A method for synergistically enriching platinum and palladium from silver anode slime nitric acid leaching solution and preparing sponge palladium by green short process

Abstract

The present application belongs to the technical field of metal smelting, and relates to a method for cooperatively enriching platinum and palladium from silver anode slime nitric acid leaching solution and preparing sponge palladium by a green short process, which comprises the following steps: adding diacetyldioxime-thiourea alkaline solution to the silver anode slime nitric acid leaching solution to perform co-precipitation of platinum and palladium, and obtaining co-precipitate; after slurry, adjusting to pH 9.0-10.0 to perform primary alkaline conversion; then adding hydrochloric acid to perform secondary conversion, and after the end, adding saturated chlorate solution to perform reaction, and obtaining platinum and palladium-containing filtrate; then adding oxidizing agent and ammonium chloride to perform reaction to obtain platinum and palladium co-precipitate; after slurry, adding reducing agent to perform reaction to obtain palladium-containing solution; then adding oxidizing agent and ammonium chloride to perform reaction to obtain high-purity palladium-containing residue; mixing the residue with ammonia water to perform complexation, adding hydrazine hydrate after reaction, and obtaining sponge palladium through reduction. The method has mild reaction conditions, realizes multi-component recovery of platinum, palladium and silver in the silver anode slime, and has high resource utilization rate.

Description

Technical Field

[0001] This invention belongs to the field of metal smelting technology, specifically relating to a method for the synergistic enrichment of platinum and palladium and the preparation of sponge palladium from nitric acid leachate of silver anode mud using a green and short process. Background Technology

[0002] Currently, common processes for recovering platinum and palladium from nitric acid leachate of silver anode mud mainly include solvent extraction, ion exchange, and selective precipitation. However, these methods all have significant limitations in practical applications: solvent extraction, while having high separation efficiency, relies on large amounts of organic reagents, resulting in high costs and potential secondary pollution; ion exchange is constrained by issues such as resin aging and regeneration difficulties, leading to poor operational stability; selective precipitation, such as using diacetyl dioxime to precipitate palladium, while exhibiting good selectivity for palladium, results in low platinum recovery, and the obtained precipitate requires high-temperature calcination, leading to a long process, high energy consumption, and a heavy environmental burden. Furthermore, while the traditional ammonium chloride co-precipitation method can achieve initial enrichment of platinum and palladium, subsequent platinum and palladium separation is difficult, often requiring multiple precipitation and dissolution cycles, resulting in a complex process and low direct metal recovery rate.

[0003] The aforementioned methods generally suffer from problems such as complex processes, numerous operational steps, high energy consumption, and unsatisfactory recovery rates. In particular, the organic co-precipitates generated by some processes require high-temperature calcination for further processing, further increasing production costs and time, and hindering the improvement of production efficiency. Therefore, developing a new process that can achieve efficient synergistic enrichment of platinum and palladium, and on this basis, directly prepare high-purity sponge palladium through a green and short process, is of significant industrial importance for improving the comprehensive recovery level of platinum group metals from silver anode slime. Summary of the Invention

[0004] To address the shortcomings of the existing technology, this invention provides a method for the synergistic enrichment of platinum and palladium from nitric acid leachate of silver anode mud and the preparation of sponge palladium using a green and short process.

[0005] The specific plan is as follows:

[0006] A method for the synergistic enrichment of platinum and palladium and the preparation of sponge palladium from nitric acid leachate of silver anode mud includes the following steps:

[0007] S1 Add diacetyl dioxime-thiourea alkaline solution to the nitric acid leachate of silver anode mud to perform platinum-palladium coprecipitation, and obtain coprecipitate and silver-containing filtrate;

[0008] S2 After the coprecipitate obtained in step S1 is pulped, the pH is adjusted to 9.0~10.0 for a first alkaline conversion; then hydrochloric acid is added for a second conversion, and after the conversion is completed, a saturated chlorate solution is added to react and obtain a platinum-palladium filtrate and filter residue.

[0009] S3 Add an oxidant and ammonium chloride to the platinum-palladium filtrate obtained in step S2, and after the reaction, obtain platinum-palladium coprecipitate and filtrate;

[0010] S4 After the platinum-palladium coprecipitate obtained in step S3 is slurried, a reducing agent is added and reacted to obtain platinum-rich material and palladium-containing solution;

[0011] S5. Add oxidant and ammonium chloride to the palladium-containing solution obtained in step S4, and after the reaction, obtain high-purity palladium-containing slag.

[0012] S6 The high-purity palladium-containing slag obtained in step S5 is mixed with ammonia water for complexation. After the reaction, hydrazine hydrate is added and reduced to obtain sponge palladium.

[0013] The nitric acid leachate from silver anode mud is the filtrate produced from the nitric acid impurity removal process of silver anode mud. Its main components include: Cu content of 4.0~6.0 g / L, Au content of 0.2~0.5 mg / L, Ag content of 20~50 g / L, Pt content of 80~120 mg / L, and Pd content of 600~1200 mg / L.

[0014] In step S1, a self-prepared diacetyldioxime-thiourea-sodium hydroxide composite precipitant is used to achieve efficient synergistic co-precipitation of platinum and palladium in the nitric acid leachate, with a platinum enrichment rate >95% and a palladium enrichment rate >99%. The main reaction is as follows:

[0015] Pd 2+ +2C4H8N2O2→Pd(C4H7N2O2)2↓+2H + ;

[0016] Pd(NO3)2+2SC(NH2)2=[Pd(SC(NH2)2)2](NO3)2↓;

[0017] Pt(NO3)2+2SC(NH2)2=[Pt(SC(NH2)2)2](NO3)2↓.

[0018] Further, in step S1, the mass ratio of diacetyldioxime, sodium hydroxide, thiourea, and deionized water in the diacetyldioxime-thiourea alkaline solution is 1.0:1.0:(0.1~0.2):5.0. The preparation method is as follows: diacetyldioxime, sodium hydroxide, and thiourea are dissolved in deionized water at 30~40℃ to obtain the diacetyldioxime-thiourea alkaline solution.

[0019] Furthermore, in step S1, the molar ratio of the total amount of Pd and Pt in the nitric acid leaching solution of silver anode mud to the total amount of diacetyl dioxime and thiourea in the diacetyl dioxime-thiourea alkaline solution is 1.0:(2.0~3.0).

[0020] Furthermore, in step S1, the platinum-palladium co-precipitation temperature is 80~85℃ and the time is 1.0~2.0h.

[0021] Preferably, in step S1, the temperature of the nitric acid leaching solution of the silver anode mud is 60~70℃.

[0022] Preferably, the silver-containing filtrate obtained in step S1 is reacted with a chloride to obtain silver chloride, which is then returned to the converter pyrometallurgical process to recover silver; wherein, the chloride is preferably sodium chloride.

[0023] In step S1, the coprecipitate contains 1.0wt% to 3.0wt% platinum and 15wt% to 29wt% palladium; the silver-containing filtrate contains 1.0 to 5.0 mg / L platinum and ≤10 mg / L palladium.

[0024] In step S2, the co-precipitate after pulping undergoes a primary alkaline conversion. By precisely controlling pH and temperature, the thiourea in the precipitate is decomposed, breaking down the organic matter and facilitating subsequent chlorination leaching of the platinum-palladium precipitate. Since a small amount of nitrate remains in the co-precipitate, hydrochloric acid is added for a secondary hydrochloric acid conversion to prevent accidental reactions caused by the coexistence of nitrate and chlorate ions. After the secondary hydrochloric acid conversion, a saturated chlorate solution is added. The chlorine / hypochlorous acid generated by the saturated chlorate in the acidic environment provided by hydrochloric acid acts as a strong oxidant, soluble sulfides (PdS and...) are then leached. The sulfur in PtS is oxidized to sulfate, which oxidizes palladium and platinum to form soluble chloropalladic acid and chloroplatinic acid, respectively, which then enter the filtrate. The saturated chlorate solution ensures a more complete and thorough reaction, breaking down the sulfur oxides and allowing the platinum and palladium to leach out fully. However, when solid chlorate is added, it reacts immediately upon contact with the solution surface, generating chlorine gas that escapes, preventing direct reaction with the coprecipitate and resulting in reagent waste. Furthermore, adding the saturated chlorate solution after a secondary hydrochloric acid conversion ensures a stable reaction and avoids the risk of overheating caused by simultaneous addition of both. The main reactions are:

[0025] Pd(C4H7N2O2)2↓+2OH - →Pd(OH)2↓+2C4H7N2O2 - ;

[0026] Pd(OH)2 + Na2S → PdS + 2NaOH;

[0027] [Pd(SC(NH2)2)2](NO3)2+4NaOH=2NaNO3+PdS+2CO2↑+4NH3↑+Na2S;

[0028] [Pt(SC(NH2)2)2](NO3)2+4NaOH=2NaNO3+PtS+2CO2↑+4NH3↑+Na2S;

[0029] 3PdS+5NaClO3+18HCl=3H2PdCl6+3H2SO4+5NaCl+3H2O;

[0030] 3PtS+5NaClO3+18HCl=3H2PtCl6+3H2SO4+5NaCl+3H2O;

[0031] 3PdS+5KClO3+18HCl=3H2PdCl6+3H2SO4+5KCl+3H2O;

[0032] 3PtS+5KClO3+18HCl=3H2PtCl6+3H2SO4+5KCl+3H2O.

[0033] Preferably, in step S2, sodium hydroxide is added to perform an alkaline conversion.

[0034] Preferably, in step S2, the temperature of the first alkaline conversion is 80~85℃, and the time is 0.5~1.0h; the mass ratio of slurry liquid to solid is (4.0~5.0):1.0.

[0035] Preferably, in step S2, the concentration of hydrochloric acid is 2.5~3.0 mol / L.

[0036] Furthermore, in step S2, the molar ratio of the total amount of Pd and Pt in the coprecipitate obtained in step S1 to hydrochloric acid is 1.0:(10~15); the secondary conversion temperature is 80~85℃, and the time is 0.5~0.6h.

[0037] Furthermore, in step S2, the molar ratio of the total amount of Pd and Pt to chlorate in the coprecipitate obtained in step S1 is 1.0:(4.0~5.0); the reaction temperature is 80~85℃, and the reaction time is 1.5~2.0h.

[0038] Preferably, in step S2, the chlorate is sodium chlorate and / or potassium chlorate.

[0039] Preferably, in step S2, the filter residue is returned to the converter pyrometallurgical process to recover silver.

[0040] In step S3, an oxidant and ammonium chloride are added to the platinum-palladium filtrate. The ammonium chloride provides a large amount of NH4. + The oxidant combines with the aforementioned chloropalladic acid and chloroplatinic acid to form ammonium salt precipitates, achieving palladium and platinum co-precipitation, thereby separating palladium and platinum from other impurities in the platinum-palladium-containing filtrate. The oxidant ensures the formation of an oxidizing environment, preventing the decomposition of chloropalladic acid and chloroplatinic acid into chloropalladic acid and chloroplatinic acid at high temperatures. Chloropalladic acid and chloroplatinic acid do not form precipitates with ammonium chloride. The main reaction is as follows:

[0041] H2PdCl6+2NH4Cl=(NH4)2PdCl6↓+2HCl;

[0042] H2PtCl6+2NH4Cl=(NH4)2PtCl6↓+2HCl.

[0043] Furthermore, in step S3, the reaction temperature is 75~80℃ and the time is 0.5~1.0h.

[0044] Furthermore, in step S3, the molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to the oxidant is 1.0:(0.2~0.5); the molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to ammonium chloride is 1.0:(2.4~4.0).

[0045] Preferably, in step S3, the oxidant is sodium chlorate.

[0046] Preferably, in step S3, the filtrate is reacted with iron powder to recover the precious metal.

[0047] In step S4, a low-temperature selective reduction method is proposed for the first time. The reaction temperature is precisely controlled at 20-30℃. A reducing agent is added to the platinum-palladium co-precipitate to selectively dissolve palladium, allowing ammonium chloropalladate to be preferentially reduced and leached. Due to temperature limitations, ammonium chloroplatinate is not easily reduced and remains in the insoluble enrichment residue, resulting in platinum-rich material. This completes the initial separation of platinum and palladium, avoiding platinum co-reduction loss. The resulting platinum-rich precipitate can be used as a raw material for further platinum purification. The main reactions are:

[0048] (NH4)2PdCl6 + reducing agent → (NH4)2PdCl4.

[0049] Furthermore, in step S4, the reaction temperature is 20~30℃ and the time is 0.5~1.0h.

[0050] Furthermore, in step S4, the molar ratio of Pd to reducing agent in the platinum-palladium coprecipitate is 1.0:(1.05~1.2).

[0051] Preferably, in step S4, the liquid-to-solid mass ratio of the slurry is (3.0~4.0):1.0; the reducing agent is at least one of sodium sulfite, sodium bisulfite, potassium bisulfite, potassium sulfite, or potassium oxalate.

[0052] In step S5, an oxidant and ammonium chloride are added to the palladium-containing solution (containing HCl) for reprecipitation-reduction purification, yielding high-purity palladium-containing slag, thus completely separating platinum and palladium. The main reactions are as follows:

[0053] (NH4)2PdCl4+NaClO3+2HCl+NH4Cl=(NH4)2PdCl6↓+NaCl+H2O;

[0054] (NH4)2PdCl4+H2O2+2HCl+2NH4Cl=(NH4)2PdCl6↓+2H2O.

[0055] Furthermore, in step S5, the reaction temperature is 75~80℃ and the time is 0.5~1.0h.

[0056] Preferably, in step S5, the oxidant is sodium chlorate and / or hydrogen peroxide.

[0057] Furthermore, in step S5, the molar ratio of Pd to oxidant in the palladium-containing solution is 1.0:(0.2~0.5); the molar ratio of Pd to ammonium chloride in the palladium-containing solution is 1.0:(2.4~4.0).

[0058] Preferably, step S5 further includes the purification of high-purity palladium-containing slag, namely, slurrying the high-purity palladium-containing slag, adding a reducing agent to carry out a reduction reaction to obtain a solution; subsequently, an oxidizing agent and ammonium chloride are added to the solution to carry out a precipitation reaction. The above process can be repeated multiple times, preferably 2-3 times, and the purified high-purity palladium-containing slag is then subjected to step S6 to obtain sponge palladium with a purity of SM-Pd99.95. The main reaction is as follows:

[0059] (NH4)2PdCl6 + reducing agent → (NH4)2PdCl4;

[0060] (NH4)2PdCl4+NaClO3+2HCl+NH4Cl=(NH4)2PdCl6↓+NaCl+H2O;

[0061] (NH4)2PdCl4+H2O2+2HCl+2NH4Cl=(NH4)2PdCl6↓+2H2O.

[0062] Preferably, the reduction reaction temperature is 20~30℃ and the time is 0.5~1.0h; the molar ratio of Pd to reducing agent in the high-purity palladium-containing slag is 1.0:(1.05~1.2); the liquid-solid mass ratio of the slurry is (3.0~4.0):1.0; and the reducing agent is at least one of sodium sulfite, sodium bisulfite, potassium bisulfite, potassium sulfite, or potassium oxalate.

[0063] Preferably, the precipitation reaction temperature is 75~80℃ and the time is 0.5~1.0h; the oxidant is sodium chlorate and / or hydrogen peroxide; the molar ratio of Pd to oxidant in the solution is 1.0:(0.2~0.5); the molar ratio of Pd to ammonium chloride in the solution is 1.0:(2.4~4.0).

[0064] In step S6, high-purity palladium-containing slag is reduced with hydrazine hydrate via complexation with ammonia water to obtain high-purity sponge palladium. The main reaction is as follows:

[0065] (NH4)2PdCl6+4NH3·H2O=[Pd(NH3)4]Cl2+2NH4Cl+Cl2↑+2H2O;

[0066] 2[Pd(NH3)4]Cl2+N2H4·H2O=2Pd↓+4NH4Cl+4NH3↑+N2↑+H2O.

[0067] Further, in step S6, the high-purity palladium-containing slag is mixed with ammonia water and the pH is adjusted to 8.0~9.0; wherein, the molar ratio of Pd to ammonia water in the high-purity palladium-containing slag is 1.0:(5.0~8.0); the molar ratio of Pd to hydrazine hydrate in the high-purity palladium-containing slag is 1.0:(0.6~1.0).

[0068] Furthermore, in step S6, the complexation temperature is 50~60℃ and the time is 0.5~1.0h; the reduction temperature is 60~70℃ and the time is 0.5~1.0h.

[0069] Preferably, in step S6, the ammonia concentration is 25wt%~28wt%.

[0070] Compared with existing technologies, the beneficial effects are as follows:

[0071] This invention overturns the traditional technical route for platinum-palladium separation, which requires lengthy processes such as high-temperature ashing, nitric acid dissolution, and hydrochloric acid denitration. It innovatively constructs a green, short-process technology of "co-precipitation enrichment - alkaline conversion - chlorination leaching - selective reduction". The reaction conditions are mild, the working environment is friendly, and there is no need for subsequent high-temperature calcination. The entire process is wet operation, which greatly shortens the process flow, reduces energy consumption, and realizes the comprehensive recovery of multiple components of platinum, palladium, and silver from silver anode mud. It has a high resource utilization rate and has good prospects for industrial application. Detailed Implementation

[0072] The embodiments of the present invention will be described in further detail below. The following examples are for illustrative purposes only and should not be construed as limiting the scope of the invention. Unless otherwise specified, the experimental methods used in the following examples are conventional methods. Unless otherwise specified, the materials and reagents used in the following examples are commercially available.

[0073] Example 1

[0074] A method for the co-enrichment of platinum and palladium and the preparation of sponge palladium from nitric acid leachate of silver anode mud, wherein the nitric acid leachate of silver anode mud is the filtrate produced from the nitric acid impurity removal process of silver anode mud, and the main components are: Cu 4.52 g / L, Au 0.31 mg / L, Ag 31.25 g / L, Pt 89.45 mg / L, and Pd 738.7 mg / L.

[0075] Includes the following steps:

[0076] S1. Diacetyl dioxime, sodium hydroxide, thiourea, and deionized water were prepared in a mass ratio of 1.0:1.0:0.1:5.0 and mixed thoroughly at 40°C to obtain a diacetyl dioxime-thiourea alkaline solution. The diacetyl dioxime-thiourea alkaline solution was then slowly added to a nitric acid leaching solution of silver anode mud at 60°C. The temperature was raised to 80°C for platinum-palladium co-precipitation. After reacting for 2.0 h, the mixture was filtered while hot to obtain the coprecipitate and a silver-containing filtrate. The silver-containing filtrate was then... Sodium chloride is added to the solution, and the reaction yields silver chloride, which is then returned to the converter pyrometallurgical process to recover silver. The molar ratio of the total Pd and Pt in the nitric acid leaching solution of the silver anode mud to the total diacetyldioxime and thiourea in the diacetyldioxime-thiourea alkaline solution is 1.0:2.0. The coprecipitate contains 1.25 wt% platinum and 22.4 wt% palladium. The silver-containing filtrate contains 3.3 mg / L platinum and 5.5 mg / L palladium.

[0077] S2. The coprecipitate obtained in step S1 is slurried at a liquid-to-solid mass ratio of 4.0:1.0. Sodium hydroxide is added to adjust the pH to 9.0, and the temperature is raised to 85℃. The reaction is carried out for 0.5 h for a first alkaline conversion. Then, 3.0 mol / L hydrochloric acid is slowly added, and a second conversion is carried out at 80℃ for 0.5 h. After the reaction is completed, a saturated sodium chlorate solution is added and the reaction is carried out at 80℃ for 2.0 h. The solid and liquid are separated while hot to obtain a platinum-palladium filtrate and a filter residue. The filter residue is returned to the converter pyrometallurgical process to recover silver. The molar ratio of the total Pd and Pt in the coprecipitate to hydrochloric acid is 1.0:12, and the molar ratio of the total Pd and Pt in the coprecipitate to sodium chlorate is 1.0:4.5.

[0078] S3. The platinum-palladium filtrate obtained in step S2 is heated to 75°C, and sodium chlorate and ammonium chloride are added to react. After 1.0 h of reaction, platinum-palladium coprecipitate and filtrate are obtained. The filtrate is then reacted with iron powder to recover the precious metals. The molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to sodium chlorate is 1.0:0.4, and the molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to ammonium chloride is 1.0:3.0.

[0079] S4 The platinum-palladium coprecipitate obtained in step S3 is slurried at a liquid-solid mass ratio of 4.0:1.0, and sodium sulfite is slowly added. After reacting at 30°C for 0.5 h, platinum-rich material and palladium-containing solution are obtained by solid-liquid separation. The platinum-rich material is used to purify platinum. The molar ratio of Pd to sodium sulfite in the platinum-palladium coprecipitate is 1.0:1.05.

[0080] S5 Add sodium chlorate and ammonium chloride to the palladium-containing solution obtained in step S4, and react at 75°C for 1.0 h to obtain a high-purity palladium-containing slag with a Pd content of 24.2 wt%; wherein, the molar ratio of Pd to sodium chlorate in the palladium-containing solution is 1.0:0.3, and the molar ratio of Pd to ammonium chloride in the palladium-containing solution is 1.0:3.0;

[0081] S6 The high-purity palladium-containing slag obtained in step S5 is mixed with ammonia water with a concentration of 25wt%, and complexed at 60℃ for 1.0h under the condition of pH 8.0. After the reaction, hydrazine hydrate is added, and the mixture is reduced at 60℃ for 0.5h to obtain sponge palladium with a purity of 99.6wt%. The molar ratio of Pd to ammonia water in the high-purity palladium-containing slag is 1.0:6.0, and the molar ratio of Pd to hydrazine hydrate in the high-purity palladium-containing slag is 1.0:0.8.

[0082] Example 2

[0083] A method for the co-enrichment of platinum and palladium and the preparation of sponge palladium from nitric acid leachate of silver anode mud, wherein the nitric acid leachate of silver anode mud is the filtrate produced from the nitric acid impurity removal process of silver anode mud, and the main components are: Cu 5.24 g / L, Au 0.28 mg / L, Ag 38.20 g / L, Pt 92.05 mg / L, and Pd 806.35 mg / L.

[0084] Includes the following steps:

[0085] S1. Diacetyl dioxime, sodium hydroxide, thiourea, and deionized water were prepared in a mass ratio of 1.0:1.0:0.2:5.0 and mixed thoroughly at 30°C to obtain a diacetyl dioxime-thiourea alkaline solution. The diacetyl dioxime-thiourea alkaline solution was then slowly added to a silver anode mud nitric acid leaching solution at 65°C. The temperature was raised to 85°C for platinum-palladium co-precipitation. After reacting for 1.0 h, the mixture was filtered while hot to obtain the coprecipitate and a silver-containing filtrate. The silver-containing filtrate was then... Sodium chloride is added to the solution, and the reaction yields silver chloride, which is then returned to the converter pyrometallurgical process to recover silver. The molar ratio of the total Pd and Pt in the nitric acid leaching solution of the silver anode mud to the total diacetyldioxime and thiourea in the diacetyldioxime-thiourea alkaline solution is 1.0:2.0. The coprecipitate contains 1.58 wt% platinum and 24.6 wt% palladium. The silver-containing filtrate contains 2.8 mg / L platinum and 4.2 mg / L palladium.

[0086] S2. The coprecipitate obtained in step S1 is slurried at a liquid-to-solid mass ratio of 4.5:1.0. Sodium hydroxide is added to adjust the pH to 9.5, and the temperature is raised to 80℃. The reaction is carried out for 1.0 h for a first alkaline conversion. Then, 2.5 mol / L hydrochloric acid is slowly added, and a second conversion is carried out at 85℃ for 0.55 h. After the reaction is completed, a saturated sodium chlorate solution is added and the reaction is carried out at 85℃ for 1.5 h. The solid and liquid are separated while hot to obtain a platinum-palladium filtrate and a filter residue. The filter residue is returned to the converter pyrometallurgical process to recover silver. The molar ratio of the total Pd and Pt in the coprecipitate to hydrochloric acid is 1.0:10, and the molar ratio of the total Pd and Pt in the coprecipitate to sodium chlorate is 1.0:4.0.

[0087] S3 The platinum-palladium filtrate obtained in step S2 is heated to 80°C, and sodium chlorate and ammonium chloride are added to react. After 1.0 h of reaction, platinum-palladium coprecipitate and filtrate are obtained. The filtrate is then reacted with iron powder to recover the precious metals. The molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to sodium chlorate is 1.0:0.2, and the molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to ammonium chloride is 1.0:2.4.

[0088] S4 The platinum-palladium coprecipitate obtained in step S3 is slurried at a liquid-solid mass ratio of 3.0:1.0, and sodium sulfite is slowly added. After reacting at 25°C for 0.5 h, platinum-rich material and palladium-containing solution are obtained by solid-liquid separation. The platinum-rich material is used to purify platinum. The molar ratio of Pd to sodium sulfite in the platinum-palladium coprecipitate is 1.0:1.05.

[0089] S5 Add sodium chlorate and ammonium chloride to the palladium-containing solution obtained in step S4, and react at 75°C for 1.0 h to obtain a high-purity palladium-containing slag with a Pd content of 26.8 wt%; wherein, the molar ratio of Pd to sodium chlorate in the palladium-containing solution is 1.0:0.2, and the molar ratio of Pd to ammonium chloride in the palladium-containing solution is 1.0:2.4.

[0090] S6 The high-purity palladium-containing slag obtained in step S5 is mixed with ammonia water with a concentration of 25wt%, and complexed at 60℃ for 1.0h under the condition of pH 9.0. After the reaction, hydrazine hydrate is added, and the mixture is reduced at 65℃ for 0.8h to obtain sponge palladium with a purity of 99.5wt%. The molar ratio of Pd to ammonia water in the high-purity palladium-containing slag is 1.0:5.0, and the molar ratio of Pd to hydrazine hydrate in the high-purity palladium-containing slag is 1.0:0.6.

[0091] Example 3

[0092] A method for the co-enrichment of platinum and palladium and the preparation of sponge palladium from nitric acid leachate of silver anode mud, wherein the nitric acid leachate of silver anode mud is the filtrate produced from the nitric acid purification process of silver anode mud, and the main components are: Cu 4.58 g / L, Au 0.32 mg / L, Ag 34.83 g / L, Pt 86.50 mg / L, and Pd 920.53 mg / L.

[0093] Includes the following steps:

[0094] S1. Diacetyl dioxime, sodium hydroxide, thiourea, and deionized water were prepared in a mass ratio of 1.0:1.0:0.15:5.0 and mixed thoroughly at 35°C to obtain a diacetyl dioxime-thiourea alkaline solution. The diacetyl dioxime-thiourea alkaline solution was then slowly added to a silver anode mud nitric acid leaching solution at 70°C. The temperature was raised to 83°C for platinum-palladium co-precipitation. After reacting for 1.5 hours, the mixture was filtered while hot to obtain the coprecipitate and a silver-containing filtrate. The silver-containing filtrate was then... Sodium chloride was added to the solution, and the reaction yielded silver chloride, which was then returned to the converter pyrometallurgical process to recover silver. The molar ratio of the total amount of Pd and Pt in the nitric acid leaching solution of the silver anode mud to the total amount of diacetyldioxime and thiourea in the diacetyldioxime-thiourea alkaline solution was 1.0:2.0. The coprecipitate contained 1.60 wt% platinum and 24.3 wt% palladium. The silver-containing filtrate contained 2.9 mg / L platinum and 3.8 mg / L palladium.

[0095] S2. The coprecipitate obtained in step S1 is slurried at a liquid-to-solid mass ratio of 5.0:1.0. Sodium hydroxide is added to adjust the pH to 10.0, and the temperature is raised to 83℃. The reaction is carried out for 0.7 h for a first alkaline conversion. Then, 2.6 mol / L hydrochloric acid is slowly added, and a second conversion is carried out at 83℃ for 0.6 h. After the reaction is completed, a saturated sodium chlorate solution is added, and the reaction is carried out at 83℃ for 1.8 h. The solid and liquid are separated while hot to obtain a platinum-palladium filtrate and a filter residue. The filter residue is returned to the converter pyrometallurgical process to recover silver. The molar ratio of the total Pd and Pt in the coprecipitate to hydrochloric acid is 1.0:15, and the molar ratio of the total Pd and Pt in the coprecipitate to sodium chlorate is 1.0:5.0.

[0096] S3 The platinum-palladium filtrate obtained in step S2 is heated to 77°C, and sodium chlorate and ammonium chloride are added to react. After 1.0 h of reaction, platinum-palladium coprecipitate and filtrate are obtained. The filtrate is then reacted with iron powder to recover the precious metals. The molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to sodium chlorate is 1.0:0.5, and the molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to ammonium chloride is 1.0:4.0.

[0097] S4 The platinum-palladium coprecipitate obtained in step S3 is slurried at a liquid-solid mass ratio of 3.5:1.0, and sodium sulfite is slowly added. After reacting at 20°C for 0.5 h, platinum-rich material and palladium-containing solution are obtained by solid-liquid separation. The platinum-rich material is used to purify platinum. The molar ratio of Pd to sodium sulfite in the platinum-palladium coprecipitate is 1.0:1.05.

[0098] S5 Add sodium chlorate and ammonium chloride to the palladium-containing solution obtained in step S4, and react at 75°C for 1.0 h to obtain a high-purity palladium-containing slag with a Pd content of 25.3 wt%; wherein, the molar ratio of Pd to sodium chlorate in the palladium-containing solution is 1.0:0.5, and the molar ratio of Pd to ammonium chloride in the palladium-containing solution is 1.0:4.0;

[0099] S6 The high-purity palladium-containing slag obtained in step S5 is mixed with ammonia water with a concentration of 25wt%, and complexed at 60℃ for 1.0h under the condition of pH 8.5. After the reaction, hydrazine hydrate is added, and the mixture is reduced at 70℃ for 1.0h to obtain sponge palladium with a purity of 99.8wt%. The molar ratio of Pd to ammonia water in the high-purity palladium-containing slag is 1.0:8.0, and the molar ratio of Pd to hydrazine hydrate in the high-purity palladium-containing slag is 1.0:1.0.

[0100] Example 4

[0101] Referring to Example 1, the difference is:

[0102] S5 Add sodium chlorate and ammonium chloride to the palladium-containing solution obtained in step S4, and react at 75°C for 1.0 h to obtain a high-purity palladium-containing slag with a Pd content of 24.2 wt%. The molar ratio of Pd to sodium chlorate in the palladium-containing solution is 1.0:0.3, and the molar ratio of Pd to ammonium chloride in the palladium-containing solution is 1.0:3.0.

[0103] The purification process then proceeded by slurrying the high-purity palladium-containing slag at a liquid-to-solid mass ratio of 4.0:1.0, slowly adding sodium sulfite, and reacting at 20°C for 1.0 h. After solid-liquid separation, a solution was obtained, in which the molar ratio of Pd to sodium sulfite in the high-purity palladium-containing slag was 1.0:1.2. Subsequently, sodium chlorate and ammonium chloride were added to the solution, and the reaction was carried out at 75°C for 1.0 h to obtain a first-stage purified high-purity palladium-containing slag with a Pd content of 26.5 wt%, in which the molar ratio of Pd to sodium chlorate in the solution was 1.0:0.3, and the molar ratio of Pd to ammonium chloride in the solution was 1.0:3.0.

[0104] The high-purity palladium-containing slag after primary purification was slurried at a liquid-to-solid mass ratio of 4.0:1.0, and sodium sulfite was slowly added. After reacting at 20℃ for 1.0 h, a solution was obtained by solid-liquid separation. In this solution, the molar ratio of Pd to sodium sulfite in the high-purity palladium-containing slag was 1.0:1.05. Subsequently, sodium chlorate and ammonium chloride were added to the solution, and the reaction was carried out at 75℃ for 1.0 h to obtain a second-purity high-purity palladium-containing slag with a Pd content of 27.8 wt%. In this solution, the molar ratio of Pd to sodium chlorate was 1.0:0.3, and the molar ratio of Pd to ammonium chloride was 1.0:3.0.

[0105] S6 The high-purity palladium-containing slag obtained after secondary purification in step S5 is mixed with ammonia water with a concentration of 25wt%, and complexed at 60℃ for 1.0h under the condition of pH 8.0. After the reaction, hydrazine hydrate is added, and the mixture is reduced at 60℃ for 0.5h to obtain sponge palladium with a purity of 99.95wt%. The molar ratio of Pd to ammonia water in the high-purity palladium-containing slag is 1.0:6.0, and the molar ratio of Pd to hydrazine hydrate in the high-purity palladium-containing slag is 1.0:0.8.

[0106] test

[0107] The recovery rates of Pd, Pt, and Ag in Examples 1-4 were calculated, and the results are shown in Table 1.

[0108] Table 1. Recovery rates of Pd, Pt, and Ag in Examples 1-4

[0109]

[0110] The above four case studies demonstrate that by preparing a novel platinum-palladium coprecipitation reagent, the average enrichment rates of palladium and platinum in the nitric acid leaching solution of silver anode mud reached 99.40% and 96.56%, respectively, showing significant enrichment and recovery effects. Furthermore, after performing two reduction-reprecipitation purification operations on the high-purity palladium-containing slag obtained in step S5, the quality of the reduced sponge palladium product was significantly improved, reaching the SM-Pd99.95 product standard.

[0111] The above description is only a preferred embodiment of the present invention and is not intended to limit the present invention. Any modifications, equivalent substitutions, improvements, etc., made within the spirit and principles of the present invention should be included within the protection scope of the present invention.

Claims

1. A method for synergistic enrichment of platinum and palladium from silver anode slime nitric acid leach liquor and green short process preparation of sponge palladium, characterized in that, Includes the following steps: S1 Add diacetyl dioxime-thiourea alkaline solution to the nitric acid leachate of silver anode mud to perform platinum-palladium coprecipitation, and obtain coprecipitate and silver-containing filtrate; S2 After the coprecipitate obtained in step S1 is pulped, the pH is adjusted to 9.0~10.0 for a first alkaline conversion; then hydrochloric acid is added for a second conversion, and after the conversion is completed, a saturated chlorate solution is added to react and obtain a platinum-palladium filtrate and filter residue. S3 Add an oxidant and ammonium chloride to the platinum-palladium filtrate obtained in step S2, and after the reaction, obtain platinum-palladium coprecipitate and filtrate; S4 After the platinum-palladium coprecipitate obtained in step S3 is slurried, a reducing agent is added and reacted to obtain platinum-rich material and palladium-containing solution; S5 Add an oxidant and ammonium chloride to the palladium-containing solution obtained in step S4, and after the reaction, obtain a high-purity palladium-containing slag. S6 The high-purity palladium-containing slag obtained in step S5 is mixed with ammonia water for complexation. After the reaction, hydrazine hydrate is added and reduced to obtain sponge palladium. In step S1, the mass ratio of diacetyldioxime, sodium hydroxide, thiourea, and deionized water in the diacetyldioxime-thiourea alkaline solution is 1.0:1.0:0.1~0.2:5.

0. The preparation method is as follows: diacetyldioxime, sodium hydroxide, and thiourea are dissolved in deionized water at 30~40℃ to obtain the diacetyldioxime-thiourea alkaline solution. In step S1, the platinum-palladium coprecipitation temperature is 80~85℃ and the time is 1.0~2.0h. In step S2, the primary alkaline conversion temperature is 80~85℃ and the time is 0.5~1.0h, and the liquid-solid mass ratio of the slurry is 4.0~5.0:1.

0. In step S4, the reaction temperature is 20~30℃. In step S4, the reducing agent is at least one of sodium sulfite, sodium bisulfite, potassium bisulfite, potassium sulfite, or potassium oxalate.

2. The method of claim 1, wherein, In step S2, the molar ratio of the total amount of Pd and Pt in the coprecipitate obtained in step S1 to hydrochloric acid is 1.0:10~15; the molar ratio of the total amount of Pd and Pt in the coprecipitate obtained in step S1 to chlorate is 1.0:4.0~5.

0.

3. The method of claim 1, wherein, In step S3, the molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to the oxidant is 1.0:0.2~0.5; the molar ratio of the total amount of Pd and Pt in the platinum-palladium filtrate to ammonium chloride is 1.0:2.4~4.

0.

4. The method according to claim 1, characterized in that, In step S5, the molar ratio of Pd to oxidant in the palladium-containing solution is 1.0:0.2~0.5; the molar ratio of Pd to ammonium chloride in the palladium-containing solution is 1.0:2.4~4.

0.

5. The method according to claim 1, characterized in that, In step S6, the molar ratio of Pd to ammonia in the high-purity palladium-containing slag is 1.0:5.0~8.0; the molar ratio of Pd to hydrazine hydrate in the high-purity palladium-containing slag is 1.0:0.6~1.

0.

6. The method according to claim 1, characterized in that, Step S5 also includes the purification of high-purity palladium-containing slag, which involves slurrying the high-purity palladium-containing slag, adding a reducing agent to carry out a reduction reaction to obtain a solution; subsequently, an oxidizing agent and ammonium chloride are added to the solution to carry out a precipitation reaction.

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