Recovery Of Platinum Group Elements Research Articles

Experimental and Modelling Study of Pt, Pd, and 2E+Au Flotation Kinetics for Platreef Ore by Exploring the Influence of Reagent Dosage Variations

This study investigates the flotation kinetics of individual platinum-group elements (PGEs) and gold, namely Pt, Pd, and 2E+Au (i.e., Pt+Pd+Au), in the context of Platreef ore flotation. Experimental tests were conducted on a Platreef ore feed using various dosages of depressants, frothers, and collectors under controlled agitation and pH conditions. The recoveries of the individual PGEs were analysed using six kinetic models, with the modified Kelsall model identified as the most suitable for accurately describing the flotation kinetics and predicting elemental recovery. Notably, the model incorporates two rate constants (kfast and kslow) to account for the distinct flotation behaviours of the PGEs. The results indicate that Pt has the fastest floatability, followed by Pd and 2E+Au. The modified Kelsall model demonstrates high effectiveness in predicting the recovery of these PGEs. Three empirical correlations for Pt, Pd, and 2E+Au recoveries based on the modified Kelsall model are proposed, enhancing the understanding and optimisation of PGE recovery in Platreef ore flotation.

Adsorption of platinum group elements (Pd(II), Ir(III) and Rh(III)) from aqueous solutions using 3-aminopropyl(diethoxy)methylsilane (APDEMS) functionalised bentonite

The recovery of platinum group elements (PGEs) from waste is of high importance due to losses incurred during their processing, high input costs and diminishing primary resources. The recovery of Pd(II), Ir(III) and Rh(III) by bentonite functionalised with 3-aminopropyl(diethoxy)methylsilane (APDEMS) was investigated. Optimum conditions (adsorbent dosage, pH, concentration and competing ions) for the maximum uptake of PGEs by APDEMS-functionalised bentonite were determined. The maximum recovery of Ir(III) and Rh(III) was attained at pH 2, 10 g L-1 adsorbent dosage and 2 mg L-1 initial concentration. However, lower adsorbent dosage (5 g L-1) was required for maximum adsorption of Pd(II) at all concentrations. The adsorption efficiency of Ir(III) decreased significantly (p < 0.05) as pH increased, but was fairly constant for Pd(II) and Rh(III). The adsorbed amount of Ir(III) and Rh(III) increased when competing ions were incorporated. APDEMS-functionalised bentonite proved to be a potential adsorbent for effective recovery of Pd(II), Ir(III) and Rh(III) from mining and industrial wastewaters in order to curb the losses incurred during processing and application.

Platinum Group Elements Recovery from Used Catalytic Converters by Acidic Fusion and Leaching

The recovery of platinum group elements (PGE (platinum group element coating); Pd, Pt, and Rh) from used catalytic converters, using low energy and fewer chemicals, was developed using potassium bisulfate fusion pretreatment, and subsequently leached using hydrochloric acid. In the fusion pre-treatment, potassium bisulfate alone (without the addition of an oxidant) proved to be an effective and selective fusing agent. It altered PGE into a more soluble species and did not react with the cordierite support, based on X-Ray Diffraction (XRD) and metallographic characterization results. The fusion efficacy was due to the transformation of bisulfate into pyrosulfate, which is capable of oxidizing PGE. However, the introduction of potassium through the fusing agent proved to be detrimental, in general, since potassium formed insoluble potassium PGE chloro-complexes during leaching (decreasing the recovery) and required higher HCl concentration and a higher leaching temperature to restore the solubility. Optimization on the fusion and leaching parameter resulted in 106% ± 1.7%, 93.3% ± 0.6%, and 94.3% ± 3.9% recovery for Pd, Pt, and Rh, respectively. These results were achieved at fusion conditions: temperature 550 °C, potassium bisulfate/raw material mass ratio 2.5, and fusion time within 30 min. The leaching conditions were: HCl concentration 5 M, temperature 80 °C, and time within 20 min.

Mineral processing: leaching process of chromium and recovery of platinum group elements from Northwestern Saudi Arabian Ophiolitic chromite

The current strategy reports the recovery of precious group metals (PGMs) from Na2CO3 roast of Saudi Arabian ophiolitic chromite ore at 1160 °C under atmospheric O2. The influence of roasting temperature (960–1200 °C) and leaching agents (H2O, HCl, H2SO4) and their concentrations on the recovery of PGMs was studied. Chromite roast concentrates a reasonable percentage of approximately 45% with an overall 43% Cr2O3. PGMs in the original ore was petitioned and ends up in Na2CrO4 liquor whereas the liquor and a solid roast residues contained the rest of chromite components (Fe2O3, Al2O3, others). Na2CrO4 in the water or H2SO4 leachate was reduced with SO2 gas to give crystalline basic chrome sulfate with excellent yield (> 99%) and purity (Fe < 0.1%). PGMs in the Na2CrO4 liquor were extracted into methylisobutylketone (MIBK) after shaking with ammonium pyrrolidine dithiocarbamate (APDC). In MIBK extract, the PGMs species was then stripped by shaking with HgCl2–HNO3 (1.0 mol L−1). Ir and Pt in the produced PGEs cake were found up to 4.0 and 0.5 µg/g, respectively whereas the majority of Ni (0.5% as NiO) was retained in the solid roast residues. The proposed method offers a simple approach with good cost-effectiveness for PGEs recovery and production of basic chrome sulfate with Fe content less than 0.1%. Leaching with water offers a simple, an eco-friendly and cost-effectiveness approach for the production of a basic chromium sulfate product with high quality yield of PGEs as a byproduct with minimum Fe content (< 0.1% m/m) required for the tanning industry compared to the method reported.

The use of assay data as a foundation for a geometallurgical model—The case of the Thaba Chromite Mine, South Africa

The Lower and Middle Group chromitites of the Bushveld Igneous Complex in South Africa are the source of a very large portion of the global chrome supply. The recovery of platinum group elements and base metals (Ni, Cu) as by-products has the potential to add value to these chrome resources. Yet, the effectiveness of chromite and platinum-group element beneficiation circuits is highly sensitive to variations in feed composition. Of particular relevance is the abundance of alteration silicates, which have a negative impact on recoveries and concentrate grade. The use of geochemical proxies, based on data acquired routinely during the exploration and mining process may provide a cost- and time-efficient alternative to more time-consuming and expensive mineralogical analyses. Such an approach is presented in this study, which focuses on the LG-6, LG-6A, MG-1 and MG-2 chromitite seams at the Thaba mine located on the western limb of the Bushveld Complex. An extensive assay data set comprising of data for Cr2O3, FeO, Al2O3, MgO, SiO2, CaO and P provided by the mine owner was evaluated and corrected for systematic bias between different sample batches. A statistical assessment was performed to discuss the variability within and between the chromitite seams and to separate the mine lease area into distinct geochemical clusters. The distribution of the samples belonging to the different geochemical clusters was then transposed onto the geology of the mine lease area. This allowed the definition of spatial domains. These spatial domains, recognized by the assessment of assay data only, are then validated by mineralogical attributes; implications for mineral beneficiation are critically assessed. According to this assessment, the chromitites of the Thaba mine area can be subdivided into three distinct domains, domains that constitute the suitable fundament for a geometallurgical model. An extensive supergene altered domain is distinguished from a domain affected by hydrothermal alteration. The latter domain occurs below the depth of modern weathering, but in obvious proximity to faults and around a prominent dunite pipe. The third domain is represented by ores least affected by post-magmatic alteration processes. This domain occupies the centre of fault blocks below the extent of modern weathering.

The recovery of Platinum (IV) from aqueous solutions by hydrazine-functionalised zeolite

The recovery of platinum group elements (PGEs) which are normally lost during processing, can add value especially in times of declining commodities prices, high operational costs and weak economic growth. The recovery of platinum (Pt(IV)) from aqueous solutions with zeolite functionalised with hydrazine monohydrochloride (NH2NH2·HCl) was studied. The zeolite was characterised with powder X-ray diffraction (PXRD), X-ray fluorescence (XRF) and Fourier transform infrared (FTIR) spectroscopy. The effect of pH (2–9), concentration (0.5–10 mg L−1), contact time (0–540 min), adsorbent dosage (50–500 mg), competing ions (Fe3+, Ca2+, Mg2+, K+, Co2+, Ni2+, Hf4+, Zn2+ and PGEs) on the recovery of Pt(IV) was studied. The recovery of Pt(IV) was found to be more effective in acidic pH (2–5). The adsorption capacity increased with concentration from 0.5 mg L−1 to 5 mg L−1 and remained constant beyond 5 mg L−1. A slight increase (9%) in the adsorption capacity of Pt(IV) was observed in the presence of competing ions which could be due to the synergistic Pt-Fe interaction. The experimental data was best described by Langmuir isotherm model (R2 > 0.99), indicating a monolayer adsorption. The adsorption behaviour was best described by pseudo second-order kinetic model (R2 > 0.99), implying a chemisorption process through Pt-N interaction. A cost-benefit analysis was done and indicated that ∼120 g Pt can be recovered per kg adsorbent and this would amount to around $3355 profit under the optimum conditions. These findings demonstrate the potential of this adsorbent for use in recovering low level concentrations of Pt and other PGEs.

A mineralogical perspective on the recovery of platinum group elements from Merensky Reef and UG2 at the Two Rivers mine on the Eastern limb of the Bushveld Complex in South Africa

Published studies dealing with the process mineralogy of Pt mines on the Bushveld Complex is generally limited to the Western Bushveld. The recognition by mine management that another resource, in addition to the Upper Group 2 (UG2) reef currently being mined at the Two Rivers platinum mine (TRP), is urgently required in order to extend the life of mine, presented an opportunity to conduct such a study on the Eastern Limb of the Bushveld Complex. A process mineralogical investigation was undertaken on ore from the Merensky Reef (MR) and the UG2 at TRP. This was conducted on a suite of geological samples (channel samples) collected from the underground workings, as well as metallurgical samples obtained from the rougher circuits at the concentrator plant during the processing of MR and UG2 ore. The geological and metallurgical samples were analysed for bulk composition and quantitative mineralogy, while the geological samples were also subjected to laboratory-scale milling and flotation tests. This study shows that, although mineralogically distinct, the MR and UG2 behave similarly in terms of metallurgical performance. This holds promise for the proposed blending of MR and UG2 ores at TRP. An evaluation of the bulk rock (ore) Pt/Pd ratio as a possible indicator of the level of hydrothermal alteration of the ore, demonstrates that this may be of use in predicting recovery plant performance.

Geometallurgical characterisation of Merensky Reef and UG2 at the Marikana Mine, Bushveld Complex, South Africa

The influence of mineralogy on the milling performance and the flotation-based recovery of Au, the platinum group elements (PGE), Co, Cr, Cu, Ni and S was investigated for three samples of the Merensky Reef (BK, RPM and WP facies types) and one sample of the UG2 at the Marikana mine, using a mineral liberation analyser (MLA). The samples differ in their milling behaviour in that a range of milling times are required in order to produce a grind of 60% passing 75μm. This is primarily controlled by the abundance of plagioclase, orthopyroxene and chromite.The size distribution of the base metal sulfides (BMS; pyrrhotite, pentlandite and chalcopyrite) is similar for the three samples of Merensky Reef, and is significantly coarser than for BMS in the sample of the UG2. Upon milling to 60% passing 75μm, the best BMS liberation is achieved for the BK facies type of Merensky Reef, relative to RPM and WP. The degree of BMS liberation in the sample of UG2 is lower than that for samples of the Merensky Reef.Cumulative mass pull during flotation is higher for the sample of the WP facies of Merensky Reef than for the rest of the samples examined. This is due to the higher abundance of orthopyroxene in this sample, which is known to be naturally floating, and which reports to concentrate.A high flotation-based recovery of PGE, Cu and S is observed for all four samples, with the highest recovery associated with the sample of the BK facies type of Merensky Reef. Ni recoveries are generally poor, suggesting that Ni is associated with gangue minerals, in addition to that in pentlandite.Of the three facies types of Merensky Reef examined, the overall characteristics of the BK facies type i.e. a high PGE grade, low abundance of orthopyroxene, a high abundance of BMS, and a higher degree of liberation of the BMS on milling of the ore, represent the most favourable set of characteristics for the efficient recovery of PGE. It is therefore the best quality ore of the three samples of Merensky Reef examined.

Leaching of Ni–Cu–Fe–S converter matte at varying iron endpoints; mineralogical changes and behaviour of Ir, Rh and Ru

The Bushveld Igneous Complex in South Africa is a rich source of platinum group elements (PGEs), as well as base metal sulphides. Typical beneficiation routes entail milling and flotation, smelting, matte converting and hydrometallurgical processing of converter matte to selectively recover metal values. The primary focus is on PGE recovery, with the secondary objective of base metals recovery. At the Western Platinum Pty. Ltd. Base Metals Refinery, converter matte typically contains 48% Ni, 29% Cu, 21% S and 0.6–5% Fe by weight, and the major mineral phases have compositions approaching heazlewoodite (Ni3S2), chalcocite (Cu2S) and a Ni–Cu alloy. It has been observed that the iron content of converter matte can have a significant impact on performance of the first stage atmospheric leach, in which the primary goal is to precipitate copper and PGEs from spent electrolyte (via metathesis and cementation reactions) and the secondary goal is the dissolution of nickel. In this work, laboratory scale batch leach tests were carried out with samples of converter matte with varying Fe-endpoints (0.53%, 0.83% and 5.72%), under oxygenated and non-oxygenated conditions, to improve understanding of the effect of the Fe-endpoint. X-ray diffraction and scanning electron microscopy techniques were employed to understand mineralogical changes that take place during leaching. Matte with a higher iron content was characterised by a lower Ni–Cu alloy content than matte with a lower iron content, and the presence of relatively inert pentlandite (NiFe)9S8. Under oxygenated conditions, nickel was leached more readily from low iron matte and copper and PGEs were precipitated more readily onto low iron matte than high iron matte. Under non-oxygenated conditions, copper was precipitated more readily onto the low iron matte and more nickel was leached from a low-iron matte than high iron matte, this can probably be attributed to the presence of the alloy phase in the low iron matte, which galvanically inhibits the metathesis reaction.

Gravity separation of a UG-2 ore secondary sample for the reduction of chromite minerals

The efficiency of secondary grinding of UG-2 (Upper Group 2) platinum ores is currently limited by the use of open-circuit grinding in many plants, which is inherently inefficient, due to short-circuiting of coarse particles and possible separation by gravity in the mill. The chromium oxide (Cr2O3) grade in UG-2 flotation concentrates is generally above 3% and this contributes to low efficiencies during PGM smelting.Gravity separation of the chromite and silicate particles prior to secondary grinding was investigated with a spiral concentrator. Laboratory cleaner spiral tests were conducted on a plant sample from a secondary ball mill feed, to determine whether a chromite-rich stream could be removed and discarded from the process prior to milling. Analysis of the chromite-rich stream showed that it could not be discarded, as a significant proportion of the Platinum Group Elements (PGEs) in the feed to secondary milling (∼16%) would be discarded with about 55% of the chromite minerals. Pilot plant cleaner spiral test work confirmed the trends observed.An alternative strategy of separate coarse grinding of the gravity concentrate, followed by combined flotation, resulted in a 32% reduction in the Cr2O3 entrainment to secondary rougher flotation concentrate, and most significantly, an increase of 3.7% in secondary recovery of PGE’s with the same overall grinding energy as the standard process.

Flotation model based on floatability component approach – PGE minerals case

This paper discusses how modelling and simulation can be used in predicting the recovery of platinum group elements (PGE) in flotation process. In platinum group ores, commonly a floatability component approach is used where each element is divided in model fitting to two or three components. This approach has the problem of breaking the natural association of platinum group element in minerals, floatability components do not have any physical meaning and as the model fitting problem is ill-posed, comparison of different tests by floatability component parameters is difficult. To avoid this modelling and simulation here is based on minerals. Additionally, the mass proportion of fast floating material was fixed on the basis of mineralogical study and results achieved in laboratory flotation tests. Finally as the rougher and cleaning flotation was performed in fixed chemical regime the model fitting was done for both rougher and cleaning stages simultaneously using cascade model. It was found in this way a number of model fitted parameters decreased giving possibility to compare tests results easier but still reaching good fit between observed, modelled and simulated data. The developed model was used in screening the best flotation conditions and to forecast the metallurgical performance in full scale process.

Dynamic modelling of E-waste recycling system performance based on product design

E-waste covers a wide range of products as well as dismantled and/or sorted components originating from these. Being able to predict the flow of materials and recycling performance for different E-waste types requires a fundamental and flexible basis in which E-waste design properties are linked to liberation and separation performance of recycling. This paper discusses the authors’ design-determined liberation and dynamic models to predict and monitor E-waste recycling technologically, economically and environmentally. The uniqueness of this work lies in the modelling of product design characteristics in terms of design tables that define the mass and material connections derived from the design in real-time. The shredding process is modelled by defining shredder connection, and shredder liberation tables, based on heuristic derived from extensive data collection on design and shredder experiments. This allows for the design-driven modelling of material liberation in the shredding process. The heuristic modelling of liberation behaviour and the prediction of particle composition and degree of liberation after shredding based on design-driven shredder modelling are a novel approach to link design to recyclate quality and recycling performance. Key to this work is that the multi- and mono-material composition of particles after liberation are not pre-defined as in earlier work, but are predicted as a function of design choices and can vary accordingly. The evolution of these models is underpinned by extensive industrial data collection on product design of various WEEE categories, which was used to define and calibrate the models. The time dependent characteristic of design and recycling technology requires the adoption of dynamic modelling to predict recycling performance over time. The discussed dynamic and predictive framework provides a first principles basis for the calculation of the dispersion of harmful/valuable elements and environmental impact. This optimizes the spending of money on large scale trials and monitoring test to establish these results. At the same time legislation and policy on recyclability of products can be tested if it stands the test of time or if new designs comply with environmental legislation in the future. Prediction of recycling performance, recyclate quality and toxicity as a function of product design, shredding and (future) recycling system configurations will be demonstrated in various industrial examples, which includes an evaluation if shredding is advisable or not for precious and platinum group elements recovery during recycling. Also the so important balance between energy recovery and feedstock recycling is shown.

Platinum-group element distribution in the oxidized Main Sulfide Zone, Great Dyke, Zimbabwe

In the Great Dyke mafic/ultramafic layered intrusion of Zimbabwe, economic concentrations of platinum-group elements (PGE) are restricted to sulfide disseminations in pyroxenites of the Main Sulfide Zone (MSZ). Oxidized ores near the surface constitute a resource of ca. 400 Mt. Mining of this ore type has so far been hampered due to insufficient recovery rates. During the oxidation/weathering of the pristine ores, most notably, S and Pd are depleted, whereas Cu and Au are enriched. The concentrations of most other elements (including the other PGE) remain quite constant. In the oxidized MSZ, PGE occur in different modes: (1) as relict primary PGM (mainly sperrylite, cooperite, and braggite), (2) in solid solution in relict sulfides (dominantly Pd in pentlandite, up to 6,500 ppm Pd and 450 ppm Pt), (3) as secondary PGM neoformations (i.e., Pt–Fe alloy and zvyagintsevite), (4) as PGE oxides/hydroxides that replace primary PGM as the result of oxidation, (5) hosted in weathering products, i.e., iron oxides/hydroxides (up to 3,600 ppm Pt and 3,100 ppm Pd), manganese oxides/hydroxides (up to 1.6 wt.% Pt and 1,150 ppm Pd), and in secondary phyllosilicates (up to a few hundred ppm Pt and Pd). In the oxidized MSZ, most of the Pt and Pd are hosted by relict primary and secondary PGM; subordinate amounts are found in iron and manganese oxides/hydroxides. The amount of PGE hosted in solid solution in sulfides is negligible. Considerable local variations in the distribution of PGE in the oxidized ores complicate a mineralogical balance. Experiments to evaluate the PGE recovery from oxidized MSZ ore show that using physical concentration techniques (i.e., electric pulse disaggregation, hydroseparation, and magnetic separation), the PGE are preferentially concentrated into smaller grain size fractions by a factor of 2. Highest PGE concentrations occur in the volumetrically insignificant magnetic fraction. This indicates that a physical preconcentration of PGE is not feasible and that chemical, bulk-leaching methods need to be developed in order to successfully recover PGE from oxidized MSZ ore.

A new approach to precious metals recovery from brown coals: Correlation of recovery efficacy with the mechanism of metal–humic interactions

The presence of gold and platinum group elements (PGE) in low-rank brown coals around the world has promoted interest in the industrial exploitation of this alternative source of precious metals. However, due to low efficacy of the methods traditionally used for the processing of mineral ores, there exists a high demand for new strategies of precious metal recovery from refractory carbonaceous materials that could significantly increase the economic potential of gold- and PGE-bearing organic resources. Here we discuss the possibility of gold and PGE recovery from alkaline extracts of brown coals using the difference in colloidal stability of bulk organic matter and its fractions enriched with precious metals. This approach enables one to avoid complete oxidation or combustion of brown coals prior to gold recovery, to minimize organic content in gold concentrate, and to obtain a valuable by-product – humic extracts. Using gold-bearing brown coals from several deposits located in the South Far East of Russia, we show that up to 95% of gold can be transferred to alkaline extracts of humic acids (HA) and up to 85% of this gold can be recovered by centrifugation at pH 4.0–6.0, when only 5–15% of HA precipitated simultaneously. We have shown that the high efficacy of gold recovery can be attributed to the occurrence of fine-dispersed elemental gold particles stabilized by HA, which differ significantly in colloidal stability from the bulk organic matter and, thus, can be separated by centrifugation.

Determination of platinum group elements (PGE) in environmental samples by ICP-MS: a critical assessment of matrix separation for the mitigation of interferences

The accurate determination of Pt, Pd and Rh (platinum group elements, PGE) in environmental samples by inductively coupled plasma–mass spectrometry (ICP-MS) is complicated by molecular ion and doubly charged ion interferences (e.g. CuAr+, HfO+, SrO+, YO+ and Pb2+). Two matrix separation techniques, tellurium coprecipitation and ion exchange, are compared in terms of PGE recovery and separation efficiency by application to standard solutions and the new reference material BCR-723, a road dust material collected from the Tanzenberg Tunnel in Austria. Ion exchange using Dowex AG50W-X8 strong cation exchange resin is highly effective in terms of both separation efficiency (&gt;99.7%) and PGE recovery (93.8 to 101.6%) and allows accurate, interference-free determination of PGEs. PGE recoveries by Te coprecipitation are incomplete, ranging between 84 and 93%. Significant amounts of copper (up to c. 30%) are also recovered, which, due to CuAr + formation, causes interferences on both 103 Rh (monoisotopic) and 105 Pd. The generally high levels of interferent elements in environmental matrices (e.g. Cu, Pb etc.) and the magnitude of Cu recovery by Te coprecipitation limit the applicability of data correction using interference equations. Reductions in Cu yield by Te coprecipitation, to &lt;4%, can be achieved through the addition of potassium iodide prior to matrix separation. However, due to the incomplete recovery of PGEs, ion exchange provides a more robust method for the determination of low levels of PGEs.

Effects of frother type and froth height on the flotation behaviour of chromite in UG2 ore

About 70% of the UG2 reef consists of the gangue mineral chromite (FeO · Cr 2O 3). In the processing of UG2 ore by flotation for the recovery of platinum group elements (PGEs) the presence of chromite in the concentrates can cause serious downstream processing problems and a grade of less than 3% Cr 2O 3 is sought. This constrains operating procedures and compromises optimum recovery of the PGEs. In this study, the influence of the froth phase on the recovery of chromite was investigated by changing both frother type and dosage and froth height in batch scale flotation tests. The results obtained showed that it was possible to obtain concentrates with less than 3% Cr 2O 3 content by increasing the froth height, allowing for better drainage of both entrained gangue particles and coarse particles with low hydrophobicity. At a 3 cm froth height, very low water and mass recovery were obtained and thus low entrainment. Nevertheless a small amount of chromite particles coarser than 45 μm was persistently recovered which may be attributed to the true flotation of these particles. The mechanism of chromite recovery was discussed on the basis of the difference in the appearance of the froth structure and water recovery.

Equilibrium Distributions of Actinides and Fission Products in Pyrochemical Separation Systems, (I)

A thermodynamic model has been described for predicting the equilibrium distributions of solute elements in pyrochemical separation systems of molten salts and liquid metals. The model assumes that the solutes are present in the form of complex compounds in the salt phase and of intermetallic compounds in the metal phase. The thermodynamic data of these compounds are available in the literatures and also can be estimated theoretically. Some predictions have been made by the present model to discuss the feasibility of liquid metal extraction system for the recovery of platinum group elements from spent fuel.

Equilibrium distributions of actinides and fission products in pyrochemical separation systems. (I). Thermodynamic model.

A thermodynamic model has been described for predicting the equilibrium distributions of solute elements in pyrochemical separation systems of molten salts and liquid metals. The model assumes that the solutes are present in the form of complex compounds in the salt phase and of intermetallic compounds in the metal phase. The thermodynamic data of these compounds are available in the literatures and also can be estimated theoretically. Some predictions have been made by the present model to discuss the feasibility of liquid metal extraction system for the recovery of platinum group elements from spent fuel.