AU2013263848B2 - Treatment process for extraction of metals from ores - Google Patents

Abstract

TREATMENT PROCESS FOR EXTRACTION OF METALS FROM ORES This invention relates to a hydrometallurgical process for extracting one or more saleable elements, in particular platinum group metals (PGMs), gold, silver and base metals from a sulphur-depleted or otherwise poorly-flotable ore (10). The process comprises either a hot sulphuric acid leaching step (12) to selectively separate base metals (16) into sulphate medium (14), from a solid residue containing the PGMs, gold and silver or a chloride leaching step (20) for recovery of PGMs, gold and silver (24) into chloride leaching medium (22), or both steps to recover the saleable elements. The process may further include a thermal treatment step (18) to condition the PGMs, gold and silver in the solid residue to be soluble in chloride medium. Denotes sulphate media .. so.utn . f. s Atmospheric or Pressure H Solui flows Sulphuric Acid Leach H S 14 Base metal recovery -Solid liquid separation g 16 *Co, REE,~ Y Thermal Treatment 18 U Th 2nuMnuCdiMoAciorLNacC Chion.atio ~ HCI and/or NaCI 22 Precious metal recovery Solid -liquid separation PGM, A...,..i.usT..n

Description

TREATMENT PROCESS FOR EXTRACTION OF METALS FROM ORES 5 BACKGROUND OF THE INVENTION This invention relates to a hydrometallurgical process for treating low-sulphur polymetallic ores for the recovery and separation of valuable metals. 10 Polymetallic orebodies containing multiple valuable metals at lower grades are becoming increasingly attractive for resource companies to assess their potential for exploitation, despite the greater metallurgical challenge in the recovery and separation of such elements into salable concentrates or products. This is particularly the case for ores containing platinum group metals (PGMs), gold or 15 silver and other valuable metals such as nickel, cobalt, copper, rare earth elements (REE) including yttrium and scandium, as well as uranium, thorium, manganese, zinc, cadmium, molybdenum, vanadium, titanium and other minor elements. Current hydrometallurgical process routes for extraction of valuable metals from 20 polymetalic orebodies are described in international patent publication no. WO 99/60178, known as the "Kell Process" (see Fig. 1) and international patent application no PCT/IB2013/055760 (the contents of which are incorporated herein by reference). Both of these processes require as the starting material a flotation concentrate of the ore. 25 The basic Kell Process route comprises the steps of: (i) leaching a flotation concentrate made from an ore in a pressure oxidation sulphate leach to dissolve base metal sulphides contained in the flotation 30 concentrate and forming a sulphate leach filtrate containing base metals and a residue containing platinum group metals (PGMs); (ii) separating the sulphate leach filtrate from the residue; (iii) roasting the residue to form a calcine; and - 1 - -2 (iv) leaching the calcine in a chloride leach to dissolve the PGMs into solution forming a chloride leach filtrate for PGM recovery and a solid waste residue. However, the current processes are not able to be used in sulphide-depleted ores such as 5 laterised and weathered ores, as the quantity and type of minerals in these ores are typically not readily amenable to flotation processing to produce the concentrate containing the value elements as the starting point for these processes. A process for the extraction of valuble minerals from low-sulphur polymetallic ores, including 0 laterised and weathered ores, which are less amenable to flotation concentration techniques as well as otherwise poorly-flotable ores is therefore needed. SUMMARY OF THE INVENTION 5 According to an aspect of the invention, there is provided a hydrometallurgical process for extracting any one or more platinum group metals (PGMs) and optionally any one or more additional valuable elements from the group consisting of gold (Au), silver (Ag), nickel (Ni), cobalt (Co), copper (Cu), chromium (Cr), rare earth elements (REE), yttrium (Y), scandium (Sc), uranium (U), thorium (Th), zinc (Zn), manganese (Mn), cadmium (Cd), molybdenum, .0 (Mo), vanadium (V), titanium (Ti), sulphur (S), sulphuric acid (H 2

SO

4 ) and iron (Fe) from a sulphur-depleted polymetallic ore containing one or more PGMs wherein the ore is a weathered ore, a laterite ore, a goethite ore, an oxidised ore, or a sulphur-depleted polymetallic ore that is not weathered but the nature of the PGMs in the ore are such that low recovery yields are obtained or concentrate grades are too low to meet smelter 25 requirements when subjected to flotation, the process comprising: (i) providing the ore to a reaction vessel; and a) subjecting the ore to a hot sulphuric acid leaching step under pressure or atmospheric conditions wherein the hot sulphuric acid leaching step is any one of or a combination of: 30 A. a conventional atmospheric temperature and pressure leach at 20-100 C, B. a low temperature and pressure leach at 100-130 2C, C. a medium temperature and pressure leach at 130-220 2C, or 10830973 1 - 2a D. a high temperature and pressure leach at 220-260 2C using sulphuric acid to generate a product slurry comprising a solid residue comprising PGMs and optionally one or more of gold, and silver and a sulphate pregnant leach solution (PLS) comprising one or more valuable metal sulphates in solution: Ni, 5 Co, Cu, Cr, REE, Y, Sc, U, Th, Zn, Mn, Cd, Mo, V, Ti; and b) subjecting the solid residue from a) to a chloride leaching step in a chloride leaching medium wherein the chloride leaching medium comprises hydrochloric acid or saline brine in conjunction with an oxidising agent including chlorine, hypochlorite or hydrogen peroxide, and the solid residue of step a) is leached under oxidising 0 conditions, thereby to generate a chloride pregnant leach solution (PLS) comprising any one or more soluble PGMs and optionally any one or more soluble additional valuable elements in the chloride leaching medium. According to a first aspect of the invention, there is provided a hydrometallurgical process for extracting one or more saleable elements from a sulphur-depleted or otherwise poorly 5 flotable ore, the process comprising (i) providing to a reaction vessel the sulphur-depleted or otherwise poorly-flotable ore; (ii) subjecting the ore to one or more leaching steps comprising: a) subjecting the ore to a hot sulphuric acid leaching step under pressure or .0 atmospheric conditions to produce a product slurry comprising saleable metal sulphates in solution and a solid residue containing saleable metals; and/or b) subjecting the ore or solid residue from a) to a chloride leaching step in a chloride leaching medium to produce soluble saleable metals in the chloride leaching medium. 25 The process may further comprise a step of separating the solid residue containing saleable metals from the saleable metal sulphates in solution from the product 10830973 1 slurry of step a) and then providing the resultant separated solid residue to step b). The step of separating may be performed by filtration, or by any other solid/liquid separation means known to those skilled in the art. The process may further comprise a step of recovery of saleable metals from the 5 metal sulphates in solution and/or from the chloride leaching medium, by means of techniques such as solvent extraction, ion exchange, precipitation using hydroxides, carbonates or sulphides, electrowinning, reduction and other techniques known to those skilled in the art based on techno-economic considerations. 10 The sulphur-depleted or otherwise poorly-flotable ore of (i) may be initially processed by crushing, milling or may be as-mined. Alternatively, or in addition, the ore may be subjected to a benefication step to produce an intermediate ore product for providing to the reaction vessel. The benefication step may be performed by screening, sizing, classification, magnetic separation, electrostatic 15 separation or gravity separation thereby to concentrate the valuable metals or reject a gangue component, or by other means of beneficiation known to those skilled in the art. The ore, intermediate ore product or the solid residue from step a) may be subjected to a thermal treatment to produce a thermally treated calcine before 20 subjecting it to step b). The thermal treatment may be performed at about 80 - 750 2C for up to 120 minutes, typically about 400 - 650 2C for 10 to 30 minutes, under oxidizing, neutral or reducing conditions, to remove volatile components from the solid residue. Additionally, the thermal treatment may be performed at about 500 - 1000 2C for 25 up to 120 minutes, typically about 700 - 1000 2C for 30 to 120 minutes, under oxidizing, neutral or reducing conditions, to condition the saleable metals to be soluble in chloride leaching medium. The thermal processes may be performed as individual steps of a sequential thermal treatment process, or as one combined step. -3- The product slurry of step a) may optionally be subjected to a hot acidic conditioning step or an atmospheric leach step, to effect the removal of iron as well as potentially aluminium and magnesium sulphates, to the solution phase, and then subjected to step b). 5 The sulphur-depleted or otherwise poorly-flotable ore, including a poorly-flotable polymetallic ore, may include a weathered ore, including a laterite or goethiite or other oxidised ore, such as weathered PGM-bearing ores that lie close to the surface and give low PGM recoveries when concentrated by flotation. Additionally, the poorly-flotable ore or poorly-flotable polymetallic ore may include a PGM ore 10 that is not weathered but where the nature of the PGMs in the ore is such that they yield a low recovery or concentrate grades are too low to meet smelter requirements when subjected to flotation. The saleable element(s) may comprise one or more platinum group metals (PGMs). In addition, the saleable elements may comprise any one or more of the 15 following: gold (Au), silver (Ag), nickel (Ni), cobalt (Co), copper (Cu), rare earth elements (REE), yttrium (Y), scandium (Sc), uranium (U), thorium (Th), zinc (Zn), manganese (Mn), cadmium (Cd), molybdenum, (Mo), vanadium (V), titanium (Ti), and the like. Further additionally, the saleable elements may comprise sulphur (S), sulphuric acid (H 2 SO4) and/or iron (Fe). 20 According to a second aspect of the invention, the solid residue of step a) may be subjected, either directly or after a milling process, to a concentration or gangue rejection process such as conventional screening, hydrocyclone or other classification means based on size, magnetic separation under either or both high and low-intensity magnetic fields, gravity separation, electrostatic separation, or 25 other means to either concentrate the valuable metals or reject gangue components. According to a third aspect of the invention, the sulphuric acid leaching step a) may comprise subjecting the ore or intermediate ore product to either or a combination of, a conventional atmospheric (20-100 0C), low (100-130 0C) medium 30 (130-220 0C) or high (220-260 0C) temperature and pressure leach using sulphuric acid to generate a product slurry comprising a sulphate pregnant leach solution (PLS) comprising one or more valuable metal sulphates: Ni, Co, Cu, REE, Y, Sc, -4- U, Th, Zn, Mn, Cd, Mo, V, Ti and a solid residue comprising one or more of the elements PGMs, gold, silver and other valuable metals. The solid residue may be separated from the metal sulphates in the sulphate PLS via filtration or other suitable means for subsequent recovery of the elements from 5 the solid residue. The sulphate PLS may be subjected to separation and/or recovery of the one or more valuable metal sulphates: Ni, Co, Cu, REE, Y, Sc, U, Th, Zn, Mn, Cd, Mo, V, Ti by means of techniques such as solvent extraction, ion exchange, precipitation using hydroxides, carbonates or sulphides, electrowinning, reduction and other 10 techniques known to those skilled in the art based on techno-economic considerations. In a particular embodiment of the invention, the sulphate PLS may be oxidised by sparging the reaction vessel with oxygen or by addition of a chemical oxidant such as hydrogen peroxide, sulphur dioxide, or other means known to those skilled in 15 the art. The sparging may be performed under either atmospheric or pressure conditions. In a further particular embodiment of the invention, the techniques described in international patent application no. PCT/IB2013/055760 (incorporated herein by reference) may be applied to the process of the invention, thereby to recover 20 sulphuric acid and precipitate a potentially saleable or storable iron product. Furthermore, technologies such as precipitation or crystallization may be employed in the process to produce a potentially saleable or storable ferric or ferrous hydroxide or sulphate product while recovering sulphuric acid into a stream suitable for recycling. 25 In an alternative embodiment of the invention, a conventional atmospheric or low pressure leach using sulphuric acid may be applied directly to the product slurry comprising the sulphate PLS and solid residue, thereby to remove excess iron sulphates from the solid residue into the sulphate PLS. The sulphate PLS may then be subjected to air- or oxygen-sparged ferric hydroxide precipitation under 30 atmospheric or pressurized conditions for removal of excess iron sulphates. -5- Alternatively, or in addition to the conventional atmospheric or low-pressure leach using sulphuric acid applied directly to the product slurry comprising the sulphate PLS and solid residue, where the solid residue is subjected to thermal treatment, excess soluble iron may be removed from the thermally treated calcine. 5 The chloride leaching medium of step b) may contain iron chloride and may be treated by pressure, precipitation or crystallization, concentrated by evaporation, reverse osmosis, nanofiltration or other membrane technology, or treated by sparging/rectification, pyrohydrolysis or other technology known to those skilled in the art to produce an iron-bearing product. 10 According to a fourth aspect of the invention, at step b) of the process, the chloride leaching medium may comprise hydrochloric acid or saline brine in conjunction with an oxidising agent such as chlorine, hypochlorite, hydrogen peroxide or other oxidising agents known to those skilled in the art, and the ore of (i) or solid residue of step a) is leached under oxidising conditions, thereby to generate a chloride 15 pregnant leach solution (PLS) comprising one or more saleable elements including PGMs, Au, Ag, as well as Ni, Co, Cu, REE, Y, Sc, U, Th, Zn, Mn, Cd, Mo, V and Ti. The chloride PLS may be subjected to separation and/or recovery of the one or more saleable elements by means of techniques such as solvent extraction, ion 20 exchange, precipitation using hydroxides, carbonates or sulphides, electrowinning, reduction and other techniques known to those skilled in the art based on techno economic considerations. In a particular alternative embodiment of the invention, step a) of the process is omitted and the sulphur-depleted or otherwise poorly-flotable ore of step (i), or the 25 intermediate ore product is separated into a solid residue containing valuable metals and a filtrate, followed by leaching of the solid residue by the oxidising chloride leaching step above and recovery of the saleable elements from the chloride PLS. The solid residue may be subjected to a thermal treatment prior to the oxidising 30 chloride leaching to produce a calcine residue. -6- The thermal treatment may be performed at about 80 - 750 2C for up to 120 minutes, typically about 400 - 650 2C for 10 to 30 minutes, under oxidizing, neutral or reducing conditions, to remove volatile components from the solid residue. Additionally, the thermal treatment may be performed at about 500 - 1000 2C for 5 up to 120 minutes, typically about 700 - 1000 2C for 30 to 120 minutes, under oxidizing, neutral or reducing conditions, to condition the valuable metals to be soluble in chloride leaching medium. The thermal processes may be performed as individual steps of a sequential thermal treatment process, or as one combined step. 10 The oxidising chloride leaching medium above, or the chloride leaching medium of step b) of the first aspect of the invention may contain iron chloride and may be treated by pressure, precipitation or crystallization, concentrated by evaporation, reverse osmosis, nanofiltration or other membrane technology, or treated by sparging/rectification, pyrohydrolysis or other technology known to those skilled in 15 the art to produce an iron-bearing product. In a further particular embodiment, the chloride leaching step may comprise a less acidic chloride leaching medium having a pH of between about 2.5 and 7.5. In a further particular embodiment, the chloride leaching step may comprise a chloride leaching medium with a free acidity of between about 50 to 300 g/L HCI. 20 According to a further particular embodiment, the chloride leaching step may be performed by atmospheric or pressure autoclave leaching with saline brine under oxidising conditions. According to another aspect of the invention, there is provided a hydrometallurgical process for extracting one or more saleable elements from a sulphur-depleted or 25 otherwise poorly-flotable ore, the process comprising: (i) providing to a reaction vessel a sulphur-depleted or otherwise poorly-flotable ore; and a) subjecting the ore to a hot sulphuric acid leaching step under pressure or atmospheric conditions to produce a -7product slurry comprising saleable metal sulphates in solution and a solid residue containing saleable metals; and/or b) subjecting the ore of (i) or solid residue from a) to a 5 chloride leaching step in a chloride leaching medium to produce soluble saleable metals in the chloride leaching medium. BRIEF DESCRIPTION OF THE DRAWINGS 10 Figure 1 is a simplified block flowsheet diagram of Kell process; Figure 2 is a simplified block flowsheet diagram showing the modified Kell process of the present invention. 15 DETAILED DESCRIPTION OF THE INVENTION The current invention provides a hydrometallurgical process for treating low sulphur polymetallic ores, including laterised and weathered ores and other poorly flotable ores, by concentration and gangue separation techniques and with 20 recovery and separation of saleable metals. The terms "element", "mineral" and "metal" are used interchangeably in this specification. 25 "PGMs" mean ruthenium (Ru), rhodium (Rh), palladium (Pd), osmium (Os), iridium (Ir), and platinum (Pt). Low-sulphur ores and poorly-flotable ores, including poorly-flotable polymetallic ores, may include weathered ores, including a laterite or goethiite or other oxidised 30 ore such as weathered PGM-bearing ores that lie close to the surface and give low PGM recoveries when concentrated by flotation. Additionally, poorly-flotable ores -8or poorly-flotable polymetallic ores may include PGM ores that are not weathered but the nature of the PGMs in the ore is such that they yield a low recovery or concentrate grades are too low to meet smelter requirements when subjected to flotation. 5 As illustrated in Figure 2, the main embodiment of the invention is a hydrometallurgical process for extracting saleable elements including valuable metals from low-sulphur or other poorly-flotable polymetalic ores (10) without the requirement of prior flotation concentration of the ore. The polymetalic ores (10) input into the process can be as-mined, or may be comminuted to a finer grain 10 size. The input ores may also have been subjected to a previous benefication step such as screening, sizing, classification, magnetic separation, electrostatic separation, dense media sepration, radiometric sorting, optical sorting, or gravity separation to concentrate the valuble metals or reject a gangue component. However, the necessity and type of comminution or benefication at this stage 15 would be dictated by the ore characteristics. In some cases, particularly weathered lateritic and goethitic ores, a substantial proportion of the ore is often present as extremely fine particles, and in these cases simple size classification by screening or hydrocycloning is employed, typically in conjunction with light crushing or milling, scrubbing or attritioning, 20 thereby obviating the need for a full comminution circuit comprising multi-stage crushing and grinding. The process may comprise a first step where the low-sulphur or other poorly flotable polymetalic ore (10) is subjected to a hot sulphuric acid leaching step under pressure or atmospheric conditions (12) to produce a product slurry 25 comprising valuable metal sulphates in solution and a solid residue containing valuble metals. The solid residue would then be separated from the metal sulphates in solution by a means of sold/liquid separation, such as by filtration or other means known to those skilled in the art. The product slurry can also be pre-processed by hot acidic conditioning or 30 atmospheric leach to remove iron and potentially also aluminium and magnesium sulphates into solution prior to separation into the metal sulphate solution and the solid residue. The operating conditions of such an acid conditioning step is -9dictated by the mineralogy of the ore being processed and are well known to those skilled in the art. An optional modification to the sulphuric acid leaching process depending on the requirement for an optimal economic outcome involves the use of either, or a 5 combination of conventional atmospheric (20-100 0C), low (100-130 0C) medium (130-220 0C) or high (220-260 0C) temperature and pressure leach using sulphuric acid. The pressure is selected for this process is based on the metal dissolution at each pressure and the economics of the project and would therefore be selected accordingly by a person skilled in the art based on the ore to be processed. This 10 leaching process leaches valuble elements including Ni, Co, Cu, REE, Y, Sc, U, Th, Zn, Mn, Cd, Mo, V, Ti into a sulphate pregnant leach solution (PLS), whilst the solid residue containing PGMs, gold, silver and other valuable metals is separated via filtration or other means for subsequent recovery. The selection of the specific unit processes for separation and/or recovery of value elements such as Ni, Co, 15 Cu, REE, Y, Sc, U, Th, Zn, Mn, Cd, Mo, V, Ti is made based on techno-economic considerations. Techniques such as solvent extraction, ion exchange, precipitation using hydroxides, carbonates or sulphides, electrowinning, reduction and others may be used to achieve separation and/or recovery of these elements from the sulphate PLS. 20 The sulphuric acid leach product slurry comprising the sulphate PLS and solid residue may be further treated by a conventional atmospheric or low-pressure leach using sulphuric acid, allowing removal of excess iron sulphates from the solid residue into the sulphate PLS for removal by air- or oxygen-sparged ferric hydroxide precipitation under atmospheric or pressurized conditions. 25 Whilst sulphide-depleted ores typically do not require oxidation, in some cases, where minor residual sulphide minerals are present, the sulphuric acid leaching step may require oxidation, via sparging the reaction vessel (atmospheric or pressure) with oxygen or by addition of a chemical oxidant such as hydrogen peroxide, sulphur dioxide, or other means. 30 Additionally, techniques described in international patent publication no. PCT/IB2013/05576 may also be applied to such polymetallic and low-sulphur ore processing, allowing recovery of sulphuric acid and precipitation of a potentially -10saleable or storable iron product. Specifically the solid residue of step a) above may, after being subjected to thermal treatment produce off-gases comprising sulphur, and the off-gases may be subjected to a step of recovering sulphur: a. by condensation as an elemental sulphur intermediate product; 5 b. into a sulphuric acid intermediate product; or c. from dryer off-gases into a sulphide, polysulphide, polythionate, thiosulphate or similar intermediate product. In an alternative embodiment, the metal sulphates in solution of step a) may be 10 treated by atmospheric or low-pressure atmospheric leach using sulphuric acid, allowing recovery of sulphuric acid and precipitation of iron product. Other technologies known to those skilled in the art such as precipitation or crystallization may be employed to produce a potentially saleable or storable ferric 15 or ferrous hydroxide or sulphate product while recovering sulphuric acid into a stream suitable for recycling. Following separation of the solid residue obtained from the sulphuric acid leach step from the sulphate PLS, the solid residue is subjected to recovery of PGMs and other saleable metals by leaching (20) in a chloride leaching medium. 20 In an optional step, and depending on the mineralogical nature of the feed material, the solid residue from the sulphuric acid leach step may be subjected to a thermal treatment (18) to condition the PGMs or other saleable metals to be soluble in the chloride leaching medium. Typically, the thermal treatment (18) would be performed at about 80 - 750 LC for 25 up to 120 minutes, preferably at about 400 - 650 LC for 10 to 30 minutes, under oxidizing, neutral or reducing conditions, to remove volatile components from the solid residue. An additional thermal treatment would typically be performed at about 500 - 1000 LC for up to 120 minutes, preferably at about 700 - 1000 LC for 30 to 120 minutes, 30 under oxidizing, neutral or reducing conditions, to condition the valuable metals to be soluble in chloride leaching medium. However, if an economically sufficient - 11 proportion of the PGMs is already soluble without requirement of the thermal treatment, then this step may be omitted. The thermal processes (18) may be performed as individual steps of a sequential thermal treatment process, or as one combined step. 5 Either instead of, or in addition to further treatment of the sulphuric acid leach product slurry comprising the sulphate PLS and solid residue by a conventional atmospheric or low-pressure atmospheric leach using sulphuric acid to remove excess iron sulphates from the solid residue, the solid residue produced from the thermal process (calcine) may be subjected to treatment to remove excess soluble 10 iron in the form of iron chloride in the calcine. The iron chloride may be treated by pressure, precipitation or crystallization, concentrated by evaporation, reverse osmosis, nanofiltration or other membrane technology, or treated by sparging/rectification, pyrohydrolysis or other technology known to those skilled in the art to produce an iron-bearing saleable product. 15 The chloride leach step (20) may be optimised for effective recovery and/or separation of some of the broader range of saleable elements that are often present in polymetalic ores generally, and particularly in lateritic, weathered, oxide and other ore types such as weathered PGM-bearing ores that lie close to the surface and give low PGM recoveries when concentrated by flotation. 20 Specifically, the chloride leaching step (20) may be performed under oxidising conditions using hydrochloric acid or saline brine in conjunction with an oxidising agent such as chlorine, hypochlorite, hydrogen peroxide or others known to those skilled in the art. Value elements such as Pt, Pd, Rh, Ru, Ir, Os (i.e. PGMs), Au, Ag, as well as Ni, Co, Cu, REE, Y, Sc, U, Th, Zn, Mn, Cd, Mo, V, Ti (24) are 25 leached into the chloride pregnant leach solution (PLS). Selection of specific unit processes for separation and/or recovery (22) of these value elements is made based on techno-economic considerations. Techniques such as solvent extraction, ion exchange, precipitation using hydroxides, carbonates or sulphides, electrowinning, reduction and others may be used to achieve separation and/or 30 recovery of these elements from the chloride PLS. -12- In one alternative embodiment of the main invention, the economic optimum may be to avoid the sulphuric acid leaching step (12) altogether, and recover the saleable elements directly in the chloride leach (20). The ore feed material (10) for the chloride leach step (20) may be first subjected to a thermal treatment (18). 5 Typically, the thermal treatment (18) would be performed at about 80 - 750 2C for up to 120 minutes, preferably at about 400 - 650 2C for 10 to 30 minutes, under oxidizing, neutral or reducing conditions, to remove volatile components from the solid residue. An additional thermal treatment would typically be performed at about 500 - 1000 10 2C for up to 120 minutes, preferably at about 700 - 1000 2C for 30 to 120 minutes, under oxidizing, neutral or reducing conditions, to condition the valuable metals to be soluble in chloride leaching medium. However, if an economically sufficient proportion of the PGMs is already soluable without requirement of the thermal treatment, then this step may be omitted. 15 The thermal processes (18) may be performed as individual steps of a sequential thermal treatment process, or as one combined step. In this alternative embodiment, the thermally treated calcine residue or the direct ore feed material (10) is leached in chloride medium under oxidising conditions using hydrochloric acid or saline brine in conjunction with an oxidising agent such 20 as chlorine, hypochlorite, hydrogen peroxide or other. Saleable elements such as Pt, Pd, Rh, Ru, Ir, Os, Au, Ag, as well as Ni, Co, Cu, REE, Y, Sc, U, Th, Zn, Mn, Cd, Mo, V, Ti (24) are leached into the chloride pregnant leach solution (PLS). As in the case of the sulphate pregnant leach solution, the selection of specific unit processes for separation and/or recovery of these value elements (22) are made 25 based on techno-economic considerations. Techniques such as solvent extraction, ion exchange, precipitation using hydroxides, carbonates or sulphides, electrowinning, reduction and others may be used to readily achieve separation and/or recovery of these elements from the chloride PLS. As in the case of the sulphate leaching step, excess soluble iron in the chloride 30 PLS may be removed. In this case, iron chloride from the chloride PLS is treated by pressure, precipitation or crystallization, concentrated by evaporation, reverse -13osmosis, nanofiltration or other membrane technology, or treated by sparging/rectification, pyrohydrolysis or other technology known to those skilled in the art to produce an iron-bearing saleable product. In a further alternative embodiment, less acidic (pH 2.5 - 7.5) chloride leaching 5 may be employed in any of the chloride leaching steps. For some materials or process residues an economically attractive recovery of value elements may be achieved by atmospheric or pressure autoclave leaching with saline brine under weak-acid oxidising conditions. In this embodiment iron is preferentially maintained in the solid leach residue and iron leaching is minimised. Sufficient value element 10 leaching is realised and further processing for iron removal is reduced. The process may be further modified by subjecting the solid residues obtained from any step of the process, either directly or after a milling process, to a concentration or gangue rejection process such as conventional screening, sizing, hydrocyclone or other classification means based on size, magnetic separation 15 under either or both high and low-intensity magnetic fields, gravity separation, electrostatic separation, or other means of physical beneficiation to either concentrate the valuable metals or reject gangue components. Selection of appropriate method or a combination of methods is dependent on the nature of the bulk and trace mineralogy, mineral associations and lock-liberation 20 characteristics. In some instances the economic optimum may entail the partial deportment of some value elements to the rejects stream, for disposal or later processing. While low-grade minerals may be targeted for rejection at any stage in the process, in some cases particularly acid-consuming components such as carbonates, serpentinites and other weathered phases, may be specifically 25 targeted with an aim at reducing acid consumption in either the sulphate or chloride leaching steps. The saleable elements would in particular, include platinum group metals (PGMs), gold and silver but may additionally include other valuable metals (24). These 30 metals are separated by means of the process from other valuable metals such as nickel, cobalt and copper, and additionally, rare earth elements, including yttrium and scandium, and uranium, thorium, vanadium, titanium, manganese, zinc and -14cadmium (16), whilst iron components may also be extracted as potentially saleable products. The chloride and/or sulphate acids from the process may be recycled and additional amounts of metals may be recovered during this recycling process. 5 Base metals such as nickel, copper and cobalt may be recovered as sulphates in wash waters from final residues and are recycled along with sulphuric acid to sulphate streams earlier in the process. Any minor fugitive PGMs or other value metals are likewise recovered as chlorides in wash waters from final residues and are recycled along with hydrochloric acid to chloride streams earlier in the process. 10 A combination of iron and hydrochloric acid recovery techniques including iron chloride requiring removal from the process mentioned above may be treated by pressure, precipitation or crystallization, concentrated by evaporation, reverse osmosis, nanofiltration or other membrane technology, or treated by sparging/rectification, pyrohydrolysis or other technology known to those skilled in 15 the art to produce an iron-bearing product. Additional amounts of soluble iron may be removed in the chloride circuit by conventional atmospheric or low-pressure atmospheric leach using hydrochloric acid, either during PGM dissolution or separately. After optional removal of some free acidity by split recycle, precipitation, reverse osmosis, nanofiltration or other 20 conventional means, iron may be recovered by use of established solvent extraction, sparging/rectification, pyrohydrolysis at ~700 OC in a fluid bed or spray roast reactor or by means of pressure precipitation techniques at ~160-1900C. This yields a potentially saleable hematite, maghemite, magnetite or goethite product, depending on the chemical and physical conditions employed. In this process, 25 further free acidity and water, as well as valuable base metals, such as residual nickel, copper and cobalt, are recovered for recycle to the process. In many cases there are potential technical, economic and environmental benefits that may arise from implementation of the process described. Techno-Economic 30 - Processing of ores that are unresponsive to concentration techniques to produce a concentrate that has suitable characteristics for their smelting -15- - Processing of ores that are unresponsive, or alternatively only responsive in a limited manner, to physical beneficiation techniques to upgrade the metal grade or reject gangue minerals - substantial decrease in energy consumption, water use and gaseous 5 emissions compared with smelting; - processing of products from low-grade ores and tailings, without constraints on grade and gangue impurities imposed by smelter terms; - recovery of reagents usable within the process, such as precipitants, coagulants and acids; 10 - recovery of further free acidity and water, as well as valuable base metals, such as residual nickel, copper and cobalt, for recycle to the process thereby increasing overall metal recoveries and reducing water, energy and reagent consumptions. Environmental 15 - capture of iron as a potentially saleable or storable by-product, substantially reducing gaseous, water-borne or land-based emissions; - recycle of water and other components, minimizing use of make-up water; - large reduction in energy consumption compared with conventional smelting, with associated substantial decrease in C02 emissions. 20 Safety and Operability - avoidance of the high temperature rock melting conditions required in traditional smelting, - minimisation of noxious fumes. EXAMPLES 25 Typically, different combinations of techniques are required, depending on specific ore mineralogy, chemistry, particle size distribution and metallurgical response of -16the specific low-sulphur polymetallic ores, the type and grade of the value metals present in the ore and their market prices. The following examples are provided to demonstrate the efficacy of individual techniques that have been brought to bear in various combinations on specific feed materials, resulting in potentially economic 5 recovery and/or separation of multiple value elements and possible reuse or regeneration of reagents. These examples, however, are not to be construed as limiting in any way either the spirit or scope of the invention. EXAMPLES 1 - 4 10 Recovery of Platinum, Palladium, Nickel, Cobalt, Copper and Scandium from a Lateritic Polymetallic Ore In these four examples a lightly milled lateritic polymetallic ore is subjected to a hot sulphuric acid leaching step to effect the removal of iron as well as other elements 15 such as aluminium, magnesium, scandium, zinc, nickel, copper, cobalt, manganese, etc. Table 1 shows the head assay of the material and results are shown in Table 2. Table 1 - Example 1 Head Assays Element Unit Assay Pt g/t 0.97 Pd g/t 0.015 Au g/t 0.008 Sc g/t 525 Fe % 36.1 Al % 5.70 Mg % 0.63 Ni g/t 1410 Cu g/t 126 Co g/t 605 Cr g/t 1620 Mn g/t 5330 Zn g/t 170 -17- Table 2 - Summary of Sulphuric Acid Leach Elemental Extractions Temperature, 'C Extraction, % Sc Pt Pd Au Ni Cu Co Fe Al Mg Mn Cr Zn 95 61 0.0 0.0 0.0 72 53 60 62 38 88 68 71 63 200 82 0.0 0.0 0.0 90 84 95 2.7 68 93 70 26 89 250 91 0.0 0.0 0.0 95 87 95 3.3 24 94 86 6.7 96 Acid leach testwork achieved high recoveries of nickel, copper, cobalt, scandium, 5 manganese and zinc of ~60%, ~80% and ~90% for leaches conducted at 95 0C (atmospheric), 200 0C (pressure) and 250 0C (pressure), respectively. No deportment of Pt, Pd, or Au into the pregnant leach solution (PLS) was observed. Low iron dissolution (< 5%) was observed for pressure leaching while ~60% Fe extraction was observed for the atmospheric leach. Chromium was effectively 10 leached at 95 0C conditions and less at higher temperatures. In a comparative study an unleached milled sample and the residues from sulphuric acid leaches conducted at 95 C, 200 C and 250 0C were each subjected to a two-stage chlorination using chlorine gas controlling oxidation-reduction 15 potential (ORP) to ~1,000 mV (vs Ag/AgCI) to dissolve the Pt and residual Sc. Results are shown in Table 3. Chlorination leach achieved up to 98% Pt extraction from both the milled, unleached sample and sulphuric acid leach residues. Recovery of Sc from the 20 milled, unleached sample was observed to be 96%. High recovery of remaining Sc (~90%) was achieved by chlorination leach of the three acid leach residues, and this serves to increase overall Sc recovery. High Fe dissolution and a high extent of mass loss was observed for all four tests, presenting potential for production of a value-added iron-containing by-product via one of the techniques described 25 above. Moreover, optimisation of leach conditions may show acceptable value element recoveries at significantly lower iron dissolutions. The results on the chlorination leach of the directly milled unleached sample show Pt recovery of ~98% from the sample and scandium recovery of 96%, -18demonstrating the possibility of recovery of all value elements in a single chlorination leach step. Table 3 - Summary of Chlorination Results Extraction, % Mass Stage Lass, Sc Pt Pd Au Ni Cu Co Fe Al Mg Mn Cr Zn %ow/w Chlorination 1 (95 C Leach Residue) 91 93 100 78 84 95 99 98 35 88 99 93 95 59.6 Chlorination 2 91 96 91 93 88 94 98 99 54 83 99 96 95 71.3 (200 0C Leach Residue) Chlorination 4 90 98 98 95 73 96 89 98 46 58 95 94 51 67.3 (250 0C Leach Residue) Chlorination 4 (Milled Sample, Not 96 98 100 92 95 97 99 99 48 98 100 96 94 73.4 Leached) 5 EXAMPLE 5 Concentration and Gangue Removal with Recovery of Platinum, Palladium, Nickel, Cobalt, Copper and Scandium from a Lateritic Polymetallic Ore 10 In this second example the milled sample and the residues from sulphuric acid leaches conducetd at 95 and 200 0C were subjected to physical beneficiation by gravity separation and magnetic separation for upgrading of Pt. Results are shown in Table 4. Gravity separation enrichment ratios (grade over head) were observed 15 to be 4.7 and 2.8, and 3.1 for the milled sample, using V-Table and Flat table, respectively. Magnetic separation enrichment ratios (grade over head) were observed to be 2.7, 1.7, and 3.1 for the milled sample, 95 0C leach residue and 200 0C leach residue, respectively. 20 -19- Table 4 - Summary of Gravity and Magnetic Separation Results Mass Pt Recovery Milled Sample Dist., Assay, Enrich. Dist.,% % ppm Ratio Feed (Assay) - 0.93 - V-Table Con 3.7 4.33 4.7 14 Flat table Con 3.3 2.64 2.8 7.5 TOTAL Conc 7.0 - - 21 V-Table Tail 17 1.02 1.1 15 Flat table Tail 76 0.98 1.1 64 TOTAL Tail 93 - - 79 Milled Mags 11 3.52 2.7 29 Milled Non-Mags 47 0.80 0.6 29 95 *C Leach Residue Mags 4.4 3.59 1.7 7.4 95 *C Leach Residue Non-Mags 70 1.98 0.9 65 200 *C Leach Residue Mags 4.0 5.75 3.1 12 200 *C Leach Residue Non-Mags 47 1.05 0.6 27 EXAMPLE 6 5 Recovery of Ti, V, Zn, Ni, Co and Mn from a Polymetallic Ore In this example a polymetallic ore was milled and calcined in a rotary kiln at 950 0C under controlled atmospheric conditions. The calcine was subjected to a chlorination leach. Head assays and results are summarized in Tables 5 and 6, 10 respectively, showing high recoveries of value metals. 15 - 20 - Table 5 - Head Assay of Polymetallic Ore for Example 6 Element Assay, g/t Al 9600 Ca 25,000 Co 200 Cr 50 Cu 10 Fe 527,000 K 250 Mg 3,200 Mn 3,000 Ni 160 Pb 60 Si 18,700 Ti 88,000 V 13,120 Zn 500 Table 6 - Leach Recoveries from Polymetallic Ore for Example 6 Extraction, % Ti V Mn Zn Fe Co Ni 81.7 93.5 94.3 82.6 89.6 100.0 87.6 5 REFERENCES Liddell, K.S. Hydrometallurgical treatment process for extraction of platinum group metals obviating the matte smelting process, R.S.A. Pat. 2000/6600: Appl. 19 May 10 1998: Acc. July 25 2001; U.S. Pat. 6,579,504: Appl. 19 May 1999: Acc. June 17, 2003; Canada Pat. 2,332,520,: Appl. 19 May 1999: Acc. January 27, 2009. Liddell, K.S. and Adams, M.D. Kell hydrometallurgical process for extraction of platinum group metals and base metals from flotation concentrates, J. S. Afr. Inst. 15 Min. Metall. Trans., vol. 112, January 2012, pp. 31-36. Liddell, K.S., Newton, T., Adams, M.D. and Muller, B. Energy consumptions in Kell hydrometallurgical refining versus conventional pyrometallurgical smelting of PGM - 21 concentrates, J. S. Afr. Inst. Min. Metall. Trans., vol. 111, February 2011, pp. 127 132. Liddell, K.S. and Adams, M.D. Hydrometallurgical Treatment Process for 5 Extraction of Metals from Concentrates, South African Provisional Patent Application No. 2012/05222; PCT - Patent Application No. PCT/1B2013/055760 in the name of Lifezone Limited, 12 July 2013.. - 22 -

Claims (15)

1. A hydrometallurgical process for extracting any one or more platinum group metals (PGMs) and optionally any one or more additional valuable elements from the group consisting of gold (Au), silver (Ag), nickel (Ni), cobalt (Co), copper (Cu), chromium (Cr), rare earth elements (REE), yttrium (Y), scandium (Sc), uranium (U), thorium (Th), zinc (Zn), manganese (Mn), cadmium (Cd), molybdenum, (Mo), vanadium (V), titanium (Ti), sulphur (S), sulphuric acid (H 2 SO 4 ) and iron (Fe) from a sulphur depleted polymetallic ore containing one or more PGMs wherein the ore is a weathered ore, a laterite ore, a goethite ore, an oxidised ore, or a sulphur-depleted polymetallic ore that is not weathered but the nature of the PGMs in the ore are such that low recovery yields are obtained or concentrate grades are too low to meet smelter requirements when subjected to flotation, the process comprising: (i) providing the ore to a reaction vessel; and a) subjecting the ore to a hot sulphuric acid leaching step under pressure or atmospheric conditions wherein the hot sulphuric acid leaching step is any one of or a combination of: A. a conventional atmospheric temperature and pressure leach at

20-100 QC, B. a low temperature and pressure leach at 100-130 9C, C. a medium temperature and pressure leach at 130-220 9C, or D. a high temperature and pressure leach at 220-260 9C using sulphuric acid to generate a product slurry comprising a solid residue comprising PGMs and optionally one or more of gold, and silver and a sulphate pregnant leach solution (PLS) comprising one or more valuable metal sulphates in solution: Ni, Co, Cu, Cr, REE, Y, Sc, U, Th, Zn, Mn, Cd, Mo, V, Ti; and b) subjecting the solid residue from a) to a chloride leaching step in a chloride leaching medium wherein the chloride leaching medium comprises hydrochloric acid or saline brine in conjunction with an 10830973 1 - 24 oxidising agent including chlorine, hypochlorite or hydrogen peroxide, and the solid residue of step a) is leached under oxidising conditions, thereby to generate a chloride pregnant leach solution (PLS) comprising any one or more soluble PGMs and optionally any one or more soluble additional valuable elements in the chloride leaching medium. 2. The hydrometallurgical process of claim 1, further comprising a step of separating the solid residue from the metal sulphates in solution from the product slurry of step a) and then providing the resultant separated solid residue to step b). 3. The hydrometallurgical process of claim 2, wherein the separation is performed by any solid/liquid separation means. 4. The hydrometallurgical process of claim 3, wherein the means of solid/liquid separation is filtration. 5. The hydrometallurgical process of any one of claims 1 to 4, wherein the process further comprises a step of recovery of the any one or more PGMs and optionally any one or more additional valuable elements from the metal sulphates in solution and/or from the chloride leaching medium. 6. The hydrometallurgical process of claim 5, wherein the recovery is by means of any one or more of the techniques including solvent extraction, ion exchange, precipitation using hydroxides, carbonates or sulphides, electrowinning, or reduction. 7. The hydrometallurgical process of any one of claims 1 to 6, wherein the solid residue from step a) is subjected to a thermal treatment to produce a thermally treated calcine before subjecting it to step b). 8. The hydrometallurgical process of claim 7, wherein the thermal treatment is performed at 80 - 750 LC for up to 120 minutes, under oxidizing, neutral or reducing conditions, to remove volatile components from the solid residue. 9. The hydrometallurgical process of either claim 7 or 8, wherein the thermal treatment is performed at 400 - 650 LC for 10 to 30 minutes, under oxidizing, neutral or reducing conditions, to remove volatile components from the solid residue. 10. The hydrometallurgical process of any one of claims 7 to 9, wherein an additional thermal treatment is performed at 500 - 1000 LC for up to 120 minutes, under 10830973 1 - 25 oxidizing, neutral or reducing conditions, to condition the any one or more PGMs and optionally any one or more additional valuable elements to be soluble in chloride leaching medium. 11. The hydrometallurgical process of claim 10, wherein the additional thermal treatment is performed at 700 - 1000 2C for 30 to 120 minutes, under oxidizing, neutral or reducing conditions, to condition the any one or more PGMs and optionally any one or more additional valuable elements to be soluble in chloride leaching medium. 12.The hydrometallurgical process of either claims 10 or 11 wherein the thermal treatments are performed as individual steps of a sequential thermal treatment process, or as one combined step. 13. The hydrometallurgical process of any one of claims 1 to 12, wherein the product slurry of step a) is subjected to a hot acidic conditioning step or an atmospheric leach step to effect the removal of iron sulphates to the solution phase, and then subjected to step b). 14. The hydrometallurgical process of any one of claims 1 to 13, wherein the ore of (i) is initially processed by crushing, milling or is as-mined. 15. The hydrometallurgical process of claim 14, wherein the ore is subjected to a benefication step performed by methods including screening, sizing, classification, magnetic separation, electrostatic separation, dense media sepration, radiometric sorting, optical sorting, or gravity separation thereby to concentrate the any one or more PGMs and optionally any one or more additional valuable elements or reject a gangue component. 16. The hydrometallurgical process of any one of claims 1 to 15, wherein the weathered ore is a PGM-bearing ore that lies close to the surface and gives low PGM recovery when concentrated by flotation. 17. The hydrometallurgical process of any one of claims 1 to 16, wherein the solid residue of step a) is subjected, either directly or after a milling process, to a process to either concentrate the any one or more PGMs and optionally any one or more additional valuable elements or reject gangue components. 18. The hydrometallurgical process of claim 17, wherein the concentration or gangue rejection process is conventional screening, hydrocyclone or other classification means based on size, magnetic separation under either or both high and low 10830973 1 - 26 intensity magnetic fields, gravity separation, dense media sepration, radiometric sorting, optical sorting, or electrostatic separation. 19. 20. The hydrometallurgical process of any one of claims 1 to 19, wherein the sulphate PLS generated in step a) is subjected to separation and/or recovery of the one or more metal sulphates including Ni, Co, Cu, Cr, REE, Y, Sc, U, Th, Zn, Mn, Cd, Mo, V, Ti by means of techniques including solvent extraction, ion exchange, precipitation using hydroxides, carbonates or sulphides, electrowinning, or reduction.

21. The hydrometallurgical process of any one of claims 1 to 20, wherein the sulphate PLS generated in step a) is oxidised by sparging the reaction vessel with oxygen or by addition of a chemical oxidant including hydrogen peroxide or sulphur dioxide.

22. The hydrometallurgical process of claim 21, wherein the sparging is performed under either atmospheric or pressure conditions.

23. The hydrometallurgical process of any one of claims 1 to 22, further comprising a process of precipitation or crystallization of the sulphate PLS of step a) to produce a ferric or ferrous hydroxide or sulphate product while recovering sulphuric acid into a stream suitable for recycling.

24. The hydrometallurgical process of any one of claims 1 to 23, wherein a conventional atmospheric or low-pressure leach using sulphuric acid is applied directly to the product slurry comprising the sulphate PLS and solid residue of step a), thereby to remove excess iron sulphates from the solid residue into the sulphate PLS.

25. The hydrometallurgical process of claim 24, wherein the resultant sulphate PLS is then subjected to air- or oxygen-sparged ferric hydroxide precipitation under atmospheric or pressurized conditions for removal of excess iron sulphates.

26. The hydrometallurgical process of claim 25, wherein the solid residue has further been subjected to thermal treatment to generate a calcine, and the process further comprises removal of excess soluble iron from the thermally treated calcine.

27. The hydrometallurgical process of any one of claims 1 to 26, wherein the chloride leaching medium of step b) comprises iron chloride and is treated by any one or more methods including pressure, precipitation or crystallization, concentrated by 10830973 1 -27 evaporation, reverse osmosis, nanofiltration or other membrane technology, sparging/rectification or pyrohydrolysis to produce an iron-bearing product.

28. The hydrometallurgical process of any one of claims 1 to 27, wherein the chloride PLS is subjected to separation and/or recovery of the any one or more PGMs and optionally any one or more additional valuable elements by means of any one or more methods including solvent extraction, ion exchange, precipitation using hydroxides, carbonates or sulphides, electrowinning, or reduction.

29. The hydrometallurgical process of any one of claims 1 to 28, wherein the chloride leaching medium of step b) contains iron chloride and is treated by any one or more methods including pressure, precipitation or crystallization, concentrated by evaporation, reverse osmosis, nanofiltration or other membrane technology, sparging/rectification or pyrohydrolysis to produce an iron-bearing product.

30. The hydrometallurgical process of any one of claims 1 to 29, wherein the chloride leaching step comprises a chloride leaching medium having a pH of between about 2.5 and 7.5.

31. The hydrometallurgical process of any one of claims 1 to 29, wherein the chloride leaching step comprises a chloride leaching medium with a free acidity of between about 50 to 300 g/L HCI.

32. The hydrometallurgical process of any one of claims 1 to 31, wherein the chloride leaching step is performed by atmospheric or pressure autoclave leaching with saline brine under oxidising conditions.

33. The hydrometallurgical process of any one of claims 1 to 32, wherein the solid residue of step a) has been subjected to the thermal treatment of any of claims 7 to 12, thereby producing off-gases comprising sulphur and the process further comprises a step of recovering sulphur from the off-gases: a. by condensation as an elemental sulphur intermediate product; b. into a sulphuric acid intermediate product; or c. from dryer off-gases into a sulphide, polysulphide, polythionate, thiosulphate or similar intermediate product. Lifezone Limited Patent Attorneys for the Applicant/Nominated Person SPRUSON & FERGUSON 10830973 1