AU2016201662A1 - Method of extracting metal values from ores - Google Patents

Abstract

A process for selectively extracting scandium from scandium-containing material to produce a scandium-containing leach liquor, the process including the steps of (a) mixing the material with a fluxing salt to produce a mixture; (b) preheating the mixture in a furnace open to the air to a 5 first predetermined temperature in order to remove moisture and create a dry product; (c) heating the mixture in said furnace open to the air to a second predetermined temperature, said second predetermined temperature being higher than said first predetermined temperature, in order to expel one or more volatile gases and to generate a reaction between said material and said fluxing salt; (d) leaching the heated product from step (c) with an aqueous liquor under ambient 10 pressure or elevated pressure to produce a slurry and effect dissolution of soluble salts; (e) filtering the slurry from step (d) to obtain a solution containing scandium values; and (f) storing the solution for further treatment. 101 103 100 Limonitic Laterite Mixing -0.5mm Ore with powdery ammonium FLUX: Ammonium sulphate flux and Sulphur containing salt Preheating up to 102 3300C and Scrubbing or subsequent high diverting NH3(g) Precipitating 104 temperature and SO2 (g) to a ammonium heating chamber sulphate flux Stirred hot water 107 tank leach at ambient pressure Process water 109 111 110 Leach slurry Filtration Solids washing Clear filtrate fil Conventional alkall-added precipitation, SX or IX extraction

Description

METHOD OF EXTRACTING METAL VALUES FROM ORES 2016201662 16 Mar 2016

Field of the Invention 5 The present invention relates to a method of extracting scandium from scandium-bearing ores. More particularly, the invention relates to a method of selectively extracting scandium from scandium-bearing ores by baking with a solid form of ammonium and sulphur containing salt and leaching with water. The present invention also relates to a method of producing a scandium-containing leach liquor that contains substantially less impurities, so that the 10 subsequent steps of making high purity scandium compounds become simpler and less complicated.

Background of the Invention 15 Scandium (Sc) is a silvery-white metallic element with rather similar chemistry to titanium, yttrium and zirconium. It is often classified as a rare earth element, together with yttrium and the lanthanides. Its uniqueness lies in the fact that it has a low density of 2.985-3.000 g/cm3, but its melting point (1,541 °C) and boiling point (2,830 °C) are quite high. It readily reacts with mineral acids. At room temperature, it does not react with the oxygen in the 20 air, but when ignited, it burns readily in the air. Interestingly, scandium is the 23rd most common element in the sun, and it is believed that the combustion of scandium is the primary contributor to what we call sunlight.

The amount of scandium found in the earth's crust varies from 18 to 25 ppm (parts per million), which is comparable to the abundance of cobalt (20-30 ppm). Among the rocks, the 25 basic and ultrabasic rocks contain the highest amounts of scandium (pyroxenite = 70 ppm, gabbro = 30 ppm, basalt = 10 ppm SC2O3). Intermediate, acidic and alkaline rocks contain significantly less scandium than the basic and ultrabasic rocks. Scandium occurs mainly in the form of isomorphic substitution in vanadic titanomagnetite ores, rare-earth and uranium ores, tin ore, tungsten ore, bauxites and coal, and minerals like wolframite, columbite-tantalite, cassiterite, 30 ilmenite, zircon, beryl, garnet, muscovite and rare-earths. These ores and minerals contain scandium typically ranging from 5 to 100 ppm equivalent SC2O3. Scandium is an essential constituent of very few minerals (e.g., thortveitite and kolbeckite) and these minerals are very rare.

In the eastern part of Australia, running roughly north to south from Queensland to New 35 South Wales, there are a number of lateritic deposits that contain scandium alone as the major valuable element or in association with nickel, cobalt and platinum. Laterites are geologically weathered ultrabasic rocks where the primary scandium occurs mainly in pyroxenes in the form of solid solutions typically ranging from 50 to 100 ppm SC2O3. The weathering concentrates scandium in goethite, limonite and clay minerals where the ore grade sometimes reaches as high 40 as 1500 ppm and if physically upgraded over 2000 ppm SC2O3.

The worldwide demand for scandium compound is relatively low and is in the order of several tonnes of scandium oxide. However, it is expected to grow significantly in the future. The principal uses for scandium are in high-strength aluminum ± magnesium alloys, high-intensity metal halide lamps, electronics, laser research and solid oxide fuel cells. The use of 45 scandium as an alloying material with Al, Mg, Ti, Cu and Zr is well established. For example, when alloyed with aluminium, scandium produces one of the lightest, strongest, most durable, weldable and corrosion resistant metal in the world. The addition of scandium to aluminium limits the excessive grain growth, where the precipitated AfiSc forms smaller crystals than are formed in other aluminium alloys. Scandium in aluminium alloy Easton Sc7005 gives a weld 50 without heat cracks and improves both thermal and corrosion resistance. 2

The aluminium-scandium alloys are often used by the aerospace industry, especially in the making of Russian fighter planes such as MiG-21 and MiG-29. In the A1 alloys, the amount of scandium used is between 0.1% and 0.5% of scandium. Aerospace could be the biggest demand driver because anything that adds strength while saving weight is invaluable. Airbus 5 estimates that aircraft made from a welded aluminium scandium zirconium (AlScZr) alloy would be 15% lighter and 15% cheaper to build compared to present materials [RichardKarn, 2011. Australian scandium coidd create new market, Critical Metals Report, Mining.com, August 9, 201 ij. 2016201662 16 Mar 2016

For the time being, there is no mine in the world specifically producing a scandium ore. The existing scandium production is as byproduct from some operating mines. 10 Information about the production of scandium is scarce. Its extraction is proposed as a byproduct of other metal recovery processes. Among those extraction from vanadic titanomagnetite ores, uranium ores, tin tungsten ores, bauxites and coal, and mineral concentrates like wolframite, columbite-tantalite, cassiterite, ilmenite, zircon and rare-earths are suggested. In these applications, an enriched scandium product is extracted by employing hydrometallurgical 15 methods such as ion exchange, solvent extraction or multistage precipitation and re-leaching techniques.

In US Patent No. 4816233, a hydrometallurgical method for recovering scandium from a tungsten ore residue is described. The method involved: (1) leaching of the tungsten residue with an acidic solution containing a reducing agent to obtain a solution containing mainly the 20 scandium, iron, manganese and other impurities, and (2) contacting the solution with an ion exchange resin at a pH of 1.9 to 2.1 where scandium is adsorbed onto the ion exchange resin.

The loaded resin is then washed with dilute acid to remove any base metals and rare earth metals on the resin, without removing scandium from the resin.

In US Patent No. 5787332 (Black et.al) there is claimed a method for recovering 25 constituents from a composite material which comprises tantalum oxides and niobium oxides, and other metals including scandium digested in hydrogen fluoride to form metal fluoride compounds of said composite material, and reacting the metal fluoride compounds of said composite material with sulphuric acid to form a variety of metal sulphate compounds, including scandium sulphate. This process requires the handling of large amounts of highly dangerous 30 acids like hydrogen fluoride (HF) and sulphuric acid to extract a small amount of scandium as a byproduct.

In US Patent Application No. 20120207656 (Duyvesteyn) there is claimed a method for extracting scandium values from scandium-containing ores, the method comprising: providing an ore which contains scandium; mixing the ore with an acid; baking the ore; and leaching 35 scandium from the baked ore under conditions which are not conducive to the formation of a scandium containing precipitate. However, in his claims Willem P.C. Duyvesteyn did not specify the baking temperature nor gave any details about the performance of their method.

Our investigations involving the baking of a lateritic ore with concentrated sulphuric acid illustrated that although this is a good approach, it required a bake mixture of at least 1:1 wt/wt 40 acid per tonne of ore to extract a commercially viable amount. Additionally, the leach step, which followed the baking step, extracted substantial amounts of iron, titanium, magnesium, manganese, silica and lesser amounts of other impurities. The baking with sulphuric acid generated copious amounts of poisonous SO2 gas.

In general, the removal of impurities from the leach liquor during subsequent solvent 45 extraction (or ion exchange) stage becomes costly and time consuming. Among the impurities, especially the removal of titanium is often not complete and its stripping from the solvent requires very strong acids such as 6.0N hydrochloric acid.

In another example, in US Patent No. 5039336 (Feuling) it is suggested the use of a fluid bed chlorination process to extract scandium from zircon ores as a byproduct. This process 50 involves feeding the zircon sand to a fluidized bed chlorinator and heating it to about 1000° C. This produces vaporous chlorides of zirconium, silicon, iron and leaves behind a solid residue 3 that contains scandium chloride besides other alkali chlorides, uranium and thorium. The separation of scandium from radioactive compounds requires skills and significant effort. There is no known commercial chlorinator designated to chlorinate zircon sands and to extract scandium. In fact, it is well known that not every zircon sand contains appreciable amounts of 5 scandium. Furthermore, chlorination is a high OH&S risk processing method and separating a pure scandium compound from other compounds of zircon is not a simple and straightforward process. 2016201662 16 Mar 2016

Our investigations as described in one of the examples below show that direct acid leaching dissolves large amounts of undesirable impurities into the pregnant leach liquor. This 10 makes the separation of scandium from the other dissolved metals difficult as the removal of these impurities introduces a number of additional steps to obtain a substantially pure scandium-containing solution. The removal of such impurities requires additional effort, energy and expense.

High pressure sulphuric acid leaching in autoclaves improves the scandium extraction 15 rate, and reduces the processing time as well minimizes the extraction of undesirable components but leaves behind a residue stuck on the walls and at the bottom of the autoclave. This is understandable as the lateritic ores are made of 30 to 60 wt% clay. The crust formation in the autoclaves could choke the material flow and create serious OH&S problems.

Furthermore, acid baking or direct acid leaching in tanks or pressure vessels requires the 20 supply of large volumes of hazardous acids like concentrated sulphuric acid. This creates logistic, environmental and safety problems. Transportation of hazardous material from port to inland is more costly than non-hazardous material and has to follow stringent rules and regulations. Building an acid production plant on site requires additional capital investment which may not be justifiable as the present worldwide scandium consumption is relatively low. 25 There is thus a need for a process that avoids the use of large amounts of hazardous liquid acids, selectively recovers scandium values with minimum recovery of metal impurities, allows the recovery and recycling of at least part of the fluxing salt and overcomes, or at least alleviates, one or more of the disadvantages of the prior art processes. 30 Summary of the Invention

According to an aspect of the invention, there is provided a process for selectively extracting scandium from scandium-containing material to produce a scandium-containing leach liquor, the process including the steps of: 35 (a) mixing the material with a fluxing salt to produce a mixture; (b) preheating the mixture in a furnace open to the air to a first predetermined temperature in order to remove moisture and create a dry product; (c) heating the mixture in said furnace open to the air to a second predetermined temperature, said second predetermined temperature being higher than said first predetermined 40 temperature, in order to expel one or more volatile gases and to generate a reaction between said material and said fluxing salt; (d) leaching the heated product from step (c) with an aqueous liquor under ambient pressure or elevated pressure to produce a slurry and effect dissolution of soluble salts; (e) filtering the slurry from step (d) to obtain a solution containing scandium values; 45 and (f) storing the solution for further treatment.

Accordingly, the present invention, in an embodiment, provides a method for extracting scandium values from scandium-containing ores. The method comprises (1) mixing the 50 scandium containing ore with a flux at a certain ratio; (2) baking the ore in a furnace; (3) leaching scandium from the baked ore using water in a heated and stirred vessel under ambient 4 pressure, (4) and diverting the gaseous effluents of heating to a scrubber or a chamber to reconstitute the ammonium sulphate used in mixing; and (5) using the reconstituted ammonium sulphate in a second cycle of baking. 2016201662 16 Mar 2016

In another embodiment, the present invention provides a method for extracting scandium 5 values from scandium-containing lateritic ores. The method comprises (6) mixing the ore which contains scandium with a flux at a certain ratio; (7) baking the ore in a furnace; (8) leaching scandium from the baked ore using water in a heated and stirred pressure vessel (9) separating the liquid from solid leach residue by filtration; (10) saving the filtrate fraction of filtration for further treatment and (11) mixing the washed solids fraction of filtration with finely ground 10 limestone and disposing it to a land fill on a mine site.

In another embodiment, the present invention provides a scandium loaded solution for further treatment by hydrometallurgical methods such as ion exchange, solvent extraction or multistage precipitation and re-leaching techniques. 15 Brief Description of the Drawings

Preferred embodiments of the invention will hereinafter be described, by way of example only, with reference to the drawings in which:

Figure 1 is a flow diagram showing the process of extracting scandium values from 20 scandium-containing ore under ambient pressure; and

Figure 2 is a flow diagram showing the process of extracting scandium values from scandium-containing ore using a pressure vessel. 25 Detailed Description of the Preferred Embodiments

This invention relates to a process where the finely divided scandium containing ore is mixed with a flux at a certain ratio and heated to a specific preheating temperature and then heated to a higher temperature in either one or two steps. The heating could be done at ambient 30 pressure or in a pressurized vessel. The process further involves the leaching of the heated product with water. The flux or fluxing agent used is ammonium and sulphur-containing salts.

In these embodiments, baking the ore with ammonium and sulphur containing salts is proposed to aid extraction of scandium from scandium bearing ores. The most common ammonium and sulphur containing salt is ammonium sulphate. The use of ammonium and 35 sulphur containing salts to aid extraction of scandium from scandium bearing ores is not known.

Ammonium sulphate decomposes upon heating above 250 °C, first forming ammonium bisulphate, Heating at higher temperatures results in decomposition into ammonia, nitrogen, sulphur dioxide, and water. Ammonium sulphate, which is a relatively cheap solid product, is used mainly by the fertiliser industry to supply nitrogen and sulphur to plants, especially in arid 40 countries where the soil is alkaline and lacks sulphur and nitrogen.

Ammonium sulphate is produced from ammonia - a byproduct of coke ovens - and sulphuric acid according to: 2 NH3 + H2S04 -+ (NH4)2S04

Dry, powdered ammonium sulphate is produced by spraying sulphuric acid into a 45 reaction chamber filled with ammonia gas. The heat of reaction evaporates the water and leads to the formation of a powdery product.

Ammonium sulphate can also be produced from crushed pyrite (-6+10 mesh B.S) by reacting it with ammonia, steam and air at temperature of 450°C and pressure of 30 atmospheres 5 for 4 hours, rs. K. Bhattacharwa. S. Dutta (1974). PRODUCTION OF AMMONIUM SULFATE FROM PYRITE, Ind. Eng. Chem. Process Des, Dev., 1974,13 (3),pp. 215-218, DOI: 10.1021/126005la004]. 2016201662 16 Mar 2016

The other ammonium sulphate production method is from mixing relatively cheap and finely divided gypsum (CaS04'2H20) with ammonium carbonate solution. Ammonium 5 carbonate is produced often by sequestration of carbon dioxide into aqueous ammonia. In this case, calcium carbonate precipitates as a solid and ammonium sulphate remains in the solution according to: (NH4)2.C03 + CaS04 -> (NH4)2S04 + CaC03

The above information shows that ammonium sulphate is made from relatively cheap 10 material. It is produced as a byproduct of large industries at a reasonably low cost. 1. Ambient Pressure Leach of Baked Products

Referring to Figure 1, there is shown a process flow diagram according to a first 15 embodiment (100) that summarizes some of the general steps in a particular, non-limiting embodiment of a process for recovering scandium values from feedstock in accordance with the teachings herein. The process begins at step (101) with the mining and crushing of the limonitic laterite ore to a size to pass at least a 2 mm sieve, in this case 0.5 mm sized ore particles. A fluxing agent of ammonium and sulphur containing salt is prepared at step (102). The ore from 20 step (101) is then mixed at step (103) with the fluxing agent from step (102) in powder form. At step (104) the mixture is then preheated first to a temperature up to 330 °C to expel moisture and some of the ammonia, followed by heating to a higher temperature, in one or two steps up to 499 °C or to 800 °C, in order to expel other volatile gases. The volatile gases NH3 and S02 are then scrubbed at step (105) to a suitable liquid or diverted to a precipitation chamber at step (106), 25 where the ammonium and sulphur-containing salt flux is precipitated. The solid form of the precipitated flux is then recycled back to step (103) to react with fresh feedstock or laterite ore.

The heated ore from step (104) is then subjected to a water leaching process at step (107) in a stirred hot water tank where the leaching takes place at ambient pressure, to extract the scandium values from it in a leach slurry at step (108). The leachate from step (108) is then 30 separated from the solids through a solid liquid separation or filtration process at step (109) to obtain a clear liquid filtrate at step (113). The solid residue at step (110), resulting from the filtration at step (109) is washed and filtered. The wash water at step (111) is recycled to be used as process water in step (107). The solids from step (110) are sent to a mine site for disposal as landfill at step (112). The scandium-loaded liquid fraction (pregnant leach liquor) resulting at 35 step (113) is sent for further treatment to make a high purity (>99.0%) Sc203, by conventional alkali-added precipitation, SX or IX extraction at step (114).

In this first embodiment, the process for selectively extracting scandium from a scandium-containing material or feedstock to produce a scandium-containing leach liquor involves the following broad steps: 40 (a) mixing the material with a fluxing salt to produce a mixture; (b) preheating the mixture in a furnace open to the air to a first predetermined temperature in order to remove moisture and part of ammonia, and create a dry product; (c) heating the mixture in said furnace open to the air to a second predetermined temperature, said second predetermined temperature being higher than said first predetermined 45 temperature, in order to expel one or more volatile gases and to generate a reaction between said material and said fluxing salt; (d) leaching the heated product from step (c) with an aqueous liquor under ambient pressure to produce a slurry and effect dissolution of soluble salts; (e) filtering the slurry from step (d) to obtain a solution containing scandium values; 50 and 6 (f) storing the solution for further treatment. 2016201662 16 Mar 2016

Also in the first embodiment: the leaching can be done in conventional stirred reactors operating at ambient pressures 5 and the slurry heated to the boiling point of the slurry; the leaching temperature for the conventional stirred reactors operating at ambient pressures can be at the boiling point of the slurry or at 90-95 °C to minimise the loss of water to evaporation. 10 2. High Pressure Leach of Baked Products

Referring to Figure 2, there is shown a process flow diagram (200) according to a second embodiment that summarizes some of the general steps in a particular, non-limiting embodiment of a process for recovering scandium values from feedstock in accordance with the teachings 15 herein. The process begins at step (201) with the mining and crushing of the limonitic laterite ore to a size to pass at least a 2 mm sieve, in this case 0.5mm sized ore particles. A fluxing agent of ammonium and sulphur containing salt is prepared at step (202). The ore from step (201) is then mixed at step (203) with the fluxing agent from step (202) in powder form. At step (204) the mixture is then pre heated first to a temperature up to 330 °C to expel moisture and some of the 20 ammonia followed by heating to a higher temperature in one or two steps, up to 499 °C or to 800 °C, in order to expel other volatile gases. The volatile gases NH3 and SO2 are then scrubbed at step (205) to a suitable liquid or diverted to a precipitation chamber at step (206), where the ammonium and sulphur-containing salt flux is precipitated. The solid form of the precipitated flux is then recycled back to step (203) in order to react with fresh feedstock or laterite ore. 25 The heated ore from step (204) is then subjected to a pressurized water leaching process in a pressure vessel or autoclave at about 250 °C at step (207) to extract the scandium values into water. This is done without any residence time. A leach slurry results at step (208). The leachate from step (208) is then separated from the solids through a solid liquid separation or filtration process at step (209) to obtain a clear liquid filtrate at step (213). The solid residue at 30 step (210), resulting from the filtration at step (209) is washed and filtered. The processed wash water at step (211) is recycled to be used as process water in step (207). The solids from step (210) are sent to a mine site for disposal as landfill at step (212). The scandium-loaded liquid fraction (pregnant leach liquor) resulting at step (213) is sent for further treatment to make a high purity (>99.0%) SC2O3, by conventional alkali-added precipitation, SX or IX extraction at step 35 (214).

In this second embodiment, the process for selectively extracting scandium from a scandium-containing material or feedstock to produce a scandium-containing leach liquor involves the following broad steps: (al) mixing the material with a fluxing salt to produce a mixture; 40 (bl) preheating the mixture in a furnace open to the air to a first predetermined temperature in order to remove moisture and part of ammonia, and create a dry product; (cl) heating the mixture in said furnace open to the air to a second predetermined temperature, said second predetermined temperature being higher than said first predetermined temperature, in order to expel one or more volatile gases and to generate a reaction between said 45 material and said fluxing salt; (dl) leaching the heated product from step (cl) with an aqueous liquor under elevated pressure to produce a slurry and effect dissolution of soluble salts; (el) filtering the slurry from step (dl) to obtain a solution containing scandium values; and 50 (fl) storing the solution for further treatment. 7

In the second embodiment, the leaching step can be performed in a pressure vessel and the slurry heated at a temperature between 110 °C and 500 °C, preferably between 200 °C and 300 °C, and most preferably at 250 °C. The leaching step can be performed in a pressure vessel for between 0 and 10 hours, more preferably for between 0 and 2 hours after the second 5 predetermined temperature (target temperature) and pressure has been reached. 2016201662 16 Mar 2016

In both of the embodiments, during the heating of the mixture to a higher temperature at steps (104), (204) up to 499 °C, NH3 is expelled. The sulphate fraction of ammonium sulphate reacts with goethite (FeO(OH) to produce iron sulphate which is water soluble. In a similar fashion, along with iron sulphates, Sc2(SC>4)3 and various other sulphated metals also form. 10 Further heating to between 500 °C and 800 °C expels part of the sulphurous gases, SO2 and SO3, The ammonia (NH3) at steps (105), (205) can be scrubbed into dilute H2SO4 to produce (NH4)2S04 which can be reused in the process at steps (103), (203). In the process of heating to 499 °C or higher to 800 °C, both the expelled SO2 and SO3 can be used to react with (NFL^OFl to produce (NFL^SC^ which can again be reused in the process at steps (103), (203). 15 In both embodiments, the following can apply in the process: the leach liquor preferably contains substantially less impurities, including at least one or more of iron, aluminium, magnesium, calcium, sodium, potassium, silica, titanium and manganese; collecting the expelled one or more volatile gases in a chamber to regenerate a portion of 20 the fluxing salt; washing a solid residue from step (f) or step (fl), selectively separating the solids from the residue, mixing the solids with an amount of finely ground calcium carbonate and disposing the resultant mixing of solids with said calcium carbonate; collecting wash water from the washing step and storing the wash water for reuse as 25 process water; the material includes any one of the ore types such as an ore, an ore concentrate or a mineral concentrate containing scandium; the material includes a scandium laterite ore or a material derived from limonite and/or saprolite zones of the laterite and/or fresh or altered serpentinite or pyroxenite rocks; 30 the material is ground to a size of 5 mm or less alone or with said fluxing salt; the fluxing salt is chosen from one or more of sulphate or sulphide or sulphite salts of ammonium such as ammonium sulphate, ammonium persulphate, ammonium peroxydisulphate, ammonium bisulphate, ammonium monosulphide, ammonium hydrosulphide, ammonium pentasulphide, ammonium sulphite and ammonium hydrogen sulphite or any combination 35 thereof;

Preferably, the fluxing salt is chosen from one or more of the ammonium salts of sulphate or bisulphate or persulphate. Most preferably, the fluxing salt is ammonium sulphate or ammonium bisulphate; the mixing of the material with the fluxing salt is conducted at a wt/wt mixing ratio of 40 1.0:10 to 10:1.0. Most preferably, the mixing of the material with the fluxing salt is conducted at a wt/wt mixing ratio of 1.0:1.0; the first predetermined temperature is at least 100°C, preferably between 100 °C and 330 °C, more preferably between 250° and 330 °C; the second predetermined temperature is below 800 °C, preferably between 330 °C and 45 700 °C; the heating to the second predetermined temperature is performed in a stepwise manner, preferably in two steps. The first step preferably is heating the mixture in step (c) or step (cl) to between 250 °C and 499 °C. The second step is preferably heating the mixture in step (c) or step (cl) to above 500 °C, preferably to between 500 °C and 800 °C, more preferably to between 550 50 °C and 700 °C; 8 the heating is conducted for a period of time of any one of up to 80 hours, for more than 1 hour or for more than 0.1 hours; 2016201662 16 Mar 2016 the furnace is any one of the types such as static, rotating, muffle or gas injected, electrical or combustion; 5 heating the mixture to the predetermined temperature generates ammonia, steam, sulphur dioxide and sulphur trioxide containing gaseous byproducts that are diverted to a chamber for production and recycling of ammonium sulphate; the chamber is injected with a sufficient amount of moisture and air to initiate the reaction: 10 4NH3(g) + 2S02(g) + 2H20(aq.) + 02(g) = 2(NH4)2SC>4(s) the Gibbs free energy of which is AG = negative 159.886 kcal at 20°C temperature; gaseous ammonia that is generated as a result of heating to the predetermined temperature is scrubbed into a first solution of dilute sulphuric acid; 15 gaseous sulphur dioxide gas that is generated as a result of heating to the second predetermined temperature is scrubbed into a second solution of dilute ammonium hydroxide, ammonium carbonate or ammonia gas; further include joining the first and second solutions and subjecting the joined solution to a concentration treatment to obtain a solid form of the fluxing salt; 20 further include recycling the solid form of the fluxing salt to prepare a new batch of mixture with the material in step (a) or step (al); step (d) or step (dl) is performed in heated and stirred reactors under ambient pressure or elevated pressure; the leaching in step (d) or step (dl) of the product with the aqueous liquor is capable of 25 being done in the presence of a scandium complexing agent. The scandium complexing agent is a compound that selectively combines with dissolved scandium and prevents it being reprecipitated. The scandium complexing agent is any compound that selectively combines with scandium ions in the leach liquor. The scandium complexing agent is preferably diglycolic acid or similar compounds. The mass of the scandium complexing agent added to said leach solution 30 is 0.01 to 10.0% of the baked ore by weight; the aqueous liquor is any one of plain water, recycled process water, reverse osmosis or demineralised water; the solid content of the slurry is between 5% and 55%, preferably between 15% and 35%; further including Stirling the slurry at a rate within the range of 2 rpm to 5000 rpm, 35 preferably between 100 rpm and 1000 rpm, in order to provide a sufficient contact between the aqueous and solid phases.

Compared with ambient pressure leaching of the baked products, the autoclave pressure leaching extracted approximately the same amount of scandium. However, advantages of the pressure leaching were: (i) the leaching process consumed substantially less time, (ii) the leach 40 liquor contained less iron and titanium impurities, (iii) the treatment produced a 100% of tapable slurry and (iv) the walls of the emptied autoclave were clean and contained no sticky material, and (v) it allowed operation at higher pulp densities.

The invention is described in detail using the following examples. The person skilled in the art will appreciate from these examples that the optimal operating conditions may vary from 45 one implementation to another, and may depend on a variety of factors, It should therefore be understood that the following descriptions are illustrative only and should not be taken in any way as a restriction on the generality of the invention described above and in the claims of this invention. 50 9 2016201662 16 Mar 2016 10

EXAMPLES

Example 1: This example shows 72.5% of the scandium would be extracted when 500g as-received -1.0 mm composite feedstock was mixed with an equal amount of ammonium sulphate 5 and preheated; first to 300 °C for 1 hour and then to 450 °C at a rate of 15°C/h (lOh) and resided at that temperature for 14 hours. The water leaching was done under ambient pressure in a heated and stirred vessel for 24 hours at 95 °C. Table 1 shows the amount of iron and titanium dissolved as a result of this treatment were low.

Table 1:

Single step bake and water leach: 24 hours baking from 300 °C to 450 °C at a rate of 15C/h and 14h dwell time and standard water leach. % Sc extracted % Fe extracted % Ti extracted 72.5 18.0 4.11

Example 2: This example is given to show that a significant improvement is gained when the baked sample in Example 1 was further heated to 590 °C at a rate of 14°C/h (lOh) and resided at 15 that temperature for 6 hours. Table 3 shows 85.6% of the scandium was extracted into the leach water. The water leaching was done under ambient pressure in a heated and stirred vessel for 24 hours at 95 °C. The amount of iron and titanium dissolved as a result of this treatment were generally low although higher than those given in Example 2.

Tab le 2: % Sc % Fe % Ti 1st 24 hours bake at 450 °C then 2nd 16 hours bake at 590 °C and extracted extracted extracted standard water leach. 85.6 21.5 8.5

Example 3: This example is given to show a marginal improvement in scandium extraction to 87.5% when the first baking from 300 to 450 °C was performed longer (30 hours plus a dwell 25 time of 14 hours). The rate of the first heating was 5 °C per hour, The rate of the second heating was the same as in Example 2. The water leaching was done under ambient pressure in a heated and stirred vessel for 24 hours at 95 °C.

Tab Le 3: 1st 45 hours bake at 450 °C then 2nd 16 hours 590 °C bake and % Sc extracted % Fe extracted % Ti extracted standard water leach. 87.5 28.3 18.2

Table 3 shows that the increase in the scandium extraction rate was only marginally higher when the baking time was longer. Table 3 also shows the amount of iron and titanium dissolved as a result of this treatment was higher than those given in Examples 1 and 2. 10 30

Example 4: This example is given to show a further deterioration in scandium extraction if the baking was done at a fast rate. In this case, 500g composite sample mixed with the same amount of ammonium sulphate was baked to 450 °C for 12 hours, at an average rate of 37.5 °C per hour, 5 and leached with water at 95 °C for 24 hours. Table 4 shows the amount of scandium dissolved was 55.1% of the total scandium in the feed. The amount of dissolved iron and titanium was nil.

Table 4:

Fast heating at a rate of 37.5 °C /h % Sc extracted % Fe extracted % Ti extracted and standard water leach 55.1 nil nil 2016201662 16 Mar 2016 10

Example 4 shows that fast baking with ammonium sulphate to 450 °C is detrimental for the dissolution of scandium during the subsequent water leach stage.

Example 5: This example shows almost no scandium is extracted when lOOg as-received -1.0 15 mm composite feedstock was leached under ambient pressure in a heated and stirred vessel for 24 hours at 95 °C with 400 mL water containing lOOg of dissolved ammonium sulphate. Table 5 shows the amount of iron and titanium dissolved as a result of this treatment were also negligible. 20 Table 5: % Sc extracted % Fe extracted % Ti extracted Direct Leach with Ammonium Sulphate 4.3 3.2 4.2

Example 5 shows that baking with flux is essential for any substantial dissolution of scandium during the subsequent water leach stage. 25

Example 6: This example is given to show further improvement in scandium extraction to 88.2% when the feed was ground to a size of 5 microns. Previously mixed with ammonium sulphate and baked at 480 °C for 24 hours with 1:1 wt/wt ammonium sulphate. The baked sample was leached with water at 95 °C for 24 hours. Table 6 shows the amount of iron and 30 titanium dissolved as a result of this treatment were reasonably low.

Table 6:

Micronized feed baked with flux and heated at 480°C for24 hand standard leached with water. % Sc extracted % Fe extracted % Ti extracted 88.2% 24.6 9.3 11

Example 7: This example is given to show the benefit of extracting more scandium when a small amount of scandium complexing agent, in this case diglycolic acid, is added to the leach liquor. Other complexing agents that can be used include carboxylic acid substituted with a hydroxy group. In this test, 350g of pre-ground sample feedstock is mixed thoroughly with 175g 5 of ammonium sulphate and heated first at 330 °C for 1 hour then continued heating in the same furnace to 480 °C at a rate of 10 °C per hour. The baked product kept at this temperature for further 8 hours and cooled down to room temperature in a switched off mode. This baked product is then split into two equal fractions (A and B) for water leaching. 2016201662 16 Mar 2016 10 Leach A: 201.74g of the baked product is leached with 605g of reverse osmosis (RO) water at 95 °C for 24 hours by stirring at a rate of 500 r.p.m. The pale greenish colour filtrate is analysed by ICP (Inductively Coupled Plasma) method.

Leach B: 201,75g of the baked product is leached with 605g of reverse osmosis (RO) water 15 containing 4.0 g (0.66% wt/vol%) diglycolic acid. The leaching was carried out at 95 °C for 24 hours by stirring at a rate of 500 r.p.m. The pale greenish colour filtrate is analysed by ICP method.

The filtrate analyses for extracted scandium, iron and titanium are shown in Table 7.

Table 7:

Sc Fe Ti mg/L g/L mg/L Leach A with RO water 110 0.17 <0.5 Leach B with RO water+diglycolic acid 130 0.37 <0.5 20

Example 8: This example is given for comparison and to show leaching using only sulphuric 25 acid extracts less scandium and higher amounts of iron and titanium (Table 8) when the leaching was earned out for 24 hours at 95 °C and using 806 g concentrated sulphuric acid per kilogram of sample diluted to 3.98 moles with water.

Table 8: % Sc % Fe % Ti No Baking: Direct Sulphuric Acid Leach at 33.1% Solids at 95 °C for 24 h. 806 g concentrated sulphuric acid per kilogram of extracted extracted extracted sample is used. 63.7 53.4 19.8 30

Under similar conditions when 1:1 wt:wt sulphuric acid to ore ratio was used scandium extraction increased to 85.8% but dissolved iron and titanium contents are also increased to 84.2 and 61.5%, respectively. 35

Example 9

The next example is given to demonstrate that water leaching of the baked ore may be done in a pressure vessel. In Example 9, -1.00 mm size composite ore was mixed with the same amount of 40 ammonium sulphate and heated at 480 °C for 24 hours and then cooled. 421.4g of this sample 12 2016201662 16 Mar 2016 in Table 9.

Sample is added with 1:1 wt:wt with flux and baked at 480 °C /24 h. The baked sample is then leached in an autoclave at 250 °C . was leached with 1685g water in an autoclave heated to 250 °C. The pressurised heating was stopped as soon as the temperature reached 250 °C and cooling switched on. The slurry was filtered and filter cake washed. The dry residue weighed 259g. The extracted amounts are shown 5

Table 9: % Sc extracted % Fe extracted % Ti extracted 86.8 2.7 nil

Example 10 10

The next example is given to demonstrate that pressure vessel leaching of the ore without baking is not as effective as the pressure leaching of the baked ore. In this example, 391g of-1.00 mm size composite ore was mixed with an equal amount of ammonium hydrogen sulphate and 1564g H2O in an autoclave, and heated to 250°C. The pressurised heating was stopped as soon as the 15 temperature was reached 250°C, and cooling switched on. The slurry was filtered and the filter cake washed. The dry residue weighed 435g. The extracted amounts are shown in Table 10.

Table 10: % Sc % Fe % Ti Autoclave leach on as-received sample. extracted extracted extracted 45.5 4.7 1.3 20

Example 10 shows that baking with flux is also essential for any substantial dissolution of scandium during the autoclave leaching stage.

The composition of the composite (head/feed) sample used in tests described above is shown in 25 Table 11. The remaining 12.1% comprised water.

Table 11: ai2o3 (%) CoO (%) Cr203 (%) Fe203 (%) MgO (%) Mn30 4 (%) NiO (%) Sc/Sc203 (ppm) Si02 (%) Ti02 (%) Sum (%) AVERAGE 15.2 0.092 0.404 56.0 0.61 0.872 0.129 620/954 12.1 1.42 87.9 30

Claims (39)

1. The claims defining the invention are as follows:

   1. A process for selectively extracting scandium from scandium-containing material to produce a scandium-containing leach liquor, the process including the steps of: (a) mixing the material with a fluxing salt to produce a mixture; (b) preheating the mixture in a furnace open to the air to a first predetermined temperature in order to remove moisture and create a dry product; (c) heating the mixture in said furnace open to the air to a second predetermined temperature, said second predetermined temperature being higher than said first predetermined temperature, in order to expel one or more volatile gases and to generate a reaction between said material and said fluxing salt; (d) leaching the heated product from step (c) with an aqueous liquor under ambient pressure or elevated pressure to produce a slurry and effect dissolution of soluble salts; (e) filtering the slurry from step (d) to obtain a solution containing scandium values; and (f) storing the solution for further treatment.
2. 2. The process according to claim 1 further including collecting said expelled one or more volatile gases in a chamber to regenerate a portion of said fluxing salt.
3. 3. The process according to claim 1 or claim 2 further including washing a solid residue from step (f), selectively separating the solids from the residue, mixing the solids with an amount of finely ground calcium carbonate and disposing the resultant mixing of solids with said calcium carbonate.
4. 4. The process according to claim 3 further including collecting wash water from said washing step and storing said wash water for reuse as process water.
5. 5. The process according to any one of the preceding claims wherein the leach liquor contains one or more of the following impurities: iron, aluminium, magnesium, calcium, sodium, potassium, silica, titanium and manganese.
6. 6. The process according to any one of the preceding claims wherein the material includes any one of the ore types such as an ore, an ore concentrate or a mineral concentrate containing scandium.
7. 7. The process according to any one of claims 1 to 6 wherein the material includes a scandium laterite ore or a material derived from limonite and/or saprolite zones of the laterite and/or fresh or altered serpentinite or pyroxenite rocks.
8. 8. The process according to any one of the preceding claims wherein said material is ground to a size of 5mm or less alone or with said fluxing salt.
9. 9. The process according to any one of the preceding claims wherein said fluxing salt is chosen from one or more of sulphate or sulphide or sulphite salts of ammonium such as ammonium sulphate, ammonium persulphate, ammonium peroxydisulphate, ammonium bisulphate, ammonium monosulphide, ammonium hydrosulphide, ammonium pentasulphide, ammonium sulphite and ammonium hydrogen sulphite or any combination thereof.
10. 10. The process according to any one of the preceding claims wherein said fluxing salt is chosen from one or more of the ammonium salts of sulphate or bisulphate or persulphate.
11. 11. The process according to any one of the preceding claims wherein the fluxing salt is ammonium sulphate or ammonium bisulphate.
12. 12. The process according to any one of the preceding claims wherein the mixing of the material with the fluxing salt is conducted at a wt/wt mixing ratio of 1.0:10 to 10:1.0.
13. 13. The process according to claim 12 wherein the mixing of the material with said fluxing salt is conducted at a wt/wt mixing ratio of 1.0:1.0.
14. 14. The process according to any one of the preceding claims wherein said first predetermined temperature is at least 100°C, preferably between 100 °C and 330 °C, more preferably between 250° and 330 °C.
15. 15. The process according to any one of the preceding claims wherein said second predetermined temperature is below 800 °C, more preferably between 330 °C and 700 °C.
16. 16. The process according to any one of the preceding claims wherein the heating to the second predetermined temperature is performed in a stepwise manner, preferably in two steps.
17. 17. The process according to claim 16 wherein the first step is heating the preheated mixture in Claim 14 to a temperature below 500 °C, preferably between 330 °C and 499 °C.
18. 18. The process according to claim 16 wherein the second step is heating the preheated mixture in Claim 14 to a temperature above 500 °C, preferably to between 500 °C and 800 °C, more preferably to between 550 °C and 700 °C.
19. 19. The process according to any one of claims 15 to 18, wherein said heating is conducted for a period of time of any one of up to 80 hours, for more than 1 hour or for more than 0.1 hours.
20. 20. The process according to any one of the preceding claims wherein said furnace is any one of the types such as static, rotating, muffle or gas injected, electrical or combustion.
21. 21. The process according to any one of the preceding claims wherein heating generates ammonia, steam, sulphur dioxide and sulphur trioxide containing gaseous byproducts that are diverted to a chamber for production and recycling of ammonium sulphate.
22. 22. The process according to claim 21 wherein said chamber is injected with a sufficient amount of moisture and air to initiate the reaction: 4NH3(g) + 2S02(g) + 2H20(aq.) + 02(g) = 2(NH4)2SC>4(s) the Gibbs free energy of which is AG = negative 159.886 kcal at 20°C temperature.
23. 23. The process according to claim 21 or claim 22 wherein gaseous ammonia generated as a result of heating is scrubbed into a first solution of dilute sulphuric acid.
24. 24. The process according to any one of claims 21 to 23 wherein gaseous sulphur dioxide gas generated as a result of heating is scrubbed into a second solution of dilute ammonium hydroxide, ammonium carbonate or ammonia gas.
25. 25. The process according to claim 24 further including joining said first and second solutions and subjecting the joined solution to a concentration treatment to obtain a solid form of said fluxing salt.
26. 26. The process according to claim 25 further including recycling said solid form of said fluxing salt to prepare a new batch of mixture with said material in step (a).
27. 27. The process according to any one of the preceding claims wherein step (d) is performed in heated and stirred reactors under ambient pressure or elevated pressure.
28. 28. The process according to claim 27 wherein said leaching is done in conventional stirred reactors operating at ambient pressures and said slurry is heated to the boiling point of said slurry.
29. 29. The process according to claim 28 wherein the leaching temperature for said conventional stirred reactors operating at ambient pressures is at the boiling point of said slurry or preferably at 90-95 °C to minimise the loss of water to evaporation.
30. 30. The process according to any one of the preceding claims wherein the leaching in step (d) of the product with said aqueous liquor is capable of being done in the presence of a scandium complexing agent.
31. 31. The process according to claim 30 wherein said scandium complexing agent is a compound that selectively combines with dissolved scandium and prevents it being reprecipitated.
32. 32. The process according to claim 30 or claim 31 wherein said scandium complexing agent is any compound that selectively combines with scandium ions in the leach liquor.
33. 33. The process according to any one of claims 30 to 32 wherein said scandium complexing agent is diglycolic acid or similar compounds.
34. 34. The process according to any one of claims 30 to 33 wherein the mass of said scandium complexing agent added to said leach solution is 0.01 to 10.0% of the baked ore by weight.
35. 35. The process according to any one of the preceding claims wherein said leaching step is performed in a pressure vessel and said slurry is heated at a temperature between 110 °C and 500 °C, preferably between 200 °C and 300 °C, and most preferably at 250 °C.
36. 36. The process according to any one of the preceding claims wherein said leaching step is performed in a pressure vessel for between 0 and 10 hours, more preferably for between 0 and 2 hours after the second predetermined temperature and pressure has been reached.
37. 37. The process according to any one of the preceding claims wherein said aqueous liquor is any one of plain water, recycled process water, reverse osmosis or demineralised water.
38. 38. The process according to any one of the preceding claims wherein the solid content of the slurry is between 5% and 55%, preferably between 15% and 35%.
39. 39. The process according to any one of the preceding claims further including stirring the slurry at a rate within the range of 2 rpm to 5000 rpm, preferably between 100 rpm and 1000 rpm, in order to provide a sufficient contact between the aqueous and solid phases.