WO2008139412A1

WO2008139412A1 - Producing a metal like zinc from sulphide ores by chloride leaching and electrowinning

Info

Publication number: WO2008139412A1
Application number: PCT/IB2008/051873
Authority: WO
Inventors: Johann Du Toit Steyl; Jan Tjeerd Smit
Original assignee: Anglo Operations Pty Ltd
Current assignee: Anglo Operations Pty Ltd
Priority date: 2007-05-10
Filing date: 2008-05-12
Publication date: 2008-11-20
Anticipated expiration: 2009-11-10

Abstract

A method for producing a metal from a metal sulphide ore or mineral is claimed. The metal sulphide ore is leached by non-oxidative dissolution in a chloride solution to produce a metal chloride solution and hydrogen sulphide; the metal chloride solution is subjected to one or more electrolytic processes to produce chlorine gas; and the chlorine gas is then converted into hydrochloric acid. The chlorine may be produced in an electrolytic cell, such as a zinc metal electrowinning cell, a chlor-alkali cell, an iron metal electrowinning cell, a manganese metal electrowinning cell, an iron- manganese alloy electrowinning cell or any combination of these cells or any other electrolytic cell that can selectively remove a metal from solution while producing chlorine at the anode. The metal may be Zn, Cu, Mo, Ti, Al, Cr, Ni, Co, Mn, Fe, Pb, Na, K, Ca, Ag, Au or any platinum group metal.

Description

PRODUCING A METAL LIKE ZINC FROM SULPHIDE ORES BY CHLORIDE LEACHING AND ELECTROWINNING

BACKGROUND OF THE INVENTION

The invention provides a process for producing a metal from sulphide ores.

One process for producing zinc from sulphide ores is the Roast Leach Electrowinning (RLE) process, shown in simplified form in Figure 1. This process produces zinc metal and excess acid (SO₂ - equivalent to the sulphur entering with the ore), which cannot be directly discharged into the environment. Zinc ferrite (formed during the roasting step) processing is expensive and technically challenging (especially when the feed ore contains high iron), since a high acid leach (HAL) is associated with more iron leaching (which has to be neutralised in situ during hydrolysation). Although ferri- hydrite formation is effective in controlling various heavy metal impurities, this residue material is not environmentally inert, thus it has to be contained and cannot be simply used as landfill. It is also a bulky material and exhibits poor settling and filtration characteristics, which eliminates the viability of treating high iron ores using this technology route. Typically, the leach and neutralisation steps would be integrated counter-currently to minimise the use of an external neutralising agent (such as lime).

Since the ferri-hydrite would not survive HAL conditions, jarosite precipitation has been proposed as an alternative (the zinc ferrite would have to be leached under conditions where jarosite is stable). Long residence times would, however, be required to achieve this, as well as a suitable cation, such as sodium. Some sulphur is lost from the circuit as part of the jarosite phase. Similarly to ferri-hydrite, jarosite disposal is a problem (it is bulky and further treatment is required in order for it to be used as landfill). Although less bulky, goethite is not considered a good alternative because the feed solution requires pre-reduction (prior to slow oxidation) or dilution {in large tanks) in order to kinetically favour goethite formation during hydrolysis and goethite would not stand up to the harsh HAL conditions.

Similarly, hematite (despite containing the highest weight percentage iron, as compared to the other hydrolysis products) precipitation would not integrate well with a ferrite leaching step (high lime cost), not to mention the fact that pre-reduction (reductant), followed by slow oxidation (oxidant) and seeding would be required in order for it to form under atmospheric conditions.

Any other means of removing iron impurity (such as solvent extraction), would require some form of expensive (equivalent to the amount of iron leached) neutralisation step.

The Pressure Oxidation (POX) Process resolves the iron issue because hydrolysis is temperature driven, and thus there is no need for in-situ neutralisation (Figure 2). The process takes place in an autoclave at 15O⁰C and hematite is the preferred hydrolysis product. Sulphur, entering with the sulphide ore, is almost stoichiometricaily converted into elemental sulphur, exiting with the final hematite residue, i.e. it does not add to process revenue (from producing elemental sulphur). However, the POX Process is simple and acid is utilised optimally due to the rejection of iron in the main process step (the autoclave) itself, thus releasing the sulphate that was initially consumed by the iron impurity. A counter-current integration of two autoclaves is utilised to achieve both high zinc extraction and near complete iron removal. The inefficiency of the process is exclusively contained in approximately 10 g/l residual acid and 2 g/l iron in the final PLS, and less than 1% zinc in the final residue. This is exactly the aspect that would be difficult to compete with in an atmospheric process, where acid cannot be regenerated in situ.

However, the major drawback of the POX Process is its high capital cost and its inability to treat, for example, concentrates that contain high lead, gold or silver values (gold or silver would be difficult and expensive to recover, while lead would contaminate the final residue, not to mention its possible passivation effect due to the formation of insoluble PbSO₄ on the unreacted sphalerite surface). Due to the exothermic nature of the oxidation process, pulp density is another critical factor that would limit the throughput rate of a sulphide concentrate. Since the capital investment would be highly dependent on the autoclave size (or number of autoclaves), the zinc grade and sulphur content in the feed would be constraining factors to the economic viability of a specific concentrate. Bulk concentrates or concentrates that contain high pyrite concentrations (produce excess heat and acid) or a high pyrrhotite (excess heat) content, for example, would not be economically treatable in an autoclave. Also, even though the kinetics of leaching would be significantly faster during POX, as compared to atmospheric leaching (AL), the concentrate still needs to be ground to typically 98% passing 45μm.

The relatively flat response of the capital cost (in absolute terms) of AL processes to changes in Zn grade and sulphide content is why companies are pursuing alternatives to pressure leaching. The generic conceptual AL process is illustrated in Figure 3.

Various processes have been proposed to utifise the general scheme of the AL process in treating concentrates under atmospheric leaching conditions in sulphate medium, such as the MIM Albion process (Hourn et ai, 1999) This process relies on ultra fine milling (80% passing 3μm) to improve the leaching kinetics while the MIM bioleaching (also called the Biozinc process) process (Steemson et a/., 1994 and 1997) employs bacteria for the oxidative leaching of zinc sulphide.

The oxidation, neutralisation and electrowinning steps contribute significantly to the overall reagent and energy operating costs of these atmospheric processes. As an alternative to the sulphate-based processes, chloride medium may be adopted in the leaching step in order to achieve a more efficient integration with the rest of the circuit. For example, the HydroCopper Process (Hyvarinen and Hamalainen, 1999) integrates the chloride leach with the chlor-alkali cell (standard technology), to produce internally all the reagents required for regenerating the primary leaching agent (cupric ion) and to recover value metal. Another example is the lntec process (www.intec.com.au). operating in a closed loop fashion and using the electrowinning of zinc from chloride solution to recover zinc metal and to regenerate the oxidant required for leaching. A significant cost saving is realised by using electrical energy to regenerate the primary reagents and also to win the metal from chloride solution, rather than using electrical energy to win the metal from sulphate solution. The reason for this is that the chlor- alkali electrolysis cell is easily operated and more efficient than the copper electrowinning cell, using relatively inexpensive sodium chloride as the make-up reagent. Also, higher current efficiencies can be obtained when winning zinc from chloride solution compared to the sulphate solution. However, electrowinning from chloride solutions poses unique problems, such as powder production at the cathode (handling problems) and chlorine generation at the anode (corrosion).

SUMMARY OF THE INVENTION

According to the invention, there is provided a method for producing a metal from a metal sulphide ore or mineral, the method including the steps of:

(i) leaching the metal sulphide ore or mineral by non-oxidative dissolution in a chloride solution to produce a metal chloride solution and hydrogen sulphide; (ii) subjecting the metal chloride solution to one or more electrolytic processes to produce chlorine gas; and (iii) converting the chlorine gas into hydrochloric acid.

The hydrogen sulphide, or a portion thereof, may be used to convert the chlorine gas into hydrochloric acid, thereby also producing elemental sulphur. Alternatively, an additional hydrogen source, whether produced in the process or not, may be used to convert the chlorine gas into hydrochloric acid.

The hydrochloric acid may be returned to the chloride solution.

The chlorine may be produced in an electrolytic cell, such as a zinc metal electrowinning cell, a chlor-alkali cell, an iron metal electrowinning cell, a manganese metal electrowinning cell, an iron-manganese alloy electrowinning cell or any combination of these electrolytic cells or any other electrolytic cell that can be used to selectively remove a metal from solution while producing chlorine gas at the anode.

The metal may be a valuable (value) metal, such as zinc.

The method may be conducted at atmospheric pressure and without the formation of elemental sulphur.

The bulk of the sulphide sulphur may be recovered as an elemental sulphur byproduct, while harvesting the energy of the sulphide sulphur oxidation to elemental sulphur and reducing the need for fresh reagent make-up.

Impurities (e.g. iron, aluminium, silica and trace heavy metals) may be removed from soiution to levels required by electrolytic processes. Preferably, chemicals produced in the process, such as chlorine gas, caustic soda, hydrogen sulphide, hydrogen gas, metal (for example zinc), sulphur dioxide, and/or sulphur dioxide/oxygen mixture, will be used to remove the impurities. Cooling, salting out or evaporative crystallisation may also be used to remove impurities. Alternatively, solvent extraction may be used to produce a solution, pure enough to be subjected to the electrolytic process. Depending on the iron and manganese impurity levels in the feed material, electrolytic processes may be included in the integrated flowsheet to directly remove iron as a metal and/or manganese as a metal or alternatively in one single cell as an iron- manganese alloy.

The method may be used without first pre-concentrating the minerals in a cleaner concentrate. Thus, direct treatment of the minerals in the form of a bulk concentrate may be possible, or even direct treatment of the raw ore. The overall metal recovery may thereby be increased. The method may therefore be performed without the need for a fine grinding step of the ore.

A portion or all of the hydrogen sulphide may be converted to elemental sulphur using Claus or other technology.

A portion of the hydrogen sulphide may be burned in air or oxygen to produce sulphur dioxide, which may be converted to sulphuric acid. The sulphuric acid may then be used to replace an equivalent amount of hydrochloric acid in a mixed chloride/sulphate brine leach.

Sulphate salts of metals may be produced and removed from the leach process, thermally decomposed to oxides and sulphur dioxide and subsequently converted to sulphuric acid. BRIEF DESCRIPTION OF THE DRAWINGS

Figure 1 shows the commercial Roast-Leach-Electrowinning (RLE) process;

Figure 2 shows the commercial Pressure Oxidation (POX) process;

Figure 3 shows the generic oxidative atmospheric leach (AL) concept (near commercialisation ref. Albion Process);

Figure 4 shows a first embodiment of the present invention, integrating a non- oxidative sulphide leach step with a Chlor-alkali electrolytic cell, and regenerating hydrochloric acid via reaction of chlorine gas with hydrogen sulphide gas;

Figure 5 shows a second embodiment of the present invention, integrating the non-oxidative sulphide leach step with the Zn electrowinning cell, and regenerating hydrochloric acid via reaction of chlorine gas with hydrogen sulphide gas;

Figure 6 shows a third embodiment of the present invention, integrating the non- oxidative sulphide leach step with both the chlor-alkali electrolytic cell and the zinc electrowinning cell, and regenerating hydrochloric acid via reaction of chlorine gas with hydrogen sulphide gas, hydrogen gas from the chlor-alkali electrolytic cell and/or an alternative source of hydrogen;

Figure 7 shows a fourth embodiment of the present invention, integrating the non-oxidative sulphide leach / purification / zinc electrowinning / hydrochloric acid regeneration circuit with the chlor-alkali electrolytic cell and/or the iron-manganese electrowinning cell/s. DETAILED DESCRIPTION OF THE INVENTION

A new metal processing technology, using sulphide as the primary ore, is described herein.

The new process will be described and exemplified below with reference to zinc as the metaf to be obtained from the sulphide ore, but it is intended that the invention could apply to any metal, and in particular any value metal, i.e. any metal from which revenue can be generated. For example, the metal may be a light metal, base metal or precious metal, such as Zn, Cu, Mo, Ti, Al, Cr, Ni, Co, Mn, Fe₁ Pb, Na, K, Ca, Ag, Au and platinum group metals.

The ore may be any metal bearing ore containing sulphide (e.g. sphalerite), which may be in a disseminated form.

As mentioned above, the major drawback of operating under atmospheric pressure is the absence of a thermal driving force for the in-situ regeneration of the acid (equivalent to the amount of Fe entering the leach). If the atmospheric leach (AL) is a bio-leach in a heap, the bulk of the sulphide sulphur would oxidise to acid, requiring even more neutralising agent (high lime requirement) in order to hydrolyse the Fe. Nevertheless, recycling of some of the raffinate acid would still be required to satisfy the acid demand in the leach. A major advantage of opting for a heap leaching operation would be the direct utilisation of the ore, thus eliminating the need for upgrading (and hence the associated metal losses).

Process selection may be a playoff between POX (low risk but high capex) or, for example, heap leaching (lower capex and probably a higher overall Zn recovery, but also higher neutralisation costs and a high risk associated with controlling the heap). Local environmental factors (ferrihydrite vs. hematite formation) and the Zn grade of the feed ore would also be important. If the primary leach is to be conducted in a tank, using ferric ion {Albion approach), ultra-fine grinding (80% passing 3μm) would be required. It would also be necessary to remove the elemental sulphur (via flotation, for example), prior to disposal, to ensure that the tailings (containing the jarosite precipitate) are environmentally acceptable. All of these process flowsheets require an effluent stream and, at least, a bleed stream to be neutralised (the residual acid, present as free acid and Fe) in order to control the impurity levels and to recover the Zn. If a major portion of the sulphide sulphur is converted to elemental sulphur (stoichiometricaily more than what is recovered during the electrowinning of Zn), a deficiency of sulphate would exist, which would have to be added externally, resulting in higher operating costs. This is not desirable, especially since the feed ore introduces a potential source of sulphur that is best utilised in order to compete with POX. Heap leaching, on the other hand, produces excess amounts of acid during the oxidation of sulphide sulphur and would require a neutralising reagent to be added.

This process introduces an atmospheric leaching process, which utilises chloride brine (the term 'brine' as used herein is intended to possibly include sulphate as well as NaCI) solution as the lixiviant, ensuring comparative (cf. other atmospheric leaching processes) leaching kinetics, but without the need for fine-grinding. The applicant has shown that the presence of ZnCI₂ negatively effects zinc extraction, but this can be overcome by the addition of NaCI, which increases zinc extraction. It is therefore beneficial to conduct the leaching step in a high sodium chloride environment. The leaching process operates non-oxidatively, which eliminates the presence of elemental sulphur product layers that may adversely affect leaching kinetics under otherwise oxidative conditions:

ZnS + 2HCI → ZnCi₂ + H₂S (1 )

The absence of elemental sulphur in the leaching step alleviates the need for ultra-fine grinding of the feed material. Depending on the amount and type of impurities present in a specific deposit, a bulk concentrate or even the raw ore may be directly treatable. This holds the potential of significantly increasing the overall Zn recovery as compared to the conventional POX Process, which would require a high-grade cleaner concentrate as feed material.

The vapour stream of the non-oxidative leaching step can be integrated with an electrochemicaϋy generated chlorine gas stream in order to control the extent of elemental sulphur production. Hydrogen sulphide gas, a product of the non-oxidative Jeaching step, reacts with chlorine gas to produce elemental sulphur and hydrochloric acid (Sims and Sheinbaum): H₂S + Cl₂ → S° + 2HCI (2)

Chlorine gas production, using electrical energy, may be incorporated into the flowsheet via the conventional chlor-alkali technology, as commercially applied throughout the world, using, for example, cheap and readily available sodium chloride sa!t as make-up reagent

Anode: 2Cl^" → Cl₂ + 2e^~ (3)

Cathode: 2H₂O + 2e^" → 2OH^" + H₂ (4)

In addition, chlorine would be produced in an electrolytic cell during the winning of zinc metal from chloride solution (Zinclor; lntec Ltd ):

Anode: 2CI^" → Cl₂ + 2e^" (5)

Cathode: Zn²⁺ + 2e^" → Zn⁰ (6)

Compared with electrolytic processes using sulphate media, the chloride processes are characterised by improved current efficiency, besides the fact that chlorine gas may be used to regenerate the primary leaching agent, i.e. hydrochioric acid. The two schemes described above are independently illustrated in Figures 4 and 5, respectively. These electrolytic processes may also be combined in a novel way into one flowsheet, as depicted in Figure 6. Here, the function of the chlor-alkali cell would be primarily to regenerate the equivalent amount of chlorine (hydrochloric acid) that was originally consumed by the impurities in the primary leach, and at the same time produce the required amount of caustic to remove these impurities from the circuit. The water balance in the circuit would probably be maintained by a multiple-effect evaporation step. This is the point in the circuit where sodium chloride is likely to precipitate (due to super-saturation). This salt may then be recycled back to the chlor- alkali cell, which would substantially reduce or eliminate the sodium chloride make-up requirement.

Besides the winning of the zinc metal, impurities such as iron and manganese may also be electrowon, thereby reducing the relative size of the chlor-alkali plant. The major advantage of such a scheme wouid be associated with the potential by-product value from producing iron and/or manganese metal. This unit is expected to be a minor electrowinπing operation since the iron content in a zinc concentrate is typically a 1/5^th or less compared to the zinc content. Another advantage of producing iron metal would be its low volume compared to an iron hydrolysis residue. Very little wash-water would also be added to the circuit, making the water balance (and hence, the energy requirement) more favourable:

Anode: 2cr → a₂ + 2e^" (7)

Cathode: Fe²⁺ + 2e^" → Fe° (8)

Anode: 2cr → C⁾ ₂ + 2e~ (9)

Cathode: Mn^2÷ + 2e^" → Mn° (10)

or

Anode: 2(x+y)CI^" → Cl₂ + 2(x+y)e^" (11 )

Cathode: xFe²⁺ + yMn²⁺ + 2(x+y)e^" -> Fe_xMn_y° (12)

Figure 7 represents a schematic illustration of how this electrowinning of iron and/or manganese impurity may be integrated into the circuit. However, this step may potentially be introduced in any part of the circuit, provided the operating conditions at that specific point in the circuit are conducive to the technical and economical viability of flowsheet.

All the electrolytic processes described above could either use diaphragm or membrane cells (see for example, http://www.olinchloralkali.com for the chlor-alkafi cell options and Diaz et. a). (1993) for the Zinclor design options).

In these electrolytic processes, some degree of purification of the process solution would be required prior to the electrolytic process of choice. These purification steps may include any commercial or non-commercial method without detracting from the main process concept, as described above. Preferably, these purification steps would utilise the chemicals produced in the process itself, for example: chlorine gas, caustic soda, hydrogen sulphide, zinc metal, hydrogen gas, sulphur dioxide, and/or sulphur dioxide/oxygen mixture, and the like, although external reagents could also be added. Alternatively, cooling, evaporative or salting-out crystallisation may be utilised to remove impurities, such as Fe(II). Any of these methods may also be used to produce an intermediate zinc salt product (prior to the electrolytic step) that may be further purified outside the primary process, as described above. Solvent extraction may also be utilised to produce a pure enough solution for the electrolytic processes. Solvent extraction may be integrated into the basic concept of this invention to the best benefit of the workings of the basic flowsheet.

Besides allowing hydrogen sulphide to react with chlorine gas to regenerate the stoichiometrical equivalent hydrochloric acid, a portion of the hydrogen sulphide gas may be converted to elemental sulphur using the well-known Claus technology, or other similar technology, as, for example, frequently applied in the petrochemical industry. Alternatively, any source of hydrogen (whether a product of the chlor-alkali cell or not) may be used to covert chlorine gas into hydrochloric acid. A portion of the hydrogen sulphide gas may be burned in air or oxygen to produce sulphur dioxide, which may be converted to sulphuric acid in an acid plant. This sulphuric acid may then replace the equivalent amount of hydrochloric acid in a mixed chloride/sulphate brine leach. Alternatively, sulphate salts of any of the metals may be produced and removed from the process stream, thermally decomposed to the oxide and sulphur dioxide gas and similarly converted to sulphuric acid. Any one or combination of the above schemes may be utilised to control the reagent regeneration cycles and the management of the hydronium ion, total chloride and total sulphur balance around the circuit. This would automatically reduce the requirement of fresh reagent make-up in the control of the circuit and management of the major impurities.

The process of this invention has the potential for: a) reducing capital and operating expenditure - the reagent consumption is essentially nil (excluding reagent top-up as required). The primary leaching agent is regenerated, there is no neutralisation of the main process stream amd the neutralisation agents are regenerated internally; b) catering for ores which cannot be treated or are difficult to treat by conventional means (e.g. low grade, refractory or disseminated ores) and the ability to treat bulk concentrate and coarse material (this could also allow for the recovery of valuable by-products and a reduced leach time); and c) minimising environmental impact - elemental sulphur (from H₂S generated in the leach) and Pb and Ag could be recovered as valuable by-products and stable Fe₂O₃ and MnO₂ or a potentially saleable ferromanganese alloy could be formed.

The present invention is further described by the following examples. Such examples, however, are not to be construed as limiting in any way either the spirit or scope of the invention.

Example 1: Leaching

Leaching tests have shown in excess of 99 % zinc extraction for cleaner and rougher concentrates and ore. The percentage zinc extraction decreases with an increase in ZnCI₂ concentration in the lixiviant. This effect is illustrated in Table 1.

Table 1: Percentage zinc extracted from 1O g cleaner concentrate under non- oxidative leaching conditions with HCI and with various concentrations Of ZnCI₂ (2 % pulp density, 115 ⁰C)

The addition of NaCI to the lixiviant increases the percentage zinc extraction in the presence of ZnCI₂ as indicated by the results in Table 2. Table 2: Percentage εinc extracted from 10O g cleaner concentrate under non- oxidative leaching conditions with HCI (2 % pulp density, 180 min, 115 ⁰C) and with various concentrations of NaCI and ZnCl₂

Example 2: Solvent extraction

A variety of extractants, e.g. Alamine 336™, Aliquat 336™, Cyanex 921 ™, Cyanex 923™, isodecanol, tιϊ-/7-butylphosphate and dJbutylbutylphosphonate, have been tested in the solvent extraction of zinc from the chloride matrix. The use of these extractants has shown that zinc can be quantitatively extracted using a smail number of extraction stages. This would assist in limiting the ZnCI₂ returned to the leach.

Example 3: HCI regeneration

The reaction of hydrogen sulphide gas with chlorine gas to form elemental sulphur and hydrochloric acid gas is thermodynamically favoured. Table 3 presents the results (data generated with HSC Chemistry 6 software) and confirms negative free energies for Reaction 2 over a wide temperature range. This reaction is therefore expected to proceed rapidly in an exothermic manner, providing a viable route for regenerating hydrochloric acid and producing heat from by-products (hydrogen sulphide and chlorine gas) formed in other parts of the circuit. Table 3: Free energies for the reaction of hydrogen sulphide gas with chlorine gas to form elemental sulphur and hydrochloric acid gas as a function of temperature as determined using HSC Chemistry 6 software

The plating of relatively smooth, adherent deposits of zinc and iron has been achieved by electrowinning. The energy consumption for the plating from the chloride medium is lower than that for the sulphate medium (energy consumption of 2.7 kWh/kg Zn for the deposition of zinc from the chloride system as compared to 3.2-3.3 kWh/kg Zn for the typical sulphate systems). Tests have also been conducted to codeposit iron and manganese.

While the invention has been described in detail with respect to specific embodiments thereof, it will be appreciated by those skilled in the art that various alterations, modifications and other changes may be made to the invention without departing from the spirit and scope of the present invention. References

Diaz, G., Regife, J. M., Frias, C. and Parrilla, F. (1993). Recent advances in Zinclor technology, International Symposium - World Zinc '93, October 1993, pp.341-346.

Hourn, MM., Turner, D.W. and Holzberger, I. R. (1999). Atmospheric Mineral Leaching Process. US Patent US 5993635 A1, 30 November, 1999. (Assigned to MIM Holdings, Ltd.).

Hyvarinen, O. and Hamalainen, M. (1999). Method for Producing Copper in Hydrometallurgical Process, US Patent US 6007600, December 29, 1999. (Assigned to Outokumpu Oy).

Kyuchoukov, G. D. et al (1988). Methods for the Recovery of Metals from Chloride Solutions, Bulgarian Patent Publication 030178382 & European Patent Application 88306804.1. (Filed July 25, 1988). See: Fletcher, A.W., Sudderth, R.B. and Olafson, S.M. (1991) Combining sulphate electrowinning with chloride leaching, JOM (August 1991 ), pp.57-59.

Moyes, A.J. (2007). The lntec Zinc Process (IZP), wvywjntec.com.au, pp.1-4.

Sims, A.V. and Sheinbaum, I. (1983). The Direct Chlorination Process for geothermal power plant off-gas - hydrogen sulphide abatement, Work performed for the Department of Energy, I. Sheinbaum Co. Inc., Monrovia, California, pp.1-46.

Spanish Patent (1984). Procedimiento y aparato para Ia obtenciόn simultanea de cloro y cine a partir de disoluciones acuosas de su cloruro, No. 518 560, April 1984.

Steemson, M. L., Sheehan, GJ. , Winborne, D.A. and Wong, F.S. (1994). An Integrated Bioteach/Solvent Extraction Process for Zinc Metal Production from Zinc Concentrates. PCT World Patent WO 94/28184, 8 December, 1994. (Assigned to MIM Holdings Limited).

Steemson, M. L., Wong, F.S. and Goebel, B. (1997). The integration of zinc bioleaching with solvent extraction for the production of zinc metal from zinc concentrates. In: Internationa! Biohydrometallurgy symposium IBS97 BIOMINE97 "Biotechnology Comes of Age." Glenside, Australia: Australian Mineral Foundation, pp. M 1.4.1 -M 1.4.10.

Claims

CLAIMS:

1. A method for producing a metal from a metal sulphide ore, the method including the steps of:

(i) leaching the ore by non-oxidative dissolution in a chloride solution to produce a metal chloride and hydrogen sulphide; (ii) subjecting the metal chloride to one or more electrolytic processes to produce chlorine gas; and (iii) converting the chlorine gas to hydrochloric acid.

2. A method according to claim 1 , wherein the hydrogen sulphide produced in step (i), or a portion thereof, is used to convert the chlorine gas into hydrochloric acid.

3. A method according to either claim 1 or 2, wherein added hydrogen is used to convert chlorine gas into hydrochloric acid.

4. A method according to any one of claims 1 to 3, wherein the hydrochloric acid is returned to the chloride solution.

5. A method according to any one of claims 1 to 4, wherein the chlorine gas is produced in an electrolytic cell.

6. A method according to claim 5, wherein the electrolytic cell is selected from a zinc metal electrowinning cell, a chlor-alkali cell, an iron metal electrowinning cell, a manganese metal electrowinning cell, an iron-manganese alloy electrowinning cell, any combination of these electrolytic cells or any other electrolytic cell that can be used to selectively remove a metal from solution while producing chlorine gas at the anode.

7. A method according to any one of claims 1 to 6, wherein the metal is a value metal.

8. A method according to claim 7, wherein the value metal is selected from the group consisting of light metals, base metals and precious metals.

9. A method according to claim 8, wherein the value metal is selected from the group consisting of Zn, Cu, Ti, Al, Cr, Ni_s Co, Mn, Fe, Pb, Na, K, Ca, Ag, Au and platinum group metals.

10. A method according to any one of claims 1 to 9, wherein the metal is zinc.

11. A method according to any one of claims 1 to 10, wherein all or a portion of the sulphur in the hydrogen sulphide is recovered as elemental sulphur.

12. A method according to any one of claims 1 to 9, wherein the hydrogen sulphide or a portion thereof is burned in air or oxygen to produce sulphur dioxide.

13. A method according to claim 12, wherein the sulphur dioxide is converted to sulphuric acid.

14. A method according to claim 13, wherein the sulphuric acid is recycled to the leach solution.

15. A method according to any one of claims 1 to 14, wherein impurities from the ore are removed from the leach solution prior to the electrolytic process step.

16. A method according to any one of claims 1 to 15, which is conducted at atmospheric pressure.

17. A method according to any one of claims 1 to 16, wherein the leaching step does not result in the formation of elemental sulphur.

18. A method according to any one of claims 1 to 17, wherein the leaching step can be carried out on ore which has not been finely ground.