FIELD OF THE INVENTION The present invention relates to extraction of precious metals from scrap or from mineral earth, in particular to the extraction of such metals from natural sources as well as from scrape of electronic circuits, computer scrap and catalysts.

BACKGROUND OF THE INVENTION Precious metals which are non-ferrous metals such as gold and platinum group metals (PMG' s) exist in many industrial products, such as electronic scrap (electrical/electronic equipment), printed circuit boards and relay scrap. Gold, silver and palladium are frequently used for electroplating of connectors and contacts because they have excellent corrosion resistance and high electrical conductivity. In addition, silver and gold are used in hybrid links and solders. Such metals also exist in vehicles exhaust catalysts. Merely discarding old printed circuits, old computers or vehicle's exhaust catalysts poses an environmental problem. Furthermore, it results in a raise in their price. Electronic scrap is composed of plastics (30%), refractory oxides (30%), and metals (40%). The metals are found in many electronic components such as edge connectors, integrated circuits, and transistors. Generally, the amounts of the metals are the following (wt.%): 0.02-0.05 Au, 0.1-0.2 Ag, 0.01-0.005 Pd, 14-18 Cu5 1-2 Ni, 4-6 Sn, 2-4 Pb, 5-8 Al and 1-3 Fe. The physical forms of these precious metals may be plated gold or palladium on copper laminate and plated gold or silver on nickel or iron. Thus, the main economic driving force for the recycling of electronic scrap is the recovery of precious metals, particularly gold and palladium. Environmental considerations are another, very important consideration in electronic scrap recycling. The disposal of obsolete electronic equipment is a problem of considerable magnitude. Treating these wastes in a way that does not harm the environment is a complex process due to the heterogeneous composition of the obsolete equipment. Electronic scrap is recycled using mechanical, hydrometallurgical, and thermal processes. The mechanical processing of scrap is usually used as a pre-treatment and its main purpose is to separate compounds and components from the initial scrap. Such processing is done by comminuting the scrap, classification, and separation (by differences of density, weight, size, magnetic properties, etc.) of the different compounds included in the wastes. After each comminution step, the resulting fraction is, in general, already enriched in a certain material that can be separated from the main stream. After the last step, concentrate and wastes can be further refined by air/or hydraulic classification and density separation. Various methods are known in the art for recovering silver, gold and platinum group metals (PMG' s) from separated metal containing concentrates and semi-products. Several methods exist for obtaining back such precious metals from such used scrap. The two main approaches are either pyrometallurgy or hydrometallurgy. In the pyrometallurgy process the precious metals-bearing scrap is melted in common with copper scrap to a high temperature where the metals are maintained and the connecting matrix, e.g. polymers, is burnt. In the hydrometallurgy process, the metals are freed from the matrix by immersing in an appropriate solution, where the temperature is maintained at elevated temperatures of 100-2500C. For example, HCl boils at a temperature of about 1100C. So is HNO3. H2SO4 exists as a liquid up to 24O0C. The chemistry involved in such processes are reduction and oxidation of the metals in order to extract them and bring them into solution. The main steps in hydrometallurgical processing consist of a series of acid or caustic beads of finely divided solid material. The solutions are then subjected to separation procedures such as electrowinning, solvent extraction, precipitation, cementation, ion exchange, filtration and distillation to isolate and concentrate the metals of interest. Cyanide pressure leaching was examined (CM. Madelines et al, Hydrometallurgy, 35: (1994)) for recovering silver, gold and palladium from precious metals containing scrap. The remarkable success of cyanide as lixiviate for gold is due to the enormous stability of the dicyanoaurate ion. However, the high concentration of copper in initial raw material might interfere with the selective recovery of silver, gold and palladium. In addition, cyanide process where substantial care is needed to prevent evolution of the hazardous HCN gas and destruction of the remaining cyanide in the waste solution. Another approach by Loebel and Meisner (DD 253048, 1988) illustrated the recovery of copper, gold, silver, palladium, tin and lead from electronic scrap that also includes plastics (e.g., connectors and printed circuits). The scrap is treated with a 30-50% HNO3 solution at 35°C to yield metal nitrates in solution, solid gold and H2O. The solution is cooled, sulfates are added to water, forming a solution containing copper, palladium, silver, and the insoluble material (Au, SnO2 and H2SO4) is filtered off. The filter cake is melted with Na2CO3 to yield gold metal, while the filtrate is cemented with copper at 40-500C to yield an Ag- Pd alloy. Kang (In Rhel et al, Recycling of Metals and Engineered Materials, ed. P.B Queneau and R.D. Peterson (Warrendale, PA: + MS, 1985) pp,. 469-478), leached the metal scrap in 50% aqueous solution HNO3(?) and gold was dissolved completely at 600C within 2 hours. All the metal particles were dissolved. The resulting solution was boiled to remove hazardous nitrous compounds from the aqueous media. by J. Zakrenski et al (PL 146,269, 1989) suggested loading of scrap in diluted H2SO4 and 30% H2O2 at 50-800C to dissolve nickel, iron, and a salt. The residue containing gold was treated with aqueous medium and filtered. The gold from the final solution was precipitated by reaction with Na2SO2. AIl these processes suffer from high reagent consumption corrosion problems and the need for scrubbing emitted hazardous gases. Han et al. in U.S. Patent 5,989,311 (1999) relates to a method for the recovery of various metals by combined use of nitrate and halogen salts in the presence of oxygen. In this technology copper and precious metals are extracted from their elemental states using bromine/bromide or iodine/iodide plus sodium nitrate and oxygen. At the same time, the halogen ions, such as bromide or iodide, and nitrogen oxide generated from the pressure leading process are continuously converted from respectively, to bromide or iodide and to nitrate, which are the major reagents for the production of copper and gold ions in the leaching solution. However pressure leaching is expensive and can render the process complicated with respect to materials and handling. F. Solomon in U.S. Patent 4,997,532 (1989) suggested a process based on treating precious metal-bearing scrap with a polar organic solvent where bromine is dissolved to form non-aqueous precious metals complexes. The organic solvent is subjected to electrolysis to recover the extracted metals. This process is not suitable for use with combined scraped metal second material. This is because highly flight of polar organic solvent at temperature 80-900C which the optimal condition of platinum and rhodium dissolution in bromine/bromide system. Besides polar organic solvent so possess necessary selectivity in liquid- liquid extraction process by respective precious metals recovering from leaching solution, containing also copper, metal, iron, zinc, etc. Other methods for recovering gold, silver and platinum group metals consist of oxidation and lixiviation of the precious metals by liquid solvents such as hydrochloric acid and hydrogen peroxide mixtures, acid chlorine solutions, and hyper chlorite. These methods yield low recovery rates and are only useful for specific source materials. They also require sophisticated machinery and the process themselves are very time-consuming, and require large amounts of energy. When used with electronic scrap the active plastic and ceramic onto which the plated precious metals are impregnated presents an obstacle to the process. In liquid-phase lixiviation two competing processes always occur due to the large contact surface of crushed plastic and ceramic, the adsorption of gold ions from the crushed ceramic surface into the solution, and their re-adsorption. Because of the two reactions that constantly are occurring, it is necessary to perform repeated cycles of lixiviation and washing in order to ensure complete recovery of the metals. Thus, a large volume of acid is required, and the extra cited PMG' s is diluted due to the large amount of solution employed. This translates into a high consumption of energy and time, as well as high costs.

SUMMARY OF THE INVENTION The present invention is based on the fact that a novel approach to extract non-ferrous precious metals from mineral earth, mineral materials, and scraped metal second source including electrical and electronic scrap as well as catalysts, was developed. The process of the present invention allows for a high percentage of separate recovery of platinoids, gold and silver into solution, requiring a relatively small amount of reagent. The approach is based on the use of two complexing agents, an aqueous agent in combination with an organic agent. Thus the present invention is directed to a method for recovering non- ferrous precious metals from raw material such as mineral earth and scrap of electronic appliances and catalysts. The raw material is initially (a) mechanically crushed and shredded to obtain dust containing fine particles which is further processed by the steps of: (a) mixing the dust containing fine particles in concentrated HX solution containing MgX2, (X = halogen), heating the mixture, cooling, filtering and obtaining a solid; (b) mixing the obtained solid in an aqueous solution containing a strong acid further containing MgX2 (X = halogen) and further adding a strong oxidizing agent, heating the mixture, cooling, filtering and obtaining a solid; (c) mixing said solid in a mixture of an aqueous HX (X= halogen) solution where X" ions are added to achieve an appropriate X" concentration, adding an organic extracting and complexing agent in the form of (RO)3PO, R being C3-10-alkyl, in a C9-17- hydrocarbon and further adding an oxidizing agent capable of transmitting halogen ions into free halogen, adjusting the pH in the range of about 2-4 by addition of a strong base and separating the phases wherein from the organic phase are obtained by sedimentation all of the gold and a minor amount of palladium initially present in the dust; (d) mixing the solid residue in an aqueous solution whose pH is adjusted to 0.8 by HX, further adding X' ions to achieve appropriate X" concentration, adding an organic extracting and complexing agent in the form of NR3, R being C5-12-alkyl, in a C9-17-hydrocarbon and further adding an oxidizing agent capable of transmitting halogen ions into free halogen, heating, cooling and separating the phases, wherein from the organic phase are obtained by sedimentation essentially all of the remaining palladium present in the dust; and (e) mixing the solid residue in an aqueous solution comprising HX and alkali metal halogen to achieve appropriate X" concentration, adding an organic extracting and complexing agent in the form of triisobutylphosphinesulfoxide in (RO)3PO, R being C3-i0-alkyl, and C9-i7-hydrocarbon and further adding an oxidizing agent capable of transmitting halogen ions into free halogen, heating, cooling and separating the phases, wherein from the organic phase are obtained by sedimentation essentially all of the platinum present in the dust. Preferably, the dust is obtained from a cyclone system, the strong acid is a H2SO4 and the halogen used is bromine or chloride, most preferably bromine. The oxidizing agent should be a strong agent capable of converting halogen ions into basic halogen. Such an oxidizing agent may be H2O2 or MnO2. Preferably, the combined organic extracting and complexing agents are added in a form supported by macro-pore matrix, most preferably polypropylene.

DETAILED DESCRIPTION OF THE INVENTION The method of the present invention concerns recovering of precious, non- ferrous metals from raw material. The raw materials are preferably from electrical and electronic scrap, catalysts or from mineral earth. Non limiting examples of electronic scrap are electronic circuits such as printed circuits, computer scrap such as boards and screens, catalysts are catalytic converters, and homogeneous and heterogeneous catalysts, mineral earth are minerals of natural source which include the various non-ferrous precious metals. The method according to the present invention employs hydrometallurgical techniques, use of liquid reagents for treating and reducing of ores. Prior to commencing the hydrometallurgical process the raw material is processed by an initial mechanical processing to yield a fraction of fine particles. In case of electronic scrap, the processing comprises shredding the scrap, desk vibration separator, dust extraction system to obtain three separate groups differing in their particle size and density. The first group has a high density typically above 8 gr/c3 and this fraction contains mainly copper concentrate bearing non-ferrous precious metals. The second group has a density of less than 1.5 gr/c containing mainly plastic, rubber and residues of metals. These two fractions are removed where a third group containing dust from cyclone system containing fine particles is further processed according to the present invention. Mineral earth ores are directly subjected to the hydrometallurgical process. The process employs a complexation/oxidation leaching process where the concentration of the oxidized metals in the leach media is continuously increased through the use of acid mixtures both aqueous and organic solution for achieving precious metal oxidation. As mentioned above, in the case of recovering the metals from electronic scrap or from used catalysts, the initial raw material is first subjected to a mechanical/or chemical enrichment obtaining concentrates of precious metals including platinum group metals (PGM' s), gold and silver. The obtained concentrate is then treated in an aqueous solution containing halogen ions in an amount sufficient for forming in the system complex anions of silver, gold and PGM' s as M+11 Xpp"π. To the solution is added an oxidizing agent which is able to transfer free halogen ions (e.g. Cl", Br", I") into the neutral free halogen (e.g. Cl2, Br2 or I2) in the leach liquor. The oxidizing agent may be any strong agent capable of converting 2X" to X2 where non- limiting examples of such an oxidizing agent are H2O2, MnO2; KMnO4 and chromates. The addition of an organic solvent, serving as a complexant for halogen and halogen ions overcomes, and actually eliminates, the known problem of inadequate contact between metals and the oxidative agent resulting from diffusion and thus ensures full separate transfer of the extracted precious metals to the organic solvent, thus obviating the need for many repeated cycles. Two preferred embodiments for carrying out the invention will now be disclosed. In one embodiment, the extraction process is carried stepwise, recovering in each step a different type of precious metal or a mixture of two metals. In a second embodiment a continuous one-pot process is carried resulting in a fortified mixture of all precious metals. As mentioned, in case the non-ferrous metals to be extracted orginate from scrap of electronic appliances, the scrap is initially mechanically crushed resulting in three separate groups differing in their particle size and density where the third group containing dust from cyclone system containing fine particles is further processed according to the present invention. Turning to the first method, in order to increase the efficiency of the extraction process and reduce the consumption of reagents, the concentrate of the mechanical grinding is first treated with HX acidic medium to eliminate residual metals. Preferably, HX is hydrochloric acid solution, where a solution containing 50-200 gr/L HCl and 250-320 gr/L MgCl2 at the temperature of 85-110°C for about 1.5-3 hours followed by filtration. The solution contains aluminum, tin and lead. The obtained solid residue is cleaned, and further treated with a strong acid such as hydrosulfuric acid solution, containing 50-200 gr/L H2SO4 and 150-250 gr/L MgCl2. The pulp (solution) is now brought to the desired redox-potential in the range of 450-550 mV, by the addition to the pulp of 50% solution of hydrogen peroxide at a temperature of 60-800C. The solution is agitated until the majority of copper, nickel and zinc are removed. In order to selectively recover gold, the solid residue obtained, is transferred into an aqueous solution having a pH of 2-4, obtained by adding HX (X = halogen) to the solution and further introducing bromide ions in the form of NaBr until a total concentration of 80- 100 gr/L Br" is obtained. The pulp obtained is mixed with 20-50% solution of tributylphosphate in kerosene where the aqueous phase to organic phase ratio is 20-50:1. The combined phases having a redox-potential 850-950 mV are further vigorously agitated at a temperature not exceeding 30-400C, until complete dissolution of gold is achieved where the gold is now in the organic solvent. Phase separation yields the gold as powder. For selective recovery of palladium, the residual solid is transferred into a water solution with pH 0.5-1.5 obtained by the addition of HBr or HCl, where bromide ions are additionally introduced in the form of CaBr2 until the total concentration of bromide ions is 150-250 gr/L Br". The obtained pulp is mixed with 2-20% solution of trioctylamine in kerosene and with intensive stirring of the said pulp, a solution comprising an oxidant is added. In cae the oxidant is hydrogen peroxide, a solution of 50% hydrogen peroxide is added setting the redox-potential of the system at a level of 680-870 mV where the temperature is kept at a range of about 50-600C. The organic phase is separated and palladium is recovered from the organic phase by sedimentation using Zn in acidic medium. The acidic medium is preferably, HCl (50gr/liter), however, also H2SO4 or any other acid may yield similar results. In order to selectively recover platinum, the obtained solid residue is cleaned and transferred into the water solution having an HBr or HCl concentration of between 20 to 30 gr/liter, followed by the addition of bromide ions in the form of CaBr2 or NaBr in order to obtain a total concentration 80-120 gr/L Br". The obtained pulp is further mixed with 10-20% solution of the tertiary aliphatic amine, preferably, triisobutylphosphinesulfoxide in tributylphosphate and kerosene. A solution of a strong oxidizing agent such as MnO2 or H2O2 is added. Preferably 50% hydrogen peroxide is added to the aqueous/organic slurry while vigorous agitating the pulp is continued providing a redox-potential to the system at an optimal level of 800-950 mV where the temperature is kept such that it does not exceed 80-950C. The agitation is continued where all platinum is in the organic phase. The organic phase is separated and platinum is recovered from the organic phase by sedimentation using Zn in acidic medium. The acidic medium is preferably HCl (50 gr/liter), however, also H2SO4 or any other acid may yield similar results. It may be noted that according to the present invention, all the precious metals are extracted from the aqueous phase into the organic phase from which they are easily obtained. It should further be noted that the temperature, concentration of bromide ions and bromine, the redox-potetnial of the solution and the pH of the leach solution are important and should be maintained carefully. After each stage of leaching operation, the pulp is transferred from the reactor and filtered. The solid residue is washed thoroughly and the washed material is combined with the mother solution. The platinum group metals (PGM' s), gold and silver which are in the organic solution, can then be recovered from using known methods. Turning to the second method, initially, in order to increase the efficiency of the extraction process according to the present invention and in order to reduce the consumption of the employed oxidant, the shredded scrap is treated with an acidic solution, HX (X = halogen) preferably, hydrochloric acid solution, containing 50-220 gr/L HX and 150-320 gr/L MgX2 (X = halogen), preferably MgCl2 at the temperature of 85-1100C for 1.5-3 hrs followed by filtration. The solid residue is then mixed with 50-100 gr/L hydrosulfuric solution, containing 180-320 gr/L MgCl2 in 60-80 gr/L bromide-ions in the form of NaBr. A solution comprising 5.0% trioctylamine in tributylphosphate and kerosene at the ratio of aqueous phase to organic phase 100-200:1 is added to the above slurry, followed by the addition of a strong oxidizing agent such as MnO2 or H2O2. Preferably, 50% solution of hydrogen peroxide is added resulting in the formation of a system having a redox-potential of 650-870 mv. From the resulting leach solution, silver, gold, palladium and platinum sequentially precipitate in the form of powders of the respective precious metals, where the mother liquor is returned to fresh scarp material processing. The present invention will now be described with reference to the following examples. The examples however are not intended to be limiting in any manner. They are provided only as preferred embodiments of the process of the present invention.

EXAMPLE 1: 15 Kg of original computer scrap with initial metal content, wt. %: 17.85 Cu, 4.78 Al, 5.28 Sn, 2.0 Fe, 2.17 Zn, 4.19 Pb, 1.63 Ni, 0.13 Ag, 0.035 Au, 0.025 Pd, 0.00046 Pt was crushed and separated by differences of size and density with obtaining three separate fractions. The first fraction, weighted 4.18 kg is copper concentrate with following composition, wt. %: 54.5 Cu, 5.74 Fe, 4.66 Ni, 1.62 Al, 1.34 Sn, 2.9 Zn, 0.81 Pb, 0.27 Ag, 0.0645 Au, 0.042 Pd other plastic, rubber, glass, humidity. This particular fraction comprising a copper concentrate bearing precious metals can be sent direct to copper refinery. The second fraction, weighted 8.12 kg, contained essential amount of plastic, rubber and only 5.64% Al, 1.43% Fe, 0.49% Ni, 6.9% Sn, 2.1.% Zn, 2.28% Cu, 5.66 Pb. The third fraction is dust from cyclone system, weighted 2.56 kg, with total content, wt., %: 6.78 Cu, 7.20 Al, 6.87 Sn, 0.98 Zn, 5.23 Pb, 0.295 Ag, 0.106 Au, 0.0769 Pd, 0.00269 Pt and 78.46 plastic, rubber, humidity. The third fraction is most suitable for recovering the precious metals, was processed by a hydrometallurigical technique. The processing of the dust product was performed in a 20 L Pyrex round-bottomed flask placed in a thermostatically controlled water bath. The flask was installed with mechanical stirrer and fitted with Pyrex condenser, thermometer, electrodes for pH and Eh measurement, and three tubes for addition 40% solution of NaOH or 20% solution of hydrobromic acid and 50% hydrogen peroxide solution. Preparation of first leach solution: 5.3L of 32% hydrochloric acid solution which is 120% of stochiometric amount required to complete the dissolution of aluminum, tin, lead, and zinc was mixed with a prepared 500 gr/L solution of magnesium chloride, and the mixture was poured into a flask. 2.65 Kg of dust product (third fraction) was added to the solution and the slurry was stirred. The first leaching process was conducted for 3 hours at temperature 950C. During this period 98% of aluminum, 96% of tin, 98.6% of lead and 92% of zinc were extracted in each solution. The leaching residue was filtered, washed and then transferred to begin a second stage of processing. To recover copper, the solid residue was treated with 8 L of hydrosulfuric acid solution, containing 60gr/L H2SO4 and 150gr/L MgCl2. The redox-potential of the leach system near 45OmV was adjusted by adding to the pulp a solution of 50% solution of hydrogen peroxide. The second leach procedure was performed for a period of 3 hours at a temperature of 65°C. In result more than 92% of the copper and 94% of the nickel originally present were dissolved by this procedure. The solid residue was filtered, washed and in order to recover the precious metals present therein was treated with 3L of acidic sodium bromide solution containing 10 gr/L hydrobromic acid and 80 gr/L bromide ions. To the pulp was added a solution of 50% solution of tributylphosphate in kerosene where the phase ratio between the aqueous phase and the organic phase is 20:1. Thereafter, with intensive contact of phases, to the pulp are added a 50% solution of hydrogen peroxide providing in the system redox-potential 850 mV. The acidity of the leach system was reduced to pH 4 by addition of 40% solution of NaOH. In result more than 99% of gold and 20% of initial amount of palladium were transferred to the organic phase within a period of 3 hours at ambient temperature. The organic phase was separated and the solid residue was used for further separation. In order to selective recover palladium, the solid residue was transferred into a water solution with pH 0.8, achieved by the addition of HBr, and to the said solution bromide ions were introduced in the form of CaBr2 until the total concentration of bromide ions is 180 gr/L Br'. The obtained pulp is further mixed with 10% solution of trioctylamine in kerosene and with intensive stirring of the pulp it was further treated by 50% H2O2 providing redox-potential of the system at the level of 750 mV and the temperature not more than 600C. The pulp was stirred for 2 hours, where all the palladium was transferred to the organic solution. The organic phase is separated and palladium is recovered from the organic phase by sedimentation using Zn in acidic medium. Preferably, in HCl (50 gr/liter), however, also H2SO4 or any other acid may yield similar results. The residual solid was further processed in order to selective recover platinum. The residual solid was washed and transferred into a water solution, containing 50gr/L HBr and 100gr/L CaBr2. To the obtained pulp were added a 5% solution of the triisobutylphosphinesulphoxide in tributylphosphate and kerosene and while the pulp is vigorously stirred, there was added rapidly a solution of 50% solution of H2O2 providing a redox-potential of the system at the level of 850 mV and the temperature not more than 750C. Stirring was continued where all the platinum is transferred to the organic solution and recovered by evaporation of the solvent. Contents of gold, silver, palladium and platinum in filtrate solutions (aquaphase) and organic phases were measured using an inductially coupled plasma-atomic emission spectrometer (ICP-AES) standard solutions [1,000 ppm from SPEX (Metuchen, New Jersey)] of each metal were diluted with bi-distilled water and used for calibration of the ICP-AES. The precious metal recovering efficiency was estimated from the weight of the dissolved metal and the initial weight of the same metal in the representative sample of raw material. From the said organic solution gold and PGM' s being precipitated in the form of tradable powders by adding zinc powder. In result, about 96% of gold, 82% of silver, 96% of palladium and 94% of platinum were extracted using proposed process.

EXAMPLE 2: A 1.5 kg of representative dust fraction sample from cyclone system with total content, wt, and %: 5.64 Cu, 8.36 Al5 7.24 Sn, 0.82 Zn5 5.64 Pb5 0186 Au5 0.118 Au5 0.0484 Pd, 0.00183 Pt and plastic, rubber, humidity other was processed for recovering precious metals. The processing of the dust product was performed in 20 L Pyrex round-bottomed flask placed in a thermostatically controlled water bath. The flask was installed with mechanical stirrer (turbine type) and fitted with Pyrex condenser, thermometer, electrode for Eh measurement and a tube for addition 50% hydrogen peroxide solution. A 1.5 kg of dust product was added to 6.0 L of hydrochloric solution, containing 180gr/L HCl and 250gr/L MgCl2 and then the slurry was stirred. The first leaching process was conducted for 3 hrs at temperature 80-950C. During this period 98% of aluminum, 94% of tin, 96% of lead and 94% of zinc were extracted in each solution. The leaching residue was filtered and then transferred to thermostatically flask for a second stage of processing. To recover copper, the solid residue was treated with 4.6 L of hydrosulfuric acid solution, containing 50 gr/L H2SO4 and 200 gr/L MgCl2. The redox-potential of the leach system near 550 mv was adjusted by adding to the stirred pulp a solution of 50% solution of hydrogen peroxide. The second leach procedure was performed for a period of 2.5 hrs. at a temperature 800C. In result more than 96% of copper and 98% of nickel were dissolved by this procedure. The solid residue was filtered, washed and in order to combine recovering of precious metals Was placed in the thermostatically flask, 3 L of acidic sodium bromide solution containing 30 gr/L HCl and 180 gr/L bromide-ions were added. To the obtained pulp were added 100 mL of 10% solution of triisobutylphosphinesulfoxide in tributylphosphate and kerosene during intensive stirring of this pulp it was further treated by 50% hydrogen peroxide solution providing redox-potential of the system at the level near 850 mv and the temperature not more than 6O0C. The pulp was stirred for 3 hrs where more than 98% of gold, 84% of silver, 96% of palladium and 92% of platinum were extracted by organic solvent. The metals were extracted from the organic solution by sedimentation using Zn in an acidic medium.

EXAMPLE 3: A 1.46 kg mixed spent catalytic converters sample was found to have 0.08% platinum, 0.04% palladium and 0.008% rhodium by mass. It was treated with 6 L of acidic sodium bromide solution containing 60 gr/L HCl and 220 gr/L bromide-ions. To the obtained pulp were added 60 mL 8% solution of triisobutylphosphinesulfoxide in tributylphosphate and kerosene. The three components organic solvent was supported by 300 gr. macro pore polypropylene granules and during intensive stirring of the pulp it was further treated by 50% hydrogen peroxide solution providing the redox-potential of the system at about 870 mv and the temperature not more than 600C. The pulp was stirred for 3 hrs where more than 94% platinum, 96% palladium and 94% rhodium were recovered by supported organic solvent.

EXAMPLE 4: A 1.2 kg of platinum, palladium and gold bearing quartz-feldspar porphyry sample from the Coronation Hill deposit was treated in condition of proposed process. It was leached with 4 L of acidic sodium bromide solution containing 86 gr/L HCl and 120 gr/L bromide-ions. To the obtained pulp were added 20 mL of 12.8% solution of triisobutylphosphinesulfoxide in tributylphosphate and kerosene. To the combined organic solvent were added 60gr macropore polypropylene granules and the pulp stirred vigorously. An oxidizing agent, 20% concentrate of manganese dioxide, was added providing the redox-potential of the system at a level of more than about 850 mv and the temperature not more than 600C. The pulp was stirred for 6 hrs where more than 92% platinum, 94% palladium and 96% gold were extracted by supported organic solvent.

EXAMPLE 5: A 1.5kg of complex gold bearing concentrate composed of pyrite and stibnite was treated in condition of proposed process. It was leached with 6 L of acidic sodium bromide solution containing 60gr/L HCI and 220gr/L bromide- ions. To the obtained pulp were added 140 mL of supported liquid membrane (SLM) in form of 12.8% solution of triisobutylphosphinesulfoxide in tributylphosphate and kerosene. 160gr of Celgard-2500, a micro porous polypropylene film, was used as the solid support for the liquid organic solvent and the pulp stirred vigorously. An oxidizing agent, 20% concentrate of manganese dioxide, was added providing the redox-potential of the system at a level of more than about 850mv and the temperature not more than 30-400C. The pulp was stirred for 6hrs where more than 92% gold were extracted by supported organic solvent.