Recovering metal values from complex concentrates

Abstract

Complex materials which contain a mix of metals are processed to recover valuable metals present where the materials are roasted under conditions in which the metals form compounds that can be dissolved in dilute acid and / or water and / or ammonia containing alkaline solutions from which the metals can be recovered as marketable products while leaving unwanted impurity elements such as arsenic and radionuclides in an inert residue for disposal.

Description

Recovering Metal Values from Complex ConcentratesField of the Invention

[0001] This invention relates to a novel process for treating complex materials to recover valuable metals contained in marketable products where the valuable metals present are commonly copper and / or nickel and / or cobalt and / or molybdenum and / or lithium and / or one or more rare earth metals.Background of the Invention.

[0002] Many ores contain significant concentrations of valuable metals in addition to the primary metal being recovered. This is especially true of some of the copper and nickel sulphide orebodies. These are usually treated to recover the main metals but there is little recovery and / or no payment for other contained metals.

[0003] In addition to the primary metal many ores contain impurities such as arsenic and / or radionuclides which are difficult to separate from the target metals and report as impurities in concentrates that are produced and can be of sufficient concentration to hinder or prevent processing of the concentrates through to metal.

[0004] Arsenic is a major problem in the non-ferrous metals industry as many of the ores contain significant concentrations of arsenic which typically reports into concentrates that are produced for further treatment. The amount of arsenic in these materials far exceeds the global market for arsenic and therefore the arsenic must be disposed of safely such that it does not cause contamination of the air, soil or water at either the mine or the processing plant.

[0005] The current approach is to form arsenic compounds such as ferric arsenate and / or calcium arsenate and dispose of these into a suitable repository such as controlled landfill and / or on conjunction with other residues and tailings. The formation of these compounds is quite complex and typically adds extra processing steps and costs to the metal processing.

[0006] The arsenic typically reports in the ore as mixed metal sulphides such as enargite and / or tennantite which have proven very difficult to separate from other more valuable minerals such as chalcopyrite and bornite. There has been extensive research on separating the minerals physically but these have been only moderately successful with the most common outcome being processingthe ore to give a high value low arsenic concentrate which can readily be sold and a much lower value high arsenic concentrate which is very difficult to sell unless treated within the company’s own operations.

[0007] Radionuclides present a similar problem to arsenic in that they are very difficult to separate from the target metals by physical processing. In some cases the concentrations are high enough to warrant their recovery by adding hydrometallurgical processing but in selected cases the concentrations are too low to justify recovery but are high enough to be a barrier to selling the concentrate to an external smelter.

[0008] Many of the large copper producers have managed the issues with arsenic and / or radionuclides by building their own smelters in the same country as the mines to avoid the need to transport the concentrate internationally and avoid cost penalties from external processing. This approach is less satisfactory for miners with smaller deposits which do are not capable of supporting the CAPEX of building a new smelter.

[0009] Secondary materials such as waste Li-ion batteries and / or e-waste and / or fabrication waste are also difficult to recycle and current processes are expensive and can leave difficult to dispose of residues.

[0010] The industry has been actively searching for alternative processes capable of economically treating complex concentrates and / or those with problem impurities and / or waste materials for many years. This work has had limited success with the technology proving difficult and the need to develop a technology suitable for building a plant suitable for small to moderate tonnages with acceptable CAPEX and OPEX costs.Description of the Invention

[0011] This invention uses a novel sulphate roasting step to convert metals present in the target materials to compounds that can be processed hydrometallurgically to enable their recovery and / or conversion to inert compounds for disposal. The roasting step is suitable for metals present as sulphides and / or in minerals such as monazite and / or bastnasite and / or as carbonates and / or in metallic alloys. The roasting is carried out at temperatures between 400C and 750C with a controlled atmosphere of oxygen and sulphur dioxide. The atmosphere is controlled by balancing the amount of oxygen present to the amount of sulphur in the feed and where needed by theaddition of a calcium source most commonly lime and / or limestone to limit the sulphur dioxide and sulphur trioxide content of the gas. The sulphur dioxide content of the gas in the roaster is preferably keptbelow 3% and more preferably is less than 1% by controlling the ratio of the metals to sulphur in the feed and most commonly by adding calcium as calcium oxide and / or as calcium hydroxide and / or as calcium carbonate. .Where calcium is added the calcium reacts with part or all of the sulphur dioxide and / or sulphur trioxide resulting from the oxidation of the sulphides to control their partial pressure partial pressure and air addition is used to maintain the desired partial pressure of oxygen.

[0012] The critical reactions involved in the roasting step are:Copper Sulphides copper sulphate - copper oxide but no copper ferrite. Molybdenum Sulphide Molybdenum Trioxide Cobalt SulphideCobalt SulphateCopper arsenic sulphides Copper arsenateIron Sulphides Iron Oxide (hematite)Iron arsenic sulphidesIron arsenateNickel SulphidesNickel SulphateREE’sSulphates with limited oxide.Manganese oxide Manganese sulphate Lithium carbonates / metal Lithium sulphateCaOCalcium Sulphate

[0013] The inventors have found that by controlling the roasting step to match the composition of the feed material and achieve the desired reactions the calcine from the roast can be processed hydrometallurgically to recover the valuable metals in forms suitable for sale. This is most commonly used for copper sulphide concentrates but is also applicable to cocnentrates where nickel and / or cobalt are the most valuable metals present.

[0014] For copper ores containing significant amounts of enargite and / or tennanite the process is operated such that almost all of the copper is recovered whilst fixing the arsenic as an inert iron arsenate. In one embodiment the ore is processed using flotation separation to give a copper concentrate with low arsenic levels suitable for sale to a smelter, and a copper concentrate with high arsenic levels which is treated using this novel process. For this concentrate the roasting stage is carried out with added lime at a temperature in the range from 400C to 700C and more preferably in the range from 550C to 650C. The lime is added such that the molar ratio of Ca / S in the feed materials is >0.3 and more preferably >0.5. Sufficient oxygen is added to the roaster to oxidise the sulphur present to sulphate and the added lime and oxygen are controlled to maintain an oxidizing atmosphere such that the majority of the arsenic present does not volatilize but reacts with copper present in the minerals to form copper arsenate.

[0015] In the embodiment above the calcine typically contains a mix of copper arsenate, copper sulphate, copper oxide and iron oxide plus unreacted gangue oxides and silicates. A part of the arsenic may also be present as iron arsenate and / or calcium arsenate. The calcine is leached in an ammonia ammonium chloride liquor typically containing from 20-100g / l NH3 and from 20- 100g / l NH4C1 to dissolve the copper present as copper oxide and / or copper sulphate whilst leaving the arsenate compounds in the residue. The copper is then precipitated from the pregnant leach liquor by evaporating off some of the ammonia and / or by cooling and / or by adding an alkali such as lime.

[0016] The arsenic containing residue may contain valuable copper if part of the arsenic is present as copper arsenate and if so the residue is further processed to recover the copper and fix the arsenic as iron arsenate. The residue is leached in dilute sulphuric acid typically having from 5-50g / l H2SO4 which dissolves substantially all the copper and arsenic present and also dissolves any arsenic present as calcium arsenate. Some of the iron present also dissolves and then an alkali such as lime is added to the pregnant liquor to raise the PH to around 2 to precipitate the arsenic present as an iron arsenate. If there is insufficient iron leached from the residue to fully precipitate the arsenic in the desired form additional iron can be added as ferrous sulphate and / or ferric sulphate in either solid or liquid form such that the iron to arsenic molar ratio in the liquor is greater than 1.0.

[0017] The iron arsenate precipitate is separated from the liquor and the liquor is then further processed to recover the copper from solution. This can be done using conventional technology such as solvent extraction - electrowinning but is more is preferably done by adding more alkali to raise the PH to around 6 and precipitate the copper. The precipitate is then separated and in one embodiment is recycled within the process most commonly such that the copper is recovered through the ammonia - ammonium leaching stage either directly or by adding to the roaster. An alternative in some cases is to add the copper precipitate from this step to the low arsenic concentrate to increase the grade and allow the copper to be recovered by sale to a smelter. A further alternative is to use this precipitate as feed to produce copper chemicals such as copper sulphate which can be done through additional process steps and / or by sale to an external chemical company.

[0018] The iron arsenate precipitate can be disposed of as an inert compound but in some cases will be recycled to the roaster to improve the properties and enhance the stability when beingdisposed of. Ferric arsenate in calcine does not dissolve in either the ammonia - ammonium chloride or the dilute sulphuric acid and hence the recycle does not affect any of the process steps.

[0019] In another embodiment complex ores containing economic levels of copper, nickel, cobalt, molybdenum and rare earth elements are processed to give a mixed metal concentrate which is fed to the roaster. The feed commonly has sufficient sulphur present to enable the formation of the desired sulphates but if necessary the sulphur level can be increased by the addition of high sulphur material such as pyrite and / or conventional copper concentrate. This feed material is roasted at a temperature of between 400C and 700C with sufficient air to convert the sulphides to sulphates and if necessary lime is added to react with part of the sulphur to form calcium sulphate to both control the oxygen potential in the roaster and to minimize the amount of sulphur dioxide being given off in the flue gas.

[0020] In this embodiment of the invention where multiple metals are present the calcine from the roaster is leached at a PH in the range of 6-8 (a neutral leach) to dissolve soluble metal sulphates such as cobalt sulphate, nickel sulphate and REE sulphates such a cerium sulphate. Copper present as oxide and / or sulphate and / or arsenate is not soluble at this PH and does not leach. Molybdenum and iron oxides also do not leach under these conditions.

[0021] In one embodiment where molybdenum is present at economic levels the solid residue from the neutral leach is then leached with caustic soda which dissolves the molybdenum present but leaves the copper and iron unreacted in the solid residue. The leach liquor is separated from the solids using conventional technology, most commonly filtration, and further processed to give a suitable molybdenum product such as molybdenum trioxide and / or molybdenum metal.

[0022] In an alternative embodiment the calcine is leached with caustic soda prior to the neutral leach to dissolve the molybdenum and where present lithium and this leach liquor is separated and further processed to give separate molybdenum and lithium products.

[0023] The copper containing solid residue is then leached in an ammonia - ammonium chloride liquor using the technology described in PCT / AU2022 / 050188 to produce copper hydroxy chloride and / or copper oxide and / or copper metal and to recover precious metals such as gold and silver. If the residue contains significant levels of arsenic as copper arsenate then following theammonia - ammonium chloride leach a dilute acid leach followed by iron arsenate precipitation is included as discussed previously.

[0024] The leach liquor from the neutral leach containing nickel, cobalt and rare earth metal is then processed through one or more solvent extraction stages to separate the metals and give marketable products. Most commonly the cobalt and nickel are removed first using commercial reagents such a Cyanex-272 and / or Versatic 10 and stripped into acid liquors for subsequent recovery.

[0025] The raffinate from the cobalt and nickel extraction is then processed through several solvent extraction stages to yield individual rare earth oxide products. This technology is well established commercially so is not a novel part of this invention and is not described further.

[0026] In another embodiment used where the copper content of the concentrate is low and / or the tonnage of feed materials is low a simpler flowsheet can be used. In this embodiment the roasting stage is used to convert the metals to their leachable forms and the calcine is first subject to a caustic leach to dissolve the molybdenum for recovery. The solid residue is then leached in an alkaline ammonia liquor most commonly ammonia ammonium chloride but other ammonium salts such as ammonium sulphate or ammonium carbonate can be used. The conditions for this leach are set to dissolve all of the target metals leaving unwanted gangue, including iron oxides, for disposal. In this embodiment sulphide precipitation using a sulphur source such as H2S gas or NaSH or Na2S is used to give a mixed metal sulphide precipitate which has sufficient contained value to be sold to, and / or custom smelted, in existing smelters. The H2S required can be purchased externally or more preferably is generated within the process using one or more anaerobic bioreactors where sulphate present in the liquor reduces to form gaseous H2S.

[0027] The liquor from the sulphide precipitation stage is then passed to a rare earth recovery section where solvent extraction can be used to recover the individual metals. An alternative is to precipitate the rare earth metals as a mixed oxide by the addition of oxalic acid or a suitable alternative and to sell the mixed oxide.

[0028] Sulphation roasting can also be used for materials such a black mass from batteries and / or electronic waste and / or waste from fabrication of finished products even though these contain little or no sulphur by combining the materials with a suitable sulphur source. The sulphur source preferably contains valuable metals that can also be recovered. The preferred sulphur source is ametal sulphide concentrate containing recoverable levels of one or more of copper, nickel cobalt or molybdenum.

[0029] The sulphation roasting of the blended material converts all the metal components into compounds that can be recovered by leaching and while utilizing nonmetallic components as fuels to minimize the amount of solid residue which requires disposal. The flowsheet of the hydrometallurgical process used for the calcine varies depending upon the metals present. The simplest process is to leach the lithium and / or molybdenum and / or manganese at high PH using caustic soda and / or aqueous ammonia and then filter off the leach liquor for recovering the molybdenum and / or lithium and / or manganese then leaching the residue in ammonia-ammonium chloride to dissolve copper, nickel and cobalt and recovering these as a mixed metal sulphide by precipitation with H2S. The lithium and molybdenum are recovered from the high PH liquor most commonly by selective precipitation through neutralization and / or carbonate precipitation.

[0031] For calcine with high concentrations of copper it can be advantageous to selectively leach the nickel and cobalt present as sulphate in the calcine at a PH around 6 to separate them from the copper leaving a residue suitable for leaching in ammonia - ammonium chloride and recovery of copper and any gold and silver present.

[0032] In some situations the calcine may best be leached in dilute sulphuric acid at PH where the copper sulphate, nickel sulphate and cobalt sulphate are soluble but the iron in the calcine if present as ferric iron such as in hematite does not dissolve. The preferred PH is greater than PH of 2 and more preferably greater than 2.3. The liquor from this leach can then be processed to recover the metals by precipitation to raise the PH to around PH 8 either in a single step or in stages. The preferred chemical for increasing the PH is magnesium carbonate as the magnesium reacts to form magnesium sulphate which is soluble in the liquor and does not contaminate the metal precipitates.

[0033] The magnesium sulphate liquor can then be further processed using bioreactors to recover the magnesium carbonate and sulphuric acid for reuse. The preferred system is to first use an anaerobic bioreactor to convert the sulphate to elemental sulphur and / or hydrogen sulphide gas and / or metal sulphides and precipitate the magnesium as magnesium carbonate and then treat the hydrogen sulphide gas and if warranted any elemental sulphur or metal sulphides in an aerobic bioreactor to oxiidise them to sulphates, mainly as sulphuric acid, which can be returned to the leach.

[0034] The use of magnesium carbonate to precipitate the metals as oxide / hydroxides from dilute sulphuric acid liquor with subsequent regeneration of the magnesium carbonate and sulphuric acid using bioreactors can also be used in other situations such as for treating liquor from heap leaching, ROM dump leaching and acid main drainage.

[0035] In another embodiment where the concentrate is primarily a nickel sulphide concentrate which contains significant arsenic making it unattractive to conventional smelters the concentrate is blended with material containing elements that can readily form arsenates such as copper and / or iron and / or calcium such that the arsenic present forms the arsenate and remains in the calcine while the nickel sulphide oxidises to a leachable form preferably nickel sulphate and / or nickel oxide. The calcine is then processed using a mix of neutral, alkaline and acid leaching as described for the complex concentrate. The preferred materials added to provide the arsenate forming metals are copper concentrate and / or pyrite concentrate as these also provide sulphur to assist in nickel sulphate formation and energy for use in the process plant.Description of the Drawings

[0036] Figure 1 shows the flowsheet as described in embodiment one where the complex sulphide concentrate is sulphation roasted at 600C in an air - sulphur dioxide atmosphere to convert the metals present to oxides an / or sulphates. The calcine is then leached at PH around 7 in water and the slurry then filtered to give solids stream and a liquid stream. The solids stream is then leached with caustic soda to dissolve the molybdenum present, and the slurry again filtered to give a copper rich residue and a sodium molybdate containing liquor. The liquor is further processed to give molybdenum trioxide and / or molybdenum metal and / or other marketable molybdenum chemicals. The copper rich residue is leached with ammonia-ammonium chloride liquor and the slurry filtered to leave a residue for disposal and / or processing to recover any PGM’s present and a copper rich liquor from which copper containing crystals are precipitated by stripping off part of the ammonia. The liquor from the initial PH 7 leach passes to multiple solvent extraction stages to separate out the individual metals present. The cobalt and nickel are recovered first with the raffinate form this passing to staged solvent extraction to give individual rare earth oxide products.

[0037] Figure 2 shows the flowsheet as described in embodiment two where the complex sulphide concentrate is sulphation roasted at 600C in an air - sulphur dioxide atmosphere to convert the metals present to oxides an / or sulphates. The calcine is then leached in caustic soda and the slurry then filtered to give solids stream and a liquid stream. The solids stream is then leached withcaustic soda to dissolve the molybdenum present, and the slurry filtered to give a sodium molybdate containing liquor and a solid residue. The liquor is further processed to give molybdenum trioxide and / or molybdenum metal and / or other marketable molybdenum chemicals. The residue is leached with ammonia-ammonium chloride liquor at moderate PH and the slurry filtered to leave a residue for disposal and / or processing to recover any PGM’ s present and a metal rich liquor, from which copper containing crystals are precipitated by stripping off part of the ammonia. The copper and / or cobalt and / or nickel in the liquor are precipitated as a high grade mixed metal sulphide which is filtered off and sold. The liquor from the sulphide precipitation stage is then treated with oxalic acid or an alternative reactive chemical to precipitate the rare earth metals present as a high grade mixed oxide for sale to a rare earth’s processing plant.

[0038] Figure 3 shows the process for recovering copper from an arsenic containing copper sulphide ore. In this version of the process the copper ore is fed directly to a conventional concentrator where the sulphide minerals are separated from the gangue using a comminution - flotation circuit. An additional flotation stage is included where the sulphide minerals present are further separated into a low arsenic copper concentrate suitable for sale to a conventional smelter and a high arsenic concentrate which requires further processing. This high arsenic copper concentrate is fed to a roaster where a calcium source such as calcium oxide or more preferably calcium carbonate is added mixed with the sulphide concentrate such that as the material is roasted in an oxygen containing gas at -575C. The sulphur present oxidises to sulphates and is retained in the calcine predominantly as a mix of calcium sulphate and copper sulphate. This removes the need for an acid plant. The arsenic present is oxidized during the roast and predominantly reacts with copper in the concentrate to form copper arsenate and is retained in the calcine. The gas from the roaster is scrubbed with lime to remove any sulphur containing gases and also capture any arsenic and other metals present. The solids for the scrubber are then returned to the roaster to make use of excess lime and for any arsenic present to react with iron to convert it to ferric arsenate. The roasted material is then fed to an ammonia - ammonium chloride leach stage to dissolve the majority of the copper away from the arsenic compounds and gangue material in the residue. The slurry is then filtered to give a copper depleted solid residue and a copper rich pregnant liquor. The residue is washed using conventional means and then fed to a dilute sulphuric acid leach to dissolve the arsenic and copper present Some ferric sulphate is added to the leach to ensure there is sufficient iron present in the liquor to enable the precipitation of ferric arsenate in a later stage. The slurry from the leach is filtered and the residue washed before being fed to a conventional gold circuit where gold and silver are recovered using cyanide technology. After recovery of the gold and silver the residue is disposedof into a suitable facility such as a tailings dam. The acidic liquor containing copper, arsenic and iron is partially neutralized using lime to raise the PH to ~2 and precipitate the majority of the arsenic present as ferric arsenate. The slurry is filtered to separate the iron arsenate which is disposed of into a suitable facility such as a lined tailings dam. This leaves a high copper liquor which is then further neutralized to precipitate the copper as a copper hydroxide / oxide. This copper precipitate is filtered off and washed and returned to the ammonia - ammonium chloride leach to recover the copper. The copper rich pregnant liquor is fed to an ammonia stripping - crystallization step where the ammonia is stripped off by heating the liquor to boiling point. The pregnant liquor continues to boil as it is heated and this drives off ammonia gas in an ammonia - water gas. As the ammonia concentration drops the copper solubility decreases and copper containing crystals which contain some copper hydroxy chloride precipitate. The crystal containing liquor is then cooled to 25C to precipitate more crystals and filtered to separate the crystal from the spent liquor. The ammonia - water vapour gas passes into an absorption column where it is contacted with the cooled spent liquor from the crystallisation stage and the ammonia and water are recaptured for recycle to the leaching stage. The crystals are filtered and washed and hydrolysed by mixing with a dilute ammonium hydroxide solution to hydrolyse the copper hydroxy chloride present to cupric oxide. The dilute ammonia solution also dissolves traces of unwanted impurities that are precipitated with, or more generally entrained, in the crystals. The pure cupric oxide crystals are then filtered of, washed in pure hot water and passed to the reduction stage. In the reduction stage the crystals pass through a rotary kiln held at a temperature of -250C with hydrogen gas continually fed into the rotary kiln to reduce the cupric oxide to copper metal which is the desired product.

[0039] Figure 4 shows the process for recovering copper from an arsenic containing copper sulphide ore. In this version of the process the copper ore is fed directly to a conventional concentrator where the sulphide minerals are separated from the gangue using a comminution - flotation circuit. An additional flotation stage is included where the sulphide minerals present are further separated into a low arsenic copper concentrate suitable for sale to a conventional smelter and a high arsenic concentrate which requires further processing. This high arsenic copper concentrate is fed to a roaster where a calcium source such as calcium oxide or more preferably calcium carbonate is added mixed with the sulphide concentrate such that as the material is roasted in an oxygen containing gas at -575C. The sulphur present oxidises to sulphates and is retained in the calcine predominantly as a mix of calcium sulphate and copper sulphate. This removes the need for an acid plant. The arsenic present is oxidized during the roast and predominantly reacts with copper in the concentrate to form copper arsenate and is retained inthe calcine. The gas from the roaster is scrubbed with lime to remove any sulphur containing gases and also capture any arsenic and other metals present. The solids for the scrubber are then returned to the roaster to make use of excess lime and for any arsenic present to react with iron to convert it to ferric arsenate. The roasted material is then fed to an ammonia - ammonium chloride leach stage to dissolve the majority of the copper away from the arsenic compounds and gangue material in the residue. The slurry is then filtered to give a copper depleted solid residue and a copper rich pregnant liquor. The residue is washed using conventional means and then fed to a dilute sulphuric acid leach to dissolve the arsenic and copper present. Some ferric sulphate is added to the leach to ensure there is sufficient iron present in the liquor to enable the precipitation of ferric arsenate in a later stage. The slurry from the leach is filtered and the residue washed before being fed to a conventional gold circuit where gold and silver are recovered using cyanide technology. After recovery of the gold and silver the residue is disposed of into a suitable facility such as a tailings dam. The acidic liquor containing copper, arsenic and iron is partially neutralized using lime to raise the PH to ~2 and precipitate the majority of the arsenic present as ferric arsenate. The slurry is filtered to separate the iron arsenate which is returned to the roaster to improve the form of the iron arsenate and leave a high copper liquor which is then further neutralized to precipitate the copper as a copper hydroxide / oxide. The copper precipitate is filtered off and added to the low arsenic concentrate for sale to a smelter. The copper rich pregnant liquor is fed to an ammonia stripping - crystallization step where the ammonia is stripped off by heating the liquor to boiling point. The pregnant liquor continues to boil as it is heated and this drives off ammonia gas in an ammonia - water gas. As the ammonia concentration drops the copper solubility decreases and copper containing crystals which contain some copper hydroxy chloride precipitate. The crystal containing liquor is then cooled to 25C to precipitate more crystals and filtered to separate the crystal from the spent liquor. The ammonia - water vapour gas passes into an absorption column where it is contacted with the cooled spent liquor from the crystallisation stage and the ammonia and water are recaptured for recycle to the leaching stage. The crystals are filtered and washed and hydrolysed by mixing with a dilute ammonium hydroxide solution to hydrolyse the copper hydroxy chloride present to cupric oxide. The dilute ammonia solution also dissolves traces of unwanted impurities that are precipitated with, or more generally entrained, in the crystals. The pure cupric oxide crystals are then filtered off, washed and blended with the low arsenic concentrate for sale to a smelter.

[0040] Figure 5 shows the flowsheet where a complex sulphide concentrate containing valuable copper, nickel and cobalt is blended with black mass from recycled lithium ion batteries which also contains valuable copper, nickel and cobalt as well as lithium and the blend is sulphationroasted at 600C in an air - sulphur dioxide atmosphere to convert the metals present to oxides and / or sulphates. The calcine is then leached in caustic soda at around 4M / 1 to dissolve the lithium and leave the other metals in the solid residue. The slurry then filtered to give solids stream and a liquid stream. Magnesium carbonate is added to the liquid stream to precipitate the lithium as lithium carbonate suitable for sale. The metal rich residue is leached with ammonia-ammonium chloride liquor and the slurry filtered to leave a residue for disposal and / or processing to recover any PGM’s present. The liquid is treated with H2S gas to precipitate the valuable copper, nickel and cobalt as a mixed sulphide that can be sold to a smelter for further treatment. The H2S gas required is obtained by reacting part of the magnesium sulphate containing liquor from the lithium precipitation in an anaerobic reactor where sulphate is converted to the H2S and carbon dioxide is generated to precipitate magnesium carbonate for recycle to the lithium precipitation stage.

[0041] Figure 6 shows the flowsheet where a complex sulphide concentrate containing copper as the most valuable metal with minor, but important, amounts of nickel and cobalt is blended with black mass from recycled lithium ion batteries which also contains valuable copper, nickel and cobalt as well as lithium and the blend is sulphation roasted at 600C in an air - sulphur dioxide atmosphere to convert the metals present to oxides and / or sulphates. The calcine is then leached in caustic soda at around 4M / 1 to dissolve the lithium and leave the other metals in the solid residue. The slurry then filtered to give solids stream and a liquid stream. Magnesium carbonate is added to the liquid stream to precipitate the lithium as lithium carbonate suitable for sale. The metal rich residue is leached in water at a PH of around 6-8 (a neutral leach) to dissolve the cobalt and nickel present while leaving the copper in the residue. The liquid is treated with H2S gas to precipitate the valuable nickel and cobalt as a mixed sulphide that can be sold to a smelter for further treatment while the solid residue is leached in ammonia - ammonium chloride liquor. The copper is then recovered from this liquor by stripping off ammonia and cooling the liquor to precipitate copper oxide and / or hydroxy chloride .The residue from the ammonia -ammonium chloride leach is further processed to recover PGM’s present. The H2S gas required is obtained by reacting part of the magnesium sulphate containing liquor from the lithium precipitation in an anaerobic reactor where sulphate is converted to the H2S and carbon dioxide is generated to precipitate magnesium carbonate for recycle to the lithium precipitation stage.

[0042] Figure 7 Shows the flowsheet for processing the residue from where the metals such as molybdenum, cobalt, nickel and rare earths have been leached from the calcine and the residue is then fed to an ammonia - ammonium chloride leach stage to dissolve the majority of the copper away from the arsenic compounds and gangue material in the residue. The slurry is then filteredto give a copper depleted solid residue and a copper rich pregnant liquor. The residue is washed using conventional means and then fed to a dilute sulphuric acid leach to dissolve the arsenic and copper present. Some ferric sulphate is added to the leach to ensure there is sufficient iron present in the liquor to enable the precipitation of ferric arsenate in a later stage. The slurry from the leach is filtered and the residue washed before being fed to a conventional gold circuit where gold and silver are recovered using cyanide technology. After recovery of the gold and silver the residue is disposed of into a suitable facility such as a tailings dam. The acidic liquor containing copper, arsenic and iron is partially neutralized using lime to raise the PH to ~2 and precipitate the majority of the arsenic present as ferric arsenate. The slurry is filtered to separate the iron arsenate which is returned to the roaster to improve the form of the iron arsenate and leave a high copper liquor which is then further neutralized to precipitate the copper as a copper hydroxide / oxide. The copper precipitate is filtered off and added to the low arsenic concentrate for sale to a smelter. The copper rich pregnant liquor is fed to an ammonia stripping - crystallization step where the ammonia is stripped off by heating the liquor to boiling point. The pregnant liquor continues to boil as it is heated and this drives off ammonia gas in an ammonia - water gas. As the ammonia concentration drops the copper solubility decreases and copper containing crystals which contain some copper hydroxy chloride precipitate. The crystal containing liquor is then cooled to 25C to precipitate more crystals and filtered to separate the crystal from the spent liquor. The ammonia - water vapour gas passes into an absorption column where it is contacted with the cooled spent liquor from the crystallisation stage and the ammonia and water are recaptured for recycle to the leaching stage. The crystals are filtered and washed and hydrolysed by mixing with a dilute ammonium hydroxide solution to hydrolyse the copper hydroxy chloride present to cupric oxide. The dilute ammonia solution also dissolves traces of unwanted impurities that are precipitated with, or more generally entrained, in the crystals. The pure cupric oxide crystals are then filtered off, washed and blended with the low arsenic concentrate for sale to a smelter.

[0043] Figure 8 shows the solubility of several metals in sulphate solutions as a function of the PH. The most important features are that ferric iron and aluminium have little solubility above PH 3.6 and copper is soluble at PH 3.6 but has very low solubility above PH 6 and cobalt and nickel are soluble at PH 6 but have very low solubility above PH8.

[0044] Figure 9 shows the flowsheet for processing a complex sulphide concentrate containing valuable amounts of one or more of copper, cobalt, zinc and nickel where the cobalt, nickel and zinc and at least part of the copper are recovered from a dilute acid liquor by stagedprecipitation. The calcine from the sulphation roast is leached in dilute sulphuric acid to dissolve the readily leachable sulphates. The liquor is filtered from the solids and PH adjusted in two stages using magnesium carbonate. The first precipitation stage raises the PH to ~6 to precipitate the copper as copper hydroxide / copper oxide which is then filtered off to leave a magnesium sulphate rich liquor containing the nickel and cobalt. More magnesium carbonate is added to raise the PH to ~8.4 to precipitate the nickel, zinc and cobalt as hydroxides leaving a magnesium sulphate rich liquor with low levels of the copper, cobalt and nickel. The magnesium sulphate liquor is then passed to an anaerobic bioreactor where the sulphate is reduced to give hydrogen sulphide gas and magnesium carbonate precipitates and is recycled to the precipitation stages. The hydrogen sulphide gas transfers to an aerobic bioreactor where to the hydrogen sulphide is oxidized to generate sulphuric acid to return to the dilute acid leach stage. Where there is excess sulphate in the system and there is no other use for the sulphuric acid being generated lime and / or limestone is added to precipitate part of the sulphate to retain the sulphur balance in the system. The residue from the dilute acid leach passes to an ammonia - ammonium chloride leach to dissolve the copper that was not leachable with the dilute acid. The copper is recovered from this liquor by precipitation as described in other examples and the residue passes to a conventional precious metals circuit to recover gold and silver that is present.

[0045] Figure 10 shows the flowsheet for recovering metals from a dilute sulphuric acid solution which contains valuable amounts of one or more of copper, cobalt, nickel and zinc where the copper, cobalt, nickel and zinc are recovered from the dilute acid liquor by staged precipitation. The liquor is first oxidized using a suitable oxidant which is preferably air and / or oxygen gas to ensure all the iron present is there as ferric iron and readily precipitated at a PH of greater than 2.5. The first precipitation stage uses magnesium carbonate to raise the PH to ~4 to precipitate out any ferric iron and aluminium that are present and these are filtered off leaving a sulphate liquor containing the valuable metals. The PH is then raised to PH 6.2 to precipitate out the copper as copper oxide and / or copper hydroxide which is then filtered off to leave a magnesium sulphate rich liquor containing metals such as nickel, cobalt, and zinc. More magnesium carbonate is added to raise the PH to ~8.4 to precipitate any nickel, cobalt and zinc that are present as hydroxides leaving a magnesium sulphate rich liquor with low levels of the metals. The magnesium sulphate liquor is then passed to an anaerobic bioreactor where the sulphate is reduced to give hydrogen sulphide gas and magnesium carbonate precipitates and is recycled to the precipitation stages. The hydrogen sulphide gas transfers to an aerobic bioreactor where the hydrogen sulphide is oxidized to generate sulphuric acid to return to the dilute acid leach stage. Where there is excess sulphate in the system and there is no other use for the sulphuric acidbeing generated lime and / or limestone is added to precipitate part of the sulphate to retain the sulphur balance in the system.ExamplesExample 1

[0046] Ten gram of a copper concentrate containing 45%Cu, 18% Fe and 237ppm As was mixed with 3.05g of CaO and roasted in air at 650C for 3 hours. The calcined solids were then leached with water and the solids separated from the liquor and anlaysed. The resultant liquor had a PEI of ~3.8 and contained 65g / l Cu with the As content being below detection at <lppm indicating that there was no arsenic present in the calcine that was soluble in the acidic liquor. This confirmed that there was no calcium arsenate or arsenic trioxide in the calcine and that it was present as the much more inert ferric arsenate. The residue from the water leach were then further processed by leaching in ammonia - ammonium chloride liquor containing ~5 Og / 1 NH3 and ~50g / l NH4C1 for 2 hours at 50C and the solids and liquids again separated. The liquor from this leach was analysed and found to contain 40g / l Cu but no detectable Arsenic confirming that it was present at <lppm and not leached. The solid residue from this leach was analysed and found to contain 240ppm Arsenic confirming that the arsenic was present in an inert form.Example 2

[0047] A blend was prepared with 16gram of copper concentrate with 25.2% copper, 9.2% arsenic, 38.4% sulphur and 18.7% iron and 14.8g of lime and roasted in air for two hours at 575C. The 37.5 gram of calcine obtained contained 12% copper, 3.9% arsenic, 15% sulphur and 8.1% iron confirming that all of the copper, arsenic, sulphur and iron wewre retained in the calcine with minimal lost through volatilization.Example 3.[0048) A 31.1 g of calcine from example 2 wash washed in water and then leached in a liquor containing 50g / l NH3 and 50g / l NH4C1 at 50C for 2 hours. The liquors were analysed for copper, arsenic, and iron and a mass balance constructed which showed that 59% of the copper, 2.7% of the arsenic and 0.03% of the iron dissolved in the liquor with the balance being retained in the solid residue.Example 4

[0049] A 16.3 gram sample of the residue from Example 3 was leached in 20g / l H2SO4 and the liquor analysed. The liquor analyses was used to construct a mass balance which showed that 85.5% of the copper, 95.5% of the arsenic and 17% of the iron in the feed sample dissolved in the acid. The slurry was filtered and a 50g / l sulphuric acid liquor added to the solids residue to dissolve further copper and arsenic. The residue from the leach was analysed and found to contain 0.3% copper, 0.1% arsenic and 13% iron. The mass balance from the residue analyses showed that 97.5% of the copper, 99.1% of the arsenic and 39% of the iron was extracted by the 2 stage acid lead.

[0050] Example 5. A mass balance was constructed across the roasting, water washing, ammonia - ammonium chloride and 2 stage acid leaches as described in examples 2-4 and found that for the entire process >99% of the copper was extracted into the liquors, >99% of the arsenic reported to the dilute sulphuric acid liquor and 34% of the iron reported to the dilute acid liquor.

Claims

Claims1. Roasting of a sulphur containing material containing one or more of copper and / or cobalt and / or nickel and / or molybdenum and / or lithium and / or rare earth elements and / or arsenic and / or radionuclides at a temperature between 400C and 750C where the material has a controlled ratio of calcium to oxygen to sulphur such that the metals are converted into sulphates and / or oxides and / or arsenates suitable for subsequent hydrometallurgical recovery.

2. A process as described in Claim 1 where the feed material contains both copper and arsenic and the calcium / sulphur ratio is >0.3 / 1.

3. A process according to claim 1 where the roasting temperature is between 500C and 680C.

4. A process according to claim 1 where the material is a sulphide concentrate where the major metal present is copper.

5. A process according to claim 1 where the material is a blend of a sulphide concentrate and black mass where the black mass comes from part processing of recycled li-ion batteries.

6. A process according to claim 1 where the material includes a sulphur source and e-waste and / or black mass.

7. A process according to claims 1 to 3 where the calcine contains copper arsenate which is dissolved in dilute sulphuric acid and iron is also present in the acid solution and a calcium source such as lime or limestone is added to raise the PH above 1.5 to precipitate an iron arsenate compound.

8. A process according to claim 1 where the calcine contains copper and cobalt and nickel is leached in a solution of ammonia - ammonium chloride to dissolve the copper and cobalt and nickel.

9. A process according to claim 8 where the copper and / or cobalt and / or nickel are precipitated from the liquor as sulphides.

10. A process according to claim 1 where the calcine contains lithium and is leached in water or caustic to dissolve the lithium.

11. A process according to claim 9 where the lithium in solution is precipitated from the liquor as lithium carbonate by adding a carbonate source such as magnesium carbonate and / or sodium carbonate and / or carbon dioxide.

12. A process according to claim 1 where the calcine contains molybdenum and is leached in caustic soda to dissolve the molybdenum.

13. A process according to claim 1 where the calcine contains copper, nickel, cobalt, lithium and rare earth metals where it is leached in caustic soda to dissolve the lithium and in a water-based liquor at a PH in the range 6-8 to dissolve the nickel, cobalt and rare earths and in ammonia - ammonium chloride liquor to dissolve the copper.

14. A process in which a copper concentrate which also contains arsenic is mixed with a calcium source such as lime or limestone to give a blend which is roasted at a temperature in the range 450-680C in air to convert most of the arsenic present to copper arsenate where sufficient calcium is added to ensure the molar ratio of calcium to sulphur is greater than 0.3 and the sulphur dioxide content of the gas is maintained below 1%.

15. A process according to claim 14 where the calcine from the roaster is leached in an ammonia - ammonium chloride liquor to dissolve part of the copper.

16. A process according to claims 14 and 15 where the residue after the ammonia - ammonium chloride leaching is leached in dilute sulphuric acid to dissolve the copper and arsenic and part of any iron present.

17. A process according to claim 16 where the molar ratio of iron to arsenic in solution is >1 through dissolving iron from the calcine and if necessary through adding extra iron such as with iron sulphate and then lime is added to the solution containing arsenic, iron and copper to raise the PH to within the range 1.5 to 4 to precipitate out an iron arsenate compound.

18. A process according to claim 17 where the iron arsenate compound is returned to the roaster as part of the feed material.

19. A process according to claim 16 where copper is recovered from the liquor by adding more lime to raise the PH to above 5 to precipitate out the copper as copper oxide and / or copper hydroxide.

20. A process according to claim 16 where copper is recovered from the liquor using solvent extraction.

21. A process where valuable metals are recovered from dilute sulphuric acid solutions containing one or more of copper and / or cobalt and / or nickel and / or zinc by raising the PH in one or more stages using magnesium carbonate to precipitate the metals as their oxides and / or hydroxides and produce a magnesium sulphate solution which is then treated using bioreactors to regenerate the magnesium carbonate and sulphuric acid for recycle within the process.

22. A process according to claim 21 where the dilute sulphuric acid solution containing metals is obtained by leaching calcine from roasting of feed materials containing one or more of copper, cobalt, nickel and zinc in the presence of sulphur and air at a temperature in the range of 500-700C.

23. A process in which a nickel concentrate which also contains arsenic is mixed with a metal sulphide such as a copper concentrate and / or a pyrite concentrate and a calcium source such as lime or limestone to give a blend which is roasted at a temperature in the range 450-680C in air to convert most of the arsenic present to a metal arsenate such as copper arsenate and / or iron arsenate and / or calcium arsenate.

24. A process according to claim 23 where sufficient calcium is added to ensure the molar ratio of calcium to sulphur is greater than 0.3 and the sulphur dioxide content of the gas is maintained below 1%.

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