Figure 3 shsws a flowsheet for a variant of the process which includes an iron reduction stage and a low iron slag recycle; and Figure 4 shows a flowsheet for the process described in Example 4.
Essentially the slag comprises FeO, Fe203, CaO, sio2 A12O3, PbO, ZnO components; a typical composition would be 20% FeO, 15% CaO, 20% sio2, 5% A12O3, 10% PbO, 30% ZnO.
--- 12~79~98 The oxidation reaction is generally driven at a temperature of about 1350C. 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 2ZnS + 32 = 2ZnO + 2SO2 (1) 12FeO + 32 = 6Fe23 (2) In the bulk of the slag ZnS + 3Fe2O3 = ZnO + 6FeO + SO2 (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 1350C
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 ~L~79~i9~3 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 from the oxidation stage is separated from the SO2 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 - ¦
ZnO + C = Zn + CO (4) 2CO + 2 = 2CO2 (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 H2 to CO2 and H2O, but to still leave the gases sufficiently reduced to ensure that the zinc is not .~ ~
~2~98 oxidised to zinc oxide. Maintenance of a residual CO/CO2 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.
.. ~ . . .
-- ~279198 - 8a -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 pcints.
The slag (tail slag) produced in the reduction zone or reactor has been found to contain only a few percent o~
zinc and can be discarded.
~ r 12791g8 ' Reference is now made to Figure 2. The reactor 1 is a water-cooled refractory-lined vessel. Partition 2 in the gas space above the slag bath separates the SO2 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. Addit:ional partitions in the slag bath could be provided to give multiple reduction zone.
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. ~inc 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.
The variant of the process described in the flowsheet of Figure 3 includes an iron reduction stage and a low iron slag recycle. The oxidation and zinc reduction stages are as described for Figure 1, with the following additions: -2~ 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 slagto 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.
---` 1279198 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 (FeO)slag + C = CO + (Fe)copper (6) In the zinc reduction reactor (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 ceduction reactor slag bath.
12~9~g8 .
One aspect Inentioned 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 SO2 would be dispersed through the slag phase as discrete entities as they react with the Fe2O3 in the slag (refer Reaction ~3)).
Secondly, the equilibrium, between copper matte, copper metal and SO2 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 -inade~uate 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 2ZnS + 32 = 2(ZnO)slag + 2SO2 (8) - 1279~8 With excessive oxygen (Cu2S)matte + 2 = 2Cu + SO2 (9) 2(ZnS)matte + 32 = 2(ZnO)slag + 2SO2 (10) With insufficient oxygen 2Cu + 2ZnS + 2 = (Cu2S)matte + 2(ZnO)slag (11) 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 occurring simultaneously.
A possible disadvantage of using matt 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 lS 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 manly 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 l~ad to zinc ratios, a separate lead bullion phase will be produced which can be tapped from the reactor for further processing to refined lead.
``` ~2'79~9B
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 1350C. 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 -~L2~9198 ~n . .
_I o ~ ~o o ~ ~ ~ oo ¢
o C~ `. . .
o o~o X
.,l . . .
U~o ~CO
_I
~P
G~00 0 ~r ~~. o o o o o . .
~o o ~ ~o C 1~ .
~ .
d~
~ .
V~ oo ~ o~ , U
_~ o ~ _~
UV~
.., ~ ,o ~ _~
S ~ u~ ~ L .
.,, U~
`-` 1279198 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 1300C. Oxygen plus coal and nitrogen were injected through a lance below the surface ofthe slag.
Oxygen feed rate was 1.4 1/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. ~nder these conditions, the slag bath equilibrated at 3.4% Zn after 30 min.
~279~g8 :
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/CO2 ratio of 1:4.
Example 3 The following table gives examples of a range or high zinc slag compositions that were found to be fluid at a temperature of 1350C, 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.
. ~.. ; .
o~ ~L27~
oP
o . . . .
~D O r~
d~
Oa~ o ~,~
... . . . .
+
a +
o ~ o o o aJ
a)I_1-- c,~ ~ o ~ ~ , d~
-~ , o Q~) o ~1~ o 0~1 ~ ~ ~1 C ~ ~ ~D
C~ . .
a~ o ~ o o o Eu~u ) u~o:) In Z ~ ~ ~ ~ n Lt~
.
. ' . ' .
18 - gL279198 EXAMPLÆ 4 This example is illustrated in the attached flow sheet~ ~ig. 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 1250C.
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 1250C.
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 1250C, 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 CO2 and reverberate the heat released back into the slag bath. The CO/CO2 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.
Claims (4)
1. A zinc smelting process which comprises (a) establishing at least one oxidation zone and at least on 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;
(b) maintaining a molten matte phase rich in cuprous sulphide in the or each oxidation zone, feeding raw material comprising zinc sulphides and iron sulphides, together with fluxes, to the or each oxidation zone, and feeding oxygen or oxygen-rich gas into the molten bath in the or each oxidation zone to oxidise metal compounds therein to produce a slag bearing zinc oxide and iron oxide and to produce an off-gas containing fume and SO2, said fume containing zinc and other volatiles;
(b.1) removing from said oxidation zone the off-gas produced in step(b), removing fume from said off gas, separating volatiles other than zinc from the removed fume to produce a zinc-containing fume residue, and recycling zinc-containing fume residue to the oxidation zone;
(c) transferring zinc oxide-bearing slag from at least one oxidation zone to at least one reduction zone;
(d) introducing coal into the or each reduction zone to reduce zinc oxide present in said zinc oxide-bearing slag to produce reduced zinc, discard slag, and reducing gases containing carbon monoxide and hydrogen, said discard slag containing iron oxide;
(e) 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 said carbon monoxide and hydrogen, thereby producing heat which is transferred back to the slag bath in the reduction zone, while leaving said reducing gases sufficiently reduced to ensure that reduced zinc vapor is not re-oxidised;
(f) recovering reduced zinc from at least one reduction zone; and (g) removing discard slag from at least one reduction zone, dividing the removed discard slag into two streams, passing one of the streams to waste, passing the other stream to an iron reduction reactor to reduce iron oxide present in said discard slag to produce iron metal, off-gases and a low-iron slag, recycling iron metal to the zinc reduction zone, and recycling low-iron slag from the reduction reactor to the oxidation zone.
(b) maintaining a molten matte phase rich in cuprous sulphide in the or each oxidation zone, feeding raw material comprising zinc sulphides and iron sulphides, together with fluxes, to the or each oxidation zone, and feeding oxygen or oxygen-rich gas into the molten bath in the or each oxidation zone to oxidise metal compounds therein to produce a slag bearing zinc oxide and iron oxide and to produce an off-gas containing fume and SO2, said fume containing zinc and other volatiles;
(b.1) removing from said oxidation zone the off-gas produced in step(b), removing fume from said off gas, separating volatiles other than zinc from the removed fume to produce a zinc-containing fume residue, and recycling zinc-containing fume residue to the oxidation zone;
(c) transferring zinc oxide-bearing slag from at least one oxidation zone to at least one reduction zone;
(d) introducing coal into the or each reduction zone to reduce zinc oxide present in said zinc oxide-bearing slag to produce reduced zinc, discard slag, and reducing gases containing carbon monoxide and hydrogen, said discard slag containing iron oxide;
(e) 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 said carbon monoxide and hydrogen, thereby producing heat which is transferred back to the slag bath in the reduction zone, while leaving said reducing gases sufficiently reduced to ensure that reduced zinc vapor is not re-oxidised;
(f) recovering reduced zinc from at least one reduction zone; and (g) removing discard slag from at least one reduction zone, dividing the removed discard slag into two streams, passing one of the streams to waste, passing the other stream to an iron reduction reactor to reduce iron oxide present in said discard slag to produce iron metal, off-gases and a low-iron slag, recycling iron metal to the zinc reduction zone, and recycling low-iron slag from the reduction reactor to the oxidation zone.
2. Process according to Claim 1 further comprising providing a molten copper phase in said iron reduction reactor to dissolve the metallic iron produced therein and to convey it in molten form to the zinc reduction zone.
3. Process according to Claim 1 in which off-gases from the iron reduction reactor are cleaned and recycled to the zinc reduction zone slag bath.
4. Process according to Claim 1 in which the slag baths of the oxidation and reduction zones are separated by a partition.
Applications Claiming Priority (2)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| AUPH000285 | 1985-04-03 | ||
| AUPH0002 | 1985-04-03 |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA1279198C true CA1279198C (en) | 1991-01-22 |
Family
ID=3771015
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA000505638A Expired - Fee Related CA1279198C (en) | 1985-04-03 | 1986-04-02 | Zinc smelting process using oxidation zone and reduction zone |
Country Status (8)
| Country | Link |
|---|---|
| CN (1) | CN1023904C (en) |
| CA (1) | CA1279198C (en) |
| DE (1) | DE3611159A1 (en) |
| FR (1) | FR2579996B1 (en) |
| GB (1) | GB2173820B (en) |
| PL (1) | PL152530B1 (en) |
| YU (1) | YU44913B (en) |
| ZA (1) | ZA862380B (en) |
Families Citing this family (15)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| GB8626085D0 (en) * | 1986-10-31 | 1986-12-03 | Rtz Technical Services Ltd | Smelting complex polymetallic sulphide materials |
| DE3701846A1 (en) * | 1987-01-23 | 1988-08-04 | Metallgesellschaft Ag | DIRECT MELTING PROCESS FOR SULFIDIC ORES |
| CN102011014B (en) * | 2010-11-21 | 2012-11-14 | 中国恩菲工程技术有限公司 | Continuous lead-smelting device and continuous lead-smelting process |
| CN102268557A (en) * | 2011-08-02 | 2011-12-07 | 云南锡业集团（控股）有限责任公司 | One-step oxidizing smelting, retailoring and slag fuming volatilization method for lead sulphide concentrate |
| CN103100719B (en) * | 2012-11-12 | 2015-12-09 | 柳州百韧特先进材料有限公司 | A kind of technological process of producing zinc powder |
| CN102965510B (en) * | 2012-11-27 | 2014-12-10 | 中南大学 | Reduction sulfur-fixing bath smelting method and device of low-sulfur lead-containing secondary material and iron-rich heavy metal solid waste |
| CN102965509A (en) * | 2012-11-27 | 2013-03-13 | 中南大学 | Method and device for treating daub and iron-rich heavy metal solid waste of waste lead-acid storage battery |
| CN106222444A (en) * | 2015-08-25 | 2016-12-14 | 赖成章 | Zinc sulfide ore method is smelted with high-tech |
| CN105890362B (en) * | 2016-04-28 | 2018-11-09 | 天津闪速炼铁技术有限公司 | A kind of metallurgical reaction device that oxidation coexists with reducing atmosphere |
| CN106086463B (en) * | 2016-08-02 | 2019-04-09 | 中国恩菲工程技术有限公司 | The smelting equipment of Pb-Zn deposits |
| CN106498182B (en) * | 2016-10-31 | 2018-10-02 | 云南驰宏锌锗股份有限公司 | The compound metallurgical furnace of fluorine chlorine impurity in a kind of efficient removal zinc oxide fumes |
| CN107475522A (en) * | 2017-08-30 | 2017-12-15 | 阳谷祥光铜业有限公司 | A kind of method for reclaiming copper, selenium and lead in the relieving haperacidity process from copper blast furnace off-gas in caused high-copper selenium waste material |
| CN108277361B (en) * | 2018-04-08 | 2020-11-17 | 中国恩菲工程技术有限公司 | Crude copper pyrogenic process continuous refining furnace |
| CN108759461B (en) * | 2018-06-21 | 2023-12-15 | 赫章领轩环保科技有限公司 | Equipment and method for graded production of direct zinc oxide by tunnel kiln |
| CN114182097B (en) * | 2021-12-08 | 2024-03-12 | 西安建筑科技大学 | Method for cooperatively recycling copper-zinc-containing oxide and zinc sulfide |
Family Cites Families (10)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| AU421261B2 (en) * | 1967-01-31 | 1972-02-11 | Ministerstvo Tsvetnoimetallurgii | A method of processing ores and concentrates containing rare metals and a unit for effecting said method |
| US3892559A (en) * | 1969-09-18 | 1975-07-01 | Bechtel Int Corp | Submerged smelting |
| DE2038227C3 (en) * | 1970-07-31 | 1973-06-20 | Vni Gornometallurgitscheskij I | Process for the preparation of ores and concentrates |
| US3941587A (en) * | 1973-05-03 | 1976-03-02 | Q-S Oxygen Processes, Inc. | Metallurgical process using oxygen |
| FI56397C (en) * | 1974-07-05 | 1980-01-10 | Outokumpu Oy | OIL ANALYZING FOR SUSPENSIONSSMAELTNING AV FINFOERDELADE SULFID- OCH / ELLER OXIDMALMER ELLER -KONCENTRAT |
| DE2807964A1 (en) * | 1978-02-24 | 1979-08-30 | Metallgesellschaft Ag | METHOD FOR THE CONTINUOUS CONVERSION OF NON-METAL SULFID CONCENTRATES |
| US4252560A (en) * | 1978-11-21 | 1981-02-24 | Vanjukov Andrei V | Pyrometallurgical method for processing heavy nonferrous metal raw materials |
| FR2444721A1 (en) * | 1978-12-22 | 1980-07-18 | Mo I Stali I Splavov | Pyrometallurgical treatment of non-ferrous heavy metal ores - in furnace where gas contg. specific amt. of oxygen is blow into slag contg. sulphide(s) and oxide(s) |
| DE3101369A1 (en) * | 1981-01-17 | 1982-08-26 | Klöckner-Humboldt-Deutz AG, 5000 Köln | METHOD AND DEVICE FOR PREVENTING FUSIBLE SUBSTANCES LIKE ORE CONCENTRATE |
| DE3420372A1 (en) * | 1984-06-01 | 1985-12-05 | Klöckner-Humboldt-Deutz AG, 5000 Köln | Process and apparatus for treating molten nonferrous metals and/or their slags by reaction gas top blowing |
-
1986
- 1986-03-27 GB GB8607670A patent/GB2173820B/en not_active Expired
- 1986-04-01 ZA ZA862380A patent/ZA862380B/en unknown
- 1986-04-02 CA CA000505638A patent/CA1279198C/en not_active Expired - Fee Related
- 1986-04-03 DE DE19863611159 patent/DE3611159A1/en active Granted
- 1986-04-03 FR FR868604777A patent/FR2579996B1/fr not_active Expired
- 1986-04-03 CN CN86102264A patent/CN1023904C/en not_active Expired - Fee Related
- 1986-04-03 PL PL1986258744A patent/PL152530B1/en unknown
- 1986-04-03 YU YU529/86A patent/YU44913B/en unknown
Also Published As
| Publication number | Publication date |
|---|---|
| CN1023904C (en) | 1994-03-02 |
| GB8607670D0 (en) | 1986-04-30 |
| YU52986A (en) | 1988-10-31 |
| GB2173820B (en) | 1989-06-28 |
| GB2173820A (en) | 1986-10-22 |
| YU44913B (en) | 1991-04-30 |
| CN86102264A (en) | 1986-10-29 |
| ZA862380B (en) | 1986-12-30 |
| FR2579996B1 (en) | 1989-04-28 |
| DE3611159C2 (en) | 1989-12-28 |
| DE3611159A1 (en) | 1986-10-09 |
| PL152530B1 (en) | 1991-01-31 |
| FR2579996A1 (en) | 1986-10-10 |
Similar Documents
| Publication | Publication Date | Title |
|---|---|---|
| US8101153B2 (en) | Method for the valorisation of zinc-and sulphate-rich residue | |
| US4741770A (en) | Zinc smelting process using oxidation zone and reduction zone | |
| CA1279198C (en) | Zinc smelting process using oxidation zone and reduction zone | |
| CA1219133A (en) | Continuous direct process of lead smelting | |
| CN100392123C (en) | Recovery of non-ferrous metals from zinc residues | |
| US6699302B1 (en) | Treatment of metal sulphide concentrates by roasting and electrically stabilized open-arc furnace smelt reduction | |
| US4588436A (en) | Method of recovering metals from liquid slag | |
| Sohn et al. | Lead and zinc production | |
| US3663207A (en) | Direct process for smelting of lead sulphide concentrates to lead | |
| EP0557312B1 (en) | Direct sulphidization fuming of zinc | |
| US4487628A (en) | Selective reduction of heavy metals | |
| US4135912A (en) | Electric smelting of lead sulphate residues | |
| US3847595A (en) | Lead smelting process | |
| CA2098521C (en) | Method for producing high-grade nickel matte and metallized sulfide matte | |
| GB2196649A (en) | Smelting complex sulphidic materials containing lead, zinc and optionally copper | |
| US5178667A (en) | Dry process for refining zinc sulfide concentrates | |
| US4515631A (en) | Method for producing blister copper | |
| US2816022A (en) | Smelting of lead-containing ores | |
| EP0427699B1 (en) | Method and apparatus for treating zinc concentrates | |
| US5443614A (en) | Direct smelting or zinc concentrates and residues | |
| US3773494A (en) | Smelting of copper sulphide concentrates with ferrous sulphate | |
| US4274868A (en) | Recovery of tin from ores or other materials | |
| WO1997020958A1 (en) | Recovery of cobalt from slag | |
| CA1208444A (en) | High intensity lead smelting process | |
| AU646510C (en) | Direct sulphidization fuming of zinc |
Legal Events
| Date | Code | Title | Description |
|---|---|---|---|
| MKLA | Lapsed |