Recovery of zinc from lead slag

Abstract

A method for recovering zinc from slag derived from lead smelting comprises subjecting the slag to a leaching step under conditions in which zinc is dissolved into solution and silica present in the slag dissolves and re-precipitates in a form that is readily separable from liquid, and recovering zinc from the solution. The slag may be subjected to leaching in at least two stages in which in a first leaching stage only part of the zinc is removed from the slag and further zinc leaching from the slag occurs in a second stage to form a pregnant leaching solution and recovering zinc from the solution. The method may be used to remove SO.sub.2 from a gas stream by using the SO.sub.2-containing gas stream to leach a slurry of the slag.

Claims

1. A method for recovering zinc from slag derived from lead smelting comprising subjecting the slag to a leaching step under conditions in which zinc is dissolved into a solution, silica being added to the leaching step at a specific silicon addition rate of 10 g Si/L.Math.hour or less, separating the solution from a solid residue and recovering zinc from the solution, wherein the specific silicon addition rate is expressed as tonnes per hour divided by the live volume of the leaching step expressed as cubic meters.

2. The method of claim 1, wherein the slag is ground or milled to reduce particle size of the slag so that the particle size is reduced such that P.sub.80 is approximately from (a) 30 to 300 m, (b) 30 to 150 m, or (c) about 65 m.

3. The method of claim 1, wherein the slag is subjected to acid leaching using a leachant comprising sulphuric acid.

4. The method of claim 1, wherein the leaching step comprises a single leaching stage.

5. The method of claim 1, wherein the leaching step is conducted in a first leaching stage that removes part of the zinc from the slag and a second leaching stage that removes further zinc and the first stage of the leaching process comprises adding a leaching agent in an amount that provides only a portion of overall acid demand.

6. The method of claim 5, wherein the leaching agent comprises an acid and the acid is added in an amount such that either (a) from 40% to 70%; (b) from 50% to 65%; or (c) around 60% of the zinc is extracted from the slag in the first leaching stage.

7. The method of claim 5, wherein the pH in the first leaching stage is controlled such that the pH falls in the range of (a) from 2.0 to 5.0 or (b) from about 3.0 to about 4.0.

8. The method of claim 7, wherein the first leaching stage is operated at a temperature that is below the boiling point of the solution; or the first leaching stage is operated at temperatures falling within the range of (a) 50 C. to 100 C., (b) from 70 C. to 95 C. or (c) about 90 C.

9. The method of claim 5, wherein the residence time in the first leaching stage is (a) at least 4 hours or (b) at least 6 hours.

10. The method of claim 5, wherein in the second leaching stage, the solids are further leached with a further leaching agent.

11. The method of claim 10, wherein the further leaching agent comprises acid.

12. The method of claim 11, wherein the acid is added to the slurry from the first leaching stage such that an excess of acid beyond the acid demand of the solids is used as the further leaching agent.

13. The method of claim 5, wherein the pH in the second leaching stage is (a) less than 2, (b) from 0.5 to 2.0, (c) from 0.5 to 1.5, or (c) around 1.0 which is equivalent to 15 g/L free sulphuric acid.

14. The method of claim 5, wherein the residence time in the second leaching stage is (a) at least 4 hours or (b) at least 6 hours.

15. The method of claim 5, wherein oxygen introduced in the second leaching stage to oxidise Fe++ to Fe+++ allowing the precipitation of some of the iron ions as jarosite.

16. The method of claim 5, wherein the second leaching stage is operated at a temperature that is below the boiling point of the solution; or the second leaching stage is operated at temperatures falling within the range of (a) 50 C. to 100 C., (b) from 70 C. to 95 C. or (c) about 90 C.

17. The method of claim 1, wherein the specific silicon addition rate, as calculated across all leaching stages is (a) 5 g Si/L.Math.h. or less or (b) 3.3 g.Si/L.Math.h. or less.

18. The method of claim 5, wherein the first leaching stage and the second leaching stage are conducted in separate vessels.

19. The method of claim 5, wherein the leaching step includes a treatment step to precipitate dissolved iron.

20. The method of claim 5, wherein the first and second leaching stages are conducted in the same vessel.

Description

BRIEF DESCRIPTION OF THE DRAWINGS

(1) FIG. 1 shows a flow sheet of an embodiment of the process of the present invention where slag is leached directly with sulphuric acid

(2) FIG. 2 shows a flow sheet of an embodiment of the process of the present invention where slag is leached directly with sulphuric acid. FIG. 2 uses a different zinc recovery process to that of FIG. 1; and

(3) FIG. 3 shows a flowsheet of an embodiment of the process of the present invention where slag is leached with SO.sub.2 in a gas scrubbing system.

DETAILED DESCRIPTION OF THE DRAWINGS

(4) It will be appreciated that the attached drawing has been provided for the purposes of describing a preferred embodiment of the present invention. Thus, it will be understood that the present invention should not be considered to be limited to the features as shown in the attached drawings.

(5) Direct Sulphuric Acid Leaching of Slag

(6) In FIG. 1, a grinding circuit A is supplied with water 1, and slag resulting from lead smelting operations. This slag is supplied directly from the lead smelter via stream 2 and/or from a stockpile of slag via stream 26.

(7) A typical analysis of the slag being fed to the mill is as follows:

(8) TABLE-US-00001 Zn 11-16% Pb 2.0-2.5% Fe 20.5-25.0% Ca 16.0-18.0% Mg 0.4-1.2% Al 0.6-1.0% Si 8.5-10.5% Cu 0.1-0.4%

(9) Throughout this specification, all percentages are given in terms of weight percent.

(10) The slag that is fed to the grinding circuit A has a particle size distribution such that the P.sub.80 of the slag is approximately 1300 m. The mill grinds or crushes the slag such that the slag leaving the circuit has a particle size distribution such that the P.sub.80 is approximately 65 m.

(11) After leaving the grinding circuit A, the slurry of water and ground slag 3 is sent to the leaching process. The leaching process is conducted in two steps or two stages, namely B and C. The leaching stage may be conducted in two separate process vessels, with the first step or stage being conducted in one process vessel or vessels and the second step or stage being conducted in another process vessel or vessels. Alternatively, the leaching process B and C may take place in a single reactor, with the conditions of the leaching and residence time being changed as the process progresses from the first step or stage to the second step or stage.

(12) One possible embodiment is shown with leach stages B and C in which the leaching process is conducted in two separate process vessels. The slurry 3 from grinding circuit A is fed to the first leaching vessel B. Acid 4 is added to the first vessel B. The acid 4 may comprise sulphuric acid or an SO.sub.2 stream. Air may be added to both stages of leaching B and C for evaporative cooling. Oxygen may also be added to leach stage C for dissolved iron control resulting in jarosite precipitation.

(13) The slurry 5 from process vessel B (which contains solids and liquids) is transferred to the second process vessel C in which the second stage of the leaching process is conducted. In particular, in the second stage, further sulphuric acid 6 is added to the process vessel to cause further dissolution of zinc from the solids. The slurry 7 from the second process vessel C (and therefore from the second stage of the leaching process) is then removed from the second process vessel.

(14) The first stage of the leaching process is operated at a pH that falls between 3.0 and 4.0. In the first stage of the leaching process, about half of the acid required to meet the acid demand of the slag is added. This results in leaching of approximately 60% of the zinc contained in the slag. The acid also dissolves some of the iron in the slag and, in the first stage, around 40% extraction of iron occurs. Furthermore, some of the magnesium, aluminium and manganese in the slag also dissolves in the first stage of the leaching process. Fe.sup.3+ species dissolved in the first stage also precipitate in the first stage. Fe.sup.2+ and Zn.sup.2+ also dissolve in the first stage and remain soluble. The solute in the first stage may have a dissolved zinc content of approximately 35 g per liter and a dissolved iron content of approximately 25 g per liter.

(15) The reactions that occur in the first stage of the leaching process are highly exothermic. Therefore, it will generally be necessary to cool the first stage in order to control the temperature of the first stage. The present inventors operate the first stage of the leaching process at 90 C., although other range of temperatures may also be used.

(16) The residence time in the first stage of the leaching process is desirably at least 4 hours, but preferably 6 hours.

(17) The slurry from the first stage of the leaching process is then treated in the second stage of the leaching process. In the second stage of the leaching process, an excess of acid is used, in that more acid than is required to meet the acid demand of the slag is added. Typically, an excess of 10 to 15 g per liter of acid is used. The pH in the second stage of the leaching process is typically around 0.5 to 1.5, more desirably from 0.5 to 1.0. The temperature and the residence time in the second stage of the leaching process are generally similar to that used in the first stage of the leaching process (i.e. 90 C. and a residence time of at least 6 hours).

(18) For reasons of maximising throughput of the process, it is desirable to use a specific silicon addition rate that is compatible with process stability and robustness. It is believed that a specific silicon addition rate of up to 10 g Si/l.Math.h, more suitably up to 5 g Si/l.Math.h, will produce acceptable processing. However, a specific silicon addition rate of up to 3.3 g Si/l.Math.h is believed to produce further enhancements to process stability and robustness while still retaining a process report that is significantly higher than the prior art led slag leaching processes known to the inventor. Process residence time and slag feed rate can be adjusted to achieve the acceptable range of silicon addition rate.

(19) While a process residence times of 4 hours per leaching stage or 8 hours total are possible and do fall within the scope of the present invention, process residence times below 6 hours per leaching stage or 12 hours in total may cause unstable kinetic behaviour of silica which results in downstream problems such as solid/liquid separation or formation of silica gel. Operating below a total leaching residence time of 12 hours will reduce process robustness. In more preferred embodiments of the present invention, the specific silicon addition rate is from 3.3-5 g Si/L.Math.h and is desirably kept below this value to ensure acceptable process conditions. In both stages of leaching, air 27 may be introduced for temperature control.

(20) Sufficient copper, present mainly as a sulphide, will also be leached from the slag to act as a catalyst for the oxidation of iron species.

(21) The discharge slurry 7 from the leaching stages may have the following composition:

(22) Solids:

(23) TABLE-US-00002 Zn 0.2-0.4% Pb 1.0-2.0% Fe 5.0-8.0% Ca 12.0-16.0% Mg 0.1-0.2% Al 0.1-0.3% Si 6.0-9.0%

(24) Liquid:

(25) TABLE-US-00003 Zn 50-60 g per litre Fe.sup.2+ 5-40 g per litre Fe.sup.3+ zero Ca 0.15-0.5 g per litre Mg 1.2-2.5 g per litre Al 2.0-5.0 g per litre Mn 0.2-0.7 g per litre Cu 0.1-0.5 g per litre Si 0.1-0.3 g per litre H.sub.2SO.sub.4 10-20 g per litre pH 0.5-1.0

(26) The slurry 7 from slag leaching is subsequently fed to the treatment step D in which precipitation of dissolved Fe and Al takes place. In the process flow sheets shown in FIG. 1, the treatment step D involves a jarosite precipitation. In this process, air or oxygen 28 and lead blast furnace slag 8 is fed to the process vessel. Sodium sulphate 30 is also fed to the process vessel. The source of sodium may also be fed to leach reactor C, if required. The sodium sulphate provides a source of sodium for the precipitation of jarosite. Other sodium or ammonia sources may also be used. Addition of the blast furnace slag 8 acts to neutralise some of the acid and increases then maintains the pH to around 1.0- 2.0, which corresponds to about 5 g/L free sulphuric acid. In some embodiments, oxygen may also be introduced to leaching vessel 2 (C) to precipitate some jarosite as well. This results in the formation of jarosite having a chemical formula of NaFe.sub.3(SO.sub.4).sub.2(OH).sub.6. The reaction that forms jarosite also generates acid. The slag will neutralise acid coming from the leaching stage and also neutralise the acid generated during iron precipitation. Adding oxygen to the leaching stage may be desirable to precipitate some of the jarosite in the leaching stage because the acid generated from the precipitation reaction reduces the acid demand of the slag. It will be understood that the sodium component in this formula may be replaced by other cations, as will be well known to person is skilled in the art. The jarosite is formed as a precipitate and the jarosite precipitates on the solids in the slurry in process D.

(27) The jarosite precipitation process D results in the precipitation of substantial amounts of the Fe.sup.2+ in solution after oxidation to Fe.sup.3+. However, complete precipitation of the Fe.sup.2+ is unlikely to occur in step D. Indeed, modelling conducted by the present applicant has shown that the amount of Fe.sup.2+ remaining in the liquid following the jarosite precipitation may be up to 3 to 4 g per liter but more likely 1-2 g/L.

(28) It will be understood that the jarosite precipitation step D may be conducted in one or more separate stages.

(29) Jarosite stages may require heat addition via steam injection as excessive heat loss can occur when gases are injected for iron oxidation.

(30) The slurry 9 from the jarosite precipitation process D is subsequently sent to solid/liquid separation step E. A flocculating agent may be added prior to solid/liquid separation. In the flow sheets shown in FIG. 1, solid/liquid separation step E utilises a filter to separate the solids from the liquid. A flocculating agent 10 may be added to assist in the filtration step. Wash water 11 is also used to wash the filter cake. Approximately 1-1.5 tonne of wash water is required for every tonne of solid to ensure 99% solution recovery.

(31) The solid residue 12 from filter step E is recovered and used as mine fill. It has been found that the solid residue 12 has good qualities that enable it to be used as a mine fill. Indeed, the solids residue 12 has structural properties that are equivalent to about half those of cement. Therefore, it is believed that the solids residue has the potential to provide a replacement for approximately one third of the cement presently used for mine fill operations.

(32) Advantageously in the process of the present invention, the leaching step utilised in the process results in the formation of precipitated silica compounds that have desirable filtering qualities and therefore do not deleteriously affect the filtering operation. This is in sharp contrast to previous attempts to recover zinc from lead smelting slag in which silica containing gels or colloidal silica containing particles were formed in a precipitation step and which proved to be very difficult to separate from the liquor.

(33) The liquid separated from the solids in filter step E is transferred via stream 13 to downstream processing for polishing of iron and aluminium. The liquid stream 13 may have the following dissolved components:

(34) TABLE-US-00004 Zn 34-38 g per litre Fe.sup.2+ 0.5-1.0 g per litre Fe.sup.3+ 0.01 g per litre Ca 0.2-0.4 g per litre Mg 1.0-2.0 g per litre Al 0.5-1.0 g per litre Mn 0.1-0.5 g per litre Cu 0.1-0.5 g per litre Si 0.05-0.2 g per litre pH 1.0-3.0

(35) The liquor 13, as mentioned above, is subsequently treated to polish remaining iron and aluminium. In the flow sheets shown in FIG. 1, the subsequent treatment of the liquor 13 may be summarised as including a further Fe precipitation step, followed by zinc oxide precipitation and separation.

(36) The, further Fe precipitation step may be described as a polishing step. In the further Fe precipitation step F, the liquid 13 is mixed with slag, lime or any other neutralising agent 14 and recycle material from zinc precipitation 15 and air or oxygen 29. As the liquid also contains some sodium sulphate, further precipitation of jarosite occurs to reduce the content of dissolved Fe in the liquid to significantly less than the 1 g per liter. In this step, iron is more likely precipitated as goethite or some other iron hydroxide rather than jarositesodium is just carried through. The pH of the liquid following the further Fe precipitation step F is approximately 4.5.

(37) The slurry 16 from solution polishing F is sent to clarifier G. In clarifier G, the precipitated solids are removed in the underflow 31 and the polished liquor or Pregnant Leach Solution (PLS) 18 is removed via the overflow. A flocculating agent 17 may be added to the clarifier G in order to assist in the solid/liquid separation process.

(38) As shown in FIG. 1, the solids contained in the underflow 31 are recycled to the leach process B. These solids contain approximately 3% to 10% zinc and by returning solids 31 to the leaching process, further zinc may be extracted therefrom.

(39) The PLS 18 from clarifier G may contain dissolved zinc in an amount of up to 40 g per liter, typically in the range of 30 to 40 g per liter. This liquid has minimal dissolved iron content (10 ppm) and minimal dissolved aluminium content (5 ppm).

(40) This liquor is subsequently treated to recover zinc therefrom. Zinc may be recovered using any process known to be suitable to the person skilled in the art. There are two options provided in this instance. One is the zinc oxide precipitation method and the other is the zinc carbonate precipitation method. It will be appreciated that the present invention should not be considered to recovery of zinc using these two processes alone and that the present invention encompasses any other suitable processes known to recover zinc from solution.

(41) FIG. 1 shows zinc recovery using zinc oxide precipitation. In FIG. 1, liquid 18 may be subsequently fed to zinc oxide precipitation I as described in U.S. Pat. No. 6,726,889, the entire contents of which are herein incorporated by cross reference. However it should be noted that different process conditions to those described in U.S. Pat. No. 6,726,889 are employed to prevent the precipitation of magnesium and magnesium reporting to the concentrate. The preferred process conditions are pH 6.5 at 70-90 C. In this step, lime 19 is added to the liquid. The lime may be hydrated lime (Ca(OH).sub.2) or lime (CaO). Steam may also be required for heating, but if CaO is added, the heat released by the exothermic hydration reaction to form Ca (OH).sub.2 may be sufficient to heat the solution to the desired temperature. Addition of the lime causes precipitation of zinc oxide. Gypsum will also be formed. Careful control of the precipitation parameters results in the zinc oxide forming with a crystal structure that allows for easy separation of the zinc oxide from the precipitated gypsum by virtue of differences in the size of the zinc oxide particles and the gypsum particles. This aspect of the zinc oxide precipitation process (to produce the zinc oxide precipitate with physical properties that allow for easy separation from gypsum) is well known to person is skilled in the art, but it is believed that the operating conditions to prevent precipitation of magnesium are unique.

(42) The slurry 20 of liquid and precipitated solids from zinc oxide precipitation step I is sent to zinc oxide separator J, which suitably may be in the form of a cyclone. In this separator J, the solids are separated into a fine overflow stream (e.g. sub 30 m) 21 (which contains approximately 70% zinc and approximately 2 percent calcium, equating to 95 to 99% recovery of the zinc oxide) and a coarse underflow stream (e.g. plus 30 m stream) (which contains approximately 2% zinc and the bulk of the remainder being gypsum).

(43) The zinc oxide stream 21 is sent to zinc oxide thickener K. A flocculating agent may be added. The thickened zinc oxide stream 22 obtained therefrom is sent to the zinc oxide filter L. Wash water is used to wash the filter cake to remove any soluble contaminants such as chlorine. The filtrate 23 is subsequently returned to the zinc oxide thickener K or a portion of the flow may be bled to control accumulation of deleterious minor elements. A zinc oxide containing filter cake 24 is then sent to stockpile. The zinc oxide filter cake may be sold as a concentrate suitable for recovery of zinc therefrom.

(44) Returning now to the zinc oxide separator J, the underflow is split into recycle streams 15 and 25. Stream 15 is returned to the iron polishing stage F for recovery of contained zinc and neutralisation duty. Stream 25 is recycled to permit seeding of the gypsum crystal facilitating improved separation of zinc oxide and gypsum at the zinc oxide separator. If hydrated lime slurry is used, stream 25 is recycled to the hydrated lime stock tank. If CaO is used, stream 25 is recycled to the zinc oxide precipitator.

(45) FIG. 2 shows zinc precipitation using a process that produces a zinc carbonate precipitate. In the process shown in FIG. 2, the PLS liquid 18 may be subsequently fed to re-treated solids reactor I. In this reactor, fresh PLS feed 18 is contacted with the precipitate formed in zinc precipitator K to remove any unreacted limestone. In reactor I, the dominating reaction is that of acid in the fresh PLS feed (18) (which contains zinc sulphate) and unreacted limestone. As such very little zinc is precipitated. The process can operate at a pH range of 4.5-5.5 but preferably 5.0 and at a temperature of from 70-90 C. The slurry from I, 20, is sent forward to thickener, J. The overflow from thickener J, 21, is sent forward to zinc precipitation with limestone K. The underflow from thickener J, 23, is sent forward for gravity separation, N, which may include a preceding step to polish any zinc from solution using hydrated lime. Returning now to zinc precipitator K, overflow 21 from thickener J, is contacted with limestone, 19, to precipitate zinc carbonate and gypsum also leaving unreacted limestone in the solids. The slurry 22 from zinc precipitator K, is sent to the zinc precipitate thickener, M. The overflow 31 of thickener M, may contain a small amount of dissolved zinc and can be recovered in the abovementioned hydrated lime polishing stage. The underflow 25 from thickener M, containing zinc carbonate, gypsum and unreacted limestone is returned to the residual limestone reactor, I.

(46) The slurry 23 of liquid and precipitated solids from zinc carbonate precipitation step I is sent to zinc oxide separator N, which suitably may be in the form of a cyclone. In this separator N, the solids are separated into a fine overflow stream (e.g. sub 30 m) 26 (which contains approximately 50% zinc and approximately 4 percent calcium, equating to 90 to 95% recovery of the zinc carbonate) and a coarse underflow stream (e.g. plus 30 m stream) (which contains approximately 4% zinc and the bulk of the remainder being gypsum).

(47) The zinc carbonate stream 26 is sent to zinc oxide thickener O. A flocculating agent may be added. The thickened zinc carbonate stream 30 obtained therefrom is sent to the zinc carbonate filter L. Wash water is used to wash the filter cake to remove any soluble magnesium and zinc which can be recycled to the process. The filtrate 29 is returned to the process in the absence of zinc polishing prior to filtration as there will still be some zinc in solution or a portion of the flow may be bled to control accumulation of deleterious minor elements such as magnesium. A zinc carbonate containing filter cake 24 is then sent to stockpile. The zinc carbonate filter cake may be sold as a concentrate suitable for recovery of zinc therefrom.

(48) Returning now to the zinc carbonate separator N, the underflow is split into recycle streams 27 and 28. Stream 28 is returned to the iron polishing stage F for recovery of contained zinc and neutralisation duty. Stream 27 is recycled to permit seeding of the gypsum crystal facilitating improved separation of zinc carbonate and gypsum at the zinc carbonate separator.

(49) The present invention provides a process that allows for the recovery of significant quantities of zinc from lead blast furnace slag, material that would otherwise be a waste material. Thus, the slag becomes a valuable resource that can improve the economics of the mining and smelting operations. The zinc is recovered from the lead smelting slag at greater than 95% recovery.

(50) Scrubbing SO.sub.2 from a Gas Stream and Leaching Slag:

(51) In FIG. 3, a grinding circuit A is supplied with water 1, and slag resulting from lead smelting operations. This slag is supplied directly from the lead smelter via stream 2 and/or from a stockpile of slag via stream 26.

(52) A typical analysis of the slag being fed to the mill is as follows:

(53) TABLE-US-00005 Zn 11-16% Pb 2.0-2.5% Fe 20.5-25.0% Ca 16.0-18.0% Mg 0.4-1.2% Al 0.6-1.0% Si 8.5-10.5% Cu 0.1-0.4%

(54) Throughout this specification, all percentages are given in terms of weight percent.

(55) The slag that is fed to the grinding circuit A has a particle size distribution such that the P.sub.80 of the slag is approximately 1300 m. The mill grinds or crushes the slag such that the slag leaving the circuit has a particle size distribution such that the P.sub.80 is approximately 65 m.

(56) After leaving the grinding circuit A, the slurry of water and ground slag 3 is sent to the scrubbing system vessels via a ring main. The scrubbing process is shown in two stages, namely, B and C, for the purpose of maximising scrubbing efficiency but it can be operated in one stage.

(57) One possible embodiment is shown with leach stages B and C in which the leaching process is conducted in two separate process vessels. Gas stream 40 containing SO.sub.2 enters the scrubbing stage B where it is contacted with slag slurry where a portion of the SO.sub.2 is removed. The exit gas 41 from stage B enters scrubbing stage C for further scrubbing where the exit gas 42 from stage C is the final clean gas and is typically emitted to the atmosphere. The slurry 3 from grinding circuit A is fed to both scrubbing vessels B and C as required. Scrubbing slurry from stage B can be pumped to stage C and vice versa depending on the operating pH set-points in each scrubbing stage. SO.sub.2 laden gas 40 is added to the first scrubbing stage only. Fresh slag slurry is added to both scrubbing stages B and C to regulate pH to a pre-set value. Air 27 may be added to both stages of scrubbing B and C to oxidise sulphites to sulphates and ensure the effective operation of the scrubbing system.

(58) In this example, both scrubbing stages are operated at pH 4.0, but the two stages could be operated at different pH settings to optimise slag consumption and zinc extraction. The first stage of the scrubbing process is operated at a pH that falls between 3.0 and 4.0. This results in leaching of approximately 50 to 60% of the zinc contained in the slag. The acid also dissolves some of the iron in the slag and, in the first scrubbing stage, typically around 40% extraction of iron occurs. Furthermore, some of the magnesium, aluminium and manganese in the slag also dissolves in the first stage of the leaching process. Fe.sup.3+ species dissolved in the first stage also precipitate in the first stage. Fe.sup.2+ and Zn.sup.2+ also dissolve in the first stage and remain soluble. The operating range of pH 3.0 to 4.0 results in the iron oxidising readily through the introduction of air and precipitating as an iron oxide/hydroxide. The solute in the first stage may have a dissolved zinc content of approximately 38 g per liter and a dissolved iron content of approximately 0.5 g per liter.

(59) The reactions that occur in the scrubbing process are highly exothermic however the process will run autothermally at around 60 C.

(60) In the first scrubbing stage, depending on the equipment size and liquid to gas ratio, the SO.sub.2 scrubbing efficiency will be around 85%.

(61) The residence time in the first scrubbing stage is typically around 10 to 24 hours based on the high recirculating load of scrubbing slurry in the scrubbing vessel.

(62) The slurry from the first stage of the leaching process is then treated in the second stage of the leaching process or alternatively discharged to the thickener D. In, the second stage of the scrubbing process, the exit gas from stage B scrubbing is treated to remove further SO.sub.2 from the gas stream. The pH in the second stage of the scrubbing process is typically around 4.0 in this example but could be lower to optimise zinc extraction and slag consumption with the first scrubbing stage. The temperature and the residence time in the second stage of the scrubbing process are generally similar to that used in the first stage of the scrubbing process (i.e. 90 C. and a residence time of at least 6 hours).

(63) For reasons of maximising throughput of the process, it is desirable to use a specific silicon addition rate that is compatible with process stability and robustness. It is believed that a specific silicon addition rate of up to 10 g Si/l.Math.h, more suitably up to 5 g Si/l.Math.h, will produce acceptable processing. However, a specific silicon addition rate of up to 3.3 g Si/l.Math.h is believed to produce further enhancements to process stability and robustness while still retaining a process report that is significantly higher than the prior art led slag leaching processes known to the inventor. Process residence time and slag feed rate can be adjusted to achieve the acceptable range of silicon addition rate. It is expected that due to process design the residence time in each scrubbing stage will be between 10 to 24 hours and result in acceptable silicon addition rates to both scrubbing stages.

(64) Sufficient copper, present mainly as a sulphide, will also be leached from the slag to act as a catalyst for the oxidation of iron species.

(65) The discharge slurry 7 from the scrubbing stages may have the following composition:

(66) Solids:

(67) TABLE-US-00006 Zn 3.0-5.0% Pb 1.0-2.0% Fe 11.0-16.0% Ca 11.0-16.0% Mg 0.01-0.2% Al 0.1-1.2% Si 6.0-8.0%

(68) Liquid:

(69) TABLE-US-00007 Zn 30-45 g per litre Fe.sup.2+ 0-0.1 g per litre Fe.sup.3+ 0-1.0 g per litre Ca 0.1-0.5 g per litre Mg 1.5-2.5 g per litre Al 0.05-0.2 g per litre Mn 0.1-1.0 g per litre Cu 0.1-0.5 g per litre Si 0.1-0.3 g per litre pH 3.5-4.5

(70) The slurry 7 from scrubbing is subsequently fed to the thickener D in which the first stage of solid liquid separation occurs. The thickener is fed a flocculant to aid particle settling. The thickener under flow 9 is directed to a filter for solid liquid separation step E. A flocculating agent may be added prior to solid/liquid separation. In the flow sheets shown in FIG. 3, solid/liquid separation step E utilises a filter to separate the solids from the liquid. A flocculating agent 10 may be added to assist in the filtration step. Wash water 11 is also used to wash the filter cake. Approximately 1-1.5 tonne of wash water is required for every tonne of solid to ensure 99% solution recovery.

(71) The solid residue 12 from filter step E is recovered and used as mine fill. It has been found that the solid residue 12 has good qualities that enable it to be used as a mine fill. Indeed, the solids residue 12 has structural properties that are equivalent to about half those of cement. Therefore, it is believed that the solids residue has the potential to provide a replacement for approximately one third of the cement presently used for mine fill operations.

(72) Advantageously in the process of the present invention, the leaching step utilised in the process results in the formation of precipitated silica compounds that have desirable filtering qualities and therefore do not deleteriously affect the filtering operation. This is in sharp contrast to previous attempts to recover zinc from lead smelting slag in which silica containing gels or colloidal silica containing particles were formed in a precipitation step and which proved to be very difficult to separate from the liquor.

(73) The primary filtrate and the first wash are recycled back to the thickener D and the thickener overflow is taken as the Pregnant Leach Solution (PLS) 17 to feed the zinc precipitation stage I. The subsequent filter washes are collected and used as makeup water in the scrubbing and slag milling processes to achieve an overall higher zinc PLS tenor and reduce the downstream cost of zinc precipitation.

(74) This PLS liquor is subsequently treated to recover zinc therefrom. Zinc may be recovered using any process known to be suitable to the person skilled in the art. There are two options provided in this instance. One is the zinc oxide precipitation method and the other is the zinc carbonate precipitation method. It will be appreciated that the present invention should not be considered to recovery of zinc using these two processes alone and that the present invention encompasses any other suitable processes known to recover zinc from solution.

(75) FIG. 3 shows zinc recovery using zinc oxide precipitation. In FIG. 3, liquid 17 may be subsequently fed to zinc oxide precipitation I as described in U.S. Pat. No. 6,726,889, the entire contents of which are herein incorporated by cross reference. However it should be noted that different process conditions to those described in U.S. Pat. No. 6,726,889 are employed to prevent the precipitation of magnesium and magnesium reporting to the concentrate. The preferred process conditions are pH 6.5 at 70-90 C. In this step, lime 19 is added to the liquid. The lime may be hydrated lime (Ca(OH).sub.2) or lime (CaO). Steam may also be required for heating, but if CaO is added, the heat released by the exothermic hydration reaction to form Ca(OH).sub.2 may be sufficient to heat the solution to the desired temperature. Addition of the lime causes precipitation of zinc oxide. Gypsum will also be formed. Careful control of the precipitation parameters results in the zinc oxide forming with a crystal structure that allows for easy separation of the zinc oxide from the precipitated gypsum by virtue of differences in the size of the zinc oxide particles and the gypsum particles. This aspect of the zinc oxide precipitation process (to produce the zinc oxide precipitate with physical properties that allow for easy separation from gypsum) is well known to person is skilled in the art, but it is believed that the operating conditions to prevent precipitation of magnesium are unique.

(76) The slurry 20 of liquid and precipitated solids from zinc oxide precipitation step I is sent to zinc oxide separator J, which suitably may be in the form of a cyclone. In this separator J, the solids are separated into a fine overflow stream (e.g. sub 30 m) 21 (which contains approximately 50% to 70% zinc and approximately 2 percent calcium, equating to 95 to 99% recovery of the zinc oxide) and a coarse underflow stream (e.g. plus 30 m stream) (which contains approximately 2% zinc and the bulk of the remainder being gypsum).

(77) The zinc oxide stream 21 is sent to zinc oxide thickener K. A flocculating agent may be added. The thickened zinc oxide stream 22 obtained therefrom is sent to the zinc oxide filter L. Wash water is used to wash the filter cake to remove any soluble contaminants such as chlorine. The filtrate 23 is subsequently returned to the zinc oxide thickener K or a portion of the flow may be bled to control accumulation of deleterious minor elements. A zinc oxide containing filter cake 24 is then sent to stockpile. The zinc oxide filter cake may be sold as a concentrate suitable for recovery of zinc therefrom.

(78) Returning now to the zinc oxide separator J, the underflow is split into recycle streams 15 and 25. Stream 15 is returned to the scrubbing stage B although this could also be used in an iron and aluminium polishing stage external to the scrubbing vessels as required. Stream 25 is recycled to permit seeding of the gypsum crystal facilitating improved separation of zinc oxide and gypsum at the zinc oxide separator. If hydrated lime slurry is used, stream 25 is recycled to the hydrated lime stock tank. If CaO is used, stream 25 is recycled to the zinc oxide precipitator.

(79) The zinc carbonate precipitation process described above could also be utilised for the recovery of zinc from solution.

(80) The process of this embodiment of the present invention may also offer the following additional advantages:

(81) a) dust in the gas stream may also be removed as it passes through the scrubbing process;

(82) b) elements in the dust may also be recovered to solution for subsequent recovery. These elements may include thallium and cadmium;

(83) c) the process can deal with process surges and disruptions including changes in concentration and flow of feed gas.

(84) d) the scrubber can use, in balance, the amount of slag that is generated at the blast furnace, or can consume more slag than is produced if a stockpile needs to be consumed.

(85) Those skilled in the art will appreciate that the present invention may be susceptible to variations and modifications other than those specifically described. It will be understood that the present invention encompasses all such variations and modifications that fall within its spirit and scope.