A method for recovering scattered platinum group metals from the liquid after gold extraction from silver anode mud

Abstract

The invention discloses a method for recovering scattered platinum group metals from the liquid after gold extraction from silver anode mud. The method targets a small amount of platinum, palladium, rhodium, selenium and tellurium in the liquid after gold extraction from the silver anode mud, separates platinum and selenium from palladium, rhodium and tellurium by hydrolysis, wherein the platinum and selenium enter into the hydrolysis solution, and the palladium, rhodium and tellurium enter into the hydrolysis residue. Subsequently, the hydrolysis solution and the hydrolysis residue are recovered by adopting different treatment processes, and finally the yields of platinum, palladium, rhodium, selenium and tellurium are all greater than 90%, thereby realizing the comprehensive recovery of all scattered platinum group metals such as rhodium, selenium, tellurium, platinum and palladium in the silver anode mud. The method has a short process flow, simple required equipment, convenient operation and strong industrial applicability.

Description

Technical Field

[0001] The invention relates to a method for recovering scattered platinum group metals from liquid after gold extraction from silver anode mud, relates to the technical field of metallurgical process, and can be used for various materials with low platinum group metal content, and can also be used for other materials containing platinum, palladium, rhodium, selenium and tellurium. Background Art

[0002] The process of processing copper anode mud in large smelters at home and abroad is mainly based on pyrometallurgy, that is, the copper anode mud produced by copper electrolysis of cathode copper is then leached under pressure → smelting and blowing in a Kaldo furnace → electrolysis of gold-silver alloy plates to produce silver anode mud and silver powder, of which silver powder is used for ingot casting and silver anode mud is used for gold extraction. After the silver anode mud is chlorinated with aqueous solution or dissolved in aqua regia to make liquid, base metals, gold and platinum group metals enter the solution, and gold powder ingots are produced after being reduced twice with sodium bisulfite or sulfur dioxide by controlling the potential. After the second gold extraction, the liquid contains valuable metals selenium, tellurium and rhodium in addition to platinum and palladium.

[0003] Selenium, tellurium and rhodium are rare metals with low content in the earth's crust. They are widely used in the military, pharmaceutical, semiconductor, battery and catalyst industries. It is of great significance to recover the scattered platinum group metals in the liquid after gold extraction from silver anode mud. Therefore, it is necessary to invent a process for recovering platinum, palladium, rhodium, selenium and tellurium from the liquid after gold extraction from silver anode mud. Summary of the invention

[0004] Aiming at the small amount of platinum, palladium, rhodium, selenium and tellurium in the liquid after gold extraction from silver anode mud, in order to provide a suitable separation and purification process, the present invention proposes a method for recovering scattered platinum group metals from the liquid after gold extraction from silver anode mud based on the study of the physical and chemical properties of platinum, palladium, rhodium, selenium and tellurium compounds.

[0005] The technical solution adopted by the present invention is: platinum, selenium and palladium, rhodium and tellurium are separated by hydrolysis, platinum and selenium enter the solution, palladium, rhodium and tellurium enter the hydrolysis residue; ammonium chloride is added to the hydrolysis solution to separate platinum and selenium, platinum is precipitated as crude platinum ammonium salt, and the crude platinum ammonium salt is dissolved in aqua regia-ammonium chloride repeated precipitation method to produce 99.99% sponge platinum. Selenium enters the mother liquor of precipitated platinum, is replaced in the residue, and the appropriate acidity is adjusted after alkali dissolution, and sodium pyrosulfite is added to produce crude selenium with a grade greater than 95%; chlorine gas is passed through the hydrolysis residue after hydrochloric acid dissolution, and ammonia water is added to control the end point pH=8~9, palladium is left in the complexing solution, and after multiple complexing-acidification, calcination and hydrogen reduction are carried out to produce 99.99% sponge palladium. After the complex slag is chlorinated with aqueous solution, the pH is adjusted to 1.5 with liquid alkali, and then oxalic acid is added to precipitate copper and nickel. After filtering, solid sodium nitrite is added to the solution, and ammonium chloride is added to the sodium nitrite complex solution for precipitation. The rhodium ammonium salt is reduced with formic acid to produce rhodium black with a grade greater than 99%. The complex slag is reduced with acid leaching of sodium D-isoascorbate to obtain tellurium powder with a grade greater than 99%.

[0006] The specific processing technology of the present invention is as follows:

[0007] 1) Extraction of platinum: concentrate the liquid after secondary reduction of silver anode mud to drive out acid to a small volume. When the volume reaches one-fourth of the original volume, add water to dilute it by one half, introduce chlorine gas, and wait until the potential reaches above 600mV, then add 30%~40% sodium hydroxide solution by mass to start hydrolysis, and adjust the solution pH to 8~9; after the pH value stabilizes, cool and filter, wash the hydrolysis residue with deionized water, and combine the washing liquid and the hydrolysis liquid; concentrate the hydrolysis liquid to 1 / 2 of the original volume, add hydrochloric acid to control the pH between 0~1, and then add saturated ammonium chloride solution to precipitate platinum. After no precipitation is produced, start filtering to obtain precipitated platinum The mother liquor and the filter residue are calcined to obtain a crude platinum ammonium salt. The precipitated platinum mother liquor is replaced with zinc powder to obtain a selenium replacement residue for extracting selenium; aqua regia is slowly added to the crude platinum ammonium salt in portions according to a liquid-to-solid ratio of 4-5:1 until it is completely dissolved and then filtered; the aqua regia solution is heated to drive out the nitrate until it becomes viscous, and hydrochloric acid with an acidity of 2-6 mol / L is added to drive out the acid again, and the process is repeated three times, followed by adding hydrochloric acid with a mass fraction of 0.2%-1%, and then pouring into a saturated ammonium chloride solution to precipitate platinum, boiling after no precipitation is produced, cooling and filtering to obtain ammonium chloroplatinate, and then calcining at 400°C-580°C for 12 hours to obtain sponge platinum with a purity greater than 99.95%;

[0008] 2) Extraction of selenium: add a sodium hydroxide solution with a mass fraction of 20% to 40% to the selenium replacement residue obtained in step 1), control the liquid-to-solid ratio to be 3 to 4:1, and the reaction temperature to be 85 to 95°C. After reacting for 3 to 5 hours, filter; after the filtrate is heated to 60°C, add a sulfuric acid solution with a mass fraction of 20% to 25% to neutralize, adjust the pH to 5 to 6, react for 2 to 3 hours until the pH stabilizes between 5 and 6, filter, neutralize the residue to extract tellurium, continue to add 93% concentrated sulfuric acid to the neutralized solution to control the acidity to 1 to 1.5 mol / L, then add sodium metabisulfite to reduce to obtain crude selenium with a grade greater than 95%, and after the reduced solution is tested for selenium-free by thiourea, transfer to water treatment to meet discharge standards;

[0009] 3) Extraction of palladium: dissolve the hydrolyzed residue of the silver anode mud gold secondary reduction solution after washing in step 1) with 2-4 mol / L hydrochloric acid, control the palladium concentration of the solution to 20-40 g / L, add ammonia water for complexation, control the pH to 8-9, stabilize for 0.5-1h and then filter, wash the complex residue with water, combine the washing liquid with the complex liquid, and concentrate to control the palladium concentration to 20-40 g / L; add hydrochloric acid to the complex liquid to acidify to pH 1-1.5, filter and Obtaining crude palladium ammonium salt; adding ammonia water to the crude palladium ammonium salt for re-complexation, controlling the pH value to 8-9, stabilizing for 0.5-1h and then filtering, adding hydrochloric acid to the secondary complexation solution to acidify to pH value to 1-1.5 and then filtering to obtain crude palladium ammonium salt, calcining the crude palladium ammonium salt from room temperature to 500° C. for 12h, cutting off the power in the calcination furnace and cooling to 200° C., then entering the electric furnace at a constant temperature of 500° C. and introducing hydrogen for 1h for reduction to obtain sponge palladium with a grade greater than 99.95%;

[0010] 4) Extraction of rhodium, extraction step 3) Add 2~4mol / L hydrochloric acid to the complex residue produced by the palladium front stage, heat to 80~85℃, pass chlorine gas to dissolve and filter, concentrate the filtrate to a small volume to drive out the acid, dilute and control the rhodium concentration to 20~30g / L after driving out the acid, then add 20~40% sodium hydroxide by mass to adjust the pH to 0.5~1.5, add 1~2g oxalic acid per liter of solution to precipitate copper and nickel, filter, and add 1.2~1.5 times the theoretical amount of solid to the filtrate Sodium nitrite, after stirring evenly, continue to add 20-40% sodium hydroxide by mass to adjust the pH to 7-8, boil for 30 minutes, cool and filter to obtain complex residue and complex liquid, the complex residue is used to extract tellurium, solid ammonium chloride equivalent to 2-4 times of the rhodium complex is added to the complex liquid to generate a precipitate and filter, the precipitate is reduced with formic acid to produce rhodium black with a grade greater than 99%, the rhodium precipitation mother liquor is adjusted to pH 2-3 with 2-4 mol / L hydrochloric acid, replaced with zinc powder and discharged;

[0011] 5) Extraction of tellurium: dissolve the complex residue obtained in step 4) with 2-4 mol / L hydrochloric acid, concentrate it into a small volume to remove the acid, add water after the volume drops to 1 / 5 of the original volume, control the tellurium concentration to 35-50 g / L, add 20-40% by mass sodium hydroxide solution to adjust the pH to 0.5-1.5, raise the temperature to 80-85°C, add 1.1-1.3 times the theoretical amount of sodium D-isoascorbate, maintain the temperature at 80-85°C for half an hour, cool and filter, and produce tellurium powder with a grade greater than 99%.

[0012] In the above steps 1) to 5), the main elements and contents in the liquid after the secondary reduction of the silver anode mud are rhodium: 0.1-5 g / L, tellurium: 0.5-2 g / L, palladium: 2-5 g / L, platinum: 1-5 g / L; selenium: 1-3 g / L. The chlorine gas can be replaced by hydrogen peroxide or chlorate, the zinc powder can be replaced by iron powder or magnesium powder, the sodium pyrosulfite can be replaced by sulfur dioxide or hydroxylamine hydrochloride, and the sodium D-isoascorbate can be replaced by sodium vitamin C.

[0013] The beneficial effects of the present invention are: 1. Comprehensive recovery of all scattered platinum group metals such as rhodium, selenium, tellurium, platinum and palladium in silver anode mud is achieved, the process flow is short, the required equipment is simple, the operation is convenient, and the industrial applicability is strong; 2. The yields of platinum, palladium, rhodium, selenium and tellurium are all greater than 90%, and can be used for various solutions containing rare and precious metals such as platinum and palladium. BRIEF DESCRIPTION OF THE DRAWINGS

[0014] Figure 1 It is a process flow chart of the present invention. DETAILED DESCRIPTION

[0015] The present invention will be further explained below in conjunction with the accompanying drawings.

[0016] Example 1

[0017] The silver anode mud gold secondary reduction liquid (as shown in Table 1) is concentrated to a small volume to drive out the acid. When the volume reaches one-fourth of the original volume, it is diluted by water by one-half, and then a 40% mass fraction of sodium hydroxide solution is added to start hydrolysis, and the pH value of the solution is adjusted to 8. Half an hour after the pH value stabilizes, it is cooled and filtered, and the hydrolysis residue is washed with deionized water, and the washing liquid and the hydrolysis liquid are combined. The hydrolysis liquid is concentrated, and hydrochloric acid is added to control the pH between 0 and 1. Then a saturated ammonium chloride solution is added to precipitate platinum, and the crude platinum ammonium salt is obtained by filtration and calcination. The precipitated platinum mother liquor is replaced with zinc powder to obtain selenium replacement residue for selenium extraction.

[0018] Table 1 Liquid composition after secondary reduction of silver anode mud gold in Example 1 (g / L)

[0019]

[0020] 1) Extraction of platinum. Slowly add aqua regia to the crude platinum ammonium salt in portions according to a volume liquid-solid ratio of 4:1 until it is completely dissolved and then filtered. Heat the aqua regia solution to remove nitrate until it becomes viscous, add HCl with an acidity of 2 mol / L to remove acid again, repeat three times, add 1% HCl by mass, and pour in saturated NH 4 Cl solution to precipitate platinum, take a small amount of the solution and add solid ammonium chloride to detect until no precipitation is produced, then stop boiling, cool to below 40°C, filter to obtain ammonium chloroplatinate, and then calcine at 580°C for 12h to obtain sponge platinum with a purity of 99.95%;

[0021] 2) Extraction of selenium. Add 20% sodium hydroxide solution to the selenium replacement slag, control the liquid-to-solid ratio at 3:1, and the reaction temperature at 85°C. After reacting for 3 hours, filter. Heat the leaching solution to 60°C, add 20% sulfuric acid solution by mass to neutralize, adjust the pH to 5, react for 2 hours, and stabilize the pH at 5. Filter. Add 93% concentrated sulfuric acid to the neutralized solution to control the acidity to 1 mol / L, add sodium pyrosulfite to reduce, and obtain crude selenium with a grade of 98.5%. After the reduced solution is tested for selenium-free by thiourea, it is transferred to water treatment to meet the discharge standards.

[0022] 3) Extraction of palladium. After the secondary reduction of silver anode mud, the slag is washed with water and dissolved with 2mol / L hydrochloric acid. The palladium concentration of the solution is controlled at 20g / L. Ammonia water is added for complexation and the pH is controlled to 8. After stabilization for 30 minutes, it is filtered. The insoluble residue of ammonia complex is washed with water. After the washing liquid is combined with the ammonia complex liquid, the palladium concentration is controlled to 20g / L. The complex slag is used to extract rhodium and tellurium. Hydrochloric acid is added to the complex liquid to acidify it to pH=1, and crude palladium ammonium salt is obtained after filtration. Ammonia water is added to the crude palladium ammonium salt, pH=8, impurities are filtered, hydrochloric acid is added to the complex liquid to acidify it to pH=1, and the palladium ammonium salt is filtered, calcined, and hydrogen reduced to obtain a sponge palladium with a grade greater than 99.95%;

[0023] 4) Extraction of rhodium. Add 2mol / L hydrochloric acid to the complex residue produced in the front stage of palladium extraction, heat to 80℃, pass chlorine gas to dissolve and filter, concentrate the filtrate to a small volume to drive out the acid, dilute to the original volume after driving out the acid, control the rhodium concentration at 20g / l, then add 20% sodium hydroxide by mass to adjust the pH to 1.5, add 1-2g oxalic acid per liter of solution to precipitate copper and nickel, filter, add 1.5 times the theoretical amount of solid sodium nitrite to the filtrate, stir evenly and continue to add 20% sodium hydroxide by mass to adjust the pH to 7, boil for 30 minutes, cool and filter, and use the complex residue to extract tellurium. Add 3 times the amount of solid ammonium chloride equivalent to the rhodium complex to the complex solution to generate a precipitate that is filtered, and the precipitate is reduced with formic acid to produce rhodium black with a grade greater than 99%;

[0024] 5) Extraction of tellurium. Dissolve the complex slag with 4 mol / L hydrochloric acid, concentrate it into a small volume to remove the acid, add water to the original volume after the volume drops to 1 / 5 of the original volume, control the tellurium concentration to 35 g / L, add 20% sodium hydroxide by mass to adjust the pH to 1.5, heat to 80°C, add 1.1 times the theoretical amount of sodium D-isoascorbate, maintain the temperature at 80°C for half an hour, cool and filter, and produce tellurium powder with a grade greater than 99%. The yields of platinum, palladium, rhodium, selenium, and tellurium are all greater than 91.5%.

[0025] Example 2

[0026] The liquid after the secondary reduction of the silver anode mud gold (as shown in Table 2) is concentrated to a small volume to remove the acid. When the volume reaches one-fourth of the original volume, it is diluted by water by one-half, and then a 40% mass fraction of sodium hydroxide solution is added to start hydrolysis, and the pH value of the solution is adjusted to 9. Half an hour after the pH value stabilizes, it is cooled and filtered, and the hydrolysis residue is washed with deionized water, and the washing liquid and the hydrolysis liquid are combined. The hydrolysis liquid is concentrated, hydrochloric acid is added to control the pH at 1, and a saturated ammonium chloride solution is added to precipitate platinum. After filtering, it is calcined to obtain a crude platinum ammonium salt, and the precipitated platinum mother liquor is replaced with zinc powder to obtain a selenium replacement residue for extracting selenium.

[0027] Table 2 Liquid composition after secondary reduction of silver anode mud gold in Example 2 (g / L)

[0028]

[0029] 1) Extraction of platinum. Slowly add aqua regia to the crude platinum ammonium salt in portions according to a volume liquid-solid ratio of 4:1 until it is completely dissolved and then filtered. Heat the aqua regia solution to remove nitrate until it becomes viscous, add HCl with an acidity of 6 mol / L to remove acid again, repeat three times, add HCl with a mass fraction of 1%, and pour in saturated NH 4 Cl solution to precipitate platinum, remove a small amount of solution and add solid ammonium chloride to detect until no precipitation is produced, then stop boiling, cool to below 40°C, filter to obtain ammonium chloroplatinate, and then calcine at 580°C for 12h to obtain sponge platinum with a purity of 99.95%;

[0030] 2) Extraction of selenium. Add 20% of liquid alkali to the selenium replacement slag, control the liquid-to-solid ratio at 4:1 according to the mass ratio, and the reaction temperature is 95°C. After reacting for 5 hours, filter. After the leaching solution is heated to 60°C, add a 25% sulfuric acid solution by mass to neutralize it, adjust the pH to 6, and stabilize the pH at 6 after reacting for 3 hours. Filter, and continue to add 93% concentrated sulfuric acid to the neutralized liquid to control the acidity to 1.5 mol / L. Add sodium metabisulfite to reduce it to obtain crude selenium with a grade greater than 97.6%. After the reduced liquid is tested for selenium-free by thiourea, it is transferred to water treatment to meet the discharge standards;

[0031] 3) Extraction of palladium. After the secondary reduction of silver anode mud, the slag is washed with hydrolysis liquid and dissolved with 2mol / L hydrochloric acid. The palladium concentration of the solution is controlled at 25g / L. Ammonia water is added for complexation and the pH is controlled at 9. After stabilization for 30 minutes, it is filtered. The insoluble residue of ammonia complex is washed with water. After the washing liquid is combined with the ammonia complex liquid, the palladium concentration is controlled at 20g / L. The complex slag is used to extract rhodium and tellurium. Hydrochloric acid is added to the complex liquid to acidify it to pH=1.5, and crude palladium ammonium salt is obtained after filtration. Ammonia water is added to the crude palladium ammonium salt, pH=9, impurities are filtered, and hydrochloric acid is added to the complex liquid to acidify it to pH=1.5. The palladium ammonium salt is filtered, calcined, and hydrogen reduced to obtain a sponge palladium with a grade greater than 99.95%;

[0032] 4) Extraction of rhodium. Add 2mol / L hydrochloric acid to the complex residue produced in the front stage of palladium extraction, heat to 85℃, pass chlorine gas to dissolve and filter, concentrate the filtrate to a small volume to drive out the acid, dilute to the original volume after driving out the acid, control the rhodium concentration at 25g / L, and then add 20% by mass sodium hydroxide to adjust the pH to 1.5, add 1-2g of oxalic acid per liter of solution to precipitate copper and nickel, filter, and then add 1.5 times the theoretical amount of solid sodium nitrite, stir evenly and continue to add 20% by mass sodium hydroxide to adjust the pH to 8, boil for 30 minutes, cool and filter, and use the complex residue to extract tellurium. Add 3 times the amount of solid ammonium chloride equivalent to the rhodium complex to the complex solution to generate a precipitate and filter it. The precipitate is reduced with formic acid to produce rhodium black with a grade greater than 99%;

[0033] 5) Extraction of tellurium. Dissolve the complex slag with 3 mol / L hydrochloric acid, concentrate it into a small volume to remove the acid, add water to the original volume after the volume drops to 1 / 5 of the original volume, control the tellurium concentration to 45 g / L, add 20% sodium hydroxide by mass to adjust the pH to 1.5, heat to 85°C, add 1.3 times the theoretical amount of sodium D-isoascorbate, maintain the temperature at 85°C for half an hour, cool and filter, and produce tellurium powder with a grade greater than 99%. The yields of platinum, palladium, rhodium, selenium, and tellurium are all greater than 92.2%.

[0034] Example 3

[0035] The silver anode mud gold secondary reduction liquid (as shown in Table 3) is concentrated to a small volume to remove the acid. When the volume reaches one-fourth of the original volume, it is diluted by water by one-half, and then a 40% mass fraction of sodium hydroxide solution is added to start hydrolysis, and the pH value of the solution is adjusted to 8-9. Half an hour after the pH value stabilizes, it is cooled and filtered, and the hydrolysis residue is washed with deionized water, and the washing liquid and the hydrolysis liquid are combined. The hydrolysis liquid is concentrated, hydrochloric acid is added to control the pH at 0.5, and a saturated ammonium chloride solution is added to precipitate platinum. After filtering, it is calcined to obtain a crude platinum ammonium salt, and the precipitated platinum mother liquor is replaced with zinc powder to obtain a selenium replacement residue for extracting selenium.

[0036] Table 3 Composition of the liquid after secondary reduction of silver anode mud gold in Example 3 (g / L)

[0037]

[0038] 1) Extraction of platinum. Add aqua regia to the crude platinum ammonium salt in a volume liquid-solid ratio of 4:1 in portions until it is completely dissolved and then filter. Heat the aqua regia solution to remove nitrate until it becomes viscous, add 4 mol / L HCl to remove acid again, repeat three times, add 1% HCl by mass, and pour in saturated NH 4 Cl solution to precipitate platinum, remove a small amount of solution and add solid ammonium chloride to detect until no precipitation is produced, then stop boiling, cool to below 40°C, filter to obtain ammonium chloroplatinate, and then calcine at 580°C for 12h to obtain sponge platinum with a purity of 99.95%;

[0039] 2) Extraction of selenium. Add 20% of liquid alkali to the selenium replacement slag, control the liquid-to-solid ratio at 3:1 according to the mass ratio, and the reaction temperature is 90°C. After reacting for 4 hours, filter. After the leaching solution is heated to 60°C, add a 24% sulfuric acid solution by mass to neutralize it, adjust the pH to 5.5, and stabilize the pH at 5.5 after reacting for 2.5 hours. Filter, and extract tellurium from the neutralized slag. After neutralization, continue to add 93% concentrated sulfuric acid to the liquid to control the acidity to 1.3 mol / L. Add sodium metabisulfite to reduce it to obtain crude selenium with a grade greater than 96.9%. After the reduced liquid is tested for selenium-free by thiourea, it is transferred to water treatment to meet the discharge standards;

[0040] 3) Extraction of palladium. After the secondary reduction of silver anode mud, the slag is washed with hydrolysis liquid and dissolved with 2mol / L hydrochloric acid. The palladium concentration of the solution is controlled at 30g / L. Ammonia water is added for complexation and the pH is controlled at 8.5. After stabilization for 30 minutes, it is filtered. The insoluble residue of ammonia complex is washed with water. After the washing liquid is combined with the ammonia complex liquid, the palladium concentration is controlled at 20g / L. The complex slag is used to extract rhodium and tellurium. Hydrochloric acid is added to the complex liquid to acidify it to pH=1.2, and crude palladium ammonium salt is obtained after filtration. Ammonia water is added to the crude palladium ammonium salt, pH=8.5, impurities are filtered, hydrochloric acid is added to the complex liquid to acidify it to pH=1.2, and the palladium ammonium salt is filtered, calcined and hydrogen reduced to obtain sponge palladium with a grade greater than 99.95%;

[0041] 4) Extraction of rhodium. Add 2mol / L hydrochloric acid to the complex residue produced in the front stage of palladium extraction, heat to 83℃, pass chlorine gas to dissolve and filter, concentrate the filtrate to a small volume to drive out the acid, dilute to the original volume after driving out the acid, control the rhodium concentration at 30g / L, and then add 20% sodium hydroxide by mass to adjust the pH to 1.5, add 1-2g oxalic acid per liter of solution to precipitate copper and nickel, filter, and then add 1.5 times the theoretical amount of solid sodium nitrite, stir evenly and continue to add 20% sodium hydroxide by mass to adjust the pH to 7.5, boil for 30 minutes, cool and filter, and use the complex residue to extract tellurium. Add 3 times the amount of solid ammonium chloride equivalent to the rhodium complex to the complex solution to generate a precipitate and filter it. The precipitate is reduced with formic acid to produce rhodium black with a grade greater than 99%;

[0042] 5) Extraction of tellurium. Dissolve the complex slag with 4 mol / L hydrochloric acid, concentrate it into a small volume to remove the acid, add water to the original volume after the volume drops to 1 / 5 of the original volume, control the tellurium concentration to 50 g / L, add 20% sodium hydroxide by mass to adjust the pH to 1.5, heat to 83°C, add 1.3 times the theoretical amount of sodium D-isoascorbate, maintain the temperature at 83°C for half an hour, cool and filter, and produce tellurium powder with a grade greater than 99%. The yields of platinum, palladium, rhodium, selenium, and tellurium are all greater than 91%.

Claims

1. A method for recovering dispersed platinum group metals from the liquid after gold extraction from silver anode mud, characterized in that: The following steps are involved: 1) Extraction of platinum: concentrate the liquid after secondary reduction of silver anode mud to drive out acid to a small volume. When the volume reaches one-fourth of the original volume, add water to dilute it by half, introduce chlorine gas, and wait until the potential reaches above 600mV, then add 30%~40% sodium hydroxide solution by mass to start hydrolysis, and adjust the solution pH to 8~9; after the pH value stabilizes, cool and filter, wash the hydrolysis residue with deionized water, and combine the washing liquid and the hydrolysis liquid; concentrate the hydrolysis liquid to 1 / 2 of the original volume, add hydrochloric acid to control the pH between 0~1, and then add saturated ammonium chloride solution to precipitate platinum. After no precipitation is produced, start filtering, and filter to obtain the precipitate. Platinum mother liquor and filter residue, the filter residue is calcined to obtain crude platinum ammonium salt, the precipitated platinum mother liquor is replaced with zinc powder to obtain selenium replacement residue for extracting selenium; aqua regia is slowly added to the crude platinum ammonium salt in portions according to a liquid-to-solid ratio of 4-5:1 until it is completely dissolved and then filtered; the aqua regia solution is heated to drive out nitrate until it becomes viscous, and hydrochloric acid with an acidity of 2-6 mol / L is added to drive out the acid again, and the process is repeated three times, followed by adding hydrochloric acid with a mass fraction of 0.2%-1%, and then pouring into a saturated ammonium chloride solution to precipitate platinum, boiling after no precipitation is produced, cooling and filtering to obtain ammonium chloroplatinate, and then calcining at 400°C-580°C for 12 hours to obtain sponge platinum with a purity greater than 99.95%; 2) Extraction of selenium: add a sodium hydroxide solution with a mass fraction of 20% to 40% to the selenium replacement residue obtained in step 1), control the liquid-to-solid ratio to be 3 to 4:1, and the reaction temperature to be 85 to 95°C. After reacting for 3 to 5 hours, filter; after the filtrate is heated to 60°C, add a sulfuric acid solution with a mass fraction of 20% to 25% to neutralize, adjust the pH to 5 to 6, react for 2 to 3 hours until the pH stabilizes between 5 and 6, filter, neutralize the residue to extract tellurium, continue to add 93% concentrated sulfuric acid to the neutralized solution to control the acidity to 1 to 1.5 mol / L, then add sodium metabisulfite to reduce to obtain crude selenium with a grade greater than 95%, and after the reduced solution is tested for selenium-free by thiourea, transfer to water treatment to meet discharge standards; 3) Extraction of palladium: dissolve the hydrolyzed residue of the silver anode mud gold secondary reduction solution after washing in step 1) with 2-4 mol / L hydrochloric acid, control the palladium concentration of the solution to 20-40 g / L, add ammonia water for complexation, control the pH to 8-9, stabilize for 0.5-1h and then filter, wash the complex residue with water, combine the washing liquid with the complex liquid, and concentrate to control the palladium concentration to 20-40 g / L; add hydrochloric acid to the complex liquid to acidify to pH 1-1.5, filter and Obtaining crude palladium ammonium salt; adding ammonia water to the crude palladium ammonium salt for re-complexation, controlling the pH value to 8-9, stabilizing for 0.5-1h and then filtering, adding hydrochloric acid to the secondary complexation solution to acidify to pH value to 1-1.5 and then filtering to obtain crude palladium ammonium salt, calcining the crude palladium ammonium salt from room temperature to 500° C. for 12h, cutting off the power in the calcination furnace and cooling to 200° C., then entering the electric furnace at a constant temperature of 500° C. and introducing hydrogen for 1h for reduction to obtain sponge palladium with a grade greater than 99.95%; 4) Extraction of rhodium, extraction step 3) Add 2~4mol / L hydrochloric acid to the complex residue produced by the palladium front stage, heat to 80~85℃, pass chlorine gas to dissolve and filter, concentrate the filtrate to a small volume to drive out the acid, dilute and control the rhodium concentration to 20~30g / L after driving out the acid, then add 20~40% sodium hydroxide by mass to adjust the pH to 0.5~1.5, add 1~2g oxalic acid per liter of solution to precipitate copper and nickel, filter, and add 1.2~1.5 times the theoretical amount of solid to the filtrate Sodium nitrite, after stirring evenly, continue to add 20-40% sodium hydroxide by mass to adjust the pH to 7-8, boil for 30 minutes, cool and filter to obtain complex residue and complex liquid, the complex residue is used to extract tellurium, solid ammonium chloride equivalent to 2-4 times of the rhodium complex is added to the complex liquid to generate a precipitate and filter, the precipitate is reduced with formic acid to produce rhodium black with a grade greater than 99%, the rhodium precipitation mother liquor is adjusted to pH 2-3 with 2-4 mol / L hydrochloric acid, replaced with zinc powder and discharged; 5) Extraction of tellurium: dissolve the complex residue obtained in step 4) with 2-4 mol / L hydrochloric acid, concentrate it into a small volume to remove the acid, add water after the volume drops to 1 / 5 of the original volume, control the tellurium concentration to 35-50 g / L, add 20-40% by mass sodium hydroxide solution to adjust the pH to 0.5-1.5, raise the temperature to 80-85°C, add 1.1-1.3 times the theoretical amount of sodium D-isoascorbate, maintain the temperature at 80-85°C for half an hour, cool and filter, and produce tellurium powder with a grade greater than 99%.

2. A method for recovering dispersed platinum group metals from the liquid after gold extraction from silver anode mud as claimed in claim 1, characterized in that: In step 1), the main elements and contents in the liquid after the secondary reduction of the silver anode mud are rhodium: 0.1-5 g / L, tellurium: 0.5-2 g / L, palladium: 2-5 g / L, platinum: 1-5 g / L; Selenium: 1~3g / L.

3. A method for recovering dispersed platinum group metals from the liquid after gold extraction from silver anode mud as claimed in claim 1, characterized in that: In step 1), the chlorine gas may be replaced by hydrogen peroxide or chlorate.

4. A method for recovering dispersed platinum group metals from the liquid after gold extraction from silver anode mud as claimed in claim 1, characterized in that: In step 1), zinc powder can be replaced by iron powder or magnesium powder.

5. A method for recovering dispersed platinum group metals from the liquid after gold extraction from silver anode mud as claimed in claim 1, characterized in that: In step 2), sodium metabisulfite can be replaced by sulfur dioxide or hydroxylamine hydrochloride.

6. A method for recovering dispersed platinum group metals from the liquid after gold extraction from silver anode mud as claimed in claim 1, characterized in that: In step 5), sodium D-isoascorbate may be replaced by sodium vitamin C.

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