CA1228989A - Recovery of precious metals from materials containing same - Google Patents
Recovery of precious metals from materials containing sameInfo
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- CA1228989A CA1228989A CA000464865A CA464865A CA1228989A CA 1228989 A CA1228989 A CA 1228989A CA 000464865 A CA000464865 A CA 000464865A CA 464865 A CA464865 A CA 464865A CA 1228989 A CA1228989 A CA 1228989A
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- C—CHEMISTRY; METALLURGY
- C22—METALLURGY; FERROUS OR NON-FERROUS ALLOYS; TREATMENT OF ALLOYS OR NON-FERROUS METALS
- C22B—PRODUCTION AND REFINING OF METALS; PRETREATMENT OF RAW MATERIALS
- C22B11/00—Obtaining noble metals
- C22B11/06—Chloridising
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Abstract
"RECOVERY OF PRECIOUS METALS FROM
MATERIALS CONTAINING SAME"
Abstract of the Disclosure The recovery of precious metals from materials containing same is effected by leaching the precious metal in an acidic aqueous medium contain-ing chloride ions and a strong oxidizing agent to form a soluble chloride complex of the precious metal and thereafter separating the metal from the complex thus formed. According to the invention, the chloride ions are derived at least in part from a non-acidic chloride-containing leaching agent which is non-volatile and soluble in the acidic aqueous medium without crystallizing under the leaching conditions. The invention is particularly useful for recovering platinum and palladium from automobile catalytic converters.
MATERIALS CONTAINING SAME"
Abstract of the Disclosure The recovery of precious metals from materials containing same is effected by leaching the precious metal in an acidic aqueous medium contain-ing chloride ions and a strong oxidizing agent to form a soluble chloride complex of the precious metal and thereafter separating the metal from the complex thus formed. According to the invention, the chloride ions are derived at least in part from a non-acidic chloride-containing leaching agent which is non-volatile and soluble in the acidic aqueous medium without crystallizing under the leaching conditions. The invention is particularly useful for recovering platinum and palladium from automobile catalytic converters.
Description
~2~98~
Backqround of the Invention The present invention is concerned with a process for recovering precious metals from materials containing same. More particularly, the invention is directed towards the recovery of platinum and palladium from automobile catalytic converters Platinum and palladium are widely used for catalytic conversion of toxic exhaust gases from automobiles to non polluting products. Since 1975, about 140 tons of platinum and 47 tons of palladium have been consumed annually in the United States and Canada for this purpose, which represents a major consumption of these metals. Also in Japan, new environmental protection rules now oblige most cars to be equipped with catalytic converters. It is expected that in some European countries the same restriction will be introduced. Thus, scrapped automobile converters are becoming an important poten--tial source of recycled platinum and palladium.
In general, two types of catalyst are used in automotive emission-control equipment:
1. alumina pellets containing from 330 to 500 ppm platinum and 110-200 ppm palladium, and
Backqround of the Invention The present invention is concerned with a process for recovering precious metals from materials containing same. More particularly, the invention is directed towards the recovery of platinum and palladium from automobile catalytic converters Platinum and palladium are widely used for catalytic conversion of toxic exhaust gases from automobiles to non polluting products. Since 1975, about 140 tons of platinum and 47 tons of palladium have been consumed annually in the United States and Canada for this purpose, which represents a major consumption of these metals. Also in Japan, new environmental protection rules now oblige most cars to be equipped with catalytic converters. It is expected that in some European countries the same restriction will be introduced. Thus, scrapped automobile converters are becoming an important poten--tial source of recycled platinum and palladium.
In general, two types of catalyst are used in automotive emission-control equipment:
1. alumina pellets containing from 330 to 500 ppm platinum and 110-200 ppm palladium, and
2. silicate monolithic honeycombs contain-ing about 800 to 1,500 ppm platinum and 100 to 350 ppm palladium.
In both cases, small quantities of rhodium and other precious metals have recently been added. The amount of catalyst in one converter is from 0.4 to 208 kg.
Several processes for platinum and palladium recovery from converter beads are already known, among which the recovery methods involving precious metals dissolution are best suited for large-scale treatment of spent autornotive catalyst. In U.S. Patent No. 3,985,854, for example, precious metal values are recovered from catalysts by reacting the catalyst with a mixture of hydrochloric acid and an oxidizing agent, to leach out the precious metals in the form of their soluble chloride complexes. The leach solution must be capable of oxidizing the con-tained precious metals present in the metallic state and solubilizing the complexes thus formed. The metals are then recovered from the solution by one of several methods of precipitation.
Platinum and palladium dissolution requires a high concentration of chloride ions in acid solution and the presence of an oxidizing agent which is usually nitric acid. The reactions involved are the following:
3Pt + 4HN03 + 18HCl = 3H2PtCl6 + 4N0 + 8H20 (l) 3Pd + 2HN03 12HCl = 3H2PdC14 + 2N0 + 4H20 (2) Such conditions are assured by leaching at 70-95C
with aqua regia solution wherein the ratio of acids is HCl/HN03 = 3/l. Leaching with aqua regia at temperatures above 70C is extremely difficult because of the highly aggressive nature of both the solution and the gaseous products of its decomposi-tion. This solution instability is marked by a high partial pressure ox gaseous hydrogen chloride andsu~stantial nitric acid decomposition with evolution of nitrogen oxides. Such loss of oxidizing agent (HN03) and source of chloride ions (HCl) that are necessary to keep platinum and palladium ions in solution slows the rate of metal dissolution result-ing in an increased leaching time. This, in turn, promotes undesirable dissolution of the alumina substrate of the catalyst, with additional acids consumption. Apart from the fact that HCl regene-ration is technically difficult from environmental and corrosion view points, the presence of usually large amounts of aluminum in solution precludes the possibility of recycling the solution for additional leaching since this indicates that the amount of free HCl present in solution has been severely dimi-nished by alumina dissolution.
Summary of the Invention It is therefore an object of this invention to overcome the above drawbacks and to provide a process for recovering precious metals from materials containing same, which enables one to considerably reduce the gaseous products evolved during leaching as well as the extent of alumina dissolution, and also to substantially decrease the acid consumption.
In accordance with the present invention, there is thus provided in a process for recovering precious metals from materials containing same by leaching the precious metal with an acidic aqueous medium containing chloride ions and a strong oxidiz-ing agent to form a soluble chloride complex of the ~2~9~39 precious metal and separating the metal from the complex, the improvement wherein the chloride ions art derived at least in part from a non-acidic chloride-containing leaching agent which is non-volatile and soluble in the acidic aqueous mediumwithout crystallizing under the leaching conditions.
It has been unexpectedly found that by replacing at least part of the hydrochloric acid conventionally used for leaching precious metals by a non-acidic chloride-containing compound which is non-volatile and soluble in the leach solution without crystallizing under the leaching conditions, the partial pressure of hydrogen chloride and of other gaseous products evolved during leaching can be considerably reduced and the acid consumption significantly decreased. Since chloride is the most effective medium in which precious metals can be brought into solution, such a non-acidic chloride-containing leaching agent must have a crystallization point which is high enough to permit a sufficiently high concentration of dissolved chloride. Generally, about 10 to 100% of the hydrochloric acid can be replaced by this non-acidic chloride-containing leach-ing agent.
Description of Preferred_Embodiments Examples of suitable non-acidic chloride-containing leaching agen-ts for the practice of the invention include aluminum chloride (AlC13) and magnesium chloride (MgCl~). It is also possible to use calcium chloride (CaC12), sodium chloride (~aCl) and potassium chloride (KCl) to a certain extent as I
~228~8~3 the applicability thereof is limited due to their lower crystallization point. Aluminum ehloride is of course preferred since it provides 3 moles of chloride ions per mole of AlC13 used. Where the precious metal bearing material to be treated is an alumina-based catalyst, aluminum chloride is also preferably used since an increased concentration of aluminum chloride in the leach solution substantially reduces the rate of dissolution of the alumina substrate of the catalyst, aluminum chloride also has no leaching effect on the alumina substrate. For example, it has been observed that by inereasing the alumina chloride eoncentration in the leach solution to about 200 g/l, the acid consumption was deereased from about 9.0 to 6.7 moles of acids per kg of eatalyst pellets and the extent of alumina dissolution redueed from 20%
to 12% of the initial quantity of alumina.
Generally, the total eoneentration of ehloride ions present in the leaeh solution is at least about 2 moles/liter, preferably from about 2 to about 8 moles/liter. The preferred ehloride eoncen-tration is about 4.5 moles/liter.
Where aluminum ehloride is used as the non-acidic chloride-eontaining leaching agent, it is preferably present in the leach solution in an amount of about 0.2 to about 2.3 moles/liter. It should be noted that 2 3 moles/liter AlCl3 is the highest concentration permissible without AlCl3 undergoing crys-tallization and provides a sufficiently high concentration of chloride ions so that the ~L2289~
hydrochloric acid may be completely dispensed with.
On the other hand, where a mixture of HCl, AlC13 and oxidizing agent is used for leaching the precious metals, the molar ratio of HCl:AlC13 is preferably from ako~t 1:5 to about 3:1.
The leach solution is generally acidic to solubilize readily the chlcride complex of the precious metal formed. Generally, the concentration of hydrogen ions is at least 0.5 mole/liter.
l'he leach solution must also contain a strong oxidizing agent dissolved therein so as to oxidize the precious metal and render it soluble in the acid medium. Examples of suitable oxidizing agents include nitric acid, chlorine, chlorates (such as sodium or potassium chlorate), brornine, bromates (such as sodiurn or potassium bromate~, iodine, iodates (such as sodium or potassium iodates) and hydrogen peroxide.
Ni-tric acid is of course preferred since it also pro-vides a source of hydrogen ions necessary to maintain the acidity of the leach solution. The nitric acid is preferably used in an amount of about 0.6 to about 2.2 moles/liter.
Generally, the leaching is carried out at a temperature of about 70 to about 105C, preferably about 90-105C.
The separation of the precious metal from the chloride complex formed during leaching can be effected in known manner. Examples of suitable sepa-ration methods include electrolysis, precipitation by cementation, sorption with simultaneous reduction, ~2~8g89 reduction under an overpressure of hydrogen and reduction with hydrazine.
The process of the invention is partieu-larly applicable to the recovery of platinum and palladium from alumina-based catalysts containing same. As mentioned previously, by using aluminum ehloride, most of the alumina substrate i5 left intact.
Thus, after extraetion of the platinum and palladium, the catalyst can be further treated with a sulfurie aeid solution to leaeh out the alumina in the form of aluminum sulfate hydrate (alum). Up to 80% of the initial alumina content can therefore be recovered as a valuable alum by-produet.
A preferred process for reeovering platinum and palladium from an alumina-based eatalyst eomprises the steps of:
a) treating the eatalyst at a temperature of about 90 to about 105C with a leach solution eontaining aluminum ehloride, hydroehlorie aeid and nitric acid, in which the molar ra-tio of HCl:AlC13 is from about 1:5 to about 3:1, to leaeh out the platinum and palladium in the form of their soluble chloride eomplexes, thereby providing a leaeh liquor eontaining these complexes dissolved therein b) eleetrolytically depositing a major portion of the platinum and palladium from the leach liquor obtained in step (a) serving as electrolyte, while bleeding off a portion of spent electrolyte, c) washing the leached catalyst from step (a3 to extraet leaeh liquor entrapped therein and ~22~398~
containing dissolved chloride complexes, thereby providing a wash liquor containing same, d) neutralizing and purifying the portion of spent electrolyte obtained in step (b) and combin-ing the neutralized and purified electrolyte withthe wash liquor obtained in step (c); and e) extracting the remaining portion of the platinum and palladium from the combined electrolyte and wash liquor obtained in step (d) by sorption of the chloride complexes contained therein, with simultaneous reduction to metallic state.
Step (a) can be effected for example by feeding the leach solution to the inlet o-f a leaching reactor containing a static bed of the catalyst in crushed, bead or pellet form, and allowing the leach solution to intimately contact the catalyst. The leach liquor containing dissolved chloride complexes can then be withdrawn from the outlet of the reactor. The total time of leaching may vary from about 50 to about l90 minutes.
The electrolysis which takes place in step ~b) and involves a cathodic deposition of the pla-tinum and palladium from the leach liquor serving as electrolyte occurs in two steps:
1) palladium deposition and platinum reduction from Pt4 to pt2 :
PdCl42 + 2e = Pd + 4Cl , E = 0.59V (3) PtCl62 + 2e = PtCl42 + 2Cl , E = 0.64V (4) 2) platinum deposition:
30 PtCl42 + 2e = Pt + 4Cl , E = 0.75V (5) ~28~39 Good compact deposition from the leach liquor onto graphite or titanium cathodes takes place when the cathodic current density is less than 60 A/m2. For example, after 24 hours of electroly-sis at 30 A/m2, the platinum concentration decreasesfrom 326 mg/l to 121 mg/l, and palladium from 104 mg/l to 29 mg/l. Total deposition of platinum and sirnultaneous reduction in palladium concentration to 2.7 mg/l was observed after an additional 42 hours.
About one third of the spent electrolyte must be bled off due to accumulation of other metals in solution. The lead chloride concentra-tion can be reduced by crystallization, while metals such as iron, manganese, magnesium or aluminum are partially precipitated as hydroxides during neu-tralization in step (d). In general, only lead dissolved from used catalysts accumulates in solu-tion in the forrn of stable chloride complexes, and when recirculated solution becomes saturated in lead, crystallization oE lead chloride (PbC12) is necessary. Lead chloride may be crystallized from solutions containing 6 to 30 g/l Pb by simple cool-ing. Chloride complexes of platinum and palladium are stable in these solutions, and do not crystal-lize as chlorides if their concentrations are low -below a few grams per liter.
Two steps of washing are required for the total removal of platinum and palladium from the inte-rior of the leached pellets. During the first wash, carried out by a solution of concentrated chlorides (AlC13 or NaCl), about 90% of platinum and palladium adsorbed on the inside surface of the pellets is recovered. A second wash with water recovers the remaining platinum and palladium, the final wash water containing less than 3 mg/l platinum and palladium at a pH above 3.
The wash liquor is combined with the portion of spent electrolyte which has been neutralized and purified and the combined solutions are then treated in the extraction step (e). In step (e), the platinum and palladium are extracted from solution by sorption of their chloride complexes, with simultaneous reduc-tion to the metallic state, on AMBORAME (trade mark) resin, produced by Rohm and Ices Ltd.
AMBORANE contains the strongly reducing borane radical. The reduction reactions taking place are as follows:
4R3N- BH3 + 6PtCl6 + 12H20 = 4R3N~I (Cl ) + 6Pt + ~H3BO3 + 32Cl -I 20H+ (6) 2R3~ BH3 + 6PdC14 + 6~I20 = 2R3~H (Cl ) + 6Pd + 2H3B03 + 22Cl + 10H (7) Complete sorption of platinum from solution is more difficult than for palladium. The former requires an excess of AMBORANE relative to the platinum and palladium content of the solution, and a lengthy solution/resin contact time.
It is worth mentioning that the selectivity of reduction described by reactions (6) and (7) is less when other metal-bearing anionic complexes (e.g. PbC14 , CuC13 , FeC14 ) are present. Traces ~2~89~39 of these metals in the final product of sorption and reduction can then be expected.
Metallic platinum and palladium are produced by burning of loaned resin. The composition of final product recovered from catalyst pellet leach solution is about 72% platinum and 2~% palladium.
The yield of platinum sorption by AMBORANE
from concentrate leach solutions attains 95% and palladium 89% respectively. Nevertheless, about 17-28 mg/l platinum and 4-12 mg/l palladium remains in the spent solution after this process. Essentially complete platinum and palladium precipitation from solution after sorption is obtained by cementation with metallic aluminum powder or granules. The solution after final platinum and palladium precipi-tation contains less than 3 mg/l Pt or Pd.
l`he process of the invention enables one to recover more than 97% of the platinum and palladium from catalyst pellets. Although the invention is primarily directed to the recovery of platinum and palladium from catalytic converters, it is of course applicable to other precious metals which form soluble chloride complexes, e.g. iridium, rhodium and goldO Thus, gold may be recovered according to the process of the invention from electronic scrap material such as reject components, circuit board trim, router dust, punchings, trimmings, etc.
The following non-limiting examples illustrate the invention.
~L22E~8~3 Examples 1-8 The continuous chloride leaching of a catalyst pellets static bed in a two-liter leaching reactor was effected at a temperature of 90-104C using a leaching solution containing varying proportions of HCl, AlC13 and H~03. The composition of the pellets was ~27-441 ppm Pt, 203-211 ppm Pd, 3.1% Pb, 0.58% Mn, 0.55% Fe. The leaching conditions and results obtained are reported in Table 1 hereinafter.
Example 1 is given by way of comparison since it illustrates the conventional leaching method using only HCl and ~03.
As it is apparent from Table 1, by replacing at least part of the hydrochloric acid by aluminum chloride so that at least 10% of the chloride ions are derived from AlC13, the quan-tity of gaseous products evolved during leaching is reduced by at least 50%.
Example 8 illustrates the case where no HCl is present in the leach solution, the chloride ions thus being derived completely from AlC13.
It should also be noted in connection with ExampLes 2 and 4 to 8 that the leach solution fed to the inlet of the reactor contains some Pt and Pd as use was made of a recycled leach solution.
Examples 1 and 3 were performed using fresh leach solution.
Examples 9-15 Examples 2 to 8 were repeated, except that MgC12 was used instead of AlC13. The leaching condi-~2Z13~8~
tions and results obtained are reported in Table 2 hereinafter.
As it can be seen, the results obtained with MgC12 are essentially the same as with AlC13.
It should be noted that the concentration of 3.42 moles/liter MgC12 in Example 15 is the highest concen-tration permissible without MgC12 crystallization.
Examples 16-21 Examples 2 to 7 were repeated, with the exception that CaC12 was used instead of AlC13 The leaching conditions and results obtained are reported in Table 3 hereinafter.
As it can be seen, the results obtained with CaC12 are essentially the same as with AlC13.
lS It should be noted that the concentration ox 2.19 moles/liter CaC12 in Example 21 is the highest con-centration permissible without CaC12 crystallization Examples 22-27 Examples 2 to 7 were repeated, except that NaCl was used instead of AlC13. The leaching condi tions and results obtained are reported in Table 4 hereinafter.
As it can be seen, the results obtained with NaCl are essentially the same as with AlC13.
It should be noted that the concentration of 4.11 moles/liter NaCl in Example 27 is the highest concen-tration permissible without NaCl crystallization.
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In both cases, small quantities of rhodium and other precious metals have recently been added. The amount of catalyst in one converter is from 0.4 to 208 kg.
Several processes for platinum and palladium recovery from converter beads are already known, among which the recovery methods involving precious metals dissolution are best suited for large-scale treatment of spent autornotive catalyst. In U.S. Patent No. 3,985,854, for example, precious metal values are recovered from catalysts by reacting the catalyst with a mixture of hydrochloric acid and an oxidizing agent, to leach out the precious metals in the form of their soluble chloride complexes. The leach solution must be capable of oxidizing the con-tained precious metals present in the metallic state and solubilizing the complexes thus formed. The metals are then recovered from the solution by one of several methods of precipitation.
Platinum and palladium dissolution requires a high concentration of chloride ions in acid solution and the presence of an oxidizing agent which is usually nitric acid. The reactions involved are the following:
3Pt + 4HN03 + 18HCl = 3H2PtCl6 + 4N0 + 8H20 (l) 3Pd + 2HN03 12HCl = 3H2PdC14 + 2N0 + 4H20 (2) Such conditions are assured by leaching at 70-95C
with aqua regia solution wherein the ratio of acids is HCl/HN03 = 3/l. Leaching with aqua regia at temperatures above 70C is extremely difficult because of the highly aggressive nature of both the solution and the gaseous products of its decomposi-tion. This solution instability is marked by a high partial pressure ox gaseous hydrogen chloride andsu~stantial nitric acid decomposition with evolution of nitrogen oxides. Such loss of oxidizing agent (HN03) and source of chloride ions (HCl) that are necessary to keep platinum and palladium ions in solution slows the rate of metal dissolution result-ing in an increased leaching time. This, in turn, promotes undesirable dissolution of the alumina substrate of the catalyst, with additional acids consumption. Apart from the fact that HCl regene-ration is technically difficult from environmental and corrosion view points, the presence of usually large amounts of aluminum in solution precludes the possibility of recycling the solution for additional leaching since this indicates that the amount of free HCl present in solution has been severely dimi-nished by alumina dissolution.
Summary of the Invention It is therefore an object of this invention to overcome the above drawbacks and to provide a process for recovering precious metals from materials containing same, which enables one to considerably reduce the gaseous products evolved during leaching as well as the extent of alumina dissolution, and also to substantially decrease the acid consumption.
In accordance with the present invention, there is thus provided in a process for recovering precious metals from materials containing same by leaching the precious metal with an acidic aqueous medium containing chloride ions and a strong oxidiz-ing agent to form a soluble chloride complex of the ~2~9~39 precious metal and separating the metal from the complex, the improvement wherein the chloride ions art derived at least in part from a non-acidic chloride-containing leaching agent which is non-volatile and soluble in the acidic aqueous mediumwithout crystallizing under the leaching conditions.
It has been unexpectedly found that by replacing at least part of the hydrochloric acid conventionally used for leaching precious metals by a non-acidic chloride-containing compound which is non-volatile and soluble in the leach solution without crystallizing under the leaching conditions, the partial pressure of hydrogen chloride and of other gaseous products evolved during leaching can be considerably reduced and the acid consumption significantly decreased. Since chloride is the most effective medium in which precious metals can be brought into solution, such a non-acidic chloride-containing leaching agent must have a crystallization point which is high enough to permit a sufficiently high concentration of dissolved chloride. Generally, about 10 to 100% of the hydrochloric acid can be replaced by this non-acidic chloride-containing leach-ing agent.
Description of Preferred_Embodiments Examples of suitable non-acidic chloride-containing leaching agen-ts for the practice of the invention include aluminum chloride (AlC13) and magnesium chloride (MgCl~). It is also possible to use calcium chloride (CaC12), sodium chloride (~aCl) and potassium chloride (KCl) to a certain extent as I
~228~8~3 the applicability thereof is limited due to their lower crystallization point. Aluminum ehloride is of course preferred since it provides 3 moles of chloride ions per mole of AlC13 used. Where the precious metal bearing material to be treated is an alumina-based catalyst, aluminum chloride is also preferably used since an increased concentration of aluminum chloride in the leach solution substantially reduces the rate of dissolution of the alumina substrate of the catalyst, aluminum chloride also has no leaching effect on the alumina substrate. For example, it has been observed that by inereasing the alumina chloride eoncentration in the leach solution to about 200 g/l, the acid consumption was deereased from about 9.0 to 6.7 moles of acids per kg of eatalyst pellets and the extent of alumina dissolution redueed from 20%
to 12% of the initial quantity of alumina.
Generally, the total eoneentration of ehloride ions present in the leaeh solution is at least about 2 moles/liter, preferably from about 2 to about 8 moles/liter. The preferred ehloride eoncen-tration is about 4.5 moles/liter.
Where aluminum ehloride is used as the non-acidic chloride-eontaining leaching agent, it is preferably present in the leach solution in an amount of about 0.2 to about 2.3 moles/liter. It should be noted that 2 3 moles/liter AlCl3 is the highest concentration permissible without AlCl3 undergoing crys-tallization and provides a sufficiently high concentration of chloride ions so that the ~L2289~
hydrochloric acid may be completely dispensed with.
On the other hand, where a mixture of HCl, AlC13 and oxidizing agent is used for leaching the precious metals, the molar ratio of HCl:AlC13 is preferably from ako~t 1:5 to about 3:1.
The leach solution is generally acidic to solubilize readily the chlcride complex of the precious metal formed. Generally, the concentration of hydrogen ions is at least 0.5 mole/liter.
l'he leach solution must also contain a strong oxidizing agent dissolved therein so as to oxidize the precious metal and render it soluble in the acid medium. Examples of suitable oxidizing agents include nitric acid, chlorine, chlorates (such as sodium or potassium chlorate), brornine, bromates (such as sodiurn or potassium bromate~, iodine, iodates (such as sodium or potassium iodates) and hydrogen peroxide.
Ni-tric acid is of course preferred since it also pro-vides a source of hydrogen ions necessary to maintain the acidity of the leach solution. The nitric acid is preferably used in an amount of about 0.6 to about 2.2 moles/liter.
Generally, the leaching is carried out at a temperature of about 70 to about 105C, preferably about 90-105C.
The separation of the precious metal from the chloride complex formed during leaching can be effected in known manner. Examples of suitable sepa-ration methods include electrolysis, precipitation by cementation, sorption with simultaneous reduction, ~2~8g89 reduction under an overpressure of hydrogen and reduction with hydrazine.
The process of the invention is partieu-larly applicable to the recovery of platinum and palladium from alumina-based catalysts containing same. As mentioned previously, by using aluminum ehloride, most of the alumina substrate i5 left intact.
Thus, after extraetion of the platinum and palladium, the catalyst can be further treated with a sulfurie aeid solution to leaeh out the alumina in the form of aluminum sulfate hydrate (alum). Up to 80% of the initial alumina content can therefore be recovered as a valuable alum by-produet.
A preferred process for reeovering platinum and palladium from an alumina-based eatalyst eomprises the steps of:
a) treating the eatalyst at a temperature of about 90 to about 105C with a leach solution eontaining aluminum ehloride, hydroehlorie aeid and nitric acid, in which the molar ra-tio of HCl:AlC13 is from about 1:5 to about 3:1, to leaeh out the platinum and palladium in the form of their soluble chloride eomplexes, thereby providing a leaeh liquor eontaining these complexes dissolved therein b) eleetrolytically depositing a major portion of the platinum and palladium from the leach liquor obtained in step (a) serving as electrolyte, while bleeding off a portion of spent electrolyte, c) washing the leached catalyst from step (a3 to extraet leaeh liquor entrapped therein and ~22~398~
containing dissolved chloride complexes, thereby providing a wash liquor containing same, d) neutralizing and purifying the portion of spent electrolyte obtained in step (b) and combin-ing the neutralized and purified electrolyte withthe wash liquor obtained in step (c); and e) extracting the remaining portion of the platinum and palladium from the combined electrolyte and wash liquor obtained in step (d) by sorption of the chloride complexes contained therein, with simultaneous reduction to metallic state.
Step (a) can be effected for example by feeding the leach solution to the inlet o-f a leaching reactor containing a static bed of the catalyst in crushed, bead or pellet form, and allowing the leach solution to intimately contact the catalyst. The leach liquor containing dissolved chloride complexes can then be withdrawn from the outlet of the reactor. The total time of leaching may vary from about 50 to about l90 minutes.
The electrolysis which takes place in step ~b) and involves a cathodic deposition of the pla-tinum and palladium from the leach liquor serving as electrolyte occurs in two steps:
1) palladium deposition and platinum reduction from Pt4 to pt2 :
PdCl42 + 2e = Pd + 4Cl , E = 0.59V (3) PtCl62 + 2e = PtCl42 + 2Cl , E = 0.64V (4) 2) platinum deposition:
30 PtCl42 + 2e = Pt + 4Cl , E = 0.75V (5) ~28~39 Good compact deposition from the leach liquor onto graphite or titanium cathodes takes place when the cathodic current density is less than 60 A/m2. For example, after 24 hours of electroly-sis at 30 A/m2, the platinum concentration decreasesfrom 326 mg/l to 121 mg/l, and palladium from 104 mg/l to 29 mg/l. Total deposition of platinum and sirnultaneous reduction in palladium concentration to 2.7 mg/l was observed after an additional 42 hours.
About one third of the spent electrolyte must be bled off due to accumulation of other metals in solution. The lead chloride concentra-tion can be reduced by crystallization, while metals such as iron, manganese, magnesium or aluminum are partially precipitated as hydroxides during neu-tralization in step (d). In general, only lead dissolved from used catalysts accumulates in solu-tion in the forrn of stable chloride complexes, and when recirculated solution becomes saturated in lead, crystallization oE lead chloride (PbC12) is necessary. Lead chloride may be crystallized from solutions containing 6 to 30 g/l Pb by simple cool-ing. Chloride complexes of platinum and palladium are stable in these solutions, and do not crystal-lize as chlorides if their concentrations are low -below a few grams per liter.
Two steps of washing are required for the total removal of platinum and palladium from the inte-rior of the leached pellets. During the first wash, carried out by a solution of concentrated chlorides (AlC13 or NaCl), about 90% of platinum and palladium adsorbed on the inside surface of the pellets is recovered. A second wash with water recovers the remaining platinum and palladium, the final wash water containing less than 3 mg/l platinum and palladium at a pH above 3.
The wash liquor is combined with the portion of spent electrolyte which has been neutralized and purified and the combined solutions are then treated in the extraction step (e). In step (e), the platinum and palladium are extracted from solution by sorption of their chloride complexes, with simultaneous reduc-tion to the metallic state, on AMBORAME (trade mark) resin, produced by Rohm and Ices Ltd.
AMBORANE contains the strongly reducing borane radical. The reduction reactions taking place are as follows:
4R3N- BH3 + 6PtCl6 + 12H20 = 4R3N~I (Cl ) + 6Pt + ~H3BO3 + 32Cl -I 20H+ (6) 2R3~ BH3 + 6PdC14 + 6~I20 = 2R3~H (Cl ) + 6Pd + 2H3B03 + 22Cl + 10H (7) Complete sorption of platinum from solution is more difficult than for palladium. The former requires an excess of AMBORANE relative to the platinum and palladium content of the solution, and a lengthy solution/resin contact time.
It is worth mentioning that the selectivity of reduction described by reactions (6) and (7) is less when other metal-bearing anionic complexes (e.g. PbC14 , CuC13 , FeC14 ) are present. Traces ~2~89~39 of these metals in the final product of sorption and reduction can then be expected.
Metallic platinum and palladium are produced by burning of loaned resin. The composition of final product recovered from catalyst pellet leach solution is about 72% platinum and 2~% palladium.
The yield of platinum sorption by AMBORANE
from concentrate leach solutions attains 95% and palladium 89% respectively. Nevertheless, about 17-28 mg/l platinum and 4-12 mg/l palladium remains in the spent solution after this process. Essentially complete platinum and palladium precipitation from solution after sorption is obtained by cementation with metallic aluminum powder or granules. The solution after final platinum and palladium precipi-tation contains less than 3 mg/l Pt or Pd.
l`he process of the invention enables one to recover more than 97% of the platinum and palladium from catalyst pellets. Although the invention is primarily directed to the recovery of platinum and palladium from catalytic converters, it is of course applicable to other precious metals which form soluble chloride complexes, e.g. iridium, rhodium and goldO Thus, gold may be recovered according to the process of the invention from electronic scrap material such as reject components, circuit board trim, router dust, punchings, trimmings, etc.
The following non-limiting examples illustrate the invention.
~L22E~8~3 Examples 1-8 The continuous chloride leaching of a catalyst pellets static bed in a two-liter leaching reactor was effected at a temperature of 90-104C using a leaching solution containing varying proportions of HCl, AlC13 and H~03. The composition of the pellets was ~27-441 ppm Pt, 203-211 ppm Pd, 3.1% Pb, 0.58% Mn, 0.55% Fe. The leaching conditions and results obtained are reported in Table 1 hereinafter.
Example 1 is given by way of comparison since it illustrates the conventional leaching method using only HCl and ~03.
As it is apparent from Table 1, by replacing at least part of the hydrochloric acid by aluminum chloride so that at least 10% of the chloride ions are derived from AlC13, the quan-tity of gaseous products evolved during leaching is reduced by at least 50%.
Example 8 illustrates the case where no HCl is present in the leach solution, the chloride ions thus being derived completely from AlC13.
It should also be noted in connection with ExampLes 2 and 4 to 8 that the leach solution fed to the inlet of the reactor contains some Pt and Pd as use was made of a recycled leach solution.
Examples 1 and 3 were performed using fresh leach solution.
Examples 9-15 Examples 2 to 8 were repeated, except that MgC12 was used instead of AlC13. The leaching condi-~2Z13~8~
tions and results obtained are reported in Table 2 hereinafter.
As it can be seen, the results obtained with MgC12 are essentially the same as with AlC13.
It should be noted that the concentration of 3.42 moles/liter MgC12 in Example 15 is the highest concen-tration permissible without MgC12 crystallization.
Examples 16-21 Examples 2 to 7 were repeated, with the exception that CaC12 was used instead of AlC13 The leaching conditions and results obtained are reported in Table 3 hereinafter.
As it can be seen, the results obtained with CaC12 are essentially the same as with AlC13.
lS It should be noted that the concentration ox 2.19 moles/liter CaC12 in Example 21 is the highest con-centration permissible without CaC12 crystallization Examples 22-27 Examples 2 to 7 were repeated, except that NaCl was used instead of AlC13. The leaching condi tions and results obtained are reported in Table 4 hereinafter.
As it can be seen, the results obtained with NaCl are essentially the same as with AlC13.
It should be noted that the concentration of 4.11 moles/liter NaCl in Example 27 is the highest concen-tration permissible without NaCl crystallization.
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Claims (17)
1. A process for recovering precious metals from materials containing same, comprising the steps of treat-ing a precious metal-containing material with an acidic aqueous medium comprising chloride ions and an oxidizing agent capable of oxidizing the precious metal in the presence of said chloride ions to leach out said precious metal in the form of a soluble chloride complex and thereafter separating said metal from said complex, wherein said chloride ions are derived at least in part from a non-acidic, chloride-containing leaching agent which is non-volatile and soluble in said acidic aqueous medium without crystallizing under the leaching conditions, said non-acidic, chloride-containing leaching agent being selected from the group consisting of aluminum chloride, sodium chloride, potassium chloride, magnesium chloride and calcium chloride.
2. A process as claimed in claim 1, wherein the total concentration of said chloride ions initally present in said acidic aqueous medium is at least about 2 moles/liter.
3. A process as claimed in claim 2, wherein said chloride ion concentration is from about 2 to about 8 moles/liter.
4. A process as claimed in claim 3, wherein said chloride ion concentration is about 4.5 moles/
liter.
liter.
5. A process as claimed in claim 3, wherein from about 10 to 100% of said chloride ions are derived from said non-acidic chloride-containing leaching agent.
6. A process as claimed in claim 1, wherein said non-acidic chloride-containing leaching agent is selected from the group consisting of aluminum chloride, magnesium chloride and calcium chloride:
7. A process as claimed in claim 3, wherein said non-acidic chloride-containing leaching agent is aluminum chloride and is used in an amount of about 0.2 to about 2.3 moles/liter.
8. A process as claimed in claim 3, wherein said non-acidic chloride-containing leaching agent is sodium chloride and is used in an amount of about 0.7 to about 4.1 moles/liter.
9. A process as claimed in claim 3, wherein said non-acidic chloride-containing leaching agent is magnesium chloride and is used in an amount of about 0.3 to about 3.4 moles/liter.
10. A process as claimed in claim 3, wherein said non-acidic chloride-containing leaching agent is calcium chloride and is used in an amount of about 0.3 to about 2.2 moles/liter.
11. A process as claimed in claim 3, where-ing said acidic aqueous medium contains hydrochloric acid as additional source of said chloride ions and said non-acidic chloride-containing leaching agent is aluminum chloride and wherein the molar ratio of HCl:AlCl3 is from about 1:5 to about 3:1.
12. A process as claimed in claim 1, where-in said oxidizing agent is selected from the group consisting of nitric acid, chlorine, chlorates, bromine, bromates, iodine, iodates and hydrogen peroxide.
13. A process as claimed in claim 12, wherein said oxidizing agent is nitric acid and is used in an amount of about 0.6 to about 2.2 moles/
liter.
liter.
14. A process as claimed in claim 1, where-in said leaching is carried out at a temperature of about 70 to about 105°C.
15. A process for the recovery of at least one precious metal selected from the group consisting of platinum and palladium from a catalyst containing same, comprising the steps of treating the catalyst at a temperature of about 90 to about 105°C with an acidic aqueous medium comprising chloride ions and an oxidizing agent capable of oxidizing the precious metal in the presence of said chloride ions to leach out said precious metal in the form of a soluble chloride complex and thereafter separating said metal from said complex, wherein said chloride ions are derived at least in part from aluminum chloride.
16. A process as claimed in claim 15, wherein the total concentration of said chloride ions initially present in said acidic aqueous medium is from about 2 to about 8 moles/liter.
17. A process as claimed in claim 16, wherein said aluminum chloride is used in an amount of about 0.2 to about 2.3 moles/liter.
Priority Applications (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA000464865A CA1228989A (en) | 1984-10-05 | 1984-10-05 | Recovery of precious metals from materials containing same |
Applications Claiming Priority (1)
| Application Number | Priority Date | Filing Date | Title |
|---|---|---|---|
| CA000464865A CA1228989A (en) | 1984-10-05 | 1984-10-05 | Recovery of precious metals from materials containing same |
Publications (1)
| Publication Number | Publication Date |
|---|---|
| CA1228989A true CA1228989A (en) | 1987-11-10 |
Family
ID=4128856
Family Applications (1)
| Application Number | Title | Priority Date | Filing Date |
|---|---|---|---|
| CA000464865A Expired CA1228989A (en) | 1984-10-05 | 1984-10-05 | Recovery of precious metals from materials containing same |
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Cited By (9)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US5342449A (en) * | 1990-12-11 | 1994-08-30 | Holbein Bruce E | Process for the decontamination of toxic, heavy-metal containing soils |
| WO2002053788A1 (en) * | 2000-12-29 | 2002-07-11 | Nichromet Extraction Inc. | Method for the recovery of base and precious metals by extractive chloridation |
| WO2007081243A3 (en) * | 2006-01-10 | 2007-10-25 | Zakrytoe Akcionernoe Obshestvo | Method for extracting precious metals |
| WO2010036144A1 (en) * | 2008-09-29 | 2010-04-01 | Закрытое Акционерное Общество "Уралкалий-Технология" | Method for processing mineral raw material |
| WO2010036142A1 (en) * | 2008-09-29 | 2010-04-01 | Закрытое Акционерное Общество "Уралкалий-Технология" | Method for the recovery of noble metals |
| WO2011156861A1 (en) * | 2010-06-15 | 2011-12-22 | The University Of Queensland | Method of recovering a metal |
| RU2497961C1 (en) * | 2012-10-02 | 2013-11-10 | Федеральное государственное бюджетное учреждение науки Горный институт Уральского отделения Российской академии наук (ГИ УрО РАН) | Processing method of potassium production wastes |
| RU2530923C1 (en) * | 2013-05-13 | 2014-10-20 | Федеральное государственное бюджетное учреждение науки Горный институт Уральского отделения Российской академии наук (ГИ УрО РАН) | Method of obtaining collective concentrate |
| CN116622998A (en) * | 2023-05-31 | 2023-08-22 | 昆明理工大学 | Method for dissolving noble metal and/or noble metal alloy |
-
1984
- 1984-10-05 CA CA000464865A patent/CA1228989A/en not_active Expired
Cited By (11)
| Publication number | Priority date | Publication date | Assignee | Title |
|---|---|---|---|---|
| US5342449A (en) * | 1990-12-11 | 1994-08-30 | Holbein Bruce E | Process for the decontamination of toxic, heavy-metal containing soils |
| WO2002053788A1 (en) * | 2000-12-29 | 2002-07-11 | Nichromet Extraction Inc. | Method for the recovery of base and precious metals by extractive chloridation |
| WO2007081243A3 (en) * | 2006-01-10 | 2007-10-25 | Zakrytoe Akcionernoe Obshestvo | Method for extracting precious metals |
| US7846234B2 (en) | 2006-01-10 | 2010-12-07 | Viktor Andreevich Sinegribov | Method of precious metal recovery |
| WO2010036144A1 (en) * | 2008-09-29 | 2010-04-01 | Закрытое Акционерное Общество "Уралкалий-Технология" | Method for processing mineral raw material |
| WO2010036142A1 (en) * | 2008-09-29 | 2010-04-01 | Закрытое Акционерное Общество "Уралкалий-Технология" | Method for the recovery of noble metals |
| CN102159738A (en) * | 2008-09-29 | 2011-08-17 | 乌拉尔卡里技术联合股份有限公司 | Method for the recovery of noble metals |
| WO2011156861A1 (en) * | 2010-06-15 | 2011-12-22 | The University Of Queensland | Method of recovering a metal |
| RU2497961C1 (en) * | 2012-10-02 | 2013-11-10 | Федеральное государственное бюджетное учреждение науки Горный институт Уральского отделения Российской академии наук (ГИ УрО РАН) | Processing method of potassium production wastes |
| RU2530923C1 (en) * | 2013-05-13 | 2014-10-20 | Федеральное государственное бюджетное учреждение науки Горный институт Уральского отделения Российской академии наук (ГИ УрО РАН) | Method of obtaining collective concentrate |
| CN116622998A (en) * | 2023-05-31 | 2023-08-22 | 昆明理工大学 | Method for dissolving noble metal and/or noble metal alloy |
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