AU612120B2 - Recovery of lead, zinc and other metals from ores concentrates or residues - Google Patents

Description

i i I I I i rrrrrr~llrsrrm~rrrrrrcllll ii i ;i AU-AI-79186/87 WORLD INTELLECTUAL PROPERTY ORGANIZATION International Bureau

PCT

0 INTERNATIONAL APPLICATION PUBLIS D ND THEPATENT COOPERATION TREATY (PCT) (51) International Patent Classification 4 11) ernati al ic o Number: WO 88/ 01654 C22B 13/02, 19/04, 19/06 C22B 19/08, 19/10, 19/12 A l (43) International ica n Date: 10 March 1988 (10.03.88) C22B 19/14 (21) International Application Number: PCT/AU87/00287 (74) Agents: CORBETT, Terence, G. et al.; Davies Collison, I Little Collins Street, Melbourne, VIC 3000 (22) International Filing Date: 25 August 1987 (25.08.87) (AU).

(31) Priority Application Numbers: PH 7693 (81) Designated States: AT (European patent), AU, BE (Eu- PH 7694 ropean patent), CH (European patent), DE (European patent), FR (European patent), GB (European (32) Priority Dates: 27 August 1986 (27.08.86) patent), IT (European patent), JP, LU (European pa- 27 August 1986 (27.08.86) tent), NL (European patent), SE (European patent),

US.

(33) Priority Country: AU Published Applicant (for all designated States except US): COM- With international search report.

MONWEALTH SCIENTIFIC AND INDUSTRIAL RESEARCH ORGANISATION [AU/AU]; Limestone Avenue, Campbell, ACT 2601 (AU).

(72) Inventors; and .O.J.P 2 A APR 1988 Inventors/Applicants (for US only) CANTERFORD, John, Harmsworth [AU/AU]; 31 Glenwood Drive,

A

Croydon, VIC 3136 DENHOLM, William, AUSTRALIAN Thomas [AU/AU]; 13 Kintore Street, Camberwell, VIC 3124 RAJAKUMAR, Viruthiamparam- 2 4 MAR1988 bath [IN/AU]; 15 Lainie Court, Wantirna South, VIC 3152 PATENT OFFICE (54) Title: PROCESS FOR THE TREATMENT OF LEAD-ZINC ORES, CONCENTRATES OR RESIDUES (57) Abstract Lead-zinc complex sulphide ores or concentrates are treated by a process which comprises; a) adding the ore or concentrate to a bath of molten matte which is overlaid by a slag layer; b) agitating and heating the bath by submerged injection of a fuel and air (or other oxygen-bearing gas mixture), c) adding a reductant to the bath; thereby to convert at least a substantial proportion of the lead and zinc content of the ore or concentrate to mixed fume of lead and zinc and form a high-iron matte in the bath; and d) collecting the lead-zinc fume. The process can also be applied to the treatment of leach residues containing at least 20 by weight of sulphur and not more than 35 by weight of lead, by modifying the bath in step a) so that it consists principally of molten slag, 1 L. I WO 88/01654 PCT/AU87/00287 PROCESS FOR THE TREATMENT OF LEAD-ZINC ORES, CONCENTRATES OR RESIDUES" This invention relates to a process for the treatment of lead-zinc complex sulphide ores and concentrates. In particular, the invention is concerned with a process which involves both pyrometallurgical and hydrometallurgical treatments.

It is now well established that lead sulphide concentrates can be smelted using a variety of processes to recover lead. Various processes, other than the conventional sintering-blast furnace process, have been technologically and commercially successful. Processes for the recovery of zinc from zinc sulphide concentrates have also been commercialised.

We recognize, however, that future resources of these metals will largely come from ore bodies in which the various mineral phases occur in a complex series of intergrowths. For these so-called complex sulphide ores, beneficiation techniques such as flotation are not always I c WO 88/01654 PCT/AU87/00287, 2 satisfactory because a clean separation of the individual metal values is rarely possible. High metal recoveries at high grades is difficult to achieve because of problems connected with selectivity during flotation. It is possible, however, to produce bulk concentrates containing say 20-50% zinc and 5-15% lead at commerciably acceptable recoveries using flotation. It is also clear that it would be of great commercial benefit if flotation or other preliminary costly concentration steps can be avoided altogether by developing an appropriate process which would be able to recover the metal values for the ore.

One aspect of the present invention relates to the recovery of zinc, lead and other metal values such as silver, gold and copper from complex sulphide ores or concentrates. In contrast to previously described processes which are largely concerned with the separate recovery of zinc or lea<. from their concentrates, we have discovered that it is possible to simultaneously recover zinc and lead as a fume by using a bath smelting process in which the contents are vigorously agitated by the injection of gas. In carrying out the smelting reactions a high-iron matte is produced and the operating conditions are selected to obtain an efficient separation of the lead and zinc from the iron in the feed.

According to one aspect of the present invention, there is provided a process for the treatment of a lead-zinc complex sulphide ore or concentrate, which comprises: adding the ore or concentrate to a bath of molten matte, which is overlaid by a slag layer; WO 88/01654 PCT/AU87/00287 3 agitating and heating the bath by submerged injection of a fuel and air (or other oxygen-bearing gas mixture); adding a reductant to the bath;thereby to convert at least a substantial proportion of the lead and zinc content of the ore or concentrate to mixed fume of lead and zinc and form a high-iron matte in the bath; and collecting the lead-zinc fume.

It is already established that it is technologically and commercially feasible to react zinc sulphide in the form of a zinc sulphide concentrate with aqueous sulphuric acid in the presence of an oxygen-bearing gas to convert the insoluble zinc sulphide into soluble zinc sulphate.

After removal of the leach residue, the resulting leachate (principally a zinc sulphate solution) is further processed to recover the zinc, for example by electrowinning to obtain pure zinc metal, or precipitation of basic zinc sulphate, or spray drying to produce zinc oxide.

The leaching conditions are generally but not always such that the bulk of the sulphide sulphur associated with the zinc is converted to elemental sulphur, the remainder being oxidized to sulphate. In order to achieve an acceptable rate of dissolution of the zinc sulphide the leaching reaction is carried out in a pressurized reaction vessel so that the oxygen partial pressure is substantially greater than atmospheric pressure. Oxygen partial pressures in the range 300-2000 kPa are normally WO 88/01654 PCT/A U87/00287' 4 used. Although air can be used as the source of oxygen, it is preferable to use compressed oxygen itself. An elevated temperature is also used to increase reaction kinetics. Temperatures in the range 90-230°C can be used. The conversion of zinc sulphide to zinc sulphate by reaction with aqueous sulphuric acid in the presence of an oxygen-bearing gas is commonly termed the "oxygen pressure leach process".

In some cases it may be preferable to carry out the oxygen pressure leaching process under more aggressive conditions than those noted above so that all the sulphide sulphur is oxidized to sulphate. This normally involves leaching at temperatures in the range 170-230°C. Under these conditions much of the pyrite in the concentrate be leached.

The feeds generally used for the oxygen pressure leach process are derived from orebodies that contain zinc sulphide that can be readily recovered in a sub:tantially 20 pure form by physical beneficiation techniques such as flotation. The zinc sulphide concentrate will typically contain 45-55% zinc with less than a total of 10% of other non-ferrous metals such as lead and copper.

Attempts have been made in the past to apply the oxygen pressure leach process to the treatment of bulk concentrates of complex lad-zinc ores of the type already discussed above. In most cases the extent of zinc sulphide dissolution was generally acceptable, being 30 greater than say 90%. The lead sulphide component of the bulk concentrate was converted to a mixture of lead sulphate (anglesite) and a basic lead-iron sulphate (plumbojarosite). Silver, a common and economically W88/01654 PCT/A U87/00287 valuable assessory metal in many bulk concentrates is coprecipitated with both the anglesite and plumbojarosite. The lead-containing phases constitute a substantial portion of the insoluble leach residue. This residue also contains the elemental sulphur formed during the dissolution reaction, which for lead and zinc sulphides can be represented by the following general reaction.

2MS 2H2SO 4 0 2MSO 4 2H O 2S (M ZnPb) The residue will also contain any gangue minerals such as silicates in the bulk concentrate and any unreacted sulphide minerals. The pyrite in the bulk concentrate is generally unreactive under the conditions used and so reports in the leach residue.

Because the lead and silver components of the bulk concentrate represent a significant proportion of the total realizeable value of the concentrate, it is an economic necessity to recover the lead and silver values from the insoluble leach residue that remains after the zinc sulphate solution produced by the leaching reactions is removed by solid-liquid separation techniques. A particular difficulty in treating the leach residue produced by the oxygen pressure leaching route as applied to bulk concentrates whatever the leaching conditions is the high iron content of the residue. Recovery of the lead and silver from the leach residue by necessity involves the development of a route that has a high degree of lead versus iron selectivity. The lead must be recovered in a form that is suitable for further processing while the iron must be in a form that can be -4 I i I i 11 WO 88/01654 PCT/A U87/00287 6 readily discarded.

Various hydrometallurgical and pyrometallurgical methods have been proposed for recovering the lead and silver from the anglesite/plumbojarosite leach residue, Methods proposed include chloride and ammoniacal ammonium sulphate leaching, electric smelting and sulphidization.

However, none of tihe proposed methods have proved to be economically or technically attractive. Thus these methods do not allow ready recovery of the lead and/or silver, are capital and/or operating cost intensive, and are not compatible with the oxygen pressure leach process used to initially treat the bulk concentrate used to produce the zinc sulphate solution and the anglesite/plumbojarosite residue.

Another disadvantage of most of the proposed methods for treating the anglesite/plumbojarosite residue is that they do not allow the recovery of the small amount of gold that is often found in bulk concentrates. The gold in the bulk concentrate is usually associated with the pyrite in the concentrate. The gold is not solubilized during the oxygen pressue leaching process so that it also reports with the anglesite/plumbojarosite residue. For some bulk concentrates, it is economically advantageous to recover this gold from the oxygen pressure leach residue.

The lack of a suitable method of treating the leadand silver-containing residue is one of the reasons why processing of bulk concentrates derived from complex sulphide ore deposits by the oxygen pressure leach route has yet to reach commercial reality.

We have now discovered that it is technically 1 WO 38/01654 PCT/A U87/00287 7 feasible to process lead- and silver-contain.ing leach residues by the bath smelting technique already described above. The bath smelting technique can be applied to leach residues that are produced under a wide variety of leaching conditions. In particular, the bath smelting technique can be applied to residue that contain varrying amounts of elemental sulphur and sulphate. It is not essential to remove any elemental sulphur from the residue prior to bath smelting, neither is it essential that the residue contain a specified amount of elemental sulphur, This means that the conditions used to leach the bulk lead-zinc concentrate are essentially those that give maximum zinc dissolution under minimum capital and operating cost conditions.

According to another aspect of the present invention, there is provided a process for the treatment of a leach residue obtained from a lead-zinc complex sulphide ore or concentrate said residue containing at least 20% by weight of sulphur and not more than 35% of lead, said process comprising: adding the residue to bath of molten slag, in which matte may be present; agitating and heating the bath by submerged injection of a fuel and air (or other oxygen-bearing gas mixture); adding a reductant to the bath; thereby to convert at least a substantial proportion of the lead and zinc content of the residue to mixed fume of lead and zinc and form a high-iron matte in the bath; and w_ WO 88/01654 PCT/A U87/00287, 8 collecting the lead-zinc fume.

The leach residue is preferably obtained by: pressure leaching a lead-zinc complex sulphide ore or concentrate with sulphuric acid in the presence of an oxygen-bearing gas to convert lead and zinc values in the ore to their sulphates; (ii) separating the mixture thus produced to provide a solution containing zinc sulphate and the leach residue; and (iii) treating the said solution to recover zinc values therefrom.

The lead-zinc fume from the smelting operation (in either aspect of this invention) can be collected by conventional means and treated to separate the zinc and lead. We have found that leaching with acid results in selective dissolution of the zinc which can then be recovered by electrowinning or other known methods. The lead-rich residue which remains after leaching can be treated to recover the lead, again by known methods.

The high-iron matte from the smelting operation can be subsequently treated, if desired, to recover the gold, silver and copper values in it.

Thus preferred embodiments of the process of the invention involve one or more of the following additional steps: I I I r -e i- i I i WO 88/0 t654: w TIA 87/00287, WO 88/~ 5 PCT/AU87/00287 9 leaching the fume with acid to selectively dissolve the zinc component and leave a lead-rich residue; treating the zinc-containing solution thus obtained to recover the zinc and other metal values therein; treating the lead-rich residue from to recover the lead and other metal values therein; treating the high-iron matte to recover desirable metal values therein.

A simplified flowsheet of the first embodiment of the process is shown in Fig. 1 and described below, In the bath smelting operation, which is essentially a reductive fuming operation, the sulphide ore o concentrate containing zinc and lead, and possibly other metal values such as gold, silver and copper, is added to a molten matte bath which is contained in a refractory lined vessel, The matte bath is agitated vigorously by means of one or more submerged lances through which fuel and air are introduced in the correct proportion. A reductant may be added either separately onto the bath or through the lance. The reductant may be the fuel or another material.

The use of the "SIROSMELT" lance, as described in U,S. Patent No.4,251,271, is particularly preferred in the smelting operation.

WO 88/01654 PCT/AUL'87/00287 The temperature of the operation depends on the composition of the feed ore or concentrate and is selected to ensure high recoveries of zinc and lead in the fume and to maintain the bath in a molten state. Typically this temperature is 1350-1400°C. The gangue can be conveniently separated into a slag phase with the addition of flux if required.

The quantity of fuel coal, oil or other hydrocarbon source, injected through the lance is contr)lled so that the combustion of the fuel provides sufficient heat to allow for the heat requirements of the smelting reactions, melting of the feed, sensible heat of the products and heat losses from the reactor. The ratio of air to fuel fed to the lance is also controlled such that the oxygen potential of the gas phase in equilibrium with the matte is at the optimum value. The selected combination of air, fuel and reductant feed rates kor a specified feed rate of concentrate of ore ensures the simultaneous, high recoveries of lead and zinc in the fume, By way of example, operation at 1350oC and an oxygen potential of approximately 10 9 atmosphere for a bulk concentvate (20-30% zinc and 10-15% lead) or for an ore zinc, 7-8% lead) gives recoveries of 9$-99% lead and 90-95% zinc in the fume. Operation at 14000C will give 92-97% zinc recovery in the fume while the recovery of lead is greater than 99%, For operation at 1350 0

C

based on coal as the fuel and reductant, typicai usage based on an average concentrate composition of 13% lead and 31% zinc is 0.30-0.35 kg carbon/kg of wet feed. For a coal with 20% ash and 65% fixed carbon, the slag fall is 0.2-0.25 kg/kg of feed and the fume rate is apoxaey 0.4kg/kg of feed. Approximately 3.0m 3 of air will t L i" i I I i ~1 'IFl~a"~~ SWO 88/01654 PCT/A C87/00287 11 required and 3.5m 3 of gas containing 5% SO 2 6.6%CO and 13.6%C02 will be generated per kg of feed under these conditions.

A high-iron matte is produced from the reductive fuming operation. The distributions of silver and copper between matte, slag and gas are dependent on the feed composition and the operating conditions and their recoveries in the matte increase with the amount of matte formed which in turn increases with the reducing potential of the gas.

This aspect of the invention is further illustrated by the following non-limiting examples.

EXAMPLE 1 In this example 195 g of a complex sulphide concentrate containing 32.2% Zn, 8.85% Pb, 19.5% Fe, 0.79% Qu, 39,9% S, 240 g/t Ag, 1.78 g/t Au was pelletised and fed continuously over a period of 65 minutes into a matte bath Fe, 8.9% Zn, 0.24% Pb, 1.45% Cu, 21.1% S) maintained at 1350°C in an aluminosilicate crucible.

70.8 g of char and 617.5 of air were injected into the bath through a lance during this period. The lead and zit\c from the feed concentrate were fumed. Slag formation was negligibly small and the slag could not be separated from the final matte. The distributions of zinc, lead and silver between the phases was as follows: (iii) treating the said solution to recover zinc values therefrom.

4. A process as claimed in any one of Claims 1 to.3, /3 i- Ir WO 88/01654 PCT/A387/00287: Distribution, Zinc Lead Silver Fume 96.0 >99.5 10.6 Matte slag) trace 89.4 The recoveries of zinc and lead into the fume are high and 89.4% of the silver was recovered in the matte.

By increasing the carbon to air ratio injected into the bath for a constant feedrate of the concentrate, the amount of matte formed can be increased, thereby increasing the recovery of silver into the matte. Using a pelletised feed of this concentrate we have carried out experiments under a variety of operating conditions and the range of results obtained is summarised below: Distribution, Fume Zinc Lead Silver 93-96 98-100 9-35 2-5 0-2 60-89 Slag 2 trace More than 99% of the gold will report to the matte phase under these conditions. It will be noted that in some cases a recoverable slag was produced.

EXAMPLE 2 This example illustrates the results for a bulk concentrate containing 31% Zn, 13%Pb, 17.3% Fe, 36.6% S, 0.6% Cu, 0.5% Si02, 280 g/t Ag. The concentrate with moisture addition is fed into a matte bath held at LIr~p: I -I IA TY 1 WO 88/01654 PCT/AU87/00287 13 1350°C. Fuel plus reductant are injected into the bath through the lance at the rate of 0.328% kg per kg of the wet feed together with air at 3.025 cubic meters per kg of concentrate. Alternatively, the reductant could be added directly to the matte bath with the fuel and air being supplied through the lance. 0.26 kg of matte and 0.01.4 kg of slag are formed per kg of the wet feed. 95% of the zinc and >99% of the lead in the feed are fumed under these conditions.

Distribution, Fume Matte slag) Zinc 95.0 Lead >99.0 EXAMPLE 3 120 g of dry pellets of a concentrate containing 46.7% Zn, 2.95% Pb, 11.65% Fe, 0.23% Cu, 32.9% S, 110 g/t Ag were fed continuously over a period of one hour into a 2 matte bath (45.5% Fe, 15.2% Zn, 0.23% Pb, 22.4% S, 0.32% Cu) held in an aluminosilicate crucible at 1350 0 C. Char and air wre injected into the bath through a lance at the rate of 32.5 g of char and 267.5 2 of air (25 0 C) per 10o3 of concentr'ate. The lead/zinc fume was collected from the gas stream leaving the reactor. Slag formation was negligibly small. The distribution of zinc, lead, copper and silver between the matte (and minor amount of slag) and fume were Fume Mte slag) Zinc 93.9 6.1 Lead >99.0 trace Copper 0.6 99.4 Silver 5.0 95.0 -e i PCT/A U87/00287 WO 88/01654 94% of the zinc and all of the lead in the feed concentrate were fumed while 99.4% of the copper and of the silver were recovered in the matte.

EXAMPLE 4 In another test the concentrate in Example 3 above was mixed with char in the ratio 30.7 g char/100 g of concentrate and dry pellets were prepared. These pellets were continuously charged into the bath of molten matte and 313.7 2 of air 100/g of concentrate were injected into the bath through a lance during a period of minutes. The char and air rates were selected to provide the required heat for the smelting reactions and to maintain an oxygen potential of 10 9 atmospheres in the system. The temperature of the bath was 1350°C. Again, the slag fall was negligibly small and the distribution of the elements between the phases were: Zinc Lead Copper Silver Fume 97.1 >99.0 trace 4.0 Matte minor slag) trace >99.0 96.0 In this case 97% of the zinc and almost all of the lead was fumed with 99% of the copper and 96% of the silver being recovered in the matte phase.

EXAMPLE This example illustrates the application of the invention to the recovery of metal values from a complex sulphide ore without prior treatment to produce a bulk

F

i i LI WO 88/01654 P CT/A U87/'2n7 concentrate which was the feed in Example 1-4 above. In this direct or smelting route the feed and reductant are added to a high-iron matte bath which is agitated by the products of combustion of the fuel and air injected through the submerged lance.

156 g of complex sulphide ore particles assaying 7.15% Pb, 12.2% Zn, 0.39% Cu, 55 g/t Ag, 29.1% Fe, 40.2% S, 1.64% Silica, 3.64% CaO were added continuously over a period of 63 minutes into a high-iron matte bath (0.72% Pb, 0.41% Zn, 0.46% Cu, 7 g/t Ag, 26.7% S) maintained at 1350 0 C Char and air were injected through the lance at the rate of 45.3 g and 2.83 1 respectively per 100 g of residue. 85.7 g of matte assaying 0.28% Pb and 1.25% Zn were produced together with a minor amount of slag which not be completely separated. The slag assayed 0,21% Pb, 0.77% Zn.

The distribution of the elements between the phases is given below: Distribution, Fume Matte minor slag) Lead 97.9 2.1 Zinc 94.4 5.6 Silver 70.1 29.9 Almost 98% of the lead and 95% of the zinc in the feed is fumed under these conditions and reasonable recovery of silver is achieved.

The mixed fume of zinc and lead is recovered using conventional means, such as a bag house. The composition of the fume from the smelting operation depends on the I Id I d I did T- WO 88/01654 Pc/AU7/00287: 16 composition of the feed and the operating conditions but typically would contain 25-30% lead and 50-55% zinc assuming 100% collection efficiency.

The preferred method of processing the lead-zinc fume is by leaching in aqueous sulphuric acid. The selective dissolution of the zinc from the lead-zinc fume is typically carried out at 75-95 0 C in a conventional agitated leach tank. The reaction is generally exothermic and is carried out under controlled pH conditions such that the pH of the resultant leach liquor is typically in the range The pulp density of the slurry being leached depends on the composition of the lead-zinc fume and will usually be such that the resultant zinc sulphate solution will contain 120-170g/litre zinc. The solution of zinc sulphate formed by the leaching reaction is separated from the leach residue by conventional solid-liquid separation techniqdes such as thickening and filtration.

The leach residue remaining after removal of the zince sulphate solution consists of lead sulphate and other lead-containing solids and any unreacted sulphides that are carried over in the fume. The residue can be processed by conventional pyrometallurgical methods for the recovery of lead.

The zinc sulphate solution produced by leaching the lead-zinc fume can be used to produce various zinc-containing products. For example, the so'i Hon can be subjected to various purification procedure: emove impurities such as copper, nickel, cobalt, man chloride, cadmium, germanium, etc., prior to recovery by electrowinning. Alternatively, a hydrated zinc aujubaI.dcidLil.y YLt-edi Ludian adcu1spnerLict pressure. uxygen partial pressures in the range 300-2000 kPa are normally i WO 88/01654 PCT/AUL87/00287 17 sulphate can be crystallized from the zinc sulphate solution, basic zinc sulphate can be precipitated by controlled addition of a suitable alkaline reagent, or zinc oxide can be produced by spray drying/roasting techniques.

The following description refers to the second embodiment of the process of the invention. Figure 2 is a flowsheet that demonstrates the overall integrated route.

'A bulk concentrate is reground to 80% minus micron, preferably 80% minus 33 micron, if necessary.

In the preferred process option, the bulk concentrate is repulped in spent electrolyte from the zinc electrowinning circuit, which will typically contain 40-60 g/litre zinc and 120-160 g/litre sulphuric acid, while at the same time the free acidity is adjusted such that the total sulphate concentration in solution is equivalent to 5-10% more than that required to convert the lead and zinc in the solid feed to the respective sulphates when leaching is complete. The pulp density of the mixture to be added to the oxygen pressure leach autoclave will typically be in the range 40-70% solids and will typically yield a final leach solution 'that contains 120-170 g/litre soluble zinc and 10-50 g/litre free sulphuric acid.

Surface active reagents such as lignin, calcium lignosulphonate or quebracho are added, individually or in combination, to the leach pulp prior to leaching such that the concentration of surface active agents is typically 0.1-0.4 g/litre.

In a modification to the above feed preparation procedures, lead-zinc fume and bulk concentrate in the (plumbojarosite). Silver, a common and economically T>i- I I I I: /I-l WO 88/01654 PCT/AU87/00287; 18 appropriate proportions are mixed while being pulped in the spent electrolyte. As before, the amounts of lead-zinc fume, bulk concentrate, spent electrolyte, sulphuric acid and process water are adjusted such that composition of the resultant leach solution typically contains 120-170 g/litre zinc and 10-50 g/litre free sulphuric ac:id. Surface active agents may be added.

In another modificrtiion of the overall process, bulk concentrate is slurried in recycled process liquor from the alternative zinc recovery circuit and sufficient sulphuric acid added so that the resultant leach liquor will contain the desired amount of excess acid. The necessary amount of surface active agents is added to the leach slurry.

In another modification of the overall process, the lead-rich fume produced in the bath smelting of the leach residue is led to a separate leaching circuit where it is reacted on its own with a 5-10% excess of sulphuric acid to produce a zinc sulphate solution. Leaching is carried out at atmospheric pressure at a temperature generally in the range 75-95°C such that the pH of the resultant leach liquor is generally in the range 3.0-3.5. The leach residue remaining after removal of the zinc sulphate 2 5 solution consists predominantly of lead sulphate. This can be processed by conventional means. The zinc sulphate solution can be processed separately or, preferably, combined with that produced in the initial pressure leaching stage.

The leach pulp is added to a suitable well-agitated autoclave tnat is pressurized such that the oxygen partial pressure is maintained in the range 300-1500 kPa and recovereoa in a rorm tnat is suitable for further processing while the iron must be in a form that can be WO 88/01654 PCT/AU87/00287 19 heated to 90-230 0 C. The preferred oxygen partial pressure and temperatures are 500-1000kPa and 140-230 0

C,

respectively. These conditions are maintained for 60-150 minutes, typically 90-120 minutes. During this period more than 90% of the zinc sulphide in the bulk concentrate is converted into soluble zinc sulphate.

When leaching is complete the reacted pulp is discharged from the autoclave and the leach residue separated from the zinc sulphate solution by known solid-liquid separation techniques. Alternatively, the reacted pulp may be treated in an appropriate manner such that it is separated into three components, comprising of the bulk of the elemental sulphur formed by the leaching reactions, the lead- and silver-rich leach residue, which also contains unreacted sulphide minerals and other gangue in the leach pulp, and the zinc sulphate solution. The elemental sulphur is remioved from reacted pulp by known methods such as flotation and/or crystallization/filtration and/or dissolution in a suitable organic solvent.

The zinc sulphate solution is subjected to a range of known purification procedures to remove soluble impurities such as iron, arsenic, copper, cobalt, nickel, fluoride, antimony, manganese, etc, The purified zinc sulphate solution is then subjected to known electrowinning procedures to produce metallic zinc. Thq spent electrolyte is returned to the bulk concentrate leaching circuit where it is used to repulp fresh bulk 30 concentrata. Alternatively, the purified zinc sulphate solution is treated in such a manner that a solid zinc-containing product is obtained. For example, the zinc sulphate solution may be passed through a spray We have now discovered that it is technically '9 ~11 3 1 1 7 1 I I WO 88/01654 PCT/AU87/00287.

drier/roaster whereby solid zinc oxide is formed, or the zinc sulphate solution subjected to an evaporation process so that a hydrated zinc silphate crystallizes out of the mother liquor, or a suitable alkaline reagent, for example, lime, is added to the zinc sulphate solution to precipitate a basic zinc sulphate.

If elemental sulphur is recovered from the leach residue, then this can be subjected to a range of known purification procedures prior to sale.

The leach residue, consisting predominantly of anglesite, plumbojarosite, unreacted sulphides, other gangue minerals, as well as elemental sulphur if this has not been removed, is subjected to a bath smelting process, as described above. The operating conditions are dependent on the composition of the leach residue, but in genera-1, with high-sulphur residues (20% S or more), reductive fuming at 1250-1300°C gives a good separation between the lead and zinc. With the correct choice of the air/fuel reductant ratio, 85-98% of the lead in the leach residue feed is removed in the fume while the iron is collected in the matte phase. The lead content of the fume is high, typically 60-75%. The fume is recovered by conventional means such as a baghouse. The recovery of 2 5 silver in the matte increases with the weight of matte produced, which in turn increases with the reducing potential of the system.

This second aspect of the invention is further 30 illustrated by the following non-limiting examples.

c 1 ,1 WO 88/01654 PCT/A U87/00287 21 EXAMPLE 6 This example illustrates the recovery of lead and silver from a leach residue assaying 12.8% Pb, 2.3% Zn, 0.41% Cu, 30.2% Fe, 33% S, 340 g/t Ag, 6.0% silica. Dry pellets prepared from the feed were continuously fed into 500 g of an end point lead slag bath containing 1.0% Pb, Zn, 33.8% FeO, 29.2% SiO2, 16.8% CaO, 9,1% A1203, 1.6% MgO, 9 g/t Ag, 0.21% Cu. The bath was contained in an aluminosilicate crucible. Char and air were injected into the bath at the rate of 23 g and 172 Z per 100 g of residue. The char and air rates were selected to provide sufficient heat to carry out the smelting reactions and to maintain the required oxygen potential of the bath and the atmosphere above it, 184 g of high-iron matte assaying 2.15% Pb and 2.7 Zn were produced. The distribution of lead and silver, which are the major metl. values in the residue, between the phases resulting from smelting at 1250 0 C are given below.

Distribution, Fume Matte la Lead 87.1 10.3 2.6 Silver 8.0 92.0 tr 87% of the lead in the residue was recovered in the 2fume and 92% of the silver reported to the matte.

EXAMPLE 7 In this test a residue assaying 18.2 Pb, 4.1% Zn, 0.28% Cu, 25.2% Fe, 23.2 (14.7% sulphate), 365 g/t Ag, silica was added to the slag bath maintained at steps: c 1P 1 I I Ir I

I

WO 88/01654 PCT/AU87/00287 22 1250 0 C. 220 g of the residue pellets were fed into the bath over a period of 50 minutes. 57.5 g of char and 365 1 of air were injected through a lance to provide sufficient energy to carry out the smelting reactions and to maintain the required reducing potential in the system. ~t the end of the test 40 g of matte were produced assaying 1,2% Pb, 2/30 Zn and 1100 g/t Ag. The distributions of the elements between the phases was as follows: Distribution, Fumg Mate Slag Lead 98.8 1.2 tr Silver 43.7 54.8 Zinc 88,9 10.2 0.9 1 5 Almost 99% of the lead and 89% of the zinc were recovered in the fume while 98.5% of the silver in the residue was$ recovered into the matte and fume.

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Claims (4)

1. prcas for the treatment of a lead-zinc complex sulphide ore or concentrate, characterized in that it comprises the steps of: adding the ore or concentrate to a bath of molten matte which iLs overlaid by a slag layer; agitating and heating the bath by submerged injection of a fuel and air (or otter oxygen-bearing gas mixture); adding a redutctant to the baththereby to conve,,t at least a substantial proportion of the lead and zinc content of the ore or concentrate to mixed fume c, lead anid zinc and form a high-iron matte in the bath; and collecting the lead-zinc fume*

2, A prooess for the treatment of a leach residue obtaiined from a lead-zInc complex suiphide ol;% or concentrate said residue containing at least 20% by Weight of sulphur and not more than 35% of lead, Said process being characterized In that It comprises the steps of adding the residue to a bath of molten slag in contact with a small quantity of matte; agitating and heating the bath by submerged injection of a fuel, and air (or other oxygen-bearizig gas mixture); i WO 88/01654 PCT/AU87/00287; 24 adding a reductant to the bath; thereby to convert at least a substantial proportion of the lad and zinc content of the residue to mixed fume of lead and zinc and form a high-iron matte in the bath; and collecting the lead-zinc fume.

3. A process as claimed in Claim 2, characterized in that the leach residue is preferably obtained by: pressure leaching a lead-zinc complex sulphide ore or concentrate with sulphuric acid in the presence Sf an oxygen-bearing gas to convert lead and zinc values in the ore to their sulphates; (ii) separating the mixture thus produced to provide a solution containing zinc sulphate and the leach residue; and (iii) treating the said solution to recover zinc values therefrom,

4. A process as claimed in any one of Claims 1 to 3, characterized in that the fume from step is treated by one or more of the following additional steps: leaching the fume with acid to selectively dissolve the zinc component and leave a lead-rich residue; treating the zinc-containing solution thus obtained to recover the zinc and other metal values therein; 2 WO 88/01654 PCT/AL'87/00287 treating the lead-rich residue from to recover the lead and other metal values therein; treating the high-iron matte to recover desirable metal values therein.