CN114807619B - Method for recovering rare and precious metals in selenium steaming slag in step manner - Google Patents
Method for recovering rare and precious metals in selenium steaming slag in step manner Download PDFInfo
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Abstract
The invention discloses a method for recovering rare and precious metals in selenium steaming slag in a gradient manner, which comprises the steps of alkaline oxidizing roasting, water leaching and separating selenium and tellurium, weak acid separating copper, chlorination separating gold, sodium sulfite precipitating gold, zinc powder replacing platinum and palladium, and hydrazine hydrate precipitating silver. According to the invention, the selenium, tellurium, copper, gold, silver and the like in the selenium steaming slag can be destroyed to form intercross compounds and inclusion structures through alkaline oxidizing roasting destruction, the step selective recovery of rare and noble metals is realized, the comprehensive recovery rate of the rare and noble metals is high, the process is simple and convenient, the production cost is low, the product obtained after separation of each process can be grafted to the original rare and noble metal treatment process of an enterprise, the defect of the conventional selenium steaming slag treatment process is made up, and the method is easy to popularize and apply in a large scale.
Description
Technical Field
The invention belongs to the field of metallurgy, and particularly relates to a method for recovering rare and precious metals in selenium steaming slag in a cascade manner.
Background
The rare and precious metal elements have excellent physical and chemical properties, electrical properties, high catalytic activity, strong coordination capacity and the like, so the rare and precious metal elements are widely applied in the professional fields of aerospace materials, information technologies, laser technologies, automation technologies, energy technologies, precious metal catalysts, new materials and the like, and are in an important position in the development of high and new technologies. With the development of science and technology, the application field and the application of the medicine are expanded, and the function is more and more important. For example: selenium, tellurium and other rare and dispersive elements are known as vitamins of modern industry, national defense and advanced technology, and are support materials of modern high and new technology materials. Meanwhile, precious metals such as gold and silver are mainly used as raw materials of jewelry and currency, are important electric contact materials at present, and are suitable for electric appliances with medium load and can also be used as sliding contact materials; alloys of noble metals are often used as resistance materials, temperature measuring materials, welding materials, and the like; in addition, the catalyst is widely applied to the catalyst and the electric key industry.
While the demand of rare noble metals is increasing, the rare noble metal resources are increasingly deficient. The selenium steaming residue is a byproduct generated by processing crude selenium produced in the comprehensive recovery process of copper anode slime through a vacuum distillation pyrogenic process, and is rich in precious metals such as gold, silver, platinum and palladium, rare metal selenium and tellurium, base metals such as copper, lead and arsenic. Compared with wet materials such as copper anode mud and the like, the selenium steaming slag has the following characteristics: (1) The catalyst has the advantages of high inertness, poor activity, compactness and trace amount of melt, and influences the high-efficiency separation of base metal copper and lead and rare dispersed metal selenium and tellurium and the high enrichment of noble metal gold, silver, platinum and palladium; (2) Selenium and tellurium are high in content and exist in multiple phase structures simultaneously, and form intercrystallization and inclusion structures with copper, gold, silver and the like, so that the selective removal of the components such as copper, lead, selenium, tellurium and the like is severely limited; (3) The copper and lead contents are high, and the dispersion and loss of noble metals cannot be avoided in the separation process. Therefore, the selenium steaming slag is different from wet-process materials such as copper anode slime and the like, is extremely difficult to comprehensively recover, is mainly stockpiled at present, is rich in high-value rare and precious metals, and needs to be researched and developed to realize the commercialization, the maximization of the production benefit and the harmlessness of the selenium steaming slag.
Disclosure of Invention
The invention aims to solve the technical problem of overcoming the defects and shortcomings in the background technology and providing a method for recovering rare and precious metals in selenium steaming slag in a gradient manner.
In order to solve the technical problems, the technical scheme provided by the invention is as follows:
a method for recovering rare and precious metals in selenium steaming slag in a gradient manner comprises the following steps:
(1) Uniformly mixing the steamed selenium slag, an alkaline material and an oxidant, placing the mixture in an electrothermal roasting rotary kiln for roasting to obtain a roasted material, and then soaking the roasted material in water to obtain a selenium-tellurium-containing leachate and selenium-tellurium-separating slag;
(2) Acid leaching the selenium and tellurium separating residue obtained in the step (1) to obtain copper and silver containing leachate and copper separating residue; adding sodium chloride into the copper-containing silver leaching solution, stirring for reaction, and aging to obtain silver chloride precipitate and copper-containing supernatant;
(3) Stirring and leaching the copper separation slag obtained in the step (2) and a composite leaching agent to obtain a gold-containing platinum-palladium leaching solution and gold separation slag; wherein the composite leaching agent is prepared from the following components in a mass ratio of 1-5: 1 to 5: 1-5 parts of sodium chlorate, sodium chloride and sulfuric acid;
(4) Adding sodium sulfite into the gold-containing platinum-palladium leaching solution obtained in the step (3) to carry out stirring reaction to obtain crude gold powder and platinum-palladium-containing solution;
(5) Adding zinc powder into the platinum-palladium-containing solution obtained in the step (4) to carry out stirring reaction and ageing to obtain platinum-palladium concentrate;
(6) Slurrying the silver chloride precipitate obtained in the step (2) and the gold separation slag obtained in the step (3), adding ammonia water, and stirring and leaching to obtain a silver-containing leaching solution and leaching slag;
(7) And (4) adding hydrazine hydrate into the silver-containing leaching solution obtained in the step (6) to carry out stirring reaction to obtain the crude silver powder.
Preferably, the selenium steaming residue is generated in the crude selenium vacuum distillation process, and the selenium steaming residue comprises the following main components: 5.0 to 15.0 weight percent of Cu, 5.0 to 15.0 weight percent of Pb, 15.0 to 25.0 weight percent of Se, 5.0 to 15.0 weight percent of Te, 500 to 1500g/t of Au, 1.0 to 10.0 weight percent of Ag, 200 to 1000g/t of Pt and 200 to 1000g/t of Pd.
Preferably, in the step (1), the alkali material is sodium carbonate and/or sodium hydroxide, the oxidant is one or more of sodium chlorate, sodium nitrate, sodium persulfate and ferric sulfate, and the mass ratio of the selenium steaming residue to the alkali material to the oxidant is 10:5 to 10:1 to 5.
Preferably, in the step (1), the selenium steaming residue, the alkali material and the oxidant are mixed and then roasted at the temperature of 200-800 ℃ for 1-5 h; the liquid-solid ratio of the roasted material subjected to water leaching is 5-10 mL/g, the leaching time is 1-3 h, and the leaching temperature is 50-90 ℃.
Preferably, in the step (2), the acid adopted in the acid leaching process is a sulfuric acid solution with the concentration of 0.1-1.0 mol/L, the liquid-solid ratio in the acid leaching process is 5-10 mL/g, the leaching temperature is 50-90 ℃, and the leaching time is 1-4 h; adding sodium chloride, stirring and reacting, wherein the sodium chloride excess coefficient is 2-5, the reaction time is 0.5-2 h, the reaction temperature is 30-50 ℃, and the aging time is 1-3 h.
Preferably, in the step (3), the mass ratio of the copper-separating slag to the composite leaching agent is 1-5: 5, the liquid-solid ratio in the leaching process is 5-10 mL/g, the leaching temperature is 50-90 ℃, the leaching time is 2-6 h, and the pH value of the solution in the leaching process is not higher than 3.
Preferably, in the step (4), the excess coefficient of sodium sulfite is 5-20, the reaction temperature is 30-50 ℃, and the reaction time is 2-4 h.
Preferably, in the step (5), the excess coefficient of zinc powder is 5-20, the pH value of the platinum-containing palladium solution is adjusted to 3.0, the reaction temperature is normal temperature, the reaction time is 2-5 hours, and the aging time is 1-2 hours.
Preferably, in the step (6), in the agitation leaching process, the excess coefficient of ammonia water is 1-5, the pH of the reaction solution is adjusted to 7.0-14 by sodium hydroxide, the leaching temperature is 30-50 ℃, and the leaching time is 2-5 h.
Preferably, in the step (7), in the stirring reaction process, the excess coefficient of hydrazine hydrate is 2-10, the pH of the reaction solution is adjusted to 13-14 by sodium hydroxide, the reaction temperature is 50-70 ℃, and the reaction time is 2-5 hours.
Compared with the prior art, the invention has the advantages that:
(1) According to the invention, intercrossed substances and inclusion structures formed by selenium, tellurium, copper, gold, silver and the like in the selenium steaming slag are destroyed through alkaline oxidizing roasting, so that the step selective recovery of rare and precious metals is realized.
(2) The method for recovering rare and noble metals in the selenium steaming slag in a cascade manner can ensure high comprehensive recovery rate of the rare and noble metals, has simple and convenient process and low production cost, can graft the products obtained after the separation of all the working procedures onto the original rare and noble metal treatment process of enterprises, makes up the deficiency of the conventional selenium steaming slag treatment process, and is easy to popularize and apply in a large scale.
Drawings
FIG. 1 is a flow chart of the step recovery of rare and precious metals in selenium-steaming slag according to the embodiment of the invention.
Detailed Description
In order to facilitate an understanding of the invention, reference will now be made in detail to the present preferred embodiments of the invention, examples of which are illustrated in the accompanying drawings, and the scope of the invention is not limited to the following specific embodiments.
Unless otherwise defined, all terms of art used hereinafter have the same meaning as commonly understood by one of ordinary skill in the art. The terminology used herein is for the purpose of describing particular embodiments only and is not intended to limit the scope of the present invention.
Unless otherwise specifically indicated, various raw materials, reagents, instruments, equipment and the like used in the present invention are commercially available or can be prepared by existing methods.
The selenium steaming slag to be treated in the following examples 1 to 3 and comparative examples 1 and 2 is generated in the course of crude selenium vacuum distillation, and the content of the main components is as follows: cu 10.39wt%, pb 6.39wt%, se 19.70wt%, te 11.49wt%, au 705.73g/t, ag 6.13wt%, pt 403.46g/t, pd 517.70g/t.
Example 1:
the invention discloses a method for recovering rare and precious metals in selenium steaming slag in a gradient manner, which has the process flow shown in figure 1 and comprises the following steps:
(1) Alkaline oxidizing roasting: the method comprises the following steps of (1) mixing selenium steaming residue, sodium hydroxide and sodium chlorate according to a mass ratio of 10:7:1, then placing the mixture into an electrothermal roasting rotary kiln, and roasting the mixture for 3 hours at the roasting temperature of 600 ℃ to obtain a roasted material.
(2) Water leaching selenium and tellurium: stirring and leaching the roasted material obtained in the step (1) with water, wherein the liquid-solid ratio of leaching is 10:1 (mL/g), the leaching time is 2h, the leaching temperature is 90 ℃, and after the leaching is finished, the leaching is carried out, and then the filtration is carried out, so as to obtain a selenium-tellurium-containing leaching solution and a selenium-tellurium-separating slag, wherein the leaching rates of selenium and tellurium are respectively 98.24% and 99.05%.
(3) Copper is separated from weak acid: stirring and leaching the selenium and tellurium-separated slag obtained in the step (2) and a sulfuric acid solution with the concentration of 0.5mol/L for 2 hours, wherein the liquid-solid ratio in the leaching process is 10:1 (mL/g), the leaching temperature is 90 ℃, and after the leaching is finished, the copper-silver-containing leaching solution and the copper separating slag are obtained by filtering, wherein the leaching rates of copper and silver are respectively 98.38% and 45.61%.
(4) And (3) stirring the obtained copper-silver-containing leachate and sodium chloride for reaction, wherein the excess coefficient of the sodium chloride is 3, the reaction time is 0.5h, the reaction temperature is 30 ℃, aging is carried out for 1h after the reaction is finished, and filtering is carried out to obtain silver chloride precipitate and copper-containing supernatant, wherein the silver precipitation rate is 98.49%.
(5) Gold separation by chlorination: stirring and leaching the copper-separated slag obtained in the step (3) and a composite leaching agent (sodium chlorate, sodium chloride and sulfuric acid are mixed according to a mass ratio of 1: the mass ratio of the copper-separating slag to the composite leaching agent is 1:5, the solid-to-liquid ratio of the leaching solution is 10:1 (mL/g), the leaching temperature is 90 ℃, the pH value of the system is adjusted to be below 3, and the leaching is finished and then the system is filtered to obtain gold-containing platinum-palladium leaching solution and gold separating slag, wherein the leaching rates of gold, platinum and palladium in the process are respectively 99.58%, 92.49% and 95.26%.
(6) And (3) gold precipitation by sodium sulfite: and (4) stirring the gold-containing platinum-palladium leaching solution obtained in the step (5) and sodium sulfite to react for 4 hours, wherein the excess coefficient of the sodium sulfite is 5, the reaction temperature is 30 ℃, and filtering is carried out after the reaction is finished to obtain crude gold powder and platinum-palladium-containing solution, wherein the gold precipitation rate is 99.68% in the process.
(7) Zinc powder displacement of platinum and palladium: and (3) stirring the platinum-palladium-containing solution obtained in the step (6) and zinc powder at normal temperature for 2 hours to react, wherein the excess coefficient of the zinc powder is 10, the pH value of the platinum-palladium-containing solution is adjusted to 3.0, aging is carried out for 2 hours after the reaction is finished, and filtering is carried out to obtain platinum-palladium concentrate and supernatant, wherein the platinum precipitation rate and the palladium precipitation rate are respectively 91.60% and 95.08%.
(8) Ammonia leaching for separating silver: slurrying the silver chloride precipitate obtained in the step (4) and the gold separation slag obtained in the step (5), adding ammonia water, stirring and leaching for 2 hours, wherein the conditions of the leaching process are as follows: the ammonia water excess coefficient is 1, the pH value of the reaction solution is adjusted to 12 by adopting sodium hydroxide, the leaching temperature is 40 ℃, and the leaching is finished and then the reaction solution is filtered to obtain silver-containing leaching liquid and leaching residues, wherein the silver leaching rate in the process is 99.46 percent.
(9) Hydrazine hydrate silver precipitation: and (3) stirring and reacting the silver-containing leachate obtained in the step (8) with hydrazine hydrate for 3h, wherein the excess coefficient of the hydrazine hydrate is 10, adjusting the pH value of the reaction solution to 14 by adopting sodium hydroxide, the reaction temperature is 60 ℃, and filtering after the reaction is finished to obtain crude silver powder and supernatant, wherein the silver precipitation rate in the process is 97.78%.
Example 2:
the invention discloses a method for recovering rare and precious metals in selenium steaming slag in a gradient manner, which has the process flow shown in figure 1 and comprises the following steps:
(1) Alkaline oxidizing roasting: mixing the selenium steaming slag with sodium carbonate and sodium persulfate according to the mass ratio of 10:5:5, then placing the mixture in an electrothermal roasting rotary kiln, and roasting the mixture for 2 hours at the roasting temperature of 800 ℃ to obtain a roasted material.
(2) Water leaching selenium and tellurium: stirring and leaching the roasted material obtained in the step (1) with water, wherein the leached liquid-solid ratio is 5:1 (mL/g), the leaching time is 3h, the leaching temperature is 80 ℃, and after the leaching is finished, the leaching is carried out, and then the filtration is carried out, so as to obtain a selenium-tellurium-containing leaching solution and a selenium-tellurium-separating slag, wherein the leaching rates of selenium and tellurium are 97.92% and 97.48%, respectively.
(3) And weak acid copper separation: stirring and leaching the selenium and tellurium-separated slag obtained in the step (2) and a sulfuric acid solution with the concentration of 1.0mol/L for 4 hours, wherein the liquid-solid ratio in the leaching process is 5:1 (mL/g), the leaching temperature is 80 ℃, and after the leaching is finished, the copper-silver-containing leaching solution and the copper separating slag are obtained by filtering, wherein the leaching rates of copper and silver are 99.76 percent and 60.37 percent respectively.
(4) And (3) carrying out stirring reaction on the obtained copper-silver-containing leachate and sodium chloride, wherein the excess coefficient of the sodium chloride is 4, the reaction time is 1h, the reaction temperature is 30 ℃, aging is carried out for 1h after the reaction is finished, filtering is carried out, a silver chloride precipitate and a copper-containing supernatant are obtained, and the silver precipitation rate is 98.88%.
(5) Gold separation by chlorination: stirring and leaching the copper-separated slag obtained in the step (3) and a composite leaching agent (sodium chlorate, sodium chloride and sulfuric acid are mixed according to a mass ratio of 1: the mass ratio of the copper separating slag to the composite leaching agent is 1:3, the solid-to-liquid ratio of the leaching solution is 10:1 (mL/g), the leaching temperature is 90 ℃, the pH value of the system is adjusted to be below 3, and after the leaching is finished, the system is filtered to obtain gold-containing platinum-palladium leaching solution and gold separating slag, wherein the leaching rates of gold, platinum and palladium in the process are respectively 97.34%, 94.60% and 92.84%.
(6) And (3) gold precipitation by sodium sulfite: and (3) stirring the gold-containing platinum-palladium leaching solution obtained in the step (5) and sodium sulfite to react for 3 hours, wherein the excess coefficient of the sodium sulfite is 3, the reaction temperature is 40 ℃, and filtering is carried out after the reaction is finished to obtain crude gold powder and platinum-palladium-containing solution, wherein the gold precipitation rate in the process is 96.72%.
(7) Replacing platinum and palladium with zinc powder: and (3) stirring the platinum-palladium-containing solution obtained in the step (6) and zinc powder at normal temperature for reaction for 3 hours, wherein the excess coefficient of the zinc powder is 5, the pH value of the platinum-palladium-containing solution is adjusted to 3.0, the platinum-palladium-containing solution is aged for 4 hours after the reaction is finished, and filtering is carried out to obtain platinum-palladium concentrate and supernatant, wherein the precipitation rates of platinum and palladium in the process are 90.61% and 92.73% respectively.
(8) Ammonia leaching for silver separation: slurrying the silver chloride precipitate obtained in the step (4) and the gold separation slag obtained in the step (5), adding ammonia water, stirring and leaching for 3 hours, wherein the conditions of the leaching process are as follows: the ammonia water excess coefficient is 2, the pH value of the reaction solution is adjusted to 13 by adopting sodium hydroxide, the leaching temperature is 50 ℃, and the reaction solution is filtered after the leaching is finished to obtain silver-containing leaching liquid and leaching residues, wherein the silver leaching rate is 99.70 percent in the process.
(9) Hydrazine hydrate silver precipitation: and (3) stirring the silver-containing leachate obtained in the step (8) and hydrazine hydrate for reaction for 5 hours, wherein the excess coefficient of the hydrazine hydrate is 5, adjusting the pH value of the reaction solution to 14 by adopting sodium hydroxide, the reaction temperature is 50 ℃, and filtering after the reaction is finished to obtain crude silver powder and supernatant, wherein the silver precipitation rate in the process is 96.37%.
Example 3:
the invention relates to a method for recovering rare and precious metals in selenium steaming slag in a gradient manner, which has the process flow shown in figure 1 and comprises the following steps:
(1) Alkaline oxidizing roasting: mixing the selenium steaming slag with sodium hydroxide and sodium nitrate according to the mass ratio of 10:10:5, then placing the mixture into an electrothermal roasting rotary kiln, and roasting the mixture for 2 hours at the roasting temperature of 500 ℃ to obtain a roasted material.
(2) Water leaching selenium and tellurium: stirring and leaching the roasted material obtained in the step (1) with water, wherein the leaching liquid-solid ratio is 10:1 (mL/g), the leaching time is 3h, the leaching temperature is 90 ℃, and after the leaching is finished, the leaching is carried out, and then the filtration is carried out, so as to obtain a selenium-tellurium-containing leaching solution and a selenium-tellurium-separating slag, wherein the leaching rates of selenium and tellurium are respectively 92.61% and 91.48%.
(3) Copper is separated from weak acid: stirring and leaching the selenium and tellurium separating slag obtained in the step (2) and a sulfuric acid solution with the concentration of 0.1mol/L for 4 hours, wherein the liquid-solid ratio in the leaching process is 10:1 (mL/g) at the leaching temperature of 90 ℃, and filtering after the leaching is finished to obtain copper-silver-containing leachate and copper separating slag, wherein the leaching rates of copper and silver are 90.09 percent and 26.37 percent respectively.
(4) And (3) carrying out stirring reaction on the obtained copper-silver-containing leachate and sodium chloride, wherein the excess coefficient of the sodium chloride is 2, the reaction time is 1h, the reaction temperature is 50 ℃, aging is carried out for 2h after the reaction is finished, and filtering is carried out to obtain a silver chloride precipitate and a copper-containing supernatant, wherein the silver precipitation rate is 99.28%.
(5) Gold separation by chlorination: stirring and leaching the copper-separated slag obtained in the step (3) and a composite leaching agent (sodium chlorate, sodium chloride and sulfuric acid are mixed according to a mass ratio of 1: the mass ratio of the copper separating slag to the composite leaching agent is 1:1, the solid-to-liquid ratio of the leaching solution is 10:1 (mL/g), the leaching temperature is 80 ℃, the pH value of the system is adjusted to be below 3, and the leaching is finished and then the system is filtered to obtain gold-containing platinum-palladium leaching solution and gold separating slag, wherein the leaching rates of gold, platinum and palladium in the process are 97.21%, 93.62% and 93.49% respectively.
(6) And (3) gold precipitation by sodium sulfite: and (3) stirring the gold-containing platinum-palladium leaching solution obtained in the step (5) and sodium sulfite to react for 2 hours, wherein the excess coefficient of the sodium sulfite is 10, the reaction temperature is 40 ℃, and filtering is carried out after the reaction is finished to obtain crude gold powder and platinum-palladium-containing solution, wherein the gold precipitation rate in the process is 98.34%.
(7) Replacing platinum and palladium with zinc powder: and (3) stirring the platinum-palladium-containing solution obtained in the step (6) and zinc powder at normal temperature for reaction for 4 hours, wherein the excess coefficient of the zinc powder is 5, the pH value of the platinum-palladium-containing solution is adjusted to 3.0, the platinum-palladium-containing solution is aged for 2 hours after the reaction is finished, and filtering is carried out to obtain platinum-palladium concentrate and supernatant, wherein the precipitation rates of platinum and palladium in the process are 94.19% and 92.68% respectively.
(8) Ammonia leaching for silver separation: slurrying the silver chloride precipitate obtained in the step (4) and the gold separation slag obtained in the step (5), adding ammonia water, stirring and leaching for 3 hours, wherein the conditions of the leaching process are as follows: the ammonia water excess coefficient is 3, the pH value of the reaction solution is adjusted to 10 by adopting sodium hydroxide, the leaching temperature is 50 ℃, and the leaching is finished and then the reaction solution is filtered to obtain silver-containing leaching liquid and leaching residues, wherein the silver leaching rate in the process is 97.50 percent.
(9) Hydrazine hydrate silver precipitation: and (3) stirring and reacting the silver-containing leachate obtained in the step (8) with hydrazine hydrate for 3 hours, wherein the excess coefficient of the hydrazine hydrate is 10, adjusting the pH value of the reaction solution to 14 by adopting sodium hydroxide, the reaction temperature is 60 ℃, and filtering after the reaction is finished to obtain crude silver powder and supernatant, wherein the silver precipitation rate in the process is 96.18%.
Comparative example 1:
the method for recovering rare and precious metals from the selenium steaming slag in the comparative example comprises the following steps:
(1) Alkaline roasting: mixing the selenium steaming slag and sodium hydroxide according to the mass ratio of 10:7, then placing the mixture in an electrothermal roasting rotary kiln, and roasting the mixture for 3 hours at the roasting temperature of 600 ℃ to obtain a roasted material.
(2) Water leaching selenium and tellurium: stirring and leaching the roasted material obtained in the step (1) with water, wherein the leaching liquid-solid ratio is 10:1 (mL/g), the leaching time is 2h, the leaching temperature is 90 ℃, and after the leaching is finished, the leaching is carried out, and then the filtration is carried out, so as to obtain a selenium-tellurium-containing leaching solution and a selenium-tellurium-separating slag, wherein the leaching rates of selenium and tellurium are respectively 10.49% and 8.48%.
(3) Copper is separated from weak acid: stirring and leaching the selenium and tellurium-separated slag obtained in the step (2) and a sulfuric acid solution with the concentration of 0.5mol/L for 2 hours, wherein the liquid-solid ratio in the leaching process is 10:1 (mL/g), the leaching temperature is 90 ℃, and after the leaching is finished, the copper-silver-containing leaching solution and the copper separating slag are obtained by filtering, wherein the leaching rates of copper and silver are respectively 9.34% and 1.52%.
(4) Gold separation by chlorination: stirring and leaching the copper-separated slag obtained in the step (3) and a composite leaching agent (sodium chlorate, sodium chloride and sulfuric acid are mixed according to a mass ratio of 1: the mass ratio of the copper-separating slag to the composite leaching agent is 1:5, the solid-to-liquid ratio of the leaching solution is 10:1 (mL/g), the leaching temperature is 90 ℃, the pH value of the system is adjusted to be below 3, and after the leaching is finished, the filtration is carried out, wherein gold, platinum and palladium are not leached in the process.
Comparative example 2:
the method for recovering rare and precious metals from the selenium steaming slag in the comparative example comprises the following steps:
(1) Alkaline oxidation leaching: the method comprises the following steps of (1) mixing selenium steaming residue, sodium hydroxide and sodium chlorate according to a mass ratio of 10:7:1, and then stirring and leaching the mixture with water, wherein the leaching liquid-solid ratio is 10:1 (mL/g), the leaching time is 2h, the leaching temperature is 90 ℃, and after the leaching is finished, the leaching is carried out, and then the filtration is carried out, so as to obtain a selenium-tellurium-containing leaching solution and a selenium-tellurium-separating slag, wherein the leaching rates of selenium and tellurium are 3.94% and 4.76% respectively.
(2) Copper is separated from weak acid: stirring and leaching the selenium and tellurium-separated slag obtained in the step (1) and a sulfuric acid solution with the concentration of 0.5mol/L for 2 hours, wherein the liquid-solid ratio in the leaching process is 10:1 (mL/g), the leaching temperature is 90 ℃, and after the leaching is finished, the copper-silver-containing leaching solution and the copper separating slag are obtained by filtering, the copper leaching rate is 1.09%, and the silver is not leached.
(3) Gold separation by chlorination: stirring and leaching the copper-separated slag obtained in the step (2) and a composite leaching agent (sodium chlorate, sodium chloride and sulfuric acid are mixed according to a mass ratio of 1: the mass ratio of the copper-separating slag to the composite leaching agent is 1:5, the solid-to-liquid ratio of the leaching solution is 10:1 (mL/g), the leaching temperature is 90 ℃, the pH value of the system is adjusted to be below 3, and the leaching is finished and then the filtering is carried out to obtain gold-containing platinum-palladium leaching solution and gold separating slag, wherein gold, platinum and palladium are not leached in the process.
Therefore, the method provided by the invention has the advantages that the intercalant and inclusion structure formed by selenium, tellurium, copper, gold, silver and the like in the selenium steaming slag is firstly destroyed in an alkaline oxidizing roasting mode, so that the subsequent selective step recovery of rare and precious metals can be ensured.
The foregoing is considered as illustrative of the preferred embodiments of the invention and is not to be construed as limiting the invention in any way. Therefore, any simple modification, equivalent change and modification made to the above embodiments according to the technical essence of the present invention shall fall within the protection scope of the technical solution of the present invention, unless the technical essence of the present invention departs from the content of the technical solution of the present invention.
Claims (10)
1. A method for recovering rare and precious metals in selenium steaming slag in a gradient manner is characterized by comprising the following steps:
(1) Uniformly mixing the steamed selenium slag, an alkaline material and an oxidant, placing the mixture in an electrothermal roasting rotary kiln for roasting to obtain a roasted material, and then soaking the roasted material in water to obtain a selenium-tellurium-containing leachate and selenium-tellurium-separating slag;
(2) Acid leaching the selenium and tellurium separated slag obtained in the step (1) to obtain copper and silver containing leachate and copper separated slag; adding sodium chloride into the copper-containing silver leaching solution, stirring for reaction, and aging to obtain silver chloride precipitate and copper-containing supernatant;
(3) Stirring and leaching the copper separation slag obtained in the step (2) and a composite leaching agent to obtain a gold-containing platinum-palladium leaching solution and gold separation slag; wherein the composite leaching agent is prepared from the following components in a mass ratio of 1-5: 1 to 5: 1-5 parts of sodium chlorate, sodium chloride and sulfuric acid;
(4) Adding sodium sulfite into the gold-containing platinum-palladium leaching solution obtained in the step (3) to carry out stirring reaction to obtain crude gold powder and platinum-palladium-containing solution;
(5) Adding zinc powder into the platinum-palladium-containing solution obtained in the step (4), stirring for reaction, and aging to obtain platinum-palladium concentrate;
(6) Slurrying the silver chloride precipitate obtained in the step (2) and the gold separation slag obtained in the step (3), adding ammonia water, and stirring and leaching to obtain a silver-containing leaching solution and leaching slag;
(7) And (4) adding hydrazine hydrate into the silver-containing leaching solution obtained in the step (6) to carry out stirring reaction to obtain the crude silver powder.
2. The method for cascade recovery of rare and precious metals in selenium steaming slag as claimed in claim 1, wherein the selenium steaming slag is produced in a crude selenium vacuum distillation process, and comprises the following main components: 5.0 to 15.0 weight percent of Cu, 5.0 to 15.0 weight percent of Pb, 15.0 to 25.0 weight percent of Se, 5.0 to 15.0 weight percent of Te, 500 to 1500g/t of Au, 1.0 to 10.0 weight percent of Ag, 200 to 1000g/t of Pt and 200 to 1000g/t of Pd.
3. The method for stepwise recovering rare and precious metals in the selenium steaming slag according to claim 1, wherein in the step (1), the alkali material is sodium carbonate and/or sodium hydroxide, the oxidant is one or more of sodium chlorate, sodium nitrate, sodium persulfate and ferric sulfate, and the mass ratio of the selenium steaming slag to the alkali material to the oxidant is 10:5 to 10:1 to 5.
4. The method for step recovery of rare and precious metals in selenium steaming slag as claimed in claim 1, wherein in step (1), the roasting temperature is 200-800 ℃, and the roasting time is 1-5 h; the liquid-solid ratio of the roasted material subjected to water leaching is 5-10 mL/g, the leaching time is 1-3 h, and the leaching temperature is 50-90 ℃.
5. The method for step recovery of rare noble metals in the selenium steaming residue in accordance with claim 1, wherein in the step (2), the acid used in the acid leaching process is sulfuric acid solution with concentration of 0.1-1.0 mol/L, the liquid-solid ratio in the acid leaching process is 5-10 mL/g, the leaching temperature is 50-90 ℃, and the leaching time is 1-4 h;
in the process of adding sodium chloride for stirring reaction, the sodium chloride excess coefficient is 2-5, the reaction time is 0.5-2 h, the reaction temperature is 30-50 ℃, and the aging time is 1-3 h.
6. The method for step recovery of rare and precious metals in selenium steaming slag as claimed in claim 1, wherein in step (3), the mass ratio of copper separating slag to composite leaching agent is 1-5: 5, the liquid-solid ratio in the leaching process is 5-10 mL/g, the leaching temperature is 50-90 ℃, the leaching time is 2-6 h, and the pH value of the solution in the leaching process is not higher than 3.
7. The method for stepwise recycling rare and precious metals in the selenium steaming residue as claimed in claim 1, wherein in the step (4), the sodium sulfite excess coefficient is 5-20, the reaction temperature is 30-50 ℃, and the reaction time is 2-4 h.
8. The method for stepwise recycling rare and precious metals in selenium steaming slag as claimed in claim 1, wherein in the step (5), the zinc powder excess coefficient is 5-20, the pH of the platinum-containing palladium solution is adjusted to 3.0, the reaction temperature is normal temperature, the reaction time is 2-5 h, and the aging time is 1-2 h.
9. The method for stepwise recycling rare and precious metals in the residue after selenium steaming as claimed in claim 1, wherein in the step (6), the excess coefficient of ammonia water is 1-5 during agitation leaching, the pH of the reaction solution is adjusted to 7.0-14 by sodium hydroxide, the leaching temperature is 30-50 ℃, and the leaching time is 2-5 h.
10. The method for step recovery of rare and precious metals in selenium steaming slag as claimed in claim 1, wherein in the step (7), in the stirring reaction process, the excess coefficient of hydrazine hydrate is 2-10, the pH of the reaction solution is adjusted to 13-14 by sodium hydroxide, the reaction temperature is 50-70 ℃, and the reaction time is 2-5 h.
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