AU2004257302B2 - A process for upgrading an ore or concentrate - Google Patents

A process for upgrading an ore or concentrate Download PDF

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Publication number
AU2004257302B2
AU2004257302B2 AU2004257302A AU2004257302A AU2004257302B2 AU 2004257302 B2 AU2004257302 B2 AU 2004257302B2 AU 2004257302 A AU2004257302 A AU 2004257302A AU 2004257302 A AU2004257302 A AU 2004257302A AU 2004257302 B2 AU2004257302 B2 AU 2004257302B2
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Australia
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stage
metal
zinc
ammonia
liquid phase
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AU2004257302A1 (en
Inventor
Sally Elizabeth Bryant
Paul Christopher Freeman
Oliver Michael Griffiths Newman
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Mmg Management Pty Ltd
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Mmg Man Pty Ltd
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Priority to PCT/AU2004/000939 priority patent/WO2005007900A1/en
Priority to AU2004257302A priority patent/AU2004257302B2/en
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    • YGENERAL TAGGING OF NEW TECHNOLOGICAL DEVELOPMENTS; GENERAL TAGGING OF CROSS-SECTIONAL TECHNOLOGIES SPANNING OVER SEVERAL SECTIONS OF THE IPC; TECHNICAL SUBJECTS COVERED BY FORMER USPC CROSS-REFERENCE ART COLLECTIONS [XRACs] AND DIGESTS
    • Y02TECHNOLOGIES OR APPLICATIONS FOR MITIGATION OR ADAPTATION AGAINST CLIMATE CHANGE
    • Y02PCLIMATE CHANGE MITIGATION TECHNOLOGIES IN THE PRODUCTION OR PROCESSING OF GOODS
    • Y02P10/00Technologies related to metal processing
    • Y02P10/20Recycling

Description

A PROCESS FOR UPGRADING AN ORE OR CONCENTRATE FIELD OF THE INVENTION The present invention relates to a 5 hydrometallurgical process for upgrading a mineral ore or concentrate to a chemical intermediate as a more concentrated source of metal. In particular, the present invention relates to a process for upgrading a mineral ore, such as although not exclusively, to zinc sulphide 10 minerals. BACKGROUND OF THE INVENTION The present invention was made to further improve the recovery of zinc in the processing of an ore body at 15 Century in Northern Queensland. Most of the zinc is recovered as a zinc concentrate containing zinc sulphide. Typically the zinc sulphide is in the mineral form of sphalerite. The dominant process for the production of zinc 20 metal from zinc sulphide concentrates is the Roast-Leach Electrowinning (RLE) process. This process is conducted in large efficient smelters that are capable of producing zinc metal of high purity. The electrowinning stage is energy-intensive and, 25 as a consequence, RLE plants are located in regions that offer low cost electrical power which is typically some distance from a remote mine site. The transport costs for transferring concentrates and other materials to the RLE plants, roasting performance considerations and the need 30 to minimise the quantities of residues generated at the RLE site all encourage the use of high-grade zinc concentrates, which are correspondingly low in impurities such as iron and silica. High-grade concentrates can be produced in most 35 zinc mines by compromising metal recovery, both at the mining and concentrating stages. In some cases, despite the rich nature of the deposits, the mineral structure of 1 the ores is such that suitable concentrate grades cannot, economically, be produced. Responses to this situation have seen the development of processes, such as the Imperial Smelting 5 Process, which are capable of processing medium/low grade concentrates (in the form of mixed lead-zinc concentrates) to metals of moderate purity. Although a relatively high cost route (requiring a sinter plant, furnace, a lead refinery and a zinc refinery), it has been a successful 10 alternative and currently represents about 10% of world smelting capacity. However, due to low metal prices, a number of these smelters have recently been closed. Processes to directly leach metal sulphide ores or concentrates have been studied extensively. An oxidative 15 acid ferric leach, for example, conveniently yields a zinc sulphate solution, from which (after solution purification) zinc can be electro-won. Acid leaching of concentrates, in pressure vessels, is practiced at two plants in Canada and ambient-pressure acid leaching has 20 been introduced at another plant in Finland. There are few mine-site hydrometallurgical plants, indicating the common difficulty in obtaining low-cost power in remote locations and the understandable reluctance to invest the capital for a smelter unless a 25 long mine life is assured. An alternative approach is to use a mine-site hydrometallurgical process to produce a zinc chemical intermediate, with just the electrowinning stage to be conducted at the second location. From a zinc sulphate 30 solution, for example, a precipitate of zinc sulphate (ZnSO 4 ) or basic zinc sulphate (3Zn(OH) 2 .ZnSO 4 ) can readily be produced. Transfer of sulphate to the electrolytic plant may, however, create a sulphate disposal problem at the smelter. 35 It is an object of the present invention to provide an alternative process for separating the valuable metal and sulphur constituents of an ore or concentrate to 2 provide the more-concentrated source of valuable metal in a non-sulphate form. SUMMARY OF THE INVENTION 5 The present invention is based on the realisation that metal sulphur compounds can be dissolved away from their host mineral ore or concentrate by using an ammoniacal solution containing ammonium carbonate (AAC solution) and then selectively precipitated to make a 10 more-concentrated source of metal which is, relatively, sulphur-free. In a situation in which the mineral ore or concentrate contains a valuable metal such as zinc in the form of sphalerite, the present invention enables the zinc 15 and sulphur constituents to be separated so that the zinc constituent can form a product that is attractive to electrolytic plants. According to the present invention there is provided a process for upgrading an ore or concentrate 20 that contains metal sulphur minerals and gangue material. The process includes the stages of: a) selectively leaching the ore or concentrate using an ammoniacal solution containing ammonium carbonate that forms soluble metal ammine complexes; 25 b) separating the solid and liquid phases formed in stage a) with the liquid phase forming a solution including soluble metal ammine complexes and the solid phase including at least in part the gangue material; c) removing ammonia and carbon dioxide from the 30 liquid phase formed in step b) under conditions that are selected to facilitate the precipitation of valuable metal(s) and minimise the sulphur content in the valuable metal(s) precipitated; and d) separating the solid and liquid phases formed in 35 stage c) with the solid phase forming a more-concentrated source of valuable metal. 3 It will be appreciated by a person skilled in the art of the present invention that stages a) to d), or any of the other stages described above may be carried out consecutively or disjunctively and may, for example, be 5 carried out at different plant sites. Depending on the operating conditions under which stages a) and c) are carried out, the solids formed may preferentially comprise metal oxides, hydroxides and carbonates. 10 An advantage provided by the present invention is that valuable metals precipitated in stage c), such as zinc, silver and copper can form a metal salt with an anion other than with a sulphur containing anion such as a sulphate. We have realized that the conditions under 15 which the valuable metal(s) are precipitated impacts on whether the valuable metal(s) can be precipitated as a salt with an anion other than a sulphur containing anion. This has major ramifications because the minimization of sulphur in the precipitate provides significant benefits 20 in downstream processes. Another advantage is that very few of the major constituents of the gangue material (notably iron and silica) are soluble in an AAC solution and, therefore, will form a major portion of the solid phase formed at 25 stage b). It is preferred that the AAC solution used in stage a) have a pH ranging from 7 to 10.5. It is preferred that stage a) be carried out at a temperature ranging from 60 to 99*C when at atmospheric 30 pressure. It is possible that stage a) may be carried out at higher temperatures and pressures. It is preferred that the method includes adding to stage a) a metal oxidant that undergoes a reduction reaction to facilitate the dissolution of the metal 35 sulphur compounds. It is preferred that the metal oxidant be in the form of a cupric cation (i.e. Cu ). This copper may be 4 all sourced from the ore itself during the leach reaction, or may be supplemented by being added in the form of a copper chemical. In a situation in which the valuable metal is 5 zinc and the material being upgraded is, for example in the form of sphalerite (ZnS), the dissolution of sphalerite may be represented by the following reaction: ZnS + 8Cu (NH 3 ) 4
CO
3 + 4H 2 0 -+ Zn (NH 3 ) 4
CO
3 + 4Cu 2
(NH
3 ) 4
CO
3 + 10 (NH 4
)
2
SO
4 + 3(NH 4 )2CO 3 + 4NH 3 Reaction A An advantage in using a divalent copper cation as the metal oxidant is that it can be regenerated using 15 oxygen by the following oxidation reaction: 2Cu 2
(NH
3
)
4
CO
3 + 02 + 4NH 3 + 2(NH 4
)
2
CO
3 -+ 4Cu(NH 3
)
4
CO
3 + 2H 2 0 Reaction B 20 Although it is possible that Reaction B occur in a separate stage, it is preferred that an oxygen containing gas be supplied to stage a) such that Reactions A and B can occur simultaneously. Indeed, a difficulty that may be encountered if oxygen is not supplied to stage 25 a) is that the copper in solution may precipitate as a copper sulphide. Although air may be used as the oxygen containing gas, it is preferred that a purified oxygen source be used as it provides a faster reaction rate and reduces heat 30 losses to the associated nitrogen gas. In addition, in order to facilitate continuous operation, an amount of make-up copper will need to be added to stage a). When oxygen is supplied simultaneously, the 35 overall reaction occurring at stage a) may be represented by the following reaction: 5 ZnS + 202 + 4NH 3 + (NH 4 ) 2
CO
3 -+ Zn (NH 3 ) 4
CO
3 + (NH 4 ) 2 SO4 Reaction C It is preferred that the concentration of copper 5 cations in the ammoniacal solution used in stage a) be at least 0.15 g/L so that the copper concentration does not limit the reaction rate. It is preferred that the ammoniacal solution in stage a) contains ammonia at a concentration that is 10 sufficient to stably maintain the metal ions, that form ammine complexes, in solution. In order to do this it is envisaged that an excess of ammonia over the stoichiometric minimum will be required. As a guide, the minimum total ammonia level (for the case of zinc with 15 copper) can be calculated by the following formulae:
[NH
3 ] ([Zn] + [Cu]) x 8) + ([SO 4 ] x 2) Formulae A where the concentrations are in mol/L. 20 As an example, in a situation in which the concentration of zinc in stage a) is 30 g/L, the concentration of ammonia (total) in the solution in stage a) should be approximately no less than 80 g/L. 25 It is also desirable that an excess over stoichiometric of dissolved carbon dioxide (or carbonate/bicarbonate) also be supplied. It is preferred that stage c) be carried out under conditions to minimise the precipitation of sulphur 30 and sulphur containing compounds. More particularly, it is preferred that stage c) be carried out at a temperature ranging from 90 0 C to boiling point so as to reduce the equilibrium levels of dissolved ammonia and carbon dioxide and thereby destabilise metal amine compounds. It is 35 preferred that steam be sparged through the liquid phase of stage c) as this not only provides an efficient source of heat but also provides a carrier gas for further 6 ammonia removal. As ammonia is removed, the metals begin to precipitate as a mixture of hydroxide-carbonate compounds substantially free of sulphur and in particular sulphate 5 compounds. This was surprising to us, as the level of sulphide in solution is about 50% higher than for zinc- in terms of mass per litre. As the reaction proceeds and the concentrations of dissolved ammonia and carbon dioxide fall (a trend readily followed by monitoring the pH), the 10 metals tend increasingly to precipitate as the basic metal sulphate. This is undesirable as it effectively down grades the attractiveness of the precipitate to the smelter. It is preferred that stage c) be carried out to an end pH of 6.8 or more to avoid excessive amounts of 15 metal sulphate forming. Those skilled in the art will appreciate that other operating parameters such as temperature and residence time will also influence the properties of the precipitate. In a situation in which the valuable metal is 20 zinc, the precipitation of zinc and the evaporation of ammonia occurring in stage c) can be represented by a reaction such as: llZn(NH 3
)
4
CO
3 + 48H 2 0 -+ 8Zn(OH) 2 .3ZnCO 3 .4H 2 04 + 8(NH4) 2
CO
3 + 25 28NH 4 0H Reaction D Although Reaction D shows a zinc hydroxide carbonate precipitate, zinc may also be precipitated in other forms including the basic carbonate and basic zinc 30 sulphate. Ammonium carbonate and ammonium hydroxide is also unstable in conditions under which stage c) is preferably carried out and may break down according to the following reactions. 35
(NH
4 ) 2
CO
3 -+ H 2 0 + CO 2 t + 2NH 3 7 Reaction E 7
NH
4 0H --+ H 2 0 + NH 3 T Reaction F In order to further increase the proportion of valuable metal in the solid phase formed in stage c), it 5 is preferred that the process includes a stage of calcining the solid phase recovered in stage d). The calcination stage involves at least part of the metal carbonates and possibly hydroxides being converted to a metal oxide. 10 It is preferred that the calcining stage be carried out by heating the solid phase formed in stage c) to a temperature of 100 0 C or more to drive off water and 300 0 C or more to decompose the carbonate. The liquid phase from stage d) contains 15 significant quantities of ammonium sulphate which can be crystallised using standard equipment to form a by-product that can be used by agricultural fertiliser manufacturers. It is preferred that the liquid phase from stage d) be treated to precipitate sulphur and compounds 20 containing sulphur from the liquid phase as a salt. An advantage provided by this preferred aspect of the invention is that additional ammonia can be recovered for reuse. It is preferred that the liquid phase from stage 25 d) be treated by adding a neutralising agent to the liquid phase. An example of a suitable neutralising agent is lime (CaO) and the sulphur-containing salt produced is calcium sulphate (i.e. gypsum). It is preferred that the neutralising agent 30 maintain the pH above 7 during the sulphate precipitation stage to minimise the level of ammonia remaining as ammonium hydroxide. It is preferred that ammonia be removed from the liquid phase in stage d) by heating the liquid phase and 35 sparging with steam. This can take place simultaneously with, or subsequent to, the treatment with lime. 8 The sulphate precipitation stage may be represented by the following reaction:
(NH
4 ) 2
SO
4 + Ca (OH) 2 -+ 2NH 3 7 + CaSO 4 I (gypsum) + 2H 2 0 5 Reaction G It is preferred that the ammonia volatilised/vapourized from either stage c) and/or the stage for precipitating the sulphate ions be recovered and reused in stage a). Standard equipment and process know 10 how-involving packed towers for ammonia and carbon dioxide recovery from vapours and distillation columns for production of a concentrated ammonia/ammonium carbonate liquid for recycling - are available, for this. The present invention also encompasses a solid 15 phase made substantially of a metal oxide and any of the other solid and liquid phases including the gypsum formed in sulphate precipitation stage made according to the process of the present invention. The present invention also encompasses a plant 20 including at least two reactor vessels for carrying out stages a) and c) and at least two solid/liquid separation devices for carrying out stages b) and d) of the process. According to the present invention there is also provided a plant for upgrading an ore or concentrate that 25 contains metal(s) sulphur minerals and gangue material, the plant including: a first stage in which an ammoniacal solution containing ammonium carbonate can selectively leach metal(s) and metal compounds from the ore or concentrate 30 to form soluble metal ammine complexes; a separator for separating the solid and liquid phases formed, in which, the liquid phase includes soluble metal ammine complexes and the solid phase includes at least in part gangue material; 35 a second stage that is supplied with the liquid phase formed in the separator and from which ammonia and carbon dioxide are removed under conditions that are 9 selected to facilitate the precipitation of valuable metal(s) and minimize the sulphur content in the valuable metal(s) precipitate; and a further separator for separating the solid and 5 liquid phases formed in the second stage whereby the solid phase forms a more-concentrate source of valuable metal(s). Suitably, the pH in the first vessel ranges from 7 to 10.5. 10 In an embodiment, the temperature in the first stage ranges from 60 to a temperature just below boiling temperature. In an embodiment, a metal oxidant is supplied to the first stage which undergoes a reduction reaction to 15 facilitate the dissolution of the metal sulphur compounds. In an embodiment, the metal oxidant can be regenerated by oxidation. In an embodiment the metal oxidant is in the form of cupric cation. 20 In an embodiment, the concentration of copper cations supplied to the first stage in the ammoniacal solution is at least 0.15 g/L. In an embodiment, the concentration of ammonia in the first stage is maintained at a level in accordance 25 with the following formulae:
[NH
3 ] > ([Zn] + [Cu]) x 8) + ([SO 4 ] x 2) In an embodiment, an oxygen containing gas, 30 suitably purified oxygen, is supplied to the first stage to regenerate the metal oxidant. In an embodiment, the second stage is carried out at a temperature ranging from 900 to boiling point so as to evaporate ammonia and thereby facilitate the precipitation 35 of metal compounds. In an embodiment steam is sparged through the liquid phase of the second stage to provide heat and a 10 carrier gas for further ammonia removal. In an embodiment, the second stage is carried out to an end pH of 6.8 or more to avoid excessive amounts of metal sulphate forming. 5 In an embodiment, the plant further includes a stage of calcining the solid phase recovered in the further separator. In an embodiment, the calcination stage is carried out by heating the solid phase formed in stage c) 10 to a temperature of at least 100*C and preferably, above 300 0 C. In an embodiment, the liquid phase from the separator stage d) is treated to precipitate sulphur and compounds containing sulphur from the liquid phase as a 15 salt. In an embodiment, the liquid phase from stage d) is treated by adding a neutralising agent to the liquid phase. In an embodiment, the neutralising agent 20 maintains the pH above 7 during the sulphate precipitation stage to minimise the level of ammonia remaining as ammonium hydroxide. In an embodiment, ammonia is removed from the liquid phase in stage d) by heating the liquid phase and 25 sparging with steam. DETAILED DESCRIPTION OF THE INVENTION A detailed description of a preferred embodiment of the present invention will now be described with 30 reference to Figure 1. The description is in the context of a zinc refining plant. However, the present invention is not confined to treating this valuable metal and is equally applicable to other valuable metals, such as copper. 35 In terms of process flow, the preferred embodiment includes an ammonia leaching stage 11 that is supplied with a zinc containing feed material such as an 11 ore or concentrate, an AAC solution and oxygen. The AAC solution and feed material form a slurry in the leaching stage 11. Once reacted in the leaching stage 11, the slurry is fed to a solid/liquid separator 12 in which the 5 liquid phase is separated from the solid phase which is largely constituted by insoluble gangue material. The liquid phase is then supplied to a zinc precipitation stage 13 in which a zinc containing solid phase is precipitated and thereby forms a slurry. The slurry is 10 then fed to another solid/liquid separator 14 in which the liquid phase is separated from the solid phase. The solid zinc containing phase is then fed to an optional calcining stage 15 to yield a product that is, substantially, zinc oxide. The liquid phase formed in separator 14 is further 15 treated in an optional sulphate precipitation stage 16 to further recover ammonia and precipitate gypsum-which is a valuable by-product in some circumstances. Ammonia and carbon dioxide are evaporated in the zinc and sulphate precipitation stages 13 and 16, and are 20 recycled back to the ammonia leaching stage 11. The operational characteristics of each stage will now be described in more detail. The ore or concentrate fed to the ammonia leaching stage 11 comprises sphalerite (ZnS) and gangue 25 material including iron and silicate minerals. An ammoniacal/ammonium stream is fed to the ammonia leach stage. If the amount of soluble copper in the ore is insufficient, a source of copper ions is also added to the 30 reactor. This can conveniently be in the form of a solution of copper sulphate in water. Copper (both Cu* and Cu 2+) will form copper ammine ions in the AAC. According to Reaction A, the cupric cations function as an oxidising agent such that the zinc 35 constituent of the feed material also forms a soluble ammine complex. There are several advantages in using copper as an oxidising agent. Firstly, it forms soluble 12 ammine complexes in a pH range of 7.0 to 10.5 and at a temperature ranging from 60 to 95*C, whereas the gangue in the feed material is substantially insoluble at these conditions. Secondly, the copper oxidising agent can be 5 conveniently regenerated using oxygen according to Reaction B set out above. The overall oxidation/reduction that dissolves sphalerite in leaching stage 11 is represented by Reaction C, set out above. 10 In some instances sphalerite may be directly oxidised by oxygen according to the following reaction: ZnS + 4NH 3 + 202 -+ Zn (NH 3
)
4
SO
4 Reaction H 15 However, it will be appreciated that the "products" formed by Reactions C and H will exist in solution as disassociated ions and ammonia carbonate will exist in solution as a mixture of bicarbonate, carbonate and free ammonia. 20 In the instance when the raw material includes zinc carbonate, it can be dissolved according to the following reaction: ZnCO 3 + 4NH 3 -+ Zn(NH 3
)
4
CO
3 Reaction I 25 Ammonia is distributed in solution between the copper and zinc ammine complexes, ammonium bicarbonate, ammonium sulphate and as hydrolyse ammonia. The amount of ammonia in solution will affect the amount of zinc and 30 copper ions that can be maintained in solution. As a guideline, the minimum ammonia level required can be estimated by the following formulae in which the concentrations of zinc, copper and sulphate are the concentrations (mol/L) present in stage 11. 35 [NH3].n = ([Zn] + [Cu]) x 8) + ([SO 4 ] x 2) 13 When the concentration of zinc present in stage a) is 30g/L, the minimum recommended concentration of NH 3 in the AAC solution is 80 g/L. The rate at which zinc is leached in stage 11 is 5 temperature dependent. A temperature of between 60 and 95 0 C has been adequate for trials conducted to date. It may be beneficial to conduct the leaching stage 11 at higher temperatures and pressures to achieve a higher reaction rate. 10 The leaching stage 11 is also dependent on sufficient oxygen being available to regenerate Cu2+ ions. In principle air could be used, but purified oxygen is preferred as it gives faster reaction rates and the heat losses will be lower. 15 If the dissolved oxygen level is not maintained during the course of the leach reaction, copper is likely to be precipitated, removing it from an active role according to the following reaction: 20 2Cu 2
(NH
3
)
4 C0 3 + ZnS -- Zn (NH 3
)
4
CO
3 + Cu 2 S4 Reaction J Any gases formed, or introduced with the oxygen, will need to be vented from the ammonia leach stage 11. As ammonia and carbon dioxide are quite volatile, there 25 will be an ammonia loss with these gases, requiring offgas treatment using condensers or water scrubbers (not illustrated in Figure 1). Once the zinc has been dissolved, and un-reacted material removed in the solid/liquid separator 12, the 30 objective is to recover the zinc. The zinc ammine complex can be broken by heating the solution to (near) boiling and sparging with steam. This drives off ammonia and carbon dioxide and precipitates zinc as the hydroxide-carbonate according to 35 Reaction D set out above. Zinc carbonates may also be present in the solid phase. 14 As can be seen from Figure 1, the ammonia and carbon dioxide are recyclable back to the leaching stage 11. Makeup AAC solution may also be fed to the leach stage 11 if needed. 5 As the ammonia is removed, the zinc will precipitate, ideally as an hydroxide-carbonate according to Reaction D. The zinc may also precipitate as a basic zinc carbonate according to the following reaction: 10 5Zn (NH 3 ) 4 C0 3 + 3H 2 0 -- + 3 (ZnO. H 2 0) .2ZnCO 3 J, + 20NH 3 7 + 3CO 2 7 Reaction K While reaction K does notTcontaminate the zinc product with sulphate ions, it does reduce the overall grade of the precipitate because the zinc content of the 15 solids in the hydroxide form is about 66%, whereas the basic zinc carbonate only contains about 60% zinc. As the pH drops with the removal of ammonia and carbon dioxide, there is a greater tendency for zinc to precipitate as a basic sulphate according to the following reaction: 20 4Zn (NH 3 ) 4
CO
3 + (NH 4 ) 2
SO
4 + 2H 2 0 -+ 3Zn (OH) 2. ZnSO4 + 4CO2 + 18NH 3 7 Reaction L 25 The selected end point for the precipitation reaction in stage 13 is a trade-off between maximising the zinc precipitation and minimizing sulphate contamination of the precipitate. Alternatively, zinc can be further encouraged to precipitate in the hydroxide form by 30 addition of an alkali (e.g. caustic soda) that maintains the pH at a suitable, higher value. The slurry formed in the zinc precipitation stage 13 is then fed to a solid/liquid separator 14 and the solid phase containing the zinc constituents is fed to the 35 calcining stage 15. 15 The calcining stage 15 essentially converts the zinc hydroxide-carbonates to zinc oxide. This will reduce the mass to be transported to the electrowinning refinery and minimise contamination of the product with ammonia. 5 The calcining stage 15 is carried out by heating the precipitate to above 300*C. The liquid phase from stage 14 contains significant quantities of ammonium sulphate which can be crystallised using standard equipment to form a by-product 10 that can be used by agricultural fertiliser manufacturers. Alternatively the ammonia can be recovered. This is achieved by reacting the liquid phase in stage 16 with a reagent such as lime or limestone to form gypsum, which precipitates. Boiling and/or steam sparging the liquid is 15 used simultaneously with, or subsequent to, the treatment with lime to volatilise the dissolved ammonia. If not valued as a by-product, the resulting gypsum slurry in stage 16 may conveniently be fed directly to a tailings dam at a mine site. 20 The ammonia and carbon dioxide evaporated in stages 13 and 16 can be recovered and reused in stage 11. Standard equipment and process know-how-involving packed towers for ammonia and carbon dioxide recovery from vapours and distillation columns for production of a 25 concentrated ammonia/ammonium carbonate liquid for recycling-are available, for this. Set out below is a description of a trial carried out according to the preferred embodiment of the present invention. 30 Example 1: Ammonia leach An AAC leaching stage was conducted in a 3L reactor at 85 0 C for 5 hours, with oxygen sparging at 600 ml/min. The starting material was 200g of a low-grade 35 concentrate containing 15% Zn, in the form of sphalerite, slurred with water to a pulp density of 200g dry solids/litre solution. After heating to 85*C, 400g of 16 ammonium hydrogen carbonate was then added together with 250 ml of a 25 wt% ammonia solution. Cupric ions were added in the form of copper sulphate (3g in 30ml of water) and the reaction commenced. The pH was controlled during 5 the test at 8.7 by automatic additions of the ammonia solution. At the conclusion of the test the slurry was filtered, washed and analysed. The filtrate is feed for the zinc precipitation stage and the solid is waste gangue material. 10 Results of the analysis of the filtrate provided an assay as set out below. The zinc extraction-was 91.4% after 5 hours. Zinc in the form of zinc silicate was not extracted from the solid phase. There was extraction of other elements (i.e. 15 lead, manganese) but they are not stable in solution and precipitated (probably as carbonates) and are disposed of in the gangue. Cadmium and copper (in the feed material) are extracted and are stable in solution. 20 Table 1: Ammonia leach assays Time (hrs) Zn Cu SiO2 Ca NH3 S04 Tot S Solutions (g/L)_________ 0 0 0.28 0.003 0.009 50.3 4.5 1.8 0.5 8.1 0.13 0.002 0.012 47.6 6.2 5.;n 1 11.8 0.37 0.001 0.006 35.5 13.3 6.5 2 15.9 0.46 0.002 0.007 29.1 23.5 8.3 - 3 14.8 0.44 0.001 0.006 34.3 23.0 7.7 4 14.7 0.43 0.001 0.008 28.3 23.8 8.0 5 14.3 0.41 0.001 0.005 31.9 22.6 7.6 Solids (%) Zn Cu Pb SiO2 Ca S04 Tot S 0 15.1 0.10 0.53 48.4 0.49 0.4 9. 0.5 9.7 0.30 0.43 51.1 0.57 <0.1 7.2 1 6.6 0.13 0.45 55.5 0.66 0.3 5.5 2 2.5 0.03 0.56 57.3 0.68 0.2 3.2 3 2.0 0.03 0.56 56.3 0.64 <0.1 2.8 4 1.7 0.03 0.55 56.8 0.62 <0.1 2.5 5 1.4 0.02 0.57 56.0 0.64 <0.1 2.3 17 The solid residue containing gangue material was wash tested. The concentration of ammonia before washing was approximately 0.1% and <0.1% after three washes. This demonstrates that ammonia can be effectively recovered by 5 washing the residue. Example 2: Zinc precipitation The solution from the ammonia leach stage was heated to about 95 0 C and sparged with oxygen 10 (experimentally, a convenient carrier gas) at 400 ml/min for 3.5 hours. Over this time, a precipitate formed and the pH dropped from 8.8 to-6-.8. In a series of experiments, the reaction was halted at different final pH levels and the resulting precipitates were filtered and 15 analysed. The analysis provided the following assays. Table 2: Zinc precipitation assays Zn Cu Si02 Ca NH 3
SO
4 Tot S Solutions (g/L) T=0 (pH 14.7 0.43 <0.00 0.01 31.9 22.6 7.6 8.8) 1 PH 8.0 5.8 0.41 <0.00 0.01 13.7 22.8 7.7 1 PH 7.5 1.7 0.38 <0.00 0.01 8.8 21.9 7.3 1 PH 6.8 0.46 0.33 <0.00 0.02 7.3 21.2 7.0 Solids (%) Zn Cu Pb S102 Ca NH 3
SO
4 Tot S COQ PH 8.0 58.5 0.31 0.25 0.73 0.24 2.7 0.9 20.5 PH 7.5 58.1 0.42 0.05 0.51 0.13 0.9 4.7 1.6 19.0 PH 6.8 57.6 1.3 0.08 0.19 0.11 0.8 6.2 1.9 15.0 The purity of the zinc product can be improved by stopping the reaction at a higher pH at the expense of 45 zinc recovery as shown below. There will be an economic trade-off between these two factors. 18 Final Zinc in Zinc Zinc Basic Basic pH product hydroxide- recovery zinc copper carbonate sulphate carbonate 5 8.0 58.5% 87.8% 59.4% 7.5 58.1% 87.2% 88.1% 6.1% 0.7% 6.8 57.6% 84.8% 96.8% 7.4% 2.3% In the instance when the zinc precipitation stage was stopped at a pH of 6.8, the solid assay comprised approximately 85% zinc hydroxide-carbonate (8Zn(OH) 2 15 3ZnCO 3 ), 7% basic zinc sulphate and 2.3% basic copper carbonate. Therefore a total of 96.8% of the zinc in the liquid phase fed to the zinc precipitation (stage 2) was precipitated. In the instance when the zinc precipitation stage 20 was stopped at a pH of 7.5, the solid assay comprised approximately 87% zinc hydroxide-carbonate (8Zn(OH) 2 3ZnCO 3 ), 6% basic zinc sulphate and 0.7% basic copper carbonate. Therefore, in this instance a total of approximately 88.1% of the zinc in the feed to stage 2 was 25 precipitated. Copper precipitation commences after zinc, at approximately pH 7.5. Lead and silica appear to precipitate relatively quickly and therefore their solids assays declines subsequently over the course of the 30 experiment. Example 3: Sulphate precipitation The solution from Example 2 was again heated to about 95*C and sparged with oxygen for 2 hours. Lime was 35 added as a 500 g/L slurry to maintain the pH at approximately 7.0. Over this time, a precipitate formed and analysis of timed samples collected (Table 3) indicates that the precipitate contained a mixture of calcium carbonate and calcium sulphate. The final liquor 40 contained very low levels of zinc, copper and ammonia. 19 Table 3: Gypsum precipitation assays Time (hrs) Zn Cu Ca NH 3
SO
4 Solutions 0 0.34 0.054 0.02 7.1 19.6 0.25 0.42 0.054 0.50 5.1 14.6 1 0.08 0.026 0.48 2.8 8.5 1.5 0.08 0.018 0.45 2.3 7.2 3 0.01 0.003 0.48 1.0 3.4 3.5 <0.01 0.002 0.53 0.6 1.4 Solids 0 - - 0.25 0.1 0.01 36.2 1.8 1 0.66 0.02 -5:.6 5.2 1.5 0.27 0.01 34.9 8.42, 3 0.41 0.05 33.4 11.1 3.5 0.49 0.06 31.6 <0.1 18.4 The majority of the precipitate contain calcium compounds, 60% calcium carbonate and 26% gypsum (calcium sulphate). Approximately 85% of the sulphate was 35 precipitated, and 92% of the ammonia was volatilised from the solution. Example 4: Calcination Using a muffle furnace, 10 gram samples of the 40 precipitated zinc product were heated between 200 0 C and 500 0 C, at 100 0 C intervals, for a minimum of two hours. The results are presented below in Table 4. 45 50 20 Table 4: Calcination Results Zn NH 3
CO
3
SO
4 Cu Pb SiO 2 Cl F Sample 1 Untreated 54.4 2.0 12.1 0.64 0.08 .60 <0.01 <0.02 200 *C 54.9 1.2 7.19 11.7 0.73 0.1 0.77 <0.01 <0.01 300 *C 66.0 0.5 1.00 12.0 0.81 0.11 0.75 <0.01 <0.01 400 *C 66.0 <0.1 0.25 12.8 0.84 0.12 0.78 <0.01 <0.01 500 *C 67.8 <0.1 0.15 13.3 0.82 0.14 0.67 <0.01 <0.01 Sample 2 Untreated 52.2 1.5 14.4 1.1 0.12 0.28 <0.01 <0.01 200 *C 55.9 1.1 3.85 15.8 1.2 0.12 0.26 <0.01 <0.01 300 *C 62.8 0.5 0,75 17.5 1.4 0.09 0.31 <0.01 <0.01 400 *C 64.5 <0.1 0.35 17.9 1.4 0.04 0.27 <0.01 <0.01 500 *C 65.1 <0.1 0.15 18.0 1.4 0.05 0.29 <0.01 <0.01 At 300 0 C, the zinc content of the product had increased by 10% to 63-65% zinc. The ammonia concentration 5 had decreased from 2.0% to 0.5% at 300 0 C, and to less than 0.1% at 400 0 C. This is equivalent to 82% (sample 1) and 71% (sample 2) zinc hydroxide, with minimal amounts of zinc carbonate present. There was approximately 18-24% basic 10 zinc sulphate in the product. Calcining the product at 300 0 C increased the zinc concentration by removal of carbonate to less than 1%. After calcining the product at 400 0 C, the ammonia in the 15 product was decreased to below its detection limit. This minimises ammonia release upon dissolution of the zinc product. Calcining the product at approximately 400 0 C resulted in increased zinc concentration and complete ammonia removal. Therefore, treating the precipitated 20 product results in reducing the amount of final product to be transported and the Occupational Health and Safety issues associated with ammonia release upon dissolving the product in a hydrometallurgical circuit. 21 It will be appreciated by those skilled in the art of the present invention that modifications may be made to the preferred embodiment of the invention without departing from the spirit and scope of the invention. 22

Claims (20)

1. A process for upgrading an ore or concentrate that 5 contains metal sulphur minerals and gangue material, the process including the stages of: a) selectively leaching the ore or concentrate using an ammoniacal solution containing ammonium carbonate that forms soluble metal ammine complexes; 10 b) separating the solid and liquid phases formed in stage a) with the liquid phase forming a solution including soluble metal ammine complexes and the solid phase including at least in part the gangue material; c) removing ammonia and carbon dioxide from the 15 liquid phase formed in step b) under conditions that are selected to facilitate the precipitation of valuable metal(s) and minimise the sulphur content in the valuable metal(s) precipitated; and d) separating the solid and liquid phases formed in 20 stage c) with the solid phase forming a more-concentrated source of valuable metal.
2. The process according to claim 1, wherein stage a) is carried out at a pH ranging from 7 to 10.5. 25
3. The process according to claim 1 or 2, wherein stage a) is carried out at a temperature ranging from 60 to a temperature just below boiling point. 30
4. The process according to any one of claims 1 to 3, wherein the process includes adding to stage a) a metal oxidant that undergoes a reduction reaction to facilitate the dissolution of the metal sulphur compounds. 35
5. The process according to claim 4, wherein the metal oxidant can be regenerated by oxidation and is in the form of a cupric cation. 23
6. The process according to claim 5, wherein the concentration of copper cations supplied to stage a) in the ammoniacal solution is at least 0.15 g/L. 5
7. The process according to claim 6, whereby when the metal is zinc and the ore contains sphalerite (ZnS), leaching of sphalerite may be represented by the following reaction: 10 ZnS + 8Cu (NH 3 ) 4 CO 3 + 4H 2 0 - Zn (NH 3 ) 4 CO 3 + 4Cu 2 (NH 3 ) 4 CO- \.+ (NH 4 ) 2SO4 + 3 (NH 4 ) 2 CO 3 + 4NH 3 .
8. The process according to claim 7, wherein the 15 process includes maintaining the concentration of ammonia in stage a) at a level in accordance with the following formulae: [NH 3 ] ([Zn] + [Cu]) x 8) + ([S04] x 2) 20
9. The process according to any one of claims 5 to 8, wherein cupric copper is regenerated by oxidation with oxygen according to the following reaction: 25 2Cu 2 (NH 3 ) 4 C0 3 + 02 + 4NH 3 + 2 (NH 4 ) 2 CO 3 ) 4Cu (NH 3 ) 4 CO 3 + 2H 2 0
10. The process according to any one of claims 1 to 9, wherein an oxygen containing gas is supplied to stage a). 30
11. The process according to any one of claims 1 to 10, wherein stage c) is carried out at a temperature ranging from 90 0 C to boiling point so as to evaporate ammonia and thereby facilitate the precipitation of metal compounds. 35
12. The process according to claim 11, wherein stage c) includes sparging the liquid phase with steam so as to 24 regulate temperature and provide a carrier gas for further ammonia removal.
13. The process according to any one of claims 1 to 5 12, wherein stage c) is carried out to an end pH of 6.8 or more to minimise the precipitation of metal sulphate minerals.
14. The process according to any one of claims 11 to 10 13, whereby when the metal is zinc the precipitation of zinc and the evaporation of ammonia occurring in stage c) can be represented by the following reaction: llZn(NH 3 ) 4 CO 3 + 48H 2 0 -> 8Zn(OH) 2 .3ZnCO 3 .4H 2 0 + 15 8 (NH4) 2 CO3 + 28NH 4 0H
15. The process according to any one of claims 1 to 14, further including a stage of calcining the solid phase recovered in stage d). 20
16. The process according to claim 15, wherein the calcination stage is carried out by heating the solid phase formed in stage c) to a temperature ranging from 100*C to 500*C. 25
17. The process according to any one of claims 1 to 16, wherein the liquid phase recovered from stage d) is treated to precipitate sulphur and compounds containing sulphur from the liquid phase as a salt. 30
18. The process according to claim 17, wherein a neutralising agent is added to the liquid phase of stage d). 35
19. The process according to claim 18, wherein the neutralising agent maintains the pH above 7 during the 25 sulphate precipitation stage to minimise the level of ammonia remaining as ammonium hydroxide.
20. A plant for upgrading an ore or concentrate that 5 contains metal(s) sulphur minerals and gangue material, the plant including: a first stage in which an ammoniacal solution containing ammonium carbonate can selectively leach metal(s) and metal compounds from the ore or concentrate 10 to form soluble metal ammine complexes; a separator for separating the solid and liquid phases.formed, in which, the liquid phase includes soluble metal ammine complexes and the solid phase includes at least in part gangue material; 15 a second stage that is supplied with the liquid phase formed in the separator and from which ammonia and carbon dioxide are removed under conditions that are selected to facilitate the precipitation of valuable metal(s) and minimize the sulphur content in the valuable 20 metal(s) precipitate; and a further separator for separating the solid and liquid phases formed in the second stage whereby the solid phase forms a more-concentrate source of valuable metal(s). 25 26
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