FIELD OF THE INVENTION

The present invention relates to a method of recovering a metal from a source material. More particularly, the present invention relates to a method of recovering a metal from impure sources such as automobile catalytic converters, ores or waste materials.

BACKGROUND OF THE INVENTION

Noble metals, including the platinum group metals, are valuable commodities which have found use in a wide range of applications. In particular, platinum, palladium and rhodium are used extensively in the production of automobile catalytic converters, otherwise known as autocatalysts.

Catalytic converters are employed on vehicles to reduce their emission of hydrocarbons, carbon monoxide and other noxious substances. The catalytic properties of platinum group metals are harnessed to convert these emissions into relatively harmless substances such as carbon dioxide, nitrogen and water vapour.

In 2006, around half of the global demand for platinum was for use in the manufacture of autocatalysts with similarly high levels of demand for palladium and rhodium. The annual consumption of these metals is predicted to continue to rise as legislation is passed in developing countries requiring that noxious vehicle emissions be lowered.

Due to their inherent value, platinum, palladium and rhodium are recycled from spent automobile catalytic converters. A number of different processes have been developed to recover these metals by leaching them into solution and subsequently separating the leach solution from the ceramic support. The metals are then recovered from solution by various means.

Prior art methods typically employ high pressure autoclave conditions and/or high temperatures during the leaching process. Many also require very high acid concentrations which not only raises safety concerns but also results in high operation costs due to constant consumption of the acid by reaction with the ceramic substrate and like components. One method which avoids high acid concentrations requires the use of cyanide salts which can be undesirable from a safety perspective as hydrogen cyanide production is a possibility.

The present inventor has identified a need for a method of recovering platinum group metals from automobile catalytic converters which can be performed at relatively low temperatures and ambient pressure. It would further be desirable to reduce the need for large quantities of acid and otherwise aid in improving the safety profile and process economics of such a method. It would be useful if the method could also be applied to the recovery of one or more non-platinum group metals.

OBJECT OF THE INVENTION

The object of the invention is to overcome or at least alleviate one or more of the above problems or to at least provide for a useful commercial choice.

SUMMARY OF THE INVENTION

In one broad form the invention resides in a method of recovering a metal from a source material by leaching the metal into a leach solution containing high concentrations of a metal halide as a halide ion donor to solubilise the metal.

In a first aspect, although it need not be the only or indeed the broadest form, the invention resides in a method of recovering a metal from a source material including the steps of:

(a) producing a leach solution comprising a metal halide;

(b) contacting the source material with the leach solution to leach at least a portion of the metal into the leach solution;

(c) separating the metal-containing leach solution from the remaining source material; and

(d) treating the metal-containing leach solution with a reducing agent to thereby recover the metal as a solid, wherein the majority of halide ions present in the leach solution are donated by the metal halide. The .concentration of halide ions donated to the leach solution by the metal halide may be at least 8 moles/Litre. Preferably, the concentration of halide ions donated to the leach solution by the metal halide may be at least 9 moles/Litre. More preferably, the concentration of the metal halide in the leach solution is about 10 moles/Litre.

In one embodiment, the metal halide donates about 500 to about 5000 molar equivalents of halide ion to extracted metal present in the metal- containing leach solution. Preferably, the metal halide donates about 1000 to about 3000 molar equivalents, more preferably about 2000 molar equivalents.

Suitably, the method further includes the step of adding an oxidant and/or an acid to the leach solution.

The leach solution may have a high ratio of halide ions donated by the metal halide to hydrogen ions/protons donated by the acid.

Suitably, the molar ratio of halide ions to hydrogen ions/protons in the leach solution is at least 2:1.

Preferably, the molar ratio of halide ions to hydrogen ions/protons in the leach solution is at least 5:1.

More preferably, the molar ratio of halide ions to hydrogen ions/protons in the leach solution is at least 10:1.

Typically, the acid is present at a concentration of between about 0.001 to about 1.0 moles-proton/L.

Preferably, when acid is added to the leach solution the majority of the halide ions present in the leach solution are donated by the metal halide. In one general embodiment, greater than 90% of the halide ions present in the leach solution are donated by the metal halide.

In a preferred embodiment, the leach solution is recycled for use in recovery of further metal from the same or a different source material.

The recycled leach solution comprises a substantially similar concentration of metal halide to the leach solution prior to contact with the source material. If required, the leach solution may be recycled indefinitely.

The leach solution may be recycled before or after recovery of the metal.

Preferably, the majority of the metal to be recovered from the source material is leached into the leach solution.

Suitably, the metal is a metal capable of forming a soluble halide complex.

Preferably, the metal is a noble metal.

The metal being recovered may be selected from the group consisting of rhenium, ruthenium, rhodium, palladium, silver, osmium, iridium, platinum, and gold.

Preferably, the metal being recovered is selected from the group consisting of rhodium, palladium, platinum, silver and gold.

The source material may be an automobile catalytic converter or part thereof, an ore or waste material comprising the metal.

When the source material is an automobile catalytic converter or part thereof then the metal is a platinum group metal (PGM) metal.

Preferably, the PGM is selected from the group consisting of platinum, palladium and rhodium.

In one embodiment, cerium may be recovered from the source material in addition to the noble metal.

When the source material is an ore or waste material then the metal recovered is preferably gold, silver or palladium.

Suitably, the metal halide is an alkali metal halide or alkaline earth metal halide or combination thereof.

Preferably, the metal halide is selected from the group consisting of a magnesium, calcium and sodium halide salt.

More preferably, the metal halide is magnesium chloride or calcium chloride.

The reducing agent may be a metal and/or hydrogen gas.

The method may further include the step of comminuting the source material prior to contact with the leach solution.

In one general embodiment the source material is calcined at an elevated temperature prior to treatment with the leach solution. The leach solution may be maintained at or below 120°C, preferably at or below 100°C during contact with the source material.

The leach solution may be at ambient pressure upon contacting the source material.

The method does not involve the use of cyanide salts in the leach solution or the reducing step.

The method may further include the step of subjecting the source material .to a magnetic separation step prior to treatment with the leach solution.

The method may further include the step of contacting the metal- containing leach solution with a precipitation agent to precipitate metallic impurities prior to treatment with the reducing agent.

Preferably, the precipitation agent is a metal oxide such as magnesium oxide.

Further features of the present invention will become apparent from the following detailed description.

Throughout this specification, unless the context requires otherwise, the words "comprise", "comprises" and "comprising" will be understood to imply the inclusion of a stated integer or group of integers but not the exclusion of any other integer or group of integers.

BRIEF DESCRIPTION OF THE FIGURES

In order that the invention may be readily understood and put into practical effect, preferred embodiments will now be described by way of example with reference to the accompanying figures wherein:

FIG 1 is a flow sheet describing steps involved in recovering a noble metal from an automobile catalytic converter, according to one embodiment of the invention;

FIG 2 is a representation of the effect of pulverising an automobile catalyst on the subsequent leaching of various metals;

FIG 3 is a graphical representation of the effect of different oxidants on the recovery of platinum;

FIG 4 is a graphical representation of the effect of varying the concentration of MgC in the leach solution on the recovery of platinum;

FIG 5 is a graphical representation of the effect of varying oxidant and acid concentrations on the recovery of platinum;

FIG 6 is a graphical representation of the recovery of platinum, palladium and rhodium from an automobile catalytic converter sample;

FIG 7 is a graphical representation of the recovery of non-noble metals from the used leach solution;

FIG 8 is a graphical representation of the effect of including ferric chloride as an additional oxidant in the leach solution;

FIG 9 is a graphical representation of the effect of sequential leaching of fresh catalyst samples using a recycled leach solution on the recovery to solution of platinum, palladium and rhodium;

FIG 10 is a graphical representation of the effect of sequential leaching of fresh catalyst samples using a recycled leach solution on the recovery to solution of cerium, lead, silicon and iron;

FIG 11 is a graphical representation of the extraction of palladium, platinum and rhodium in an initial leach, re-leach, and aqua regia digestion;

FIG 12 is a graphical representation of the removal of impurities from the leach solution by precipitation;

FIG 13 is a graphical representation of the effect of varying reducing conditions on the recovery of platinum from the leach solution;

FIG 14 is a graphical representation of the effect of acid on the recovery of platinum from solution;

FIG 15 is a graphical representation of the recovery of platinum group metals from solution; and

FIG 16 is a graphical representation of the extraction of gold and silver from an impure source material into a leaching solution.

DETAILED DESCRIPTION OF THE INVENTION

The present invention is predicated, at least in part, on the development of a method of recovering valuable metals such as platinum, palladium and rhodium from automobile catalytic converters using a leach solution having a high concentration of a metal halide, such as magnesium chloride. The method further extends to the recovery of other valuable metals, including silver and gold, from ores, waste materials or other impure sources. The leaching process further employs an oxidant and, advantageously, requires relatively low acid concentrations. The present method has the further advantage that the used leach solution, following removal of the extracted noble metal, lends itself to recycling thereby greatly reducing reagent and solution preparation costs. The method described provides for a safer, simpler and more cost effective method of recovering these valuable resources.

In the embodiments described herein, the method of recovering the metals is generally described in relation to automobile catalytic converters only. However, it will be appreciated that it is possible, using the same or a substantially similar process, to recover the same or similar metals from a wide variety of impure sources.

The terms "noble metaf and "noble metals", as used herein, may refer to any one or more of a number of metallic elements which can form soluble halide complexes and may include rhenium, ruthenium, rhodium, palladium, silver, osmium, iridium, platinum, and gold. The method of the invention is particularly suited to the recovery of platinum, palladium, rhodium, silver and gold.

The terms "platinum group metaf, "platinum group metals" and "PGMs", as used herein, may refer to any one or more of the six transition metals; platinum, palladium, rhodium, ruthenium, osmium and iridium.

The term "source materia!' , as used herein, may refer to any impure source of a noble metal including natural sources such as ores and man- made sources such as automobile catalytic converters, electronic waste or components thereof.

FIG 1 is a flow sheet describing steps involved in recovering a noble metal from an automobile catalytic converter, according to one embodiment of the invention. The skilled addressee will appreciate that not every step shown is essential to the recovery of the noble metals. The various parts of the process will now be described. The automobile catalyst is first separated from its steel shell to isolate the generally honeycomb ceramic support structure. This component, which contains the noble catalytic metals, is then crushed and/or ground to reduce its particle size to increase the surface area upon which the subsequently used reagents may act.

The automobile catalyst may be ground down to an average particle size of less than 10 mm, preferably less then 6mm. The automobile catalyst may be further pulverised to as great an extent as is practicable, for example, such that all material can pass through a 212 pm sieve. This provides an optimal surface area while avoiding any noticeable reduction in the rate of subsequent solid-liquid separation.

FIG 2 is a representation of the effect of pulverising an automobile catalyst on the subsequent leaching of various metals wherein, on the Y axis, the units for aluminium are g / kg catalyst; for iron, lead, silicon and cerium they are g / 10kg catalyst; and for palladium, platinum and rhodium they are g / 100kg catalyst. The autocatalyst was pulverised such that 100% of materials passed through a 212 pm sieve. The conditions for the leaching experiment represented in FIG 2 are a 5 M MgCI ₂ aqueous solution with 0.17 g of Na ₂ S ₂ 0 ₈ /g of catalyst, 0.09 g of HCI /g of catalyst with exposure for 6 h at 80°C with approximately 4.8 % solids. It is clear from FIG 2 that reducing the particle size of the source material provides for an improvement in the ability to leach noble metals such as the PGMs.

Optionally, any magnetic, for example iron-containing, impurities may be removed during or after the particle size reduction step by way of a magnetic separation step.

The comminuted automobile catalyst may then be calcined by heating the dry material in the presence of air, preferably under a forced convection. The calcination temperature may be between about 200°C to about 1000°C, preferably about 300°C to about 700°C and more preferably about 500°C. A person of skill in the art will easily be able to determine a suitable temperature based on the knowledge that the temperature and air flow must be sufficient to oxidise carbonaceous material and oil that is found within the used automobile catalyst. Although this step may not be essential it is a preferred approach as the carbonaceous material and oil can interfere with the subsequent leaching stage by consuming reagents and/or adsorbing the noble metals thereby lowering recovery.

The calcined catalyst material is then treated with a leaching solution which contains a relatively high concentration of a metal halide which, in the embodiment shown in FIG 1 , is magnesium chloride (MgCI ₂ ), along with an oxidant and an acid. The magnesium chloride acts as a chloride ion source enabling the noble metals to form water soluble chloro complexes thereby resulting in their dissolution and leaching into the aqueous leaching solution. The preferred concentration of the halide ions donated by the metal halide in the leach solution in the embodiments described will be about 10 moles/Litre although it will be appreciated that lower concentrations will still provide for at least partial recovery of the target noble metal.

It will be appreciated that the upper limit of solubility, and hence the concentration of the halide ions donated by the metal halide in the leach solution may vary depending on the particular metal halide used. In one embodiment the concentration of the halide ions donated by the metal halide in the leach solution will be between about 4 to about 14 moles/Litre, preferably about 6 to about 12 moles/Litre. Even more preferably, the concentration of the halide ions donated by the metal halide in the leach solution is about 0 moles/Litre. Between about 4 to about 14 moles/Litre includes 4, 5, 6, 7, 8, 9, 10, 11 , 12, 13 and 14 moles/Litre.

In one embodiment, the metal halide donates between about 500 to about 5000 molar equivalents of halide ion to extracted metal present in the metal-containing leach solution. This range is inclusive of about 500, 750, 1000, 1250, 1500, 1750, 2000, 2250, 2500, 2750, 3000, 3250, 3500, 3750, 4000, 4250, 4500, 4750 and 5000 molar equivalents of halide ion to extracted metal present in the metal-containing leach solution. Preferably, the metal halide donates about 1000 to about 3000 molar equivalents, more preferably about 2000 molar equivalents.

Suitably, the metal halide is an alkali metal halide or alkaline earth metal halide or combination thereof.

Preferably, the metal halide is selected from the group consisting of a magnesium halide, a calcium halide and a sodium halide.

Preferred metal halides salts are metal chlorides or bromides. In a particularly preferred embodiment the metal halide is a metal chloride.

In a particularly preferred embodiment, the metal halide is magnesium chloride or calcium chloride.

The important noble metals in automobile catalysts are platinum, palladium and rhodium and the process chemistry occurring in the leaching step is represented by the below equations:

Pt + 6CI ^" = PtCI ₆ ^"2 + 4e ^" (1)

Pd + 3CI ^" = PdCI ₃ ^" + 2e ^" (2)

Rh + 6CI ^' = RhCle ^"3 + 3e ^" (3)

During the leaching process the conditions are such that the reactions proceed from left to right i.e. the metals are oxidised and form water soluble chloro complexes. This is achieved by the introduction of the oxidant and relatively small quantities of an acid such as hydrochloric acid (HCI), as shown in FIG 1, Any relatively strong mineral acid may be suitable.

It is an advantage of the present invention that the high levels of halide ions necessary to ensure solubilisation of substantially all of the target metals are able to be successfully supplied by the metal halide thereby greatly reducing the need for high concentrations of acid which are seen in many prior art methods. Relatively small quantities of acid are employed to prevent the dissolved noble metals from precipitating as hydroxides and, mostly, to chemically liberate the precious metals by dissolving the alumina catalyst support. This provides for a safer leaching process than many of those seen in the prior art and reduces costs since high acid concentrations result in unwanted reactions with the aluminosilicate honeycomb support and other impurities thereby consuming the acid which must be frequently replenished. An important advantage of the present method is that the metal halide employed as a halide source to form the noble metal soluble halide complexes does not, to any great extent, react with the ceramic support and other components and so is not consumed in the manner described. This enables the leaching solution to effectively be recycled indefinitely which greatly reduces process costs and provides for a more convenient iterative process. This effect has not been seen in the prior art.

Typically, the acid is present in the leach solution at a concentration less than or equal to 1.0 moles-proton/L. Preferably, the acid is present in the leach solution at a concentration of between about 0.001 to about 1.0 moles-proton/L inclusive of 0.001 , 0.005, 0.01 , 0.05, 0.1 , 0.2, 0.3, 0.4, 0.5, 0.6, 0.7, 0.8, 0.9-and 1.0 moles-proton/L.

Another way of describing the relatively low acid requirement in the present method is to look at the ratio of halide ions to hydrogen ions/protons in the leach solution. The halide ions are mostly donated to the leach solution by the metal halide and the hydrogen ions/protons are the H ⁺ ion obtained in the leach solution upon the dissociation of acid e.g. when HCI dissociates into the H ⁺ and CI ^" ions. Since many prior art,methods rely on strong acid as their source of halide ions they have a ratio of halide ions to hydrogen ions/protons that is close to 1 :1.

The present invention employs high concentrations of the metal halide in the leach solution as the primary halide ion donor and so the molar ratio of halide ions to hydrogen ions/protons is considerably higher.

Suitably, the molar ratio of halide ions to hydrogen ions/protons in the leach solution is at least 2:1 , preferably at least 5:1 , more preferably at least 10:1 , even more preferably at least 50:1 , more preferably still at least 100:1 and more preferably yet at least 200:1.

In one embodiment of the present invention the chloride ion concentration of the leach solution is about 10 mole/L while the hydrogen ions/proton concentration is about 0.03 mole/L. This provides for a molar ratio of halide ions to hydrogen ions/protons of greater than 300: 1.

A wide variety of oxidants which would be known to the skilled addressee may be suitable for use in the present method including, but not limited to, calcium hypochlorite Ca(OCI)2, hydrogen peroxide (H ₂ O ₂ ), sodium persulfate (Na ₂ S ₂ 08), ferric (Fe ³⁺ ) ions or magnesium peroxide (MgC>2). Oxidising gases such as air, oxygen enriched air, oxygen and chlorine gas may also be suitable as alternative oxidants or may be used in combination with those already mentioned.

The next step shown in FIG 1 is the solid-liquid separation. At this stage substantially all of the noble metals should be in the form of solubilised chloro complexes as indicated in equations (1) to (3). This solution is isolated from the entrained solids by filtration, preferably under vacuum or positive pressure. The solids may be washed with water or dilute acid to remove any leaching solution in contact with the solids.

The leaching solution containing the metal complexes must then be treated to recover the solid metals from solution. Process conditions are altered to enable the equations shown above to proceed from right to left i.e. reducing conditions. To achieve this, a reducing agent is added to the metal- containing leach solution. Suitable reducing agents would be known to the skilled addressee. Particularly preferred reducing agents for use in the present invention are metals such as iron or magnesium as they are more reactive than platinum and palladium although fluids such as hydrogen gas may also be useful and, indeed, hydrogen gas may be generated upon addition of the metal to the acidic noble metal-containing leach solution. Further, if magnesium chloride is used as the metal halide in the leach solution then, advantageously, the use of magnesium as a reducing agent does not result in any contamination of the system and further improves the ability to repeatedly recycle the leach solution. Suitably between about 5 to about 50 stoichiometric equivalents of reducing agent to noble metal are added to achieve recovery, preferably about 30 equivalents. This includes about, 5, 10, 15, 20, 25, 30, 35, 40, 45 and 50 stoichiometric equivalents of reducing agent.

After treatment with the reducing agent the precipitated solid metal is filtered off, washed and collected. The solution from which the metals have been recovered may now be recycled straight away, unless further cycles of reduction are required, to the leaching step i.e. added to a fresh catalyst sample or, if required, a portion of the solution may first be bled off and treated to remove any impurities, such as iron, aluminium and other metal salts before being fed back to the leaching step. Alternatively, the entire solution from which the metal has been recovered may be treated in one batch to remove these impurities before being employed in the next leaching step. The treatment may be a simple matter of contacting the solution with magnesia and an oxidant to precipitate dissolved iron and aluminium as their hydroxides.

It will be appreciated that the treatment of the recycled leaching solution may not be necessary. This is particularly so if the same metal reducing agent is used in the reduction step as forms the metal halide in the leaching step. In this case, treatment of the recycled leach solution may be unnecessary or may be required only periodically.

A number of experiments have been carried out to test variation of the process conditions in the leaching step on the recovery of platinum. All experiments were carried out on automobile catalysts and are summarised in table , below. The details of how the experiments were performed are presented in the experimental section. The terms "clean" and "carbonated" in relation to automobile catalysts refer to the physical state of the catalyst in terms of how much dirt or carbonaceous material is observed on the catalyst surface. The terms "genuine" and "non-genuine" refer to the quality of the automobile catalyst and thus the amount of platinum group metals it will contain. Genuine catalysts are more efficacious and contain higher amounts of platinum group metals than non-genuine catalysts. The leaching curves produced from the results of these experiments and discussed below were constructed from platinum solution assay measurements using Atomic Absorption Spectrometry (AAS). The platinum concentration (mg/L) was converted to extraction in mg Pt/kg of catalyst as follows:

Extraction = Pt (mg/L) * Initial Volume of leach solution (L) / Dry Mass of Catalyst (kg) Table 1 : Summary of leaching experiments

The results of each of these experiments in terms of the amount of platinum recovered are presented in table 2. All experiments employed the use of 13% solids content (in relation to the weight of the complete autocatalyst) and a temperature of 80°C in the leaching step.

The pH of a solution only provides a relative indication of acidity and the skilled addressee will appreciate that the activity of the proton is a crucial consideration. In the present concentrated magnesium chloride solutions, the activity is considerably higher than the actual concentration for both the proton and the chloride ion, this provides the strong driving force for the leaching reactions. The activities are roughly on the order of 0 to 100 times the actual concentrations.

FIG 3 is a graphical representation of the effect of different oxidants on the recovery of platinum and represents the results of experiments 1 to 3. Three different oxidants, sodium persulfate, hydrogen peroxide and calcium hypochlorite were evaluated in conditions of 13% solids, 80°C, 5M MgCI ₂ in the leach solution and low acid concentration. A stoichiometric amount of oxidant to dissolve the platinum (assuming 1g Pt/kg catalyst) was added at time = 0 and following each sampling. Further details are given in the experimental section. Table 2: Summary of results from leaching experiments of table 1

Due to the very high concentration (near saturation) of the magnesium chloride in the present solutions they demonstrate non-ideal properties. One effect of this mentioned above is that the relative activity of the proton from the small amount of acid in the leach solution is much greater than its actual concentration so even though there is a relatively small amount of acid (for example, 0.01 mole/L), because the activity in this system is very high it behaves as if it was a much stronger acid (>1 mole/L). In short, a low concentration of acid with high acid 'power' can be employed due to the high magnesium chloride concentration.

This provides a number of benefits previously discussed including the fact that if the solution were to be diluted by 20% with water the acid power would drop by orders of magnitude. In this sense it is a safer leaching solution since, if a person accidentally got some of this solution on their skin, a water rinse would reduce the acid concentration considerably in a short period of time whereas if the solution actually contained 1 mole/L acid (the actual concentration equivalent to the observed acid activity) it would be more difficult to neutralise.

Further, when precipitating any impurities out of the used leach solution for subsequent recycling this low acid concentration provides benefits in that only a small amount of a precipitation agent such as magnesium oxide (MgO) or a like metal oxide is required to neutralise the acid whereas if a concentrated acid solution is employed it could not be efficiently recycled since it would be difficult or at least prohibitively expensive to neutralise all of the acid before purifying the solution. This is a considerable advantage when compared with prior art processes.

FIG 3 shows that the leaching process using pool chlorine (containing calcium hypochlorite) is slower than that seen with hydrogen peroxide and sodium persulfate. Pool chlorine also consumes more acid as it contains calcium hydroxide. The rate of dissolution of platinum using hydrogen peroxide was very high but a small amount of the dissolved platinum is eventually lost back to the solid phase as is seen by the downward slope in the graph as time progresses. When the liquid hydrogen peroxide solution is added to the leach reactor, small gas bubbles are immediately formed (most likely a combination of oxygen and chlorine). Some of the oxidant gas thereby created is apparently bubbling out of the reactor and is lost to the atmosphere. Finally, solid sodium persulfate is seen to perform well with no noticeable gas formation and the amount of platinum in solution does not drop off with time as was observed with the peroxide.

The results of this experiment demonstrate that a number of different oxidant classes are suitable for use in the leaching step but purely in terms of efficacy and ease of use, of those tested, sodium persulfate may be preferred. The decision as to which oxidising agent is ultimately used will involve consideration of further factors such as reagent cost and the potential of the agent to introduce impurities into the system.

FIG 4 is a graphical representation of the effect of varying the concentration of MgC^ in the leach solution on the recovery of platinum and represents experiment 4, as shown in table 1. As was discussed above, it is a feature of the present invention that the metal halide, in this case magnesium chloride, is the primary source of halide ions to solubilise the target metals rather than necessitating high acid concentrations such as a HCI/HNO3 mixture (aqua regia).

In FIG 4 the effect of using magnesium chloride concentrations of 2.5 moles/L and 5.0 moles/L in the leach solution were investigated. It can be clearly seen that halving of the magnesium chloride concentration from 5.0 to 2.5 moles/L resulted in more than a 50% reduction in the amount of platinum extracted. As discussed above, the high magnesium chloride concentration (for example 5.0 moles/L) is important since the activity or effective concentration becomes much higher than the actual concentration which favours the dissolution of the platinum group metals.

Experiments 5 to 9 were designed to test the effect of the particular source material on the yield of platinum. The results in table 2 show that the non-genuine and carbonated catalysts tested did not yield any substantial platinum into the leach solutions. It is known that these automobile catalytic converters contain low levels of the platinum group metals which is reflected in their poorer performance compared with genuine catalysts. It is, therefore, likely that there was not much platinum in these samples to begin with. However, some palladium was recovered from these samples and so the present method was still effective.

Experiment 10 was designed to assess the value in performing a second cycle of leaching on the left over solid residue source material collected after the filtration step in experiment 3, subsequent to the initial leaching step. Leaching of the residue from experiment 3 resulted in the recovery of an extra 101 mg-Pt/kg-catalyst from the already 321 mg-Pt/kg catalyst for a total of 422 mg-Pt/kg catalyst. In one embodiment of the present invention the leaching step may be an at least two-stage counter current process to maximize recovery from the source material and optimise reagent usage. Further stages, such as a third, fourth, fifth stage etc may be employed if further amounts of the metal are recovered although it is expected that the majority of the metal will be recovered in the initial and/or re-leach processes.

Experiment 11 tested the effect of introducing an excess of hydrochloric acid. The results in table 2 show that this did not improve the recovery of platinum from the autocatalyst reinforcing that it is the metal halide which is key in this process rather than the acid concentration.

FIG 5 is a graphical representation of the effect of varying oxidant and acid concentrations on the recovery of platinum and represents the results of experiment 12. The traces indicate that the addition of further acid can actually lower the recovery of platinum unless it is in combination with additional oxidant. Without wishing to be bound by any particular theory, the additional acid may liberate more platinum for potential recovery by dissolving the surrounding catalyst alumina substrate. However, the acid may also dissolve other metallic impurities such as iron and these will consume the oxidant thereby leaving insufficient quantities for optimal platinum recovery. Therefore, the best result is obtained when excess oxidant is employed in combination with the additional acid.

It will be appreciated that even when additional acid is used in the leaching process the metal halide is still by far the greatest source of halide ions to solubilise the platinum group metals. Further, as mentioned, additional acid may result in greater numbers of impurities and so a balance may be struck between increasing the acid content beyond the relatively low quantity used in experiment 3 to improve recovery and the amount of effort and expense required to remove the resulting additional impurities. The inherent value of the metals recovered is considerable and so a certain level of impurities is generally acceptable in the recovery process solutions.

FIG 6 is a graphical representation of the recovery of platinum, palladium and rhodium from an automobile catalytic converter sample as ^' achieved in experiment 12 of table 1 using 13% solids, 80°C, 5M MgCI ₂ leach solution, 15 g- a2S ₂ 0 ₈ , 4 g-H ₂ 0 ₂ and 19 g-HCI per kg catalyst. It can be seen that palladium recovery follows the same trend as is seen for platinum.

FIG 7 is a graphical representation of the recovery of non-noble metals from the used leach solution from experiment 12, as described above. This indicates that iron is by far the most common impurity. Since new catalysts should not contain iron it is believed that this is a result of contamination from the catalyst casing during separation. As previously mentioned, it may be preferable, for this reason, to remove these impurities by magnetic separation prior to leaching to reduce the unwanted consumption of acid and oxidant.

FIG 7 also shows that cerium is obtained in the leach solution. Cerium is a valuable commodity itself and so recovery of this metal in addition to those already discussed would further improve the economics of the process.

An additional round of experiments to those set out in table 1 where then carried out to investigate further aspects of the leaching process. These experiments were performed on a blend obtained by combining 12 catalytic converters from a range of different automobile manufacturers, models and years. The blend will represent a typical genuine autocatalytic source material.

The use of a 'surrogate' or additional oxidant was tested via the addition of 2 g/L ferric chloride to a 5 M MgCI ₂ aqueous leach solution with 0.17 g of Na ₂ S ₂ 0 ₈ /g of catalyst or 0.02 g of H ₂ 0 ₂ /g of catalyst as base oxidant and 9% solids of less than 5.6 mm diameter with exposure for 6 h at δΟ . ^' The results of this experiment are shown in FIG 8. · FIG 8 shows that the combination of ferric chloride and hydrogen peroxide appears to be a highly effective leaching oxidant for platinum with recovery greatly increased over either the use of persulfate alone or in combination with the ferric chloride. It is postulated that the ferric ion (Fe ³⁺ ) in solution acts as an additional oxidant to aid in the leaching process.

A series of tests were also carried out to demonstrate the ability of a leach solution to be recycled for further metal extraction. A total of four experiments of 3 hours each were conducted whereby the same leach solution was used to recover metal from four separate fresh autocatalyst samples, one after the other. The details of each experiment are displayed in table 3.

Table 3: Sequential Leaching Experiments (Fresh Catalyst, Reused Solution)

The results of these experiments are shown in FIG 9 and FIG 10. FIG 9 shows the recovery to solution of palladium, platinum and rhodium while FIG 10 shows the recovery of cerium, lead, silicon and iron with cerium being of some economic interest. The leaching solution was diluted slightly for each sequential experiment which accounts for the temporary dips in concentration of the metals at approximately 180, 360 and 540 minutes.

FIG 9 clearly shows that even when employing a recycled leach solution already containing significant amounts of the PGMs, rhodium, palladium and, particularly, platinum continue to be leached into solution when contacted with fresh autocatalyst material. This confirms the viability of recycling the leach solution to minimise reagent usage and improve process economics.

FIG ^' 10 shows that cerium can also continue to be leached into a solution which has undergone a number of cycles of usage.

After the experiments set out in table 3 had been performed there remained the four catalyst samples which had each only been exposed to one round of leaching solution. In order to ascertain just to what extent the noble metals could be recovered by further leaching of these samples the experiments detailed in table 4 were carried out. This involved exposing each of the previously leached samples to a re-leach solution, made up of reagents as described in table 4, and also treating fresh samples of the same autocatalyst batch with an aqua regia digest to ascertain the quantities of platinum, palladium and rhodium which could be extracted using this harsher approach. Induction Coupled Plasma-Mass Spectrometry was used to measure the metal levels in solution.

FIG 1 1 is a graphical representation of the extraction of palladium, platinum and rhodium in an initial leach, as performed in the first set of tests set out in table 3, a re-leach, as performed according to the conditions in table 4, and an aqua regia digestion. FIG 1 1 clearly indicates that, in most cases, the majority of the leaching of the noble metals takes place in the re- leach which confirms that a counter current leaching process, or other leach solution recycling system, would be highly effective in extracting these metals from a source material. A comparison between the total amount of noble metals extracted in the combined initial leach and re-leach cycles with the amount extracted using an aqua regia digest confirms that the present method is efficacious in recovering a very significant proportion of the palladium and rhodium and almost all of the valuable platinum thereby further reinforcing the economical benefits provided.

Table 4: Conditions for further leaching of solid autocatalyst residues

Given that successive rounds of leaching with one batch of solution have been shown to be useful in extracting the maximum amount of metals per unit of reagents it is useful to know how other metallic impurities might build up in the recycled leach solution and how this might be addressed, if problematic.

An experiment was carried out to test how the most prevalent metals, other than the desired platinum, rhodium and palladium, might be precipitated out of the leach solution. Briefly, a recycled leach solution was subjected to the addition of magnesium oxide (MgO) powder, as a precipitation agent, at 80-90°C with the MgO added every 30 minutes. The results of this experiment are displayed in FIG 12. FIG 12 demonstrates that all of the major potential impurity metals, other than lead, can be precipitated out of solution as the metal hydroxide by the addition of MgO without precipitation of the PGMs to be recovered. Any precipitation agent capable of causing the metallic impurities to form their hydroxide salts may be suitable. After approximately 6 hours the majority of all impurity metals have been removed with only cerium, which has had its concentration greatly reduced, and lead at noticeable levels. The lead may be precipitated out to at least an extent as its sulphate or may be removed . via anion exchange techniques.

The next stage of the recovery process to be analysed was the reduction step to recover the solid metals. FIG 13 is a graphical representation of the effect of varying reducing conditions on the recovery of platinum from the leach solution. In general terms the noble metals in the leach solution can be recovered to the solid state by electrochemical reduction (the opposite of the oxidation that took place in the leaching stage). Certain metal powders are suitable reducing agents, such as iron and magnesium. Fluids such as hydrogen gas may also be useful. Iron and magnesium were selected for these tests as they are relatively inexpensive and, additionally, magnesium will not contaminate the system, as discussed earlier.

Similar to the leaching experiments, the efficiency of the reduction process was monitored by analysing the concentration of platinum left in the solution. Samples taken during the recovery experiments were analysed using an atomic absorbance spectrometer to determine how much platinum was recovered from solution. The % recovery of platinum to the solid phase was calculated by:

((Initial Pt (mg/L) - Pt (mg/L))/lnitial Pt (mg/L)) x 100%

It can be seen in FIG 13 that, provided sufficient metal powder is added (between 5 and 30X the stoichiometric requirement for platinum), platinum can be effectively recovered from solution using either iron or magnesium. The reduction reaction is fast, as complete recovery is seen to occur in less than 1 hour. In some cases, to facilitate the subsequent solid- liquid separation, it may be beneficial to introduce some seed particles. In a similar manner to discussion above relating to the leaching step it may be beneficial to employ a counter current reduction process which would help maximize both metal recovery and utilisation of the reducing agent as well as increasing the purity of the recovered noble metal. Detailed parameters of the reduction experiments of FIG 13 are presented in table 5.

Table 5: Results of reduction step recovery of platinum on solution from leach test 1 in table 1

To better understand the mechanism of reduction and hence recovery of the noble metals from solution an experiment was carried out to test the effect of acid on this process. Once again the test focused on the recovery of platinum from a leach solution by treating the solution with reducing conditions of 3 g Mg powder / kg solution at 80 ^° C with hydrochloric acid added or absent. The results are shown in FIG 14.

FIG 14 indicates that acid is necessary for the efficient recovery of platinum from solution. Although not wishing to be bound by any particular theory, this result is suggestive of the role of the magnesium metal being to act as a hydrogen gas source upon contact with the acid rather than as a platform for cementation of the more reactive PGMs. Thus, as previously indicated, the reducing conditions may employ a metal which will produce ^■ hydrogen gas, as the reducing agent, upon reaction with the acid, or hydrogen gas may be directly bubbled into the reducing solution or introduced into a sealed vessel containing the leach solution with the hydrogen gas applied as an overpressure.

A further experiment was carried out on a leach solution resulting from the recycled leaching performed as per the experiments in table 3 to show the recovery of platinum, palladium and rhodium from solution. The reducing conditions used were the addition of 3 g of Mg powder / kg solution, 9 g HCI/kg solution at 90°C and with hydrochloric acid added. The results are shown in FIG 15. ^'

FIG 15 demonstrates that substantially all platinum, palladium and rhodium in solution can be recovered by electrochemical reduction within a 90 minute time period. The purity of the solid PGM concentrate recovered from solution was quite high (greater than 50% PGM by mass) and may be improved by further simple dilute sulphuric acid leaching.

In summary, the results of the leaching experiments described above indicate that the highest recovery is achieved with additional oxidant, possibly ferric ions, relatively small amounts of acid and high magnesium chloride concentrations. A leaching time of 3 to 4 hours was sufficient in all cases to enable a high level of extraction at the conditions tested. The recovery of the target metals may be further improved by increased leaching time or, preferably, employing a two-stage counter current leaching process, or other leaching solution recycling process, which also maximises reagent efficiency. Significant amounts of palladium, rhodium and, particularly, platinum are seen to be leached into solution. The leach solution is favourably selective for the metals of interest as less than 4% of the total mass of the catalyst is actually dissolved and a build up of impurities in the recycled leach solution can be successfully addressed. The recovery of the metals from solution has been shown to be extremely efficient.

From the best recovery results in this series of experiments (assuming prices of $55 and $15 USD/g of platinum and palladium, respectively) it is possible to recover up to $29 USD of platinum and palladium per kg of "genuine" used automobile catalyst into a leach solution in four hours. Further incremental improvements to the leach are possible, however, at $25 to $29 USD/kg the vast majority of the platinum group metals present in the autocatalyst are being recovered.

It will be appreciated that although much of the foregoing discussion has centred around the recovery of platinum group metals from autocatalysts, the invention is not so limited. As was previously mentioned, the method described may be applicable to any metal capable of forming a water soluble halide complex when in the presence of a high concentration metal halide solution.

For example, the inherent value of gold is well known as is its application in the manufacture of jewellery and computers as well as many other electronic devices. This results in large amounts of waste material which can potentially be exposed to the present process to recover the gold and other noble metals. Gold and silver are known to form water soluble complexes with halide ions such as chloride ions. Gold will generally form chloro complexes in either its Au(l) or Au(lll) oxidation states. The gold metal can be recovered from a leach solution by reduction, as previously described.

An experiment was carried out to ascertain the extent to which gold and silver could be recovered from a low grade sample in a single leaching step. Briefly, a standard low grade sample (Rock Labs Reference Material 0x11 , 2.94 ppm Au, 25.0 ppm Ag) was exposed to a leach solution comprising 5 M MgCk, 20% solids (fine powder), 0.03 g-H ₂ 02 g solid and heated at 80°C. The results are shown in FIG 16 which indicates that most of the silver and a commercially significant amount of gold have been successfully leached into the solution for later recovery. It is expected that the extent of the leaching could be improved by optimisation of the pH and, as shown above, use of further cycles of leaching would be expected to greatly improve the overall recovery of gold.

Gold and silver are known to be dissolved by aqua regia but the present method provides a way to form the necessary soluble halide complexes under low acid concentrations providing a range of benefits already discussed. A person of skill in the art would understand that different gold or silver-containing source materials may require different processing steps before being treated with the leach solution of the present invention. For example, electronic waste such as printed circuit boards may need to be deconstructed to remove various components that will consume reagents, particularly the acid, or will result in undesirable impurities being released into the leach solution.

Experimental

Leaching step procedure

(i) The leaching experiments described in table 1 were carried out in a 1- L glass reactor (baffled) with overhead agitation supplied by an impeller constructed of a rigid temperature resistant plastic. The mass of the dry empty glass reactor was measured and recorded.

(ii) A set mass of leach solution was added to the reactor and the mass of solution added was recorded. The leach solution typically contained 5 Moles/L MgCI ₂ and 2mL 32wt% HCI/L.

(iii) A set mass of catalyst material (100% of material to pass through a 5.6 mm mesh) was added to the reactor. The catalyst material was split using a splitter to ensure that the composition of different samples of the same catalyst were homogenous.

(iv) The solids content of the leach slurry was calculated as follows % Solids: % Solids = 100% ^* mass solids / (mass solids + mass solution). This value was kept constant throughout the tests set out in table 1 , at approximately 13%.

(v) The reactor was set-up in a consistent manner with the plastic impeller being centred and positioned at a height of approximately 4-5 cm above the bottom of the reactor. The agitation speed was set to 950 RPM, ensuring all solids are mixed and none remain on the bottom of the reactor. The top of the reactor was covered to minimize evaporation and facilitate heating of the solution.

(vi) A hotplate with a temperature feedback control was used to heat the reactor. The thermocouple was covered in Teflon tape to prevent any reaction with the leaching solution. The temperature was maintained at 80°C and the thermocouple reading was periodically cross checked with a glass thermometer. Sampling and reagent dosing

(i) Solution samples for the experiments in table 1 were generally taken before any oxidant was introduced into the reactor (time - 0), and at time intervals of 10, 30, 60, 120 and 240 minutes thereafter. An additional sample was taken after the solution was vacuum filtered.

(ii) The solution was heated to 80°C in the reactor before adding. the catalyst material. After the autocatalyst is added, there was a slight decrease in temperature so the slurry was allowed to heat up to 80°C again before the beginning the experiment.

(iii) The sampling procedure involved stopping the agitation and allowing the solids to settle for 30-60 seconds. Solution was then syringed out of the reactor (approximately 5 mL) and placed into a 10mL (clean/dry) glass beaker. The solids in the sample were allowed to settle and 4 mL of the sample was removed from the 10 mL beaker. The solution sample was then filtered with a 0.45 μιη pore size syringe filter to ensure that all the solids were removed. Using an automatic pipette, exactly 1 mL of the filtered sample was removed and diluted 10X by combining the sample with exactly 9mL of dilution solution. The dilution solution was comprised of 1v% HCI, 1v%HN0 ₃ , (by volume taking into account that the HCI is 32wt.% pure and the HNO3 is 70wt.% pure). The dilution solution also contained 0.2wt.% La added as La ₂ 0 ₃ powder to assist with the atomic absorption measurement. The pH and ORP of the remaining sample was measured and solution or slurry was returned to the reactor. ORP is measured using a Pt vs. Ag/AgCI reference system (recorded as mV Ag/AgCI). The pH probe/meter was calibrated daily using pH 4.00 and 6.88 buffers.

(iv) After the catalyst was added and the sample at time 0 was taken, the appropriate amounts of reagents were added. The stoichiometry was based on the dissolution of Pt assuming 1 g Pt/kg catalyst (Pt(0) to Pt(IV)). For the case where hydrogen peroxide is the oxidant the reaction stoichiometry was: Pt + 4HCI + MgCI ₂ + 2H ₂ 0 ₂ -» MgPtCI ₆ + 4H ₂ 0

If sodium persulfate was the oxidant the reaction stoichiometry was: Pt + 4HCI+ MgCI ₂ + 2Na ₂ S ₂ 0 ₈ -» MgPtCI ₆ + 2Na ₂ S0 + 2H ₂ S0 ₄

If calcium hypochlorite was the oxidant the reaction stoichiometry was:

Pt + 4HCI + MgCI ₂ + Ca(OCI) ₂ MgPtCI ₆ + CaCI ₂ + 2H ₂ 0

Note that if pool chlorine (containing calcium hypochlorite) is used then additional acid may be required to maintain the acidic conditions as it may contain a significant amount of Ca(OH) ₂ which neutralizes the acid. Generally, the acid concentration needs to be sufficient to dissolve the alumina matrix and prevent the noble metals precipitating as hydroxides.

(v) The pH of the samples were measured and recorded to determine whether further acid should be added or not. A stoichiometric amount of oxidant was added after every sample and at 180 minutes, and HCI may be added if deemed necessary to adjust the pH of the solution.

(vi) At 240 minutes, the heating was turned off. The reactor containing the solids and solution was weighed and the total final slurry mass calculated. The 240 minute sample was then taken.

(vii) The slurry was then vacuum filtered using an 8cm diameter ceramic filter bowl with filter paper. The mass of the clean/dry vacuum conical flask was recorded prior to filtration to enable the final solution mass to be determined.

(viii) A 5mL sample of the final filtrate solution was filtered a second time using a 45 micron syringe filter and diluted to 50mL with the dilution solution. This final solution sample was sent to ICP analysis to determine the concentration of Mg, Al, Pt, Pd, Rh, Fe, Cu, Pb, Ce and Si in the leach liquor, (ix) A separate 5ml_ sample of the solution was tested for pH and ORP. (x) All of the solids were consolidated and washed with 500ml_ de-ionized water. Once filtration with the wash water was complete, the solids and filter paper were placed into the bowl that was previously weighed.

(xi) A 1mL sample of the wash solution was sampled and analysed by ICP (Mg, Al, Pt, Pd, Rh, Fe, Cu, Pb, Ce, and Si). The remaining clean solids were dried in air overnight and the final mass of the solids recorded.

(xii) The recorded information was used to determine the mass balance. The water evaporated was determined from this calculation. Reduction Experimental Procedure

Set up

(i) The mass of the empty dry glass reactor was measured and recorded.

(ii) A set mass of the leach liquor solution was then added to the reactor.

(iii) The reactor was heated using a hotplate with feedback temperature control. The thermocouple was covered in Teflon tape so that the metal does not react with the solution. The temperature was cross checked with a glass thermometer.

(iv) A mechanical stirrer provided the agitation for the reduction experiments.

Process

The dissolved (oxidized) metals were recovered from solution by adding a more reactive metal powder to the solution. The stoichiometric amount of powder required is calculated by the following reactions based on the measured solution Pt concentration.

Iron powder (assume Fe is oxidized to Fe ²⁺ ):

Pt ⁺ + 2Fe° ^' →2Fe ²⁺ + Pt°

Magnesium powder:

Pt ⁴⁺ + 2Mg°→ 2Mg ²⁺ + Pt°

It will be appreciated from discussion presented earlier that it may actually be the hydrogen gas produced by reaction of the metal with acid which causes the reduction rather than a direct cementation type reaction.

(i) The solution was heated to the required temperature (70°C) and a sample was taken at time 0. A stoichiometric amount of reductant was then added following sampling times (time 0, 10 and 20 minutes).

(ii) Samples were taken at 10 minutes, 20 minutes, 40 minutes, 60 minutes, 90 minutes and 120 minutes, and the concentration of platinum in the solution was determined using atomic absorption. The sampling procedure was the same as for the leaching tests. The difference between this number and the initial platinum concentration of the solution indicates how much platinum is recovered from the solution. (iii) After 120 minutes, the heat was turned off and the mass of the reactor, including the solution, was measured. The solids were then separated from the solution using vacuum filtration (or pressure filtration). The initial mass of the filter paper being used was recorded, along with the mass of the vacuum conical flask. After filtration, the mass, of the vacuum flask containing the filtered solution was measured and the mass of the filtered solution was determined. The solids were then washed with de- ionized water (250 mL) and dried overnight. The mass of the final solids was determined. Solution Analysis ,

Induction Coupled Plasma (ICP) was used for multi-metal solution assays. ICPOES was generally employed, however, ICPMS was used when the solution contained only small quantities of the metal.

Atomic Absorption Spectroscopy (AAS) Procedure

The AAS instrument used is a Perkin Elmer. AAnalyst 400. The standards for AAS were prepared with platinum concentrations of 0, 5, 0, 20 and 30 mg/L. A 50mL volume of each standard was prepared. The dilution solution for the standards consisted of 1 % HCI, 1 % HN0 ₃ and 0.2% La, as well as 47.6g/L of MgC . The amount of MgC was calculated by taking into account the fact that the original 5M leach solution was diluted 10x when the samples were taken and so the samples contained approximately 47.6g/L of MgCfe. The magnesium was added to matrix match with the magnesium in the samples being analysed (assumed 10X dilution). The Lanthanum was added to counter the depressing effect of magnesium on the platinum signal in order to obtain linear calibration.

Several sets of standards were prepared for use in AAS, these being standards containing 1%HCI, 0.2%La and MgCI ₂ , standards containing 1 %HCI and MgCI ₂ and standards containing just 1 %HCI.

The discussion and results presented herein have demonstrated that the present method of recovering noble metals from automobile catalysts and other impure source materials provides for a number of advantages over those seen in the prior art. In terms of leaching it has been shown that the majority of the platinum and palladium can be recovered very quickly (less than 1 hour) at the conditions tested. The recovery increases further with increasing time provided that there is a sufficient combination of acid and oxidant present to support the metal halide bringing about soluble complex formation with the target metals.

Similarly, in the recovery stage all of the platinum (within the limits of analytical detection range) was recovered from the solution in less than 1 hour provided there was sufficient reducing agent present. This provides for a method of recovering noble metals, particularly platinum, palladium, silver, gold and rhodium, which involves leaching of the source material at ambient pressure, thereby negating the need for an autoclave, and at relatively low temperatures. In this manner initial set up costs are minimised. Further, the amount of acid required is relatively low compared to processes found in the prior art which means unnecessary dissolution of the catalyst support structure is avoided thereby minimising reagent consumption and impurities. The ability to recycle the leach solution is of key importance as it allows the ongoing operating costs to be greatly reduced due to minimal reagent consumption while providing for further noble metal extraction gains. The use of these comparatively benign operating conditions also provides for a safer process for the operator.

Throughout the specification the aim has been to describe the preferred embodiments of the invention without limiting the invention to any one embodiment or specific collection of features. It will therefore be appreciated by those of skill in the art that, in light of the instant disclosure, various modifications and changes can be made in the particular embodiments exemplified without departing from the scope of the present invention.