A comprehensive treatment method of platinum-palladium slag

Abstract

The application discloses a comprehensive treatment method of platinum-palladium slag and relates to the technical field of nonferrous metal smelting. After sulfuric acid roasting of the platinum-palladium slag, flue gas is introduced into a selenium absorption tower; the roasted slag is hydrolyzed, and tellurium is recovered by reduction of a dissolved solution; after the tellurium is precipitated, the dissolved solution is introduced into a copper recovery procedure; gold is recovered by chlorination and dissolution in the dissolved slag, and the gold-recovered slag is introduced into silver recovery; gold powder is obtained by reduction of the gold-recovered solution; platinum and palladium are precipitated from the gold-reduced solution; ammonia is added to the platinum and palladium precipitate to separate platinum and palladium; the separated platinum and palladium is refined to recover platinum and palladium; platinum and palladium are reduced from the platinum and palladium precipitated solution to obtain platinum and palladium precipitate and platinum and palladium reduced solution, and the platinum and palladium precipitate is returned to the roasting procedure; tellurium is reduced from the platinum and palladium reduced solution to obtain tellurium powder and tellurium reduced solution, and the tellurium reduced solution is introduced into a wastewater treatment procedure. The application has the beneficial effects that selenium and tellurium, copper, gold, silver, platinum and palladium are stepwise separated and recovered, the recovery rate is high, the wastewater treatment difficulty is small, and environmental pollution is reduced.

Description

Technical Field

[0001] This invention relates to the field of non-ferrous metal smelting technology, specifically to a comprehensive treatment method for platinum-palladium slag. Background Technology

[0002] Platinum-palladium slag is a byproduct of non-ferrous metal smelting, primarily derived from the treatment of copper anode slime. Copper anode slime treatment processes are mainly divided into rotary kiln + hydrometallurgical processes and Kaldor furnace processes. The platinum-palladium slag enriched by the Kaldor furnace process is derived from silver electrolytic anode slime, with a relatively simple composition, high precious metal content, and a relatively simple recovery process. In contrast, the platinum-palladium slag produced by the rotary kiln + hydrometallurgical process is enriched in the gold reduction solution. Besides gold, silver, platinum, and palladium, this slag also contains copper, selenium, and tellurium, which also have some recovery value. However, the gold, platinum, and palladium grades are relatively low, making purification difficult. Current treatment processes focus only on the purification of platinum and palladium, weakening the efficient recovery of other valuable metals. Some technologies are even ill-suited for handling complex materials, especially those with high base metal content, particularly selenium and tellurium. Therefore, a comprehensive treatment method for platinum-palladium slag is urgently needed. Summary of the Invention

[0003] The purpose of this invention is to overcome the shortcomings of the prior art and provide a comprehensive treatment method for platinum-palladium slag, which achieves stepwise separation and recovery of selenium, tellurium, copper, gold, silver, platinum and palladium, with high recovery rate, low content of metal and non-metal element pollutants in wastewater, low difficulty in wastewater treatment, and the wastewater can be reused in the recycling system after simple treatment, thereby reducing environmental pollution.

[0004] The objective of this invention is achieved through the following technical measures: a comprehensive treatment method for platinum-palladium slag, comprising the following steps:

[0005] Step 1: The platinum-palladium slag is sulfated and roasted to obtain roasted slag and flue gas. The flue gas is then passed into a selenium absorption tower to recover selenium.

[0006] Step 2: Hydrolyze the roasted slag to obtain a solution and a residue. Use the solution to reduce and recover tellurium. After tellurium precipitation, the solution enters the copper recovery process to recover copper.

[0007] Step 3: Add hydrochloric acid, sulfuric acid and the first oxidant to the dissolving residue for chlorination and dissolution to obtain gold separation solution and gold separation residue. The gold separation residue enters the silver recovery process to recover silver.

[0008] Step 4: Gold powder and gold reduction solution are obtained by controlling the potential reduction of the gold solution.

[0009] Step 5: Add ammonium chloride and a second oxidant to the gold reduction solution to precipitate platinum and palladium, obtain platinum and palladium precipitate and platinum and palladium precipitation solution, add ammonia to the platinum and palladium precipitate to separate platinum and palladium, and refine the separated platinum and palladium to recover platinum and palladium.

[0010] Step 6: Add the first reducing agent to the solution after platinum and palladium precipitation to reduce platinum and palladium, and obtain platinum and palladium precipitate residue and platinum and palladium reduced solution. The platinum and palladium precipitate residue is returned to step 1.

[0011] Step 7: Add a second reducing agent to the platinum-palladium reduction solution to reduce tellurium, and obtain tellurium powder and tellurium reduction solution. The tellurium reduction solution enters the wastewater treatment process.

[0012] In some embodiments, the liquid-to-solid ratio of sulfuric acid to platinum-palladium slag in step 1 is (0.4-0.7):1, and the calcination temperature is 500℃-550℃.

[0013] In some embodiments, during the hydrolysis of the roasted residue in step 2, the liquid-to-solid ratio is (3-5):1, and the temperature is 70-90℃.

[0014] In some embodiments, copper powder is used in step 2 to reduce the solution and recover tellurium.

[0015] In some embodiments, the reaction temperature in step 3 is 75-95°C, the liquid-to-solid ratio is (4-6):1, and the first oxidant is at least one of sodium chlorate, potassium chlorate, and chlorine.

[0016] In some embodiments, the molar amount of the first oxidant added in step 3 satisfies the following ratio: total molar amount of gold + platinum + palladium : molar amount of the first oxidant = 1 : (3-5).

[0017] In some embodiments, the reduction endpoint potential of the gold separation solution in step 4 is 500-550mV, the reaction temperature is 40-60℃, and at least one of sodium sulfite and sulfur dioxide is used.

[0018] In some embodiments, the second oxidant in step 5 is at least one of sodium chlorate, potassium chlorate, and chlorine.

[0019] In some embodiments, the reduction endpoint potential in step 6 is 380-420mV, the reaction temperature is 50-80℃, and the first reducing agent is at least one of copper powder and zinc powder.

[0020] In some embodiments, the reaction temperature in step 7 is 60-90°C, and the second reducing agent is at least one of sulfur dioxide and sodium sulfite.

[0021] Compared with existing technologies, the beneficial effects of this invention are as follows: This invention achieves the stepwise separation and recovery of selenium, tellurium, copper, gold, silver, platinum, and palladium through comprehensive treatment of platinum-palladium slag. The grade of crude selenium reaches over 96%, the grade of copper telluride (copper tellurium compound and copper) is greater than 80%, the grade of gold powder reaches over 98%, and the purity of platinum and palladium is also over 99.95 wt%, meeting the SM-Pd 99.95 and SM-Pt 99.95 standards. Gold, silver, selenium, tellurium, copper, platinum, and palladium are all effectively recovered, improving the product recovery rate.

[0022] This invention involves direct sulfation roasting of platinum-palladium slag, followed by dissolution with water. This method is cost-effective, uses water as a solvent, is economical and environmentally friendly, and reduces the difficulty of wastewater treatment. Chlorination dissolution is performed using hydrochloric acid, sulfuric acid, and a primary oxidant, allowing lead in the dissolved slag to enter the gold separation slag in the form of lead sulfate. This reduces the lead content in the gold separation solution and improves the subsequent recovery rates of gold, platinum, and palladium.

[0023] This invention involves further processing the platinum-palladium precipitate after gold reduction, reducing the resulting platinum-palladium slag, and returning it to the sulfation roasting process. This improves the recovery rate of platinum and palladium while reducing the platinum and palladium content in the wastewater, thus simplifying wastewater treatment. Furthermore, the invention further recovers tellurium powder from the platinum-palladium reduction solution, increasing the tellurium recovery rate and further reducing wastewater treatment difficulty. The wastewater can be reused in the recycling system after simple treatment, reducing environmental pollution and achieving sustainable resource utilization.

[0024] The present invention will now be described in detail with reference to the accompanying drawings and specific embodiments. Attached Figure Description

[0025] Figure 1 This is a process flow diagram of the present invention. Detailed Implementation

[0026] like Figure 1 As shown, a comprehensive treatment method for platinum-palladium slag includes the following steps:

[0027] Step 1: Sulfuric acid and platinum-palladium slag are pulped at a liquid-to-solid ratio of (0.4-0.7):1. The pulped platinum-palladium slag is then subjected to sulfation roasting at a temperature of 500℃-550℃ to obtain roasted slag and flue gas. The flue gas is then passed into a selenium absorption tower to recover selenium. Preferably, the mass percentage of sulfuric acid is 90%-98%, the pulping time is 2-6 hours, and an electric furnace is preferred for the roasting equipment. During the roasting process, selenium volatilizes in the form of selenium dioxide and enters the selenium absorption tower, where it reacts with water to form selenite. The sulfur dioxide produced during roasting reduces the selenite in the liquid phase to produce crude selenium, with a grade of over 96%. The exhaust gas discharged from the selenium absorption tower enters the exhaust gas absorption system, where residual sulfur dioxide is treated before being discharged in compliance with standards. The selenium content in the roasted slag is less than 0.2%.

[0028] Step 2: Hydrolyze water and roasted slag at a liquid-to-solid ratio of 3-5:1 at a temperature of 70-90℃ for 1-10 hours. Separate the solid and liquid components to obtain a dissolved solution and a dissolved residue. Dissolve soluble salts in the roasted slag with water, allowing copper and tellurium in the slag to enter the dissolved solution. After hydrolysis, the copper content in the dissolved residue is below 0.5wt%. Recover tellurium by reducing the dissolved solution with copper powder. Tellurium in the dissolved solution precipitates as copper telluride. Separate the solid and liquid components to obtain copper telluride and filtrate. The grade of copper telluride (copper telluride compound and copper) is greater than 80%. The filtrate is used for copper recovery. Preferably, the filtrate is used for copper electrolysis to generate copper from the copper sulfate filtrate.

[0029] Step 3: Add hydrochloric acid, sulfuric acid, and a first oxidant to the dissolved residue for chlorination dissolution. Control the dissolution temperature at 75-95℃, the chlorination dissolution time at 2-5 hours, the liquid-to-solid ratio at (4-6):1, and the mixed acid concentration of hydrochloric acid and sulfuric acid at 100-150 g / L, preferably 20-40 g / L. Preferably, the first oxidant is at least one of sodium chlorate, potassium chlorate, and chlorine. Preferably, the molar amount of the first oxidant added satisfies the following ratio: total molar amount of gold + platinum + palladium : molar amount of the first oxidant = 1:(3-5). Obtain the gold separation solution and the gold separation residue. Gold, platinum, and palladium in the dissolved residue enter the gold separation solution, while silver precipitates into the gold separation residue as silver chloride, and lead precipitates into the gold separation residue as insoluble lead sulfate. The main component of the gold separation residue is silver chloride, with a gold content below 50 g / t. The gold-separating slag is fed into the silver-separating process to recover silver. This invention does not specifically limit the silver-separating process, and can adopt processing processes commonly used by those skilled in the art, such as sodium sulfite silver separation process or ammonia silver separation process.

[0030] Step 4: The reduction endpoint potential of the gold separation solution is controlled at 500-550mV, the reaction temperature is 40℃-60℃, and the reduction time is 1-5h. At least one of sodium sulfite and sulfur dioxide is used as the reducing agent to obtain gold powder and a gold reduction solution. The gold in the gold separation solution is reduced to gold powder with a purity of over 98%. The platinum and palladium in the gold separation solution are added to the gold reduction solution.

[0031] Step 5: Add ammonium chloride and a second oxidizing agent to the gold reduction solution to precipitate platinum and palladium. The reaction temperature is 85℃-95℃. The reaction endpoint is reached when no red ammonium chloropalladate is produced when sodium chlorate is added dropwise. Preferably, the second oxidizing agent is at least one of sodium chlorate, potassium chlorate, and chlorine gas. Platinum and palladium precipitate and the solution after precipitation are obtained. Specifically, platinum and palladium precipitate as ammonium chloroplatinate and ammonium chloropalladate, respectively. Ammonia is added to the platinum and palladium precipitate. Ammonium chloropalladate reacts with ammonia to form dichlorotetraammine palladium, while ammonium chloroplatinate is insoluble in ammonia, thus achieving platinum and palladium separation. The separated platinum and palladium are then refined to recover platinum and palladium. This invention does not specifically limit the refining process of platinum and palladium; any refining process well known to those skilled in the art can be used. For example, the crude platinum powder produced after calcining ammonium chloroplatinate can be transferred to a platinum refining process to obtain sponge platinum. Acidification of dichlorotetraammine-palladium solution yields dichlorodiammine-palladium precipitate, which is then reduced with hydrazine hydrate to obtain sponge palladium.

[0032] Step 6: Add the first reducing agent to the platinum-palladium precipitation solution to reduce platinum and palladium. Control the reduction endpoint potential to be 380-420mV and the reaction temperature to be 50-80℃ to obtain platinum-palladium precipitate residue and platinum-palladium reduction solution. The platinum and palladium remaining in the platinum-palladium precipitation solution are reduced to form platinum-palladium precipitate residue, which is then further extracted. Preferably, the first reducing agent is at least one of copper powder and zinc powder. The platinum-palladium precipitate residue is returned to step 1 and mixed with the platinum-palladium slag for sulfation roasting.

[0033] Step 7: Add a second reducing agent to the platinum-palladium reduction solution to reduce tellurium. The reaction temperature is 60-90℃, and the reaction time is 1-4 hours. The residual tellurium in the platinum-palladium reduction solution is reduced, and tellurium is further recovered to obtain tellurium powder and tellurium reduction solution. The tellurium reduction solution enters the wastewater treatment process. Preferably, the second reducing agent is at least one of sulfur dioxide and sodium sulfite.

[0034] Example 1

[0035] Platinum-palladium slag produced by the copper anode mud rotary kiln + wet treatment process was selected as raw material. The element content in the platinum-palladium slag is shown in Table 1 below.

[0036] Table 1: Elemental Content of Platinum-Palladium Slag

[0037]

[0038] Note: Units marked with * in Table 1 are kg / t, and the rest are in kg / t.

[0039] Step 1: 98% concentrated sulfuric acid and 100 kg (dry weight) of platinum-palladium slag are slurried at a liquid-to-solid ratio of 0.4:1 for 2 hours. After slurrying, sulfation roasting is carried out in an electric furnace at 510℃. Selenium volatilizes as selenium dioxide and enters the water in the selenium absorption tower to form selenite. The sulfur dioxide produced during roasting reduces the selenite in the liquid phase, producing crude selenium with a grade of 98.76%. The waste gas enters the waste gas absorption system and is discharged after treatment to meet standards. The selenium content in the roasted slag is 0.05%.

[0040] Table 2: Crude Selenium Composition Table

[0041]

[0042] Note: Units marked with * in Table 2 are in g / t, and the rest are in %.

[0043] The main reactions that occur during selenium recovery are as follows:

[0044] Se(s)+2H2SO4(l)=SeO2(g)+2H2O(g)+2SO2(g);

[0045] SeO2(g) + H2O(l) = H2SeO3(l);

[0046] 2SO2(g)+H2SeO3(l)+H2O(l)=Se(s)+2H2SO4(l).

[0047] Step 2: Water and roasted slag are mixed at a liquid-to-solid ratio of 3:1, reacted at 80℃, and soaked in water for 2 hours. The mixture is then filtered to obtain a dissolved slag and a dissolved solution. Easily soluble copper sulfate, tellurite, and telluric acid in the roasted slag enter the dissolved solution. After roasting, the copper content in the dissolved slag is 0.45%. Copper powder is added to the dissolved solution to reduce and precipitate tellurium. The reaction temperature is 90℃, and the reaction time is 4.5 hours. Tellurium in the dissolved solution precipitates as copper telluride. Solid-liquid separation is performed to obtain copper telluride and a filtrate. The grade of copper telluride (copper telluride compound and copper) reaches 80.88%. The tellurium content in the filtrate is controlled at 0.0034 g / L, the selenium content is <0.0001 g / L, and the chloride ion content is 0.079 g / L. The filtrate is then sent to the copper electrolysis purification workshop for electrowinning copper recovery. The composition of copper telluride is shown in Table 3 below.

[0048] Table 3: Composition of Copper Telluride

[0049]

[0050] Note: Units marked with * in Table 3 are in g / t, and the rest are in %.

[0051] The main reactions that occur during the reduction and recovery of tellurium from copper powder are as follows:

[0052] H2TeO3(l)+4Cu(s)+2H2SO4(l)=Cu2Te(s)+2CuSO4(l)+3H2O(l);

[0053] H2TeO6(l)+6Cu(s)+5H2SO4(l)=CuTe(s)+5CuSO4(l)+6H2O(l).

[0054] Step 3: Add sodium chlorate, hydrochloric acid, and sulfuric acid to the dissolving residue for chlorination dissolution in water. The reaction temperature is 80℃, and the reaction time is 3.5h. The total molar ratio of gold, platinum, and palladium to NaClO3 is 1:3.5. The concentration of the mixed acid is 110g / L, the concentration of hydrochloric acid is 20g / L, and the liquid-to-solid ratio of the mixed acid to the dissolving residue is 4:1. The gold, platinum, and palladium ions generated in the reaction enter the gold separation solution. The gold content in the gold separation residue is 47g / t. Silver and lead in the dissolving residue precipitate into the gold separation residue as silver chloride and lead sulfate, respectively. The gold separation residue is sent to the silver recovery process for silver recovery. The composition of the gold separation residue is shown in Table 4 below.

[0055] Table 4: Composition of Gold-Dividing Slag

[0056]

[0057] Note: Units marked with * in Table 4 are in g / t, and the rest are in %.

[0058] The main reactions that occur during chlorination and dissolution are as follows:

[0059] NaClO3(l)+6HCl(l)=NaCl(l)+3Cl2(g)+3H2O(l);

[0060] 2Au(s)+2HCl(l)+3Cl2(g)=2HAuCl4(l);

[0061] Pt(s)+2HCl(l)+2Cl2(g)=H2PtCl6(l);

[0062] Pd(s)+Cl2(g)+2HCl(l)=H2PdCl4(l);

[0063] Ag2Se(s)+3Cl2(g)+3H2O(l)=2AgCl(s)+H2SeO3(l)+4HCl(l);

[0064] Ag2Te(s)+3Cl2(g)+3H2O(l)=2AgCl(s)+H2TeO3(l)+4HCl(l).

[0065] Step 4: The gold separation solution is controlled at the reduction endpoint potential of 530mV, the reaction temperature is 55℃, the reaction time is 2h, sodium sulfite is added to reduce and generate gold powder with a gold powder grade of 98.3%, which is then cast into a gold anode for electrolysis to obtain the finished gold ingot.

[0066] The main reactions that occur during the reduction of gold are as follows:

[0067] 2HAuCl4(l)+3Na2SO3(s)+3H2O(l)=2Au+6NaCl(l)+2HCl(l)+3H2SO4(l).

[0068] Step 5: Add sodium chlorate and ammonium chloride to the gold reduction solution. The reaction temperature is 90℃. Ammonium chloroplatinate and ammonium chloropalladate precipitate. The endpoint is reached when no red ammonium chloropalladate is produced when sodium chlorate is added dropwise. Slurry the ammonium chloroplatinate and ammonium chloropalladate with water, controlling the reaction temperature at 65℃. Add ammonia water. Ammonium chloroplatinate is insoluble, and ammonium chloropalladate complexes to form a dichlorotetraammine palladium solution. Filter to separate the platinum and palladium.

[0069] Ammonium chloroplatinate is calcined to form crude platinum powder, which is then transferred to platinum refining. Platinum refining includes dissolving the crude platinum powder in aqua regia, hydrolyzing it, precipitating platinum with ammonium chloride, and calcining it (sequentially calcining at 200℃ for 2 hours, calcining at 400℃ for 4 hours, and then calcining at 650℃ for 3 hours) to produce qualified sponge platinum with a platinum content of over 99.95 wt%, meeting the SM-Pt99.95 standard.

[0070] Acidification with dichlorotetraammine-palladium sulfide solution yields dichlorodiammine-palladium sulfide precipitate, which is then transferred to the palladium refining workshop. The dichlorodiammine-palladium sulfide is complexed with ammonia and reduced with hydrazine hydrate to produce qualified sponge palladium with a palladium content of over 99.95 wt%, meeting the SM-Pd99.95 standard. Zinc powder is added to the filtrate to recover residual precious metals, mainly platinum, palladium, and gold. The filtrate is then sent to the wastewater treatment system for further processing and reuse.

[0071] The main reactions that occur during the precipitation of platinum and palladium by ammonium chloride and sodium chlorate are as follows:

[0072] H2PtCl6(l)+2NH4Cl(s)=(NH4)2PtCl6(s)+2HCl(l);

[0073] NaClO3(s)+6HCl(l)=NaCl(l)+3Cl2(g)+3H2O(l);

[0074] H2PdCl4(l)+2NH4Cl(s)+Cl2=(NH4)2PdCl6(s)+2HCl(l).

[0075] The main reactions that occur during the platinum-palladium separation process using ammonia are as follows:

[0076] (NH4)2PdCl6(s)+4NH3·H2O(l)=Pd(NH3)4Cl2(l)+2NH4Cl(l)+4H2O(l)+Cl2(g).

[0077] Step 6: After precipitating platinum and palladium, control the reduction endpoint potential at 400mV, the reaction temperature at 70℃, and the reaction time at 3h. Add copper powder as a reducing agent to reduce and generate platinum and palladium precipitate residue. Return the platinum and palladium precipitate residue to Step 1.

[0078] The main reactions that occur during the reduction of platinum-palladium precipitate solution by copper powder are as follows:

[0079] H2PtCl6(l)+2Cu(s)=Pt(s)+2CuCl2(l)+2HCl(l);

[0080] H2PdCl4(l)+Cu(s)=Pd(s)+CuCl2(l)+2HCl(l).

[0081] Step 7: Sulfur dioxide is bubbled into the platinum-palladium reduction solution to reduce tellurium, producing tellurium powder with a purity of 94.75%. The reaction temperature is 80℃, and the reaction endpoint is reached when no black precipitate is formed after the addition of sodium sulfite. The tellurium reduction solution is sent to a wastewater treatment system to recover trace amounts of precious metals, and the wastewater is reused after treatment to meet standards. The composition of crude tellurium is shown in Table 5 below:

[0082] Table 5: Composition of Crude Tellurium

[0083]

[0084] Note: Units marked with * in Table 5 are in g / t, and the rest are in %.

[0085] The main reactions that occur in the solution after sulfur dioxide reduces platinum and palladium are as follows:

[0086] H2TeCl6(l)+2SO2(g)+4H2O(l)=Te(s)+2SO4 2- +10H + +6Cl - ;

[0087] H2PtCl6(l)+2SO2(g)+4H2O(l)=Pt(s)+2SO4 2- +10H + +6Cl - ;

[0088] H2PdCl4(l)+2SO2(g)+4H2O(l)=Pd(s)+2SO4 2- +10H + +4Cl - .

[0089] Example 2

[0090] Platinum-palladium slag with the same composition as in Example 1 was selected as raw material and subjected to comprehensive treatment. The specific treatment steps are as follows:

[0091] Step 1: 93% concentrated sulfuric acid and 100 kg (dry weight) of platinum-palladium slag are slurried at a liquid-to-solid ratio of 0.6:1. After slurrying for 4 hours, sulfation roasting is carried out in an electric furnace at 550℃. Selenium volatilizes in the form of selenium dioxide and enters the water in the selenium absorption tower to form selenite. The sulfur dioxide produced during roasting reduces the selenite in the liquid phase, producing crude selenium with a grade of 98.40%. The waste gas enters the waste gas absorption system and is discharged after treatment to meet standards. The selenium content in the roasted slag is 0.05%. The composition of the crude selenium is shown in Table 6 below.

[0092] Table 6: Composition of Crude Selenium

[0093]

[0094] Note: Units marked with * in Table 6 are in g / t, and the rest are in %.

[0095] Step 2: Water and roasted slag were mixed at a liquid-to-solid ratio of 4:1, reacted at 80℃, and soaked for 6 hours before filtration to obtain dissolved slag and dissolved solution. Easily soluble copper sulfate, tellurite, and telluric acid in the roasted slag entered the dissolved solution. After roasting, the copper content in the dissolved slag was 0.32%. Copper powder was added to the dissolved solution to reduce and precipitate tellurium. The reaction temperature was 90℃, and the reaction time was 4.5 hours. Tellurium in the dissolved solution precipitated as copper telluride. Solid-liquid separation was performed to obtain copper telluride and filtrate. The grade of copper telluride (copper telluride compound and copper) reached 85.43%. The tellurium content in the filtrate was controlled at 0.0034 g / L, selenium content <0.0001 g / L, and chloride ion content 0.079 g / L. The filtrate was sent to the copper electrolysis purification workshop for electrowinning copper recovery. The composition of copper telluride is shown in Table 7 below.

[0096] Table 7: Composition of Copper Telluride

[0097]

[0098] Note: Units marked with * in Table 7 are in g / t, and the rest are in %.

[0099] Step 3: Add sodium chlorate, hydrochloric acid, and sulfuric acid to the dissolving residue for chlorination dissolution in water. The reaction temperature is 90℃, and the reaction time is 5 hours. The total molar ratio of gold, platinum, and palladium to NaClO3 is 1:5. The concentration of the mixed acid is 130 g / L, and the concentration of hydrochloric acid is 35 g / L. The liquid-to-solid ratio of the mixed acid to the dissolving residue is 5:1. The gold, platinum, and palladium ions generated in the reaction enter the gold separation solution. The gold content in the gold separation residue is 34 g / t. Silver and lead in the dissolving residue precipitate into the gold separation residue as silver chloride and lead sulfate, respectively. The gold separation residue is sent to the silver recovery process for silver recovery. The composition of the gold separation residue is shown in Table 8 below.

[0100] Table 8: Composition of Gold-Dividing Slag

[0101]

[0102] Note: Units marked with * in Table 8 are in g / t, and the rest are in %.

[0103] Step 4: The gold separation solution is controlled at the reduction endpoint potential of 530mV, the reaction temperature is 40℃, the reaction time is 2h, sodium sulfite is added to reduce and generate crude gold powder with a grade of 98.7%, which is then cast into a gold anode for electrolysis to obtain the finished gold ingot.

[0104] Step 5: Add sodium chlorate and ammonium chloride to the gold reduction solution. The reaction temperature is 85℃. Ammonium chloroplatinate and ammonium chloropalladate precipitate. The endpoint is reached when no red ammonium chloropalladate is produced when sodium chlorate is added dropwise. Slurry the ammonium chloroplatinate and ammonium chloropalladate with water, controlling the reaction temperature at 65℃. Add ammonia water. Ammonium chloroplatinate is insoluble, and ammonium chloropalladate complexes to form a dichlorotetraammine-palladium solution. Filter to separate the platinum and palladium.

[0105] Ammonium chloroplatinate is calcined to form crude platinum powder, which is then transferred to platinum refining. Platinum refining includes dissolving the crude platinum powder in aqua regia, hydrolyzing it, precipitating it with ammonium chloride, and calcining it (sequentially calcining at 200℃ for 2 hours, calcining at 400℃ for 4 hours, and then calcining at 650℃ for 4 hours) to produce qualified sponge platinum with a platinum content of over 99.95 wt%, meeting the SM-Pt99.95 standard.

[0106] Acidification with dichlorotetraammine-palladium sulfide solution yields dichlorodiammine-palladium sulfide precipitate, which is then transferred to the palladium refining workshop. The dichlorodiammine-palladium sulfide is complexed with ammonia and reduced with hydrazine hydrate to produce qualified sponge palladium with a palladium content of over 99.95 wt%, meeting the SM-Pd99.95 standard. Zinc powder is added to the filtrate to recover residual precious metals, mainly platinum, palladium, and gold. The filtrate is then sent to the wastewater treatment system for further processing and reuse.

[0107] Step 6: After precipitating platinum and palladium, control the reduction endpoint potential at 400mV, the reaction temperature at 65℃, and the reaction time at 5h. Add copper powder as a reducing agent to reduce and generate platinum and palladium precipitate residue. Return the platinum and palladium precipitate residue to Step 1.

[0108] Step 7: Sulfur dioxide is bubbled into the platinum-palladium reduction solution to reduce tellurium, producing tellurium powder with a purity of 92.18%. The reaction temperature is 90℃, and the reaction endpoint is reached when no black precipitate is formed after the addition of sodium sulfite. The tellurium reduction solution is sent to a wastewater treatment system to recover trace amounts of precious metals, and the wastewater is reused after treatment to meet standards. The composition of crude tellurium is shown in Table 9 below:

[0109] Table 9: Composition of Crude Tellurium

[0110]

[0111] Note: Units marked with * in Table 9 are in g / t, and the rest are in %.

[0112] In this invention, the terms "one embodiment," "some embodiments," "example," "specific example," or "some examples," etc., refer to a specific feature, material, or characteristic described in connection with that embodiment or example, which is included in at least one embodiment or example of the invention. In this specification, the illustrative expressions of the above terms do not necessarily refer to the same embodiment or example. Furthermore, the specific features, materials, or characteristics described may be combined in any suitable manner in one or more embodiments or examples. Moreover, without contradiction, those skilled in the art can combine and integrate the different embodiments or examples described in this specification, as well as the features of different embodiments or examples.

[0113] Although the above embodiments have been shown and described, it is understood that the above embodiments are exemplary and should not be construed as limiting the present invention. Any changes, modifications, substitutions and variations made to the above embodiments by those skilled in the art are within the protection scope of the present invention.

Claims

1. A comprehensive treatment method for platinum-palladium slag, characterized in that, Includes the following steps: Step 1: The platinum-palladium slag is subjected to sulfation roasting to obtain roasting residue and flue gas. The flue gas is then passed into a selenium absorption tower to recover selenium. The platinum-palladium slag is obtained by rotary kiln + wet treatment process of copper anode mud. Step 2: Hydrolyze the roasted slag at a temperature of 70-90℃ to obtain a solution and a residue. Use copper powder to reduce the solution and recover tellurium. After tellurium precipitation, the solution enters the copper recovery process to recover copper. Step 3: Add hydrochloric acid, sulfuric acid and the first oxidant to the dissolving residue for chlorination and dissolution to obtain gold separation solution and gold separation residue. The gold separation residue enters the silver recovery process to recover silver. Step 4: Gold powder and gold reduction solution are obtained by controlling the potential reduction of the gold solution. Step 5: Add ammonium chloride and a second oxidant to the gold reduction solution to precipitate platinum and palladium, obtain platinum and palladium precipitate and platinum and palladium precipitation solution, add ammonia to the platinum and palladium precipitate to separate platinum and palladium, and refine the separated platinum and palladium to recover platinum and palladium. Step 6: Add the first reducing agent to the solution after platinum and palladium precipitation to reduce platinum and palladium, and obtain platinum and palladium precipitate residue and platinum and palladium reduced solution. The platinum and palladium precipitate residue is returned to step 1. Step 7: Add a second reducing agent to the platinum-palladium reduction solution to reduce tellurium, and obtain tellurium powder and tellurium reduction solution. The tellurium reduction solution enters the wastewater treatment process.

2. The comprehensive treatment method for platinum-palladium slag according to claim 1, characterized in that: In step 1, the liquid-solid ratio of sulfuric acid to platinum-palladium slag is (0.4-0.7):1, and the calcination temperature is 500℃-550℃.

3. The comprehensive treatment method for platinum-palladium slag according to claim 1, characterized in that: In step 2, when hydrolyzing the roasted residue, the liquid-to-solid ratio is (3-5):

1.

4. The comprehensive treatment method for platinum-palladium slag according to claim 1, characterized in that: In step 3, the reaction temperature is 75-95℃, the liquid-to-solid ratio is (4-6):1, and the first oxidant is at least one of sodium chlorate, potassium chlorate, and chlorine.

5. The comprehensive treatment method for platinum-palladium slag according to claim 1, characterized in that: In step 3, the molar amount of the first oxidant added satisfies the following ratio: total molar amount of gold + platinum + palladium : molar amount of the first oxidant = 1 : (3-5).

6. The comprehensive treatment method for platinum-palladium slag according to claim 1, characterized in that: In step 4, the reduction endpoint potential of the gold separation solution is 500-550mV, the reaction temperature is 40-60℃, and at least one of sodium sulfite and sulfur dioxide is used.

7. The comprehensive treatment method for platinum-palladium slag according to claim 1, characterized in that: In step 5, the second oxidant is at least one of sodium chlorate, potassium chlorate, and chlorine.

8. The comprehensive treatment method for platinum-palladium slag according to claim 1, characterized in that: In step 6, the reduction endpoint potential is 380-420mV, the reaction temperature is 50-80℃, and the first reducing agent is at least one of copper powder and zinc powder.

9. The comprehensive treatment method for platinum-palladium slag according to claim 1, characterized in that: In step 7, the reaction temperature is 60-90℃, and the second reducing agent is at least one of sulfur dioxide and sodium sulfite.

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