Method for multi-element stepwise separation and recovery of rhodium

Abstract

The application discloses a method for multi-element step separation and recovery of rhodium, and adopts an acid leaching lead-iso-VC sodium sedimentary tellurium-sodium sulfite palladium separation-acid leaching-sodium nitrite refining-sulfuration lead precipitation-ammonium chloride precipitation-acid dissolution-hydrazine hydrate reduction process flow, valuable elements, especially other platinum group elements and rhodium, are separated step by step after leaching of hydrolysis residue, and rhodium powder is prepared after acid dissolution and hydrazine hydrate reduction, the prepared rhodium powder has a purity greater than 99.9%, and a recovery rate greater than 86%.

Description

Technical Field

[0001] This invention relates to the field of separation and preparation of associated precious metals in non-ferrous metal smelting, specifically a method for multi-element stepwise separation and recovery of rhodium. Background Technology

[0002] Rhodium is an important precious metal element, and also a crucial element in the preparation of catalysts. It plays a vital role in industries such as chemical engineering, aerospace, and electronics. In nature, rhodium coexists with other platinum group metals, but in relatively low amounts, and also contains a large amount of other base metal impurities. How to separate and recover rhodium is a challenging problem in the separation and extraction of precious metals.

[0003] Currently, there is a lot of research on the recycling and reuse of waste rhodium-containing alloys both domestically and internationally. For example, Lü Jiuji, Wang Zhanhua, et al. separated and purified metallic rhodium from rhodium-containing waste residue using an extraction separation method (Nonferrous Metals Mining and Metallurgy, 1993(2)). The process is: rhodium-containing waste residue → acidification to prepare the solution → dimethyl diketone chloroform extraction to remove palladium and nickel → methyl isobutyl ketone extraction to separate copper, iron, lead, arsenic, etc. → N235 extraction to separate platinum and iridium → TBP extraction to separate iridium → reduction → rhodium powder. Chinese patent CN104860358B discloses a method for recycling and purifying high-purity rhodium, which involves a method for separating and purifying rhodium from rhodium-containing waste leachate and other precious and base metals. Tetramethylammonium chloride is used as a precipitant, and the concentration of the rhodium solution is controlled at 30-80 g / L. A small amount of precipitant is added first to react and remove high-valence impurities such as platinum and iron. The resulting filtrate is then added with excess precipitant and boiled and refluxed at 110-120℃ for 24-48 hours to obtain rhodium salt precipitate. Rhodium salt precipitate is thoroughly washed, dissolved, passed through resin, and precipitated with potassium hydroxide to obtain high-purity rhodium hydroxide. Chinese Patent 200910075588.6 discloses a method for extracting rhodium powder from waste platinum mesh. This method addresses the issue of high rhodium content in aqua regia-insoluble materials from waste platinum mesh processing. The extraction process involves mechanical crushing, grinding, sieving, barium peroxide calcination, dissolution, precipitation, filtration, heating concentration, washing, platinum removal, sodium hexane nitrite complexation, heating to boiling, vacuum filtration, rapid cooling to prepare a rhodium complex solution, followed by heating dissolution, ion exchange, hydrazine hydrate reduction, distillation, drying, and hydrogen calcination reduction to finally produce grayish-white rhodium powder. Chinese Patent 201210213723.0 discloses a chlorination purification method for platinum, platinum-rhodium, or platinum-rhodium-palladium alloys. This method involves melting platinum, platinum-rhodium, or platinum-rhodium-palladium metals and then passing chlorine gas over the surface of the molten metal to remove base metals. However, there is limited research on rhodium-containing materials generated during non-ferrous metal smelting, especially hydrolysis slag produced from platinum-palladium concentrate refining, and further research is needed to recover metallic rhodium. Summary of the Invention

[0004] The purpose of this invention is to provide a method for separating and extracting rhodium from rhodium-containing materials produced by non-ferrous metal smelting, particularly from hydrolysis residue produced by platinum-palladium refining.

[0005] To achieve the above objectives, the present invention provides the following technical solution:

[0006] A method for multi-element stepwise separation and recovery of rhodium includes the following steps:

[0007] S1 Acid leaching for lead removal: Water and hydrochloric acid are added to the hydrolysis residue for acid leaching, and the residue is filtered to obtain lead-removed filtrate and lead-rich slag. S2 Sodium iso-VC precipitation for tellurium: Sodium iso-VC is added to the lead-removed filtrate for tellurium precipitation, and the residue is filtered to obtain tellurium-removed filtrate and tellurium-rich slag. S3 Sodium sulfite precipitation for palladium: Sodium sulfite is added to the tellurium-removed filtrate for reaction, and the residue is filtered to obtain rhodium-rich slag and palladium-rich solution. S4 Acid leaching: Water and hydrochloric acid are added to the rhodium-rich slag for acid leaching, and the residue is filtered to obtain a rhodium-containing solution and filter residue one. S5 Sodium nitrite refining: The rhodium-containing solution is concentrated, sodium hydroxide is added to adjust the pH value, sodium nitrite is added for reaction, and the residue is filtered to obtain filter residue two and rhodium-rich solution. S6 Sulfide precipitation for lead: Sodium sulfide is added to the rhodium-rich solution for lead precipitation. The sodium sulfide is added in a ratio of lead to sodium sulfide in the rhodium-rich solution of 1:1.1 to 1.5 by mass. The solution is then heated to near boiling and maintained for 3 to 15 minutes. Ferric chloride solution is added while stirring, and the pH is adjusted to 7 to 11. The solution is filtered to obtain lead slag and lead-removed rhodium-rich solution. S7: Ammonium chloride precipitates rhodium. Acetic acid is added to the lead-removed rhodium-rich solution to adjust the pH to 3.5. Ammonium chloride is added, and the solution is filtered to obtain sodium ammonium hexanonitrosorhodium complex. S8: Acid dissolution. Water and hydrochloric acid are added to the sodium ammonium hexanonitrosorhodium complex to dissolve the solution. The solution is filtered to obtain rhodium solution. S9: Hydrazine hydrate reduction. After removing the nitrate from the rhodium solution, sodium hydroxide is added to the nitrate-removed rhodium solution to adjust the pH to 6.5 to 9. Hydrazine hydrate is added to carry out the reduction reaction. The solution is filtered to obtain rhodium powder, which is then dried.

[0008] Furthermore, in the S1 acid leaching for lead removal, water is added to the hydrolysis residue at a liquid-to-solid ratio of 5:1 to slurry the hydrolysis residue. The mass ratio of the hydrolysis residue to the hydrochloric acid is 1:1 to 1.2. The reaction temperature is 70 to 80°C, and the reaction time is 1 to 2 hours.

[0009] Furthermore, in the S2 sodium isoVC precipitation of tellurium, the lead-removing filtrate is heated to 90°C before the sodium isoVC is added, the mass ratio of sodium isoVC to tellurium is 3-4:1, and the reaction time is 1-3 hours.

[0010] Furthermore, in the S3 sodium sulfite palladium fractionation, the amount of sodium sulfite added is based on a molar ratio of rhodium to sodium sulfite in the solution of 1:5, the reaction temperature is 70-80°C, and the reaction time is 1-2 hours.

[0011] Furthermore, in the S4 acid leaching process, water is added at a liquid-to-solid ratio of 5:1, the concentration of hydrochloric acid in the acid leaching solution is 3-10 mol / L, the reaction temperature is 70-90°C, and the reaction time is 1-2 hours.

[0012] Furthermore, the S5 sodium nitrite refining process also includes:

[0013] After concentrating the rhodium-containing solution to 1 / 2 to 1 / 3 of its original volume, sodium hydroxide is added to the concentrated rhodium-containing solution to adjust the pH to 1.5, and the reaction temperature is 70 to 75°C. Then, sodium nitrite is slowly added, and the mixture is boiled for 0.5 hours. The amount of sodium nitrite added is based on a mass ratio of rhodium to sodium nitrite of 1:3 in the concentrated rhodium solution. Sodium hydroxide is then added to adjust the pH to 9, and the mixture is boiled for 1 hour. The mixture is then filtered to obtain the second filter residue and the rhodium-rich solution.

[0014] Furthermore, in the S7 ammonium chloride precipitate rhodium, the amount of ammonium chloride added is 1:1 to 2.5 in the molar ratio of rhodium to ammonium chloride in the lead-removing rhodium-rich solution, and the reaction time is 1 hour.

[0015] Furthermore, in the S8 acid solution, a hydrochloric acid solution is prepared according to a volume ratio of water to hydrochloric acid of 1:1 to 1.5, and the sodium ammonium hexanonitrosorhodium complex is added to the hydrochloric acid solution with a liquid-to-solid ratio of 5 to 11:1. The reaction temperature is 75 to 80°C, and the reaction time is 1 to 1.5 hours.

[0016] Furthermore, in the S9 hydrazine hydrate reduction, the temperature for removing nitrate is 120-150°C, and the time is 1-2 hours; the hydrazine hydrate is added at a molar ratio of hydrazine hydrate to rhodium in the nitrate-removing rhodium solution of 1-1.8:1, the reaction temperature is 80-90°C, and the reaction time is 0.5-1 hours.

[0017] Compared with the prior art, the beneficial effects of the present invention are as follows: The present invention adopts a process flow of acid leaching to remove lead-sodium isocyanate precipitation of tellurium-sodium sulfite separation of palladium-acid leaching-sodium nitrite refining-sulfide precipitation of lead-ammonium chloride precipitation-acid dissolution-hydrazine hydrate reduction. After leaching the hydrolysis residue, valuable elements are separated stepwise, realizing the stepwise separation and recovery of rhodium from lead, tellurium, and palladium. In particular, other platinum group elements are fully separated from rhodium. After further acid dissolution and hydrazine hydrate reduction, rhodium powder is obtained. The purity of the obtained rhodium powder is greater than 99.9%, and the recovery rate is greater than 86%. Attached Figure Description

[0018] Figure 1 This is a process flow diagram of the present invention. Detailed Implementation

[0019] To make the objectives, technical solutions, and advantages of this invention clearer, the technical solutions of this invention will be clearly and completely described below in conjunction with embodiments. It should be understood that the specific embodiments described herein are merely illustrative of the invention and are not intended to limit the invention.

[0020] This invention provides a method for multi-element stepwise separation and recovery of rhodium, comprising the following steps:

[0021] S1 acid leaching for lead removal begins by slurrying the hydrolysis residue. Water, preferably pure water, is added to the residue at a liquid-to-solid ratio of 5:1, and the mixture is stirred to form a slurry. The slurry is then heated to 70–80°C, and hydrochloric acid is added at a mass ratio of 1:1–1.2 (hydrochloric acid to hydrolysis residue). The leaching reaction is then carried out with stirring for 1–2 hours. The resulting solution is filtered to obtain lead-removed filtrate and lead-rich slag. The lead-rich slag is then returned to the lead smelting system for refining and casting into lead ingots.

[0022] Hydrolysis residue is the residue produced during the refining of platinum and palladium concentrate. In platinum and palladium refining, especially platinum refining, oxidative hydrolysis (such as aqueous solution chlorination) is generally used to obtain platinum with relatively high purity. Hydrolysis residue is produced during the hydrolysis process. This hydrolysis residue is rich in rhodium, as well as other valuable elements such as lead, tellurium and palladium.

[0023] S2 Sodium iso-VC precipitation of tellurium: The lead-removed filtrate obtained in step S1 is heated to 90°C, and then sodium iso-VC is added. Sodium iso-VC reduces the high-valence tellurium in the solution to form a precipitate, thereby achieving the purpose of separating tellurium. The amount of sodium iso-VC added is 3 to 4 times the mass ratio of sodium iso-VC to tellurium, and the reaction time is 1 to 3 hours. Then, the solution is filtered to obtain the tellurium-removed filtrate and tellurium-rich residue.

[0024] S3 Palladium separation with sodium sulfite: Sodium sulfite is added to the tellurium removal filtrate to react and cause rhodium to precipitate and separate from other metals. The filtrate is filtered to obtain rhodium-rich slag and palladium-rich solution. Sodium sulfite is added according to the molar ratio of rhodium to sodium sulfite in the solution of 1:5. The precipitation reaction temperature is 70-80℃ and the precipitation reaction time is 1-2 hours.

[0025] S4 acid leaching involves adding pure water to the rhodium-rich slag at a liquid-to-solid ratio of 5:1, followed by the addition of hydrochloric acid to leach the rhodium back into the solution. The solution is then filtered to obtain a rhodium-containing solution and filter residue 1. The amount of hydrochloric acid added is based on the concentration of 3–10 mol / L in the leaching solution. The reaction temperature is 70–90℃, and the reaction time is 1–3 hours.

[0026] S5 sodium nitrite refining involves concentrating the rhodium-containing solution before adding it to the solution. This is done by heating and evaporating the solution to reduce its volume to 1 / 2 to 1 / 3 of its original volume. Sodium hydroxide is then added to the concentrated solution to adjust the pH to 1.5, with the temperature controlled at 70-75°C. Sodium nitrite is then slowly added, followed by boiling for 0.5 hours. The amount of sodium nitrite added is based on a 1:3 mass ratio of rhodium to sodium nitrite in the concentrated solution. After boiling, sodium hydroxide is added again to adjust the pH to 9, and the solution is boiled for another hour. Filtration yields filter residue 2 and a rhodium-rich solution.

[0027] S6 lead sulfide precipitation involves adding sodium sulfide to a rhodium-rich solution to induce a lead sulfide precipitate. The amount of sodium sulfide added is based on a lead-to-sodium sulfide mass ratio of 1:1.1 to 1.5 in the rhodium-rich solution. The temperature should be controlled to not exceed 60°C. The solution is then heated to near boiling and maintained for 3 to 10 minutes. Ferric chloride solution is added while stirring, with a pH of 7 to 11. The solution is then filtered to obtain lead slag and a lead-removed rhodium-rich solution.

[0028] To precipitate rhodium using S7 ammonium chloride, the pH of the solution needs to be adjusted before precipitation. First, acetic acid is added to the lead-removing rhodium-rich solution to adjust the pH to 3.5, and then ammonium chloride is added. After filtration, sodium ammonium hexanonitrosorhodium complex is obtained. The amount of ammonium chloride added is based on the molar ratio of rhodium to ammonium chloride in the lead-removing rhodium-rich solution of 1:1 to 2.5. The reaction time is 1 hour. After filtration, sodium ammonium hexanonitrosorhodium complex and the rhodium-precipitated solution are obtained.

[0029] S8 acid dissolution: Prepare a hydrochloric acid solution with a water-to-hydrochloric acid volume ratio of 1:1 to 1.5. Add sodium ammonium hexanonitrosorhodium complex to the hydrochloric acid for acidic dissolution. Filter to obtain rhodium solution. The liquid-to-solid ratio of the solution to sodium ammonium hexanonitrosorhodium complex is 5 to 11:1. The reaction temperature is 75 to 80°C and the reaction time is 1 to 1.5 hours.

[0030] The S9 hydrazine hydrate reduction process requires a denitrification step before reducing rhodium. The denitrification temperature is 120–150°C, and the time is 1–2 hours. Then, sodium hydroxide is added to the denitrified rhodium solution to adjust the pH to 6.5–9, and then hydrazine hydrate is added to carry out the reduction reaction. The resulting rhodium powder is filtered and dried. The ratio of hydrazine hydrate to rhodium in the denitrified rhodium solution is 1–1.8:1, the reaction temperature is 80–90°C, and the reaction time is 0.5–1 hour.

[0031] According to the multi-element stepwise separation and recovery method of rhodium provided by the present invention, the complex impurity elements in the hydrolysis residue can be separated stepwise. The separated impurity elements are easy to further purify, and the rhodium can be purified and extracted to obtain rhodium with a purity greater than 99.9% and a recovery rate greater than 86%.

[0032] Example 1:

[0033] The composition of the hydrolysis residue used in this embodiment is shown in Table 1.

[0034]

[0035] (1) Acid leaching for lead removal: Weigh 85g of hydrolysis residue and place it in a 2L beaker. Add 340ml of pure water and 85ml of 37% hydrochloric acid to the 2L beaker, heat to 78℃, and stir for 1 hour. Filter the solution to obtain 1000ml of lead-removed filtrate and 15.5g of lead-rich slag (wet weight). The lead-rich slag is returned to the lead smelting system for refining and casting lead ingots. The composition of the lead-removed filtrate and lead-rich slag is shown in Table 2.

[0036]

[0037] (2) Sodium iso-VC precipitation of tellurium: Take the lead removal filtrate and place it in a 2L glass beaker. Heat the solution to 90℃, start stirring, add the required 27g sodium iso-VC, the reaction time is 2.5h, the reaction temperature is 90℃, and after the reaction is completed, 32.11g of wet tellurium-rich residue and 950ml of tellurium removal filtrate are obtained. The tellurium-rich residue is used to extract tellurium ingots. The composition of the tellurium removal solution is shown in Table 3.

[0038]

[0039] (3) Palladium extraction with sodium sulfite: Place the tellurium-removed filtrate in a 2L glass beaker, heat the solution in an 80℃ water bath with stirring, and slowly add 47g of sodium sulfite. The reaction time is 1h at 80℃. After standing and cooling, filter to obtain rhodium-rich slag and palladium-rich solution. Wash the slag 2-3 times with hot water. The wet weight of the slag is 118g, and the palladium-rich solution is 1430ml. The palladium-rich solution is used to extract high-purity palladium. The composition of the palladium-rich solution is shown in Table 4.

[0040]

[0041] (4) Acid leaching: Take 118g of the wet weight of the rhodium-rich slag and place it in a 2L glass beaker. Add 200ml of pure water and 300ml of hydrochloric acid to adjust the concentration of hydrochloric acid in the system to 6mol / L. Place the solution in a water bath at 90℃ and heat it. Start stirring to slurry it. React at 90℃ for 3h. After the reaction is completed, 965ml of rhodium-containing solution and 2g of acid leaching slag are obtained. The composition of the rhodium-rich solution is shown in Table 5.

[0042]

[0043] (5) Refining of sodium nitrite: Take the rhodium-containing solution obtained in the previous step, concentrate it to 200 ml, place it in a 1 L glass beaker and heat it in a water bath. Start stirring and heat at a rate of 1 °C / min. Add 65.5 ml of 20% sodium hydroxide solution to adjust the pH to 1.5. Heat the solution to 75 °C. Slowly add 44 g of sodium nitrite while stirring continuously and heat to near boiling. The final pH of the solution is 7.5. Boil for 0.5 h. Adjust the pH to 9 with 20 ml of 20% NaOH. Continue boiling the solution for 1 h and then filter to obtain 459 ml of rhodium-rich solution and 22.73 g of refining residue. The composition of the rhodium-rich solution is shown in Table 6.

[0044]

[0045] (6) Sulfide precipitation of lead: Add the rhodium-rich liquid to a 2L glass beaker, stir at room temperature, add 3.5g sodium sulfide, keep for 5min, then heat to near boiling, keep for 10min, stir and add a solution containing 0.7g ferric chloride, pH 7.5, filter, and obtain 1003ml of lead-removing rhodium-rich liquid and 11g lead slag. The lead slag is returned to the lead smelting system to produce lead ingots; the composition of the lead-removing rhodium-rich liquid is shown in Table 7.

[0046]

[0047] (7) Ammonium chloride precipitation: Take the lead-removed rhodium-rich solution and place it in a 2L glass beaker. Start stirring, add 20ml of glacial acetic acid to pH=3.5, slowly add 18g of ammonium chloride, and react for 1h. Filter and dry the precipitate to obtain 70.29g of sodium ammonium hexanonitrosorhodium chelate and 1325ml of rhodium-precipitated solution. The composition of the rhodium-precipitated solution is shown in Table 8.

[0048]

[0049] (8) Acid dissolution: Place sodium ammonium hexanonitrosotrimonium rhodium in a 2L glass beaker, add 150ml of pure water and 230ml of 37% hydrochloric acid, place in a water bath, turn on the stirring, heat to 78℃, and the reaction time is 1h; to obtain rhodium solution.

[0050] (9) Reduction of hydrazine hydrate: Place the rhodium solution obtained in the previous step on an electric multi-purpose furnace and heat it to 120°C to remove the nitrate. Place the nitrate-removing solution in a water bath, turn on the stirrer, control the temperature to 90°C, and add 270 ml of 20% sodium hydroxide solution while stirring to adjust the pH to 6.5. Then add 25 ml of 40% hydrazine hydrate. The reaction time is 1 hour. After the reaction is completed, 592 ml of reduced solution and 21.53 g of wet rhodium powder are obtained. After drying, 15.27 g of rhodium powder is obtained. The composition of rhodium powder is shown in Table 9.

[0051]

[0052] In this embodiment, the purity of rhodium is 99.96% and the recovery rate is 86.07%.

[0053] Example 2:

[0054] This embodiment also uses the hydrolysis residue from Example 1.

[0055] (1) Acid leaching for lead removal: Weigh 75g of hydrolysis residue into a 2L beaker, add 300ml of pure water and 80ml of 37% hydrochloric acid to the 2L beaker, heat to 80℃, and stir for 1.5h. Cool the solution, pour out the liquid, and filter to obtain 900ml of lead-removed filtrate and 12.87g of lead-rich slag (wet weight). The lead-rich slag is returned to the lead smelting system to produce lead ingots. The composition of the lead-removed filtrate and lead-rich slag is shown in Table 10.

[0056]

[0057] (2) Sodium iso-VC precipitation of tellurium: Take the filtrate from the previous step and place it in a 2L glass beaker. Heat the solution to 90℃, start stirring, add the required 28g sodium iso-VC, the reaction time is 2.5h, the reaction temperature is 90℃, and after the reaction is completed, 35.81g of wet tellurium-rich residue and 1088ml of tellurium-removing filtrate are obtained. The tellurium-rich residue is used to extract tellurium ingots. The composition of the tellurium-removing filtrate is shown in Table 11.

[0058]

[0059] (3) Palladium extraction with sodium sulfite: Place the tellurium removal filtrate from the previous step into a 2L glass beaker, heat the solution in an 80℃ water bath, turn on the stirrer, and slowly add 46g of sodium sulfite. The reaction time at 80℃ is 1h. Allow to stand, cool, and filter. Wash the residue 2-3 times with hot water to obtain 105g of rhodium-rich residue (wet weight) and 1300ml of palladium-rich solution. The palladium-rich solution is used to extract high-purity palladium; the composition of the palladium-rich solution is shown in Table 12.

[0060]

[0061] (4) Acid leaching: Place the rhodium-rich slag from the previous step into a 2L glass beaker, add 200ml of pure water and 450ml of 37% hydrochloric acid solution, place it in a water bath, turn on the stirring, heat to 90℃, and the reaction time is 3h. Filter to obtain 765ml of rhodium-containing solution. The composition of the rhodium-containing solution is shown in Table 13.

[0062]

[0063] (5) Refining of sodium nitrite: Take the acid leaching solution from the previous step and place it in a 1L glass beaker. Concentrate the solution to 300ml and heat it in a 70℃ water bath. Start stirring and add 115ml of 20% sodium hydroxide to adjust the pH to 1.5. Slowly add 46g of sodium nitrite while stirring continuously and heat to near boiling. Boil for 0.5h. Then adjust the pH to 9 with 15ml of 20% sodium hydroxide solution. Heat the solution in a 90℃ water bath, start stirring, and continue boiling for 1h. Filter to obtain 1150ml of rhodium-rich solution. The composition of the rhodium-rich solution is shown in Table 14.

[0064]

[0065] (6) Lead precipitation with sulfide: Add rhodium-rich solution to a 2L glass beaker, stir at room temperature, add sodium sulfide containing 9g, maintain for 5min, then heat to near boiling, maintain for 8min, stir and add a solution containing 1.8g ferric chloride, pH 9, and filter. The product is 1003ml of lead-removed rhodium-rich solution and 17g of lead slag. The lead slag is returned to the lead smelting system to produce lead ingots; the composition of the lead-removed rhodium-rich solution is shown in Table 15.

[0066]

[0067] (7) Ammonium chloride precipitation: Take the filtrate from the previous step and place it in a 2L glass beaker. Start stirring, add 10ml of glacial acetic acid to adjust the pH to slightly acidic, slowly add 19.5g of ammonium chloride, and react for 1h. Filter out the precipitated rhodium salt, dry it to obtain 62.18g of sodium ammonium hexanonitrosorhodium chelate and 400ml of rhodium precipitation solution; the composition of the rhodium precipitation solution is shown in Table 16.

[0068]

[0069] (8) Acid dissolution: Place sodium ammonium hexanonitrosotrimonium rhodium in a 2L glass beaker, add 180ml of pure water and 220ml of 37% hydrochloric acid, place in a water bath, turn on the stirring, heat to 78℃, and the reaction time is 1.5h. Filter to obtain rhodium solution.

[0070] (9) Reduction of hydrazine hydrate: Place the acid solution on an electric multi-purpose furnace and heat it to 130°C to remove the nitrate. Place the nitrate-removing solution in a water bath, turn on the stirrer, control the temperature to 85°C, and add 172 ml of 20% sodium hydroxide solution while stirring to adjust the pH to 9. Then add 20 ml of 40% hydrazine hydrate. The reaction time is 0.5 h. After the reaction is complete, 1715 ml of reduced solution and 20.29 g of wet rhodium powder are obtained. After drying, 13.59 g of rhodium powder is obtained. The composition of rhodium powder is shown in Table 17.

[0071]

[0072] In this embodiment, the purity of rhodium is 99.97% and the recovery rate is 86.83%.

[0073] Example 3:

[0074] The composition of the hydrolysis residue used in this embodiment is shown in Table 18.

[0075]

[0076] (1) Acid leaching for lead removal: Weigh 100g of hydrolysis residue and place it in a 2L beaker. Add 400ml of pure water and 100ml of 37% hydrochloric acid solution to the 2L beaker. Heat to 70℃ and stir for 2 hours. After cooling, pour out the liquid and filter to obtain 1230ml of lead-removed filtrate and 52.39g of lead-rich slag (wet weight). The lead-rich slag is returned to the lead smelting system to produce lead ingots. The composition of the lead-removed filtrate and lead-rich slag is shown in Table 19.

[0077]

[0078] (2) Sodium iso-VC precipitation of tellurium: Take the lead removal filtrate from the previous step and place it in a 2L glass beaker. Heat the solution to 90℃, start stirring, add the required 31g sodium iso-VC, the reaction time is 2.5h, the reaction temperature is 90℃, and after the reaction is completed, 26.24g of wet tellurium-rich residue and 786ml of tellurium removal filtrate are obtained. The tellurium-rich residue is used to extract tellurium ingots. The composition of the tellurium removal filtrate is shown in Table 20.

[0079]

[0080] (3) Palladium extraction with sodium sulfite: Place the tellurium removal filtrate from the previous step into a 2L glass beaker, heat the solution in an 80℃ water bath, turn on the stirrer, and slowly add 97.5g of sodium sulfite. The reaction time at 80℃ is 1h. Allow to stand, cool, and filter. Wash the residue 2-3 times with hot water. The wet weight of the rhodium-rich residue is 140g, and the palladium-rich solution is 1690ml. The palladium-rich solution is used to extract high-purity palladium; the composition of the palladium-rich solution is shown in Table 21.

[0081]

[0082] (4) Acid leaching: Take the rhodium-rich slag from the previous step and place it in a 2L glass beaker. Add 200ml of pure water and 300ml of hydrochloric acid to adjust the hydrochloric acid concentration to 6mol / L. Place the solution in a 90℃ water bath and heat it. Start stirring to slurry the solution. React at 90℃ for 3 hours. After the reaction is complete, 900ml of rhodium-containing solution and 3g of acid leaching slag are obtained. The composition of the rhodium-containing solution is shown in Table 22.

[0083]

[0084] (5) Refining of sodium nitrite: Take the rhodium-containing solution from the previous step and place it in a 1L glass beaker. Concentrate the solution to 400ml and heat it in a 70℃ water bath. Start stirring and add 235ml of 20% sodium hydroxide to adjust the pH to 1.5. Slowly add 58g of sodium nitrite while stirring continuously and heat to near boiling. Boil for 0.5h. Then adjust the pH to 9 with 125ml of 20% sodium hydroxide solution. Heat the solution in a 90℃ water bath, start stirring, and continue boiling for 1h. Filter to obtain 850ml of rhodium-rich solution and 17.36g of refining residue. The composition of the rhodium-rich solution is shown in Table 23.

[0085]

[0086] (6) Lead precipitation with sulfide: Add the rhodium-rich solution from the previous step to a 2L glass beaker, stir at room temperature, add 2.1g of sodium sulfide, maintain for 5 minutes, then heat to near boiling, maintain for 15 minutes, stir and add a solution containing 0.5g of ferric chloride, pH 8, and filter. 1180ml of lead-removed rhodium-rich solution is produced, and the lead slag is returned to the lead smelting system to produce lead ingots; the composition of the lead-removed rhodium-rich solution is shown in Table 24.

[0087]

[0088] (7) Ammonium chloride precipitation: Take the lead-removing and rhodium-rich solution from the previous step and place it in a 2L glass beaker. Start stirring, add 10ml of glacial acetic acid to adjust the pH to slightly acidic, slowly add 29.8g of ammonium chloride, react for 1 hour, let stand for 20 minutes, filter out the precipitate and dry to obtain 78.92g of sodium ammonium hexanonitrosorhodium and 280ml of rhodium-precipitated solution; the composition of the rhodium-precipitated solution is shown in Table 25.

[0089]

[0090] (8) Acid dissolution: Place sodium ammonium hexanonitrosotrimonium rhodium in a 2L glass beaker, add 160ml of pure water and 240ml of 37% hydrochloric acid, place in a water bath, turn on the stirring, heat to 78℃, and the reaction time is 1.5h.

[0091] (9) Reduction of hydrazine hydrate: Place the acid solution on an electric multi-functional furnace and heat it to 145°C to remove the nitrate. Place the nitrate-removing solution in a water bath, turn on the stirrer, heat it to 90°C, and add 380ml of 20% sodium hydroxide solution while stirring to adjust the pH to 6.5. Then add 31ml of 40% hydrazine hydrate. The reaction time is 1 hour. After the reaction is complete, 580ml of reduced solution and 35g of wet rhodium powder are obtained. After drying, 17.8g is obtained.

[0092]

[0093] In this embodiment, the recovered rhodium purity was 99.94%, and the recovery rate was 89.25%.

[0094] Although the present invention has been described in detail with reference to the foregoing embodiments, those skilled in the art can still modify the technical solutions described in the foregoing embodiments or make equivalent substitutions for some of the technical features. 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 multi-element stepwise separation and recovery of rhodium, characterized in that, Includes the following steps: S1 acid leaching for lead removal involves adding water and hydrochloric acid to the hydrolysis residue for acid leaching, followed by filtration to obtain lead-removed filtrate and lead-rich slag. S2 Sodium iso-VC precipitation of tellurium: Sodium iso-VC is added to the lead-removed filtrate to carry out a tellurium precipitation reaction, and the filtrate is filtered to obtain tellurium-removed filtrate and tellurium-rich residue. S3 Palladium separation with sodium sulfite: Sodium sulfite is added to the tellurium removal filtrate to react, and then filtered to obtain rhodium-rich slag and palladium-rich solution. S4 acid leaching: Water and hydrochloric acid are added to the rhodium-rich slag for acid leaching, and the slag is filtered to obtain a rhodium-containing solution and filter residue 1. S5 sodium nitrite refining: After concentrating the rhodium-containing solution, sodium hydroxide is added to adjust the pH value, sodium nitrite is added to react, and the solution is filtered to obtain filter residue 2 and rhodium-rich solution. S6 lead precipitation by adding sodium sulfide to the rhodium-rich solution to carry out the lead precipitation reaction. The amount of sodium sulfide added is 1:1.1 to 1.5 according to the mass ratio of lead to sodium sulfide in the rhodium-rich solution. Then the temperature is raised to near boiling and maintained for 3 to 15 minutes. Ferric chloride solution is added while stirring. The pH is 7 to 11. The solution is filtered to obtain lead slag and lead-removed rhodium-rich solution. S7 ammonium chloride precipitates rhodium. Acetic acid is added to the lead-removed rhodium-rich solution to adjust the pH to 3.

5. Ammonium chloride is added and the solution is filtered to obtain sodium ammonium hexanonitrosotrimonium. S8 acid dissolution: Water and hydrochloric acid are added to the sodium ammonium hexanonitroso rhodium complex to dissolve it, and the solution is filtered to obtain rhodium solution; S9 hydrazine hydrate reduction: After removing nitrate from the rhodium solution, sodium hydroxide is added to the nitrate-removed rhodium solution to adjust the pH value to 6.5-9, hydrazine hydrate is added to carry out the reduction reaction, and the solution is filtered to obtain rhodium powder.

2. The method for multi-element stepwise separation and recovery of rhodium according to claim 1, characterized in that, In the S1 acid leaching for lead removal, water is added to the hydrolysis residue at a liquid-to-solid ratio of 5:1 to slurry the hydrolysis residue. The mass ratio of the hydrolysis residue to the hydrochloric acid is 1:1 to 1.

2. The reaction temperature is 70 to 80°C, and the reaction time is 1 to 2 hours.

3. The method for multi-element stepwise separation and recovery of rhodium according to claim 1, characterized in that, In the S2 sodium isoVC precipitation of tellurium, the lead-removing filtrate is heated to 90°C and then the sodium isoVC is added. The mass ratio of sodium isoVC to tellurium is 3-4:1, and the reaction time is 1-3 hours.

4. The method for multi-element stepwise separation and recovery of rhodium according to claim 1, characterized in that, In the S3 sodium sulfite palladium fractionation, the amount of sodium sulfite added is based on a molar ratio of rhodium to sodium sulfite in the solution of 1:5, the reaction temperature is 70-80°C, and the reaction time is 1-2 hours.

5. The method for multi-element stepwise separation and recovery of rhodium according to claim 1, characterized in that, In the S4 acid leaching process, pure water is added at a liquid-to-solid ratio of 5:1, the concentration of hydrochloric acid in the leaching solution is 3-10 mol / L, the reaction temperature is 70-90℃, and the reaction time is 1-2 hours.

6. The method for multi-element stepwise separation and recovery of rhodium according to claim 1, characterized in that, The S5 sodium nitrite refining process also includes: After concentrating the rhodium-containing solution to 1 / 2 to 1 / 3 of its original volume; The sodium hydroxide was added to the concentrated rhodium-containing solution to adjust the pH to 1.5, and the reaction temperature was 70-75°C. Then slowly add the sodium nitrite and boil for 0.5 hours. The amount of sodium nitrite added is based on a mass ratio of rhodium to sodium nitrite in the concentrated rhodium solution of 1:

3. Add sodium hydroxide to adjust the pH to 9, and boil for 1 hour; The filter residue and the rhodium-rich solution were obtained by filtration.

7. The method for multi-element stepwise separation and recovery of rhodium according to claim 1, characterized in that, In the S7 ammonium chloride precipitate rhodium, the amount of ammonium chloride added is 1:1 to 2.5 according to the molar ratio of rhodium to ammonium chloride in the lead-removing rhodium-rich solution, and the reaction time is 1 hour.

8. The method for multi-element stepwise separation and recovery of rhodium according to claim 1, characterized in that, In S8 acid solution, a hydrochloric acid solution is prepared according to a volume ratio of water to hydrochloric acid of 1:1 to 1.

5. Sodium ammonium hexanonitrosorhodium complex is added to the hydrochloric acid solution, with a liquid-to-solid ratio of 5 to 11:

1. The reaction temperature is 75 to 80°C, and the reaction time is 1 to 1.5 hours.

9. The method for multi-element stepwise separation and recovery of rhodium according to claim 1, characterized in that, In the S9 hydrazine hydrate reduction, the temperature for removing nitrate is 120-150°C, and the time is 1-2 hours; The hydrazine hydrate is prepared in a molar ratio of hydrazine hydrate to rhodium in the rhodium-removing solution of 1 to 1.8:1, the reaction temperature is 80 to 90°C, and the reaction time is 0.5 to 1 hour.

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