BACKGROUND OF THE INVENTION

This invention relates to a process for recovering platinum group metals (PGMs - ruthenium, rhodium, palladium, osmium, iridium, platinum and gold) and base metals (copper, nickel and cobalt) from ores, concentrates, tailings as well as recycled materials containing PGMs.

Chlorination of ore concentrates have been around for decades, either in the presence of carbon or carbon monoxide, in the presence of a chloride salt such as sodium chloride, in a salt bath such as sodium chloride or combinations of the aforementioned. As far as the applicant knows, the only processes operating commercially are the carbo-chlorination of titanium and zirconium concentrates and the dry chlorination of magnesium concentrates. A number of patents exist that deals with chlorination to recover PGMs from ores and concentrates.

US2007/0131058; "Process for recovering Platinum Group Metals from ores and concentrates" - Mario Bergeron & Marc Richer-Lafieche; Resources Minieres Pro-Or Inc: Ore or concentrates are mixed with 5%- 20% NaCl, dried and reacted with gaseous chlorine and CO (CI ₂ /CO 0.5- 1.5) at 240 - 800 °C for 30 - 120 minutes in a horizontal rotating / static vertical / fluidized bed reactor. The PGM salts are then leached with water or 0.1-3 M hydrochloric acid. The problem with this process is that the extraction of the PGMs, especially rhodium, is not high enough to be competitive with the conventional smelting process. Carbo-chlorination introduces further cost and safety issues, while not giving any recovery advantage. The relatively low salt addition compromises the recovery of the insoluble PGM's. In addition, if chlorine needs to be recycled, it will need to be passed through a clean-up stage to remove carbon dioxide and/or residual carbon monoxide. The higher leaching acidity has significant cost implications with regards to acid neutralization and/or acid regeneration. RSA1996/2382; "Dry chlorination of PGM-bearing chromite ores" - Mario Bergeron & Jean-Marc Lalancette; UG Plus International Inc: Ore or concentrates are dry chlorinated with or without some NaCJ at 350 - 800 °C for 30 - 120 minutes in a horizontal rotating / ftuidized bed reactor. The PGM salts are then leached with 6M HCl / Cl ₂ . The problem with this process is the volatilization of some PGM species as well as the aggressive digestion step, dissolving more than only the PGMs thus leading to an eiaborative clean-up circuit. The low or no salt addition compromises the recovery of the insoluble PGM's, thus making this process uncompetitive against the conventional smelter process. The higher leaching acidity has significant cost implications with regards to acid neutralization and/or acid regeneration.

US 1992/5104445; "Process for recovering metals from refractory ores" - Michael Oubrovski & Paul J Marcantonio; Chevron Research & Technology Co: Ore or ore concentrates are mixed with 500%-1000% NaCl and reacted with gaseous chlorine at 300 - 650 °C to form the PGM and base metal chlorides. Less than 5%-10% of the reactor content melts, leaving the reactor content as a flowable powder. The PGM chlorides are then leached with brine. The problem with this process is that the molten salt bath chlorination would be operationally very difficuit to control. The water volumes needed after leaching would be huge and the salt must be crystallized before use. Both capital and operating costs would be prohibitively high. In addition, the lower temperatures of reaction compromise the recovery of the insoluble PGM's.

US 1993/5238662; "Process for recovering precious metals" - Michael Dubrovski; Chevron Research Company: A chloride melt is prepared from chloride salts (sodium, potassium or magnesium). Ore or matte is mixed with KCI, the quantity is stoichiometric with the PGMs and base metals present. This mixture is introduced into the chloride melt and gaseous chlorine is introduced at 300 - 650 °C to form the PGM and base metal chlorides. The PGM chlorides then react with the KCI to form the corresponding PGM sails. The problem with this process is that the molten salt bath chlorination would be operationally very difficult to control. The water volumes needed after leaching would be huge and the salt must be crystallized before use. Both capital and operating costs would be prohibitively high, in addition, the lower temperatures of reaction compromise the recovery of the insoluble PG!vYs.

SUMMARY OF THE INVENTION

According to the present invention, there is provided a method of recovering platinum group metals (PGMs - ruthenium, rhodium, palladium, osmium, iridium, platinum and gold) and (if present) base metals (for example copper, nickel and cobalt) from PGM-containing material such as PGM bearing ores, concentrates, tailings as well as recycled materials containing PGMs (including but not limited to spent catalytic converters), wherein the PGM-containing material is mixed with a chloride salt such as potassium chloride or sodium chloride, preferably sodium chloride, at a ratio of 40% to 300% by mass, preferably 50% to 200%, preferably 50% to 150%, preferably 60% to 100%, more preferably 60% to 80%, most preferably 60% to 75% by mass on a dry weight basis, chloride to PGM containing material, and the mixture is then reacted with chlorine gas in a chJorination reaction, in a suitable chlorination reactor, at a temperature of 750 to 850°C, preferably 800 to 820°C, to selectively convert the PGMs to their respective chloro-salts and the base metals to their respective metal salts and provide a residue containing PGM salts and base metal salts.

The reaction can be conducted in any suitable reactor, including but not limited to a static vessel alike vertical shaft reactor, a tubular reactor, a multiple hearth reactor or a rotary device, preferably a vertical shaft reactor.

Preferably, the chloride salt is from a chloride brine, which may be supplemented with solid chloride salt. Preferably, the PGM-containtng material/chloride salt mixture is spray dried such that more than 50% of the spray dried mixture falls in the particle size range 300 [xm to 1.2 mm, and may be calcined at a temperature of 200 to 250 °C, prior to the chlorination reaction.

!n a preferred embodiment of the invention, no carbon/CO is introduced to the reactor during the chlorination reaction.

In the chlorination reaction, the mixture becomes a tacky partial melt and PGMs and base metals in the PGM-contain/ng material are selectively chlorinated, resulting in the formation of PGM chloride salts of the type Na ₂ PtC! ₆ , Na ₂ irCI _6t Na ₂ RuCi ₆ , Na ₂ PdCI ₄ , Na ₃ RhCI ₆ and NaAuCI _4t as well as the various base metal chlorides.

The chlorination reaction typically takes place under a pressure of 1 - 2 bar and the chlorine gas may be introduced at a rate of 40 to 60 g/h for 5 to 15 minutes, followed by a rate of 10 to 15 g/h for 40 to 60 minutes.

The residue may be leached with 0.1 to 1.5 M HCl to form a ieachate containing dissolved PGM chloro acids and base metal chlorides which is filtered, and PGMs and base metals may be recovered from the filtered ieachate by conventional technology, i.e. cementation, ion exchange or solvent extraction.

DESCRIPTION OF PREFERRED EMBODIMENTS

The present invention entails the chlorination of PGM bearing ores, concentrates, tailings as well as recycled materials containing PGMs (including but not limited to spent catalytic converters) in the presence of a chloride salt like but not limited to sodium chforide (or other chloride salts like potassium chloride) to selectively convert the PGMs to their respective chloro-saits. The reaction can be conducted in any suitable reactor, including but not limited to a static vessel alike a vertical shaft reactor, a tubular reactor or a rotary device. Refractory ore concentrate is mixed with sodium chloride brine (40% to 300% by mass on a dry weight basis, chloride to ore concentrate) and then dried (typically spray-dried, such that the majority of the spray dried material fails in the size range 300 pm to 1.2 mm) and may be calcined at a temperature of 130 to 250 °C. Solid sodium chloride can also be mixed and blended with the ore concentrate before calcination. This mixture is then reacted with chlorine at a temperature of 750 to 850°C in a vertical shaft reactor for 60 to 120 minutes. Typically, chlorine gas at the bottom of the reactor at a rate of 40 - 60 g/h under a pressure of 1 to 2 bar for 5 - 15 minutes, and then the chlorine flow rate is reduces to 10 - 15 g/h for a further 40 - 60 minutes. No carbon/CO is introduced to the reactor. The advantage of this is that other metal species in the feed materia! are not converted to their respective metallic forms. The reaction mass becomes a tacky partial melt. The PGMs and base metals in the ore concentrate are selectively chlorinated, resuiting in the formation of PGM chloride salts of the type Na ₂ PtCI ₆ , Na ₂ lrCI _B , Na ₂ RuCI ₆ , Na ₂ PdCI ₄ , Na ₃ RhCI ₆ and NaAuCI ₄ , as well as the various base metal chlorides. Some iron chloride and small quantities of other metal chlorides are also formed. These PGM chloride salts are highly soluble and, after cooling, are leached from the residue with weak hydrochloric acid (0.1-1.0 M HCl solution), forming the corresponding PGM ch!oro-acids of the type H ₂ PtCI ₆ (chloroplatinic acid), H ₂ lrCI ₆ (chSoroiridic acid), H ₂ RuCI ₆ , H ₂ PdCI ₄ (chloropalladic acid), H ₃ RhCI _6> and HAuCI ₄ (chloroauric acid). The barren solids are separated from the PGM- rich hydrochloric acid through filtration. PGM metal recovery can then be done by conventional technology, i.e. cementation, ion exchange or solvent extraction. Work to date includes the production of a mixed PGM metal sponge through anionic ion exchange followed by cationic ion exchange or precipitation to produce the base metals. As a second phase, individual PGM metals may be produced. After removal of the valuable species, the sodium chloride brine is cleaned up and recycled to the reaction stage. The barren soiids are water-washed and can then be returned to the mine for back-filling. Chlorine gas exiting the reactor is directed to a dry cooling step, whereafter the chlorine can be recycled to the reactor inlet thus reducing the fresh chlorine feed. A chlorine bleed is directed to a caustic soda scrubber where sodium hypochlorite is produced, which is a saleable product.

With reference to Table 1 , using the process of the present invention, it is possible to obtain a residue (tailing) containing, respectively, less than 1 g/t platinum, palladium, rhodium, gold and iridium.

Advantages of the invention: the combination of the high salt content and high temperature of the present invention lead to a higher recovery of PGMs including the rhodium, ruthenium and iridium than prior art processes. The achieved recoveries are similar to the conventional process of smelting and converting. Other advantages of this process are that it produces a mixed PGM metal concentrate (>70% PGM metals) which is significantly richer than a conventional smelter converter matte (0.2% PGM metals) which can bypass the Base Metals Refinery and be sent directly to the Precious Metals Refinery.

Example

The invention will be described in more detail with reference to the following non-limiting example:

1. Weigh off a 100 g sample of PGM concentrate.

2. Prepare a saturated NaCl brine solution with 75 g of NaCl.

3. Thoroughly mix the concentrate sample with the NaCl brine solution and spray-dry the solution, such that the majority of the spray dried material fails in the size range 300 μιη to 1.2 mm.

4. Load the dry mixture into a vertical 40 mm ID silica quartz tube. The quartz tube has a chlorine gas inlet at the bottom and a chlorine gas outlet at the top.

5. Heat the silica quartz tube up in a vertical position inside an electrical furnace from room temperature to 820°C in 25 minutes, whilst keeping the mixture under a nitrogen blanket. When the mixture reaches 820°C, introduce chlorine gas at the bottom of the quartz tube at a rate of 50 g/h under a pressure of 1 to 2 bar g for 10 minutes. Then reduce the chlorine flow rate to 13 g/h for a further 50 minutes.

The unreacted chlorine leaving the reactor is directed to a 7% - 12% NaOH scrubber to produce sodium hypochlorite.

Switch the furnace power off and, once the temperature drops below 500 °C, shut the chlorine flow off.

Prepare 1000 ml of a 1M HCl solution and heat it up to around 90°C.

Once the temperature of the reaction mass drops below 100°C, slowly wash the reaction mass from the quartz tube into a Ieach vessel with the hot 1M HCl solution.

Keep the mixture in the Ieach vessel at 75 - 80 °C for 60 minutes, whilst continuously stirring.

Filter the mixture to separate the solids from the liquids. Re-puip the solid residue with water and filter again.

Subject the water-washed residue to gravity separation to recover a portion of the Ruthenium as Ruthenium Oxide.

After removal of the Ruthenium Oxide, the PGM and base metals barren residue is ready for disposal. Analyze the barren residue for PGMs and base metals.

The Ieach liquor is analyzed for PGMs and base metals. The Ieach liquor is subjected to anionic resin exchange to recover all the PGMs in anionic form. Once the resin is sufficiently loaded with PGMs, it is treated and incinerated to recover the PGMs as a mixed PGM metal sponge containing more than 70% PGM metals.

After the PGMs have been removed from the ieach solution, the Ieach solution is subjected to cationic ion exchange to remove ail iron from the solution. Once the resin is sufficiently loaded with iron, the iron is stripped from the resin with 6M HCl as ferric chloride, where after the resin is reused.

The base metals are then precipitated from the Ieach liquor as a 50% mixed metal sulfide product, using NaHS and NaOH. 18. Finally the minor metals (Mg, Al, Ca & Cr) are precipitated with sodium carbonate.

19. The remaining liquid is NaCl brine, which is recycled to the front-end of the process.

The results are displayed in Table 1 below:

Table 1

Tests were conducted on a number of PGM concentrates covering high chrome concentrates, high sulfur concentrates, high PGM loading concentrates (>500 g/ton PGM 4E) and low PGM loading concentrates (<70 g/ton PGM 4E). in all cases, the PGM and Base Metal extractions were comparable with conventionai smelting. The high sulfur concentrate had a PGM loading of 108 g/t 4E (Pt, Pd, Rh & Au) and a base metal loading of 8.6% (Cu, Ni & Co). The process of the present invention extracted 99.7% of the PGMs and 99.7% of the base metals. The process of the present invention is applicable across the range of PGM concentrates available in industry.