Process for the extraction of platinum group metals from matte

Abstract

The application discloses a method for extracting platinum group metals from copper matte, and relates to the technical field of precious metal metallurgy. The method comprises the following steps: after tin-containing copper matte raw materials are crushed, the crushed raw materials are added with dilute hydrochloric acid and sodium chlorate according to a solid-liquid ratio of 1:5-1:8, and then leaching is carried out at 60-90 DEG C, so that primary precious metal residues are obtained through filtration; after drying, a slagging agent and a capturing agent are added, smelting is carried out in a medium-frequency furnace at 1250-1550 DEG C, so that alloy blocks containing platinum group metals are obtained through slag-gold separation; after crushing, the alloy blocks are leached with dilute hydrochloric acid, and washing is carried out, so that precious metal fine residues are obtained; the precious metal fine residues are dissolved with aqua regia, so that aqua regia liquid and aqua regia residues are obtained through filtration, the aqua regia liquid is added with a reducing agent, and gold, platinum and palladium are extracted in sections; the aqua regia residues are activated by alkali fusion, so that rhodium, ruthenium and iridium are extracted. Through coupling of multi-section wet method and fire method, the method realizes deep removal of base metals and high-multiple enrichment of platinum group metals in the tin-containing copper matte, has a short process flow, low energy consumption, and the comprehensive recovery rate of platinum group metals can reach more than 90%.

Description

Technical Field

[0001] This invention belongs to the field of precious metal metallurgy technology, specifically relating to a method for extracting platinum group metals from copper matte. Background Technology

[0002] Copper matte is an intermediate / byproduct of pyrometallurgical copper or tin smelting processes, and its main component is... -FeS eutectic often contains trace amounts of platinum group metals (PGMs) and tin. Existing processes for extracting PGMs from copper-nickel matte mostly employ conventional acid leaching or pyrometallurgical collection, but these methods are poorly suited for tin-containing matte: tin easily leads to decreased leaching selectivity, viscous slag, and dispersion of precious metals. Therefore, a highly efficient extraction method for PGMs from tin-containing matte is urgently needed. Summary of the Invention

[0003] A method for extracting platinum group metals from copper matte includes the following steps: S1. Crush the tin-containing copper matte raw material, add dilute hydrochloric acid and sodium chlorate at a solid-liquid ratio of 1:5 to 1:8, stir and leach at 60 to 90°C, and filter to obtain primary precious metal slag. S2. The primary precious metal slag is dried, slag-forming agent and scavenging agent are added, and it is smelted in a medium frequency furnace at 1250~1550℃. After the impurities are slagified and floated to the surface, the bottom is cooled to obtain an alloy block containing platinum group metals. S3. The alloy block is crushed, leached with dilute hydrochloric acid, filtered and washed to obtain precious metal slag; S4. Dissolve the precious metal slag in aqua regia, filter to obtain aqua regia solution and aqua regia slag; add a reducing agent to the aqua regia solution to extract gold, platinum and palladium in stages; perform alkaline fusion activation on the aqua regia slag to extract rhodium, ruthenium and iridium.

[0004] Preferably, in step S1, the mass concentration of the dilute hydrochloric acid is 15%~25%; the sodium chlorate is added in 3~5 batches, and the amount added is 8%~15% of the mass of the tin-containing copper matte raw material; the leaching time is 4~8 hours.

[0005] Preferably, in step S2, the collector, based on the mass of the primary precious metal slag, comprises: 5%~15% of added copper source, 0.5%~2.0% of boron source, or 0.3%~1.5% of calcium fluoride; the residual tin content in the primary precious metal slag is detected before smelting, and the tin source is controlled or supplemented according to the mass ratio of total Sn to Cu in the added copper source of 1:8~1:15.

[0006] Preferably, in step S2, the external copper source is copper powder or copper oxide; the boron source is borax or boric acid; the melting temperature is 1250~1450℃, and the melting time is 1~2 hours.

[0007] Preferably, in step S2, the slag-forming agent comprises 3% to 8% quartz sand and 2% to 6% sodium carbonate, based on the mass of the primary precious metal slag; the collecting agent is copper or lead.

[0008] Preferably, in step S3, the alloy block is crushed to 100 mesh and then leached for 2-4 hours at 50-80°C using hydrochloric acid with a mass concentration of 10%-20% at a solid-liquid ratio of 1:4-1:6.

[0009] Preferably, in step S4, the aqua regia is a mixture of concentrated hydrochloric acid and concentrated nitric acid in a volume ratio of 3:1 to 4:1; the aqua regia is added according to a solid-liquid ratio of the precious metal slag to the aqua regia of 1:3 to 1:6, and dissolved at 80 to 95°C for 3 to 6 hours.

[0010] Preferably, in step S4, the specific steps for the fractional extraction of gold, platinum, and palladium from the aqua regia are as follows: Gold extraction: Heat the aqua regia solution to remove nitrates, control the concentration of free hydrochloric acid to 1~3 mol / L, add anhydrous sodium sulfite or sodium bisulfite in batches at 40~50℃ at 1.2~1.5 times the theoretical amount of gold, control the endpoint potential to 700~750 mV, and filter to obtain gold powder. Platinum extraction: Maintain the free hydrochloric acid concentration in the filtrate at 3~6 mol / L, heat to 50~80℃, add ammonium chloride and react for 1~3 hours, let stand to precipitate to obtain ammonium chloroplatinate; Palladium extraction: Add nitric acid to the filtrate after platinum extraction to adjust the free acid concentration to 1-3 mol / L, and add ammonium chloride at 60-90℃ to precipitate and extract palladium.

[0011] Preferably, in step S4, the specific steps for extracting rhodium, ruthenium, and iridium from the aqua regia residue are as follows: The aqua regia residue is mixed with sodium hydroxide and sodium peroxide at a mass ratio of 1:(1.5~2):(0.5~1) and melted at 550~700℃ for 1~2 hours; after the melt is cooled, it is leached with water at a solid-liquid ratio of 1:5~1:10, filtered, and the water leaching solution and water leaching residue are obtained. Add concentrated sulfuric acid or concentrated hydrochloric acid to the aqueous extract to adjust the pH to 1-3, heat to 80-95℃, and then pass chlorine gas through for oxidative distillation. The distilled ruthenium tetroxide is absorbed by hydrochloric acid solution and reduced to recover the ruthenium product. The water-leached residue is dissolved in concentrated hydrochloric acid to obtain a rhodium-iridium-containing solution. Ammonium chloride is added to the rhodium-iridium-containing solution to precipitate the residue. The precipitate is then calcined and reduced to obtain rhodium-iridium-enriched gold powder or rhodium / iridium products.

[0012] Preferably, the rhodium and iridium-containing solution is precipitated with ammonium chloride and then calcined and reduced to obtain rhodium-iridium enriched gold powder, wherein the total mass percentage of rhodium and iridium is 75% to 85%.

[0013] The method provided by this invention achieves high-rate stepwise enrichment of platinum group metals in low-grade, multi-impurity tin matte through alternating wet pre-removal and pyrometallurgical enrichment processes. At the same time, by utilizing the difference in solubility of aqua regia, easily soluble PGMs (Pt, Pd) and insoluble PGMs (Rh, Ru, Ir) are decoupled at the front end, which greatly reduces the mutual interference between various precious metal elements. The overall comprehensive recovery rate of platinum group metals can reach more than 90%. Detailed Implementation

[0014] To enable those skilled in the art to better understand the technical solution of the present invention, the present invention will be described in detail below. The description in this part is only exemplary and explanatory, and should not be used to limit the scope of protection of the present invention in any way.

[0015] This embodiment provides a method for extracting platinum group metals from copper matte, including the following steps: (1) First acid leaching to remove impurities: crush the tin-containing copper matte raw material, add dilute hydrochloric acid and sodium chlorate at a solid-liquid ratio of 1:5~1:8, stir and leach at 60~90℃, and filter to obtain primary precious metal slag.

[0016] (2) Smelting and enrichment: The primary precious metal slag is dried, and slag-forming agent and scavenging agent are added. The slag is then smelted in a medium-frequency furnace at 1250~1550℃ to make the impurities float to the surface. The bottom is cooled to obtain an alloy block containing platinum group metals.

[0017] (3) Secondary acid leaching to remove impurities: The alloy block is crushed, leached with dilute hydrochloric acid, filtered and washed to obtain precious metal slag.

[0018] (4) Aqua regia dissolution and refining extraction: Dissolve precious metal residue with aqua regia, filter to obtain aqua regia solution and aqua regia residue. Add reducing agent to aqua regia solution to extract gold, platinum and palladium in stages. Alkali fusion activation of aqua regia residue to extract rhodium, ruthenium and iridium.

[0019] This method solves the technical problems of severe entrainment of impurity elements (copper, iron, tin, etc.), dispersion of sparingly soluble platinum group metals, and low refining recovery rate in the traditional copper matte / copper matte purification process by using a multi-stage wet and pyrometallurgical coupled process of "first acid leaching to remove impurities - smelting and enrichment - second acid leaching to remove impurities - aqua regia dissolution and refining extraction".

[0020] Specifically, in step (1), the tin-containing matte raw material is copper matte / copper matte produced during pyrometallurgical copper smelting, tin smelting, or comprehensive recycling of non-ferrous metals, and its main components are... The eutectic is accompanied by Sn, Pb, Zn, and trace amounts of platinum group metals (PGMs: Pt, Pd, Rh, Ru, Ir) and gold (Au). The raw material is pulverized to 80-150 mesh (preferably 100 mesh), and dilute hydrochloric acid and sodium chlorate (an oxidizing agent) are added at a solid-liquid ratio of 1:5 to 1:8. Leaching is carried out with stirring at 60-90°C. During this process, the hydrochloric acid and sodium chlorate system selectively oxidizes and dissolves base metals such as copper, iron, and tin in the sulfide into soluble chlorides (e.g., copper, iron, tin, etc.). , , Gold and platinum group metals enter the leaching solution, while gold and platinum group metals are insoluble in this dilute acid system and remain in the filter residue, forming primary precious metal slag.

[0021] In step (2), the smelting enrichment can be achieved using one of the following two capture systems: System 1 (Copper / Lead Smelting): Based on the mass of primary precious metal slag, add 3%~8% quartz sand and 2%~6% sodium carbonate as slag-forming agents, and add 5%~10% metallic copper powder or lead blocks as smelting agents. Slag is formed by high-temperature smelting in an induction furnace at 1350~1550℃. This system is suitable for ordinary high-temperature smelting enrichment.

[0022] System Two (Acid-Leached Activated Self-Reinforcing Cu-Sn Microalloying): Based on the mass of the primary precious metal slag, add 5%~15% of an external copper source (copper powder / copper oxide) and 0.5%~2.0% of a boron source (borax / boric acid) or 0.3%~1.5% of calcium fluoride. The added tin source is elemental tin powder, metallic tin granules, or metallic tin blocks with a particle size ≤100 mesh. During batching, thoroughly dry-mix the primary precious metal slag, slag-forming agent, external copper source, boron source / calcium fluoride, and added tin source before loading them into the medium-frequency furnace. Melting is carried out in a neutral or weakly reducing atmosphere covered with a graphite plate. Melting is performed at a relatively low temperature of 1250~1450℃, utilizing the trace amount of tin remaining in the slag (or by adding additional tin powder to maintain the total Sn to Cu mass ratio at 1:8~1:15). During the initial heating and smelting phase (first 30-45 minutes), under a weak reducing atmosphere, low-melting-point tin and copper source first form Cu-Sn microalloy droplets with a relatively low melting temperature. With the viscosity reduction effect of boron source / calcium fluoride, PGMs are efficiently captured. As the smelting and holding process proceeds, borate / fluoride synergistic slag-forming agent regulates the slag basicity. The tin that has not participated in the capture and the excess tin in the slag are gradually selectively oxidized by the slag atmosphere and combined into the silicate / borate slag phase to float upward, thus achieving directional slag formation and removal of excess tin. Boron source / calcium fluoride and slag surface in-situ fluxing reduce viscosity. Added copper source and a small amount of residual tin in the primary precious metal slag (or trace amounts of added tin powder) form low-melting-point Cu-Sn microalloy droplets in-situ at the beginning of the reaction. The low viscosity and high permeability of these droplets rapidly capture and dissolve dispersed platinum group metal particles. As smelting progresses, the boron / fluorine system regulates the basicity and fluidity of the silicate slag phase, causing excess tin and base metal impurities such as iron and copper in the slag to be preferentially oxidized and combined with the slag-forming agent to form silicate / borate slag phases that float to the surface. This achieves directional slag formation and removal of excess tin, ensuring that the residual tin content in the final alloy block is controlled below 1.5%.

[0023] The specific formula for calculating the quality of the added tin source is as follows:

[0024] in, The formula for calculating the total mass of residual tin in the primary precious metal slag obtained through testing is as follows:

[0025] in The total tin mass of the primary precious metal slag after hydrochloric acid-hydrogen peroxide digestion was determined by ICP-OES; η is the reduction efficiency coefficient, representing the effective proportion of tin that can be reduced and participate in the formation of Cu-Sn alloy in the primary precious metal slag under a weakly reducing smelting atmosphere. The value of η is determined according to the phase morphology of tin in the primary precious metal slag: when tin in the primary precious metal slag mainly exists in the form of free metallic tin and tin sulfide (S / Sn molar ratio ≥ 1), the value of η is 0.90~0.95 (preferably 0.92); when some oxidized tin dioxide exists in the primary precious metal slag (S / Sn molar ratio < 1), the value of η is 0.85~0.90 (preferably 0.88). The mass of effective Cu element in the added external copper source; K is the target value to be controlled. The mass ratio, ranging from 8 to 15. If the calculated... If the result is less than or equal to 0, then add tin powder according to the calculated amount; if the result is less than or equal to 0, then no additional tin powder is needed.

[0026] High-melting-point and slightly dispersed platinum group metals are rapidly captured by the molten attractant and settle to the bottom under high-temperature molten conditions. Impurities, on the other hand, form low-viscosity silicate or borate slag phases with the slag-forming agent and float to the surface. After cooling in the furnace, slag and metal are separated to obtain an alloy block with preliminary enrichment of platinum group metals.

[0027] In step (3), the obtained alloy block is crushed a second time (preferably to 100 mesh) and then subjected to a second acid leaching with dilute hydrochloric acid (mass concentration 10%~20%). This step aims to wash and leach out residual iron, copper and other base metals on the surface and grain boundaries of the smelted alloy block, so as to increase the total mass percentage of platinum group metals and gold in the filter residue (precious metal concentrate) to 30%~40%, thus creating high-grade raw material conditions for subsequent complete dissolution with aqua regia.

[0028] In step (4), aqua regia (a mixture of concentrated hydrochloric acid and concentrated nitric acid) is used to strongly oxidize and dissolve the precious metal slag. Easily soluble precious metals such as gold, platinum, and palladium form chloroauric acid (…). ), chloroplatinic acid ( ), Chloropalladium acid ( The gold powder, platinum salt, and palladium salt are completely dissolved in the aqua regia solution, while the sparingly soluble platinum group metals (rhodium, ruthenium, iridium) remain in the aqua regia residue. A reducing agent (such as anhydrous sodium sulfite / sodium bisulfite, ammonium chloride, etc.) is added to the aqua regia solution for selective fractional precipitation separation, yielding high-purity gold powder, platinum salts, and palladium salts sequentially. The aqua regia residue is then treated with high-temperature alkaline fusion. The process involves activating the open circuit, disrupting the crystal structure of insoluble substances, and then performing water immersion, acid dissolution, and precipitation reduction to achieve efficient open-circuit extraction of rhodium, ruthenium, and iridium.

[0029] Specifically, the process of distilling ruthenium from the water extract is as follows: Before alkali melting, the aqua regia residue is dried at 105-110℃ and pre-calcined in a muffle furnace at 500℃ for 1 hour to remove residual organic matter and sulfides. The pre-calcined aqua regia residue is mixed with sodium hydroxide and sodium peroxide and placed in a corundum crucible or nickel crucible, then heated to 550-700℃ at a rate of 5-10℃ / min to melt. After filtration of the melt to obtain the water extract, concentrated sulfuric acid or concentrated hydrochloric acid is slowly added to adjust the pH to 1-3 (or a free acid concentration of 0.5-1.5 mol / L). The acidified water extract is placed in a distillation flask and heated to 80-95℃. Chlorine gas is passed through at a flow rate of 20-50 mL / min for oxidation, and the distillation product is ruthenium tetroxide (Ruthenium). The ruthenium powder was obtained by absorbing it with a 4-6 mol / L hydrochloric acid-ethanol mixed absorption solution, followed by evaporation and calcination to reduce the absorption solution.

[0030] The acid dissolution and rhodium / iridium extraction process of the water-leached residue is as follows: The water-leached residue is boiled in concentrated hydrochloric acid (25%–30% by mass) at 85–95°C for 2–4 hours, and filtered to obtain a chloride solution containing rhodium and iridium. Ammonium chloride, at a mass ratio of 2.5–3.5 times the total mass of rhodium and iridium, is added to the solution, and the reaction is carried out at 80–90°C for 1–2 hours. After cooling to room temperature, a mixed precipitate of ammonium rhodium chloride and ammonium iridium chloride is precipitated. The resulting mixed precipitate is washed 2–3 times with a dilute ammonium chloride solution (5%–8% by mass) at 40–50°C (using 5–10 mL of dilute ammonium chloride solution per gram of precipitate) to remove entrained impurities. , The residual amount of base metal impurity ions was reduced to below 0.1%; the washed precipitate was placed in a hydrogen reduction furnace and Ar-4- ions were introduced at 550~650℃ in a volume ratio of (90~95):(5~10). The mixed gas is reduced and calcined for 1.5 to 2 hours. By controlling the process parameters of high acidity dissolution, excess ammonium salt precipitation, warm and dilute ammonium chloride washing to remove impurities and medium-temperature reduction, rhodium-iridium enriched gold powder with a total rhodium and iridium mass percentage of 75% to 85% can be stably obtained.

[0031] Beneficially, this entire process achieves high-rate stepwise enrichment of platinum group metals in low-grade, multi-impurity tin matte through the alternation of wet pre-removal and pyrometallurgical enrichment. At the same time, by utilizing the difference in solubility of aqua regia, easily soluble PGMs (Pt, Pd) and sparingly soluble PGMs (Rh, Ru, Ir) are decoupled at the front end, which greatly reduces the mutual interference between various precious metal elements, and the overall comprehensive recovery rate of platinum group metals can reach more than 90%.

[0032] Further, in step (1), the mass concentration of dilute hydrochloric acid is 15%~25%. Sodium chlorate is added in 3~5 batches, with the amount added being 8%~15% of the mass of the tin-containing copper matte raw material; the leaching time is 4~8 hours; a micro-oxide layer and porous structure are formed on the surface of the primary precious metal slag.

[0033] The concentration of dilute hydrochloric acid is controlled at 15%~25% (preferably 20%). If the acidity is below 15%, the dissolution rate of sulfides and iron and copper impurities is too slow, resulting in incomplete impurity removal; if the acidity is above 25%, it not only increases the volatilization of acid mist and equipment corrosion, but may also cause trace amounts of palladium to be slightly dissolved, resulting in the loss of precious metals. Sodium chlorate, as a strong oxidant, is added at 8%~15% (preferably 10%~12%) of the mass of the tin-containing copper matte raw material, and is added slowly in 3~5 batches. Batch addition can effectively control the exothermic rate and gas release, avoiding excessively vigorous reaction that could cause the slurry to overflow. The leaching time is controlled at 4~8 hours (preferably 6 hours).

[0034] Under the interaction of the aforementioned acidity, oxidant dosage, and temperature (60~90℃), the surface of copper matte particles... Selective partial oxidation and dissolution occur, resulting in a unique physicochemical state on the surface of the filter residue particles, characterized by a micro-oxidation layer (containing some insoluble sulfur and trace amounts of oxidized species) and a locally porous structure (composed of microporous channels left by the large-scale dissolution of base metals). During this process, the dilute hydrochloric acid and sodium chlorate system efficiently and selectively leaches base metals and tin from the copper matte, achieving a tin removal rate of 90%–95%. The vast majority of the tin is removed and enters the leaching solution, with only trace amounts remaining in the primary precious metal slag.

[0035] Further, in step (2), the collector is an acid-leaching activated self-reinforcing Cu-Sn microalloy collector; based on the mass of the primary precious metal slag, the collector includes: 5%~15% of added copper source, 0.5%~2.0% of boron source or 0.3%~1.5% of calcium fluoride; before smelting, the residual tin content in the primary precious metal slag is detected, and the mass ratio of total Sn (the sum of residual tin and tin in the added tin source) to Cu in the added copper source is controlled or a trace amount of tin source is added. The tin source is elemental tin powder, metallic tin granules or metallic tin blocks with a particle size ≤100 mesh; during the smelting process, the residual tin is acid-leached from the tin-containing matte raw material or the added tin source is used to form Cu-Sn microalloys in situ; the smelting temperature is 1250~1450℃.

[0036] Furthermore, in step (2), the external copper source is copper powder or copper oxide. The boron source is borax or boric acid; the smelting temperature is 1250~1450℃, and the smelting time is 1~2 hours.

[0037] In one embodiment of this method, the collector, based on the mass of the primary precious metal slag, comprises: 5% to 15% of an added copper source (copper powder, copper scrap, or copper oxide powder), and 0.5% to 2.0% of a boron source (borax). or boric acid ) or 0.3%~1.5% calcium fluoride ( Fluorite powder). During the smelting process, the trace amount of tin remaining in the slag after acid leaching of tin-containing matte raw materials is utilized (or an additional trace amount of tin powder is added according to the Sn content in the slag to maintain the mass ratio of total Sn to Cu at 1:8~1:15).

[0038] At a smelting temperature of 1250~1450℃ (preferably 1350~1400℃), the boron source or calcium fluoride undergoes an in-situ fluxing reaction with the micro-oxide layer and porous micropores on the surface of the noble metal slag obtained in step (1), significantly reducing the local melting viscosity of the particle interface and the surface tension of the slag. The added copper source rapidly microalloys with the tin remaining in the slag under a high-temperature reducing / molten atmosphere, generating low-melting-point Cu-Sn microalloyed droplets. These droplets rapidly penetrate along the porous structure, capturing and dissolving the dispersed platinum group metal particles.

[0039] Beneficially, this collection system can reduce the smelting temperature from the conventional 1500℃ or above to 1250~1450℃, significantly reducing the energy consumption of the medium-frequency furnace and extending the furnace lining life; the same process simultaneously achieves "high-efficiency collection of PGMs, directional slag separation of tin and deep purification of impurities", and the enrichment ratio of platinum group metals in the resulting alloy block can reach more than 10 times, and the residual Sn and Fe in the alloy are significantly lower than those of conventional methods.

[0040] Further, in step (2), based on the mass of the primary precious metal slag, the slag-forming agent includes 3%~8% quartz sand and 2%~6% sodium carbonate; the collector is copper or lead, and the amount added is 5%~10% of the mass of the primary precious metal slag; the smelting temperature is 1350~1550℃.

[0041] In another embodiment of this implementation, 3% to 8% of quartz sand is added based on the mass of the primary precious metal slag. ) and 2%~6% sodium carbonate ( As a slag-forming agent, the smelting temperature is 1350~1550℃. Quartz sand reacts with sodium carbonate at high temperature to produce sodium silicate and silicate composite slag, which can effectively absorb residual iron oxide, copper oxide and other impurities in the copper matte acid leaching slag. The collector can be directly added with metallic copper powder or lead blocks (the amount added is 5%~10% of the slag mass).

[0042] Further, in step (3), the alloy block is crushed to 100 mesh and then leached for 2 to 4 hours at 50 to 80°C with hydrochloric acid of 10% to 20% by mass at a solid-liquid ratio of 1:4 to 1:6.

[0043] The alloy block obtained from smelting in step (2) is crushed and ground to 100 mesh (particle size ≤ 150 μm). Industrial dilute hydrochloric acid with a mass concentration of 10%~20% is used, and the mixture is stirred and leached at 50~80℃ for 2~4 hours at a solid-liquid ratio of 1:4~1:6. After filtration, the filter residue is washed with deionized water until neutral.

[0044] During the secondary acid leaching process, since the alloy block has been ground into fine powder, dilute hydrochloric acid can efficiently leach out the residual elemental iron, free copper, and some alkaline oxides that are enriched at the grain boundaries of the alloy powder, while platinum group metals and gold are not corroded by hydrochloric acid due to their high potential.

[0045] Beneficially, through secondary acid leaching, the total mass percentage of platinum group metals and gold in precious metal slag can be increased from 8%~12% in smelted alloy blocks to 30%~40%, removing more than 90% of residual base metals and greatly reducing acid consumption and waste liquid treatment pressure in subsequent aqua regia dissolution processes.

[0046] Further, in step (4), aqua regia is a mixture of concentrated hydrochloric acid and concentrated nitric acid in a volume ratio of 3:1 to 4:1. Aqua regia is added at a solid-liquid ratio of 1:3 to 1:6 between the precious metal slag and aqua regia, and dissolved at 80 to 95°C for 3 to 6 hours.

[0047] Furthermore, in step (4), the specific steps for the fractional extraction of gold, platinum, and palladium from aqua regia are as follows: Gold extraction: Heat the aqua regia solution to remove nitrates, control the concentration of free hydrochloric acid to 1~3 mol / L, and add anhydrous sodium sulfite or sodium bisulfite in batches at 1.2~1.5 times the theoretical amount of gold at 40~50℃, control the endpoint potential to 700~750 mV, and filter to obtain gold powder.

[0048] Platinum extraction: Maintain the free hydrochloric acid concentration in the filtrate at 3~6 mol / L, heat to 50~80℃, add ammonium chloride and react for 1~3 hours, then let it stand to precipitate to obtain ammonium chloroplatinate.

[0049] Palladium extraction: Add nitric acid to the filtrate after platinum extraction to adjust the free acid concentration to 1-3 mol / L, and add ammonium chloride at 60-90℃ to precipitate and extract palladium.

[0050] The aqua regia is prepared by mixing 37% hydrochloric acid and 65% nitric acid in a volume ratio of 3:1 to 4:1. For the high-grade, low-tin-impurity precious metal slag obtained in step (3), the solid-liquid ratio of the aqua regia is reduced to 1:3 to 1:6 (preferably 1:4), and the mixture is stirred and dissolved at 80 to 95°C for 3 to 6 hours. Due to the high purity of the alloy collected upstream and the absence of a dense tin film, the dissolution rate of the aqua regia can be significantly improved, and the time can be shortened by more than 2 hours.

[0051] The specific process for fractional extraction of gold, platinum, and palladium is as follows: 1. Gold Extraction: The aqua regia solution is first heated to remove excess nitric acid and nitrogen oxides, controlling the free hydrochloric acid concentration to 1-3 mol / L. At 40-50℃, a reducing agent (anhydrous sodium sulfite) is slowly added in batches at 1.2-1.5 times the theoretical mass of gold in the solution. or sodium bisulfite ), keep stirring. Monitor the reaction using a potentiometer, and stop the drug when the endpoint potential decreases and stabilizes at 700~750 mV. Gold ions ( The gold powder was selectively reduced to sponge gold powder and precipitated. After filtration, coarse gold powder was obtained, while platinum and palladium remained in the filtrate.

[0052] 2. Platinum Extraction: After gold extraction, add hydrochloric acid to the filtrate to adjust the free acid concentration to 3-6 mol / L (higher acidity inhibits palladium precipitation). Do not add nitric acid. Heat to 50-80℃, add excess ammonium chloride (solid or saturated solution, 1.5-3 times the mass of platinum), react for 1-3 hours, cool and let stand. Platinum ions (… It combines with chloride ions and ammonium ions to form insoluble yellow ammonium chloroplatinate. Precipitate, and filter to obtain platinum salt.

[0053] 3. Palladium extraction: After platinum extraction, add an appropriate amount of concentrated nitric acid to adjust the acidity of the filtrate to 1-3 mol / L for free acid concentration, and provide an oxidizing atmosphere. Oxidize to a higher oxidation state or form a complex, heat to 60-90℃, add ammonium chloride (2-4 times the mass of palladium), stir to react and produce red ammonium chloropalladate. Precipitate (or further dissolve in ammonia water and acidify with hydrochloric acid to precipitate dichlorodiamminepalladium), and filter to obtain palladium salt.

[0054] Beneficially, due to the upstream activation and capture yielding high-purity, low-tin precious metal concentrate, the solid-liquid ratio and dissolution time in aqua regia can be significantly reduced. Simultaneously, during segmented extraction, the interference from impurity ions is minimal, the endpoint potential control for gold extraction is more precise (700~750 mV), and the acidity windows for platinum and palladium extraction are more stringent and clear. The single-element recovery rates for gold, platinum, and palladium can all exceed 98.5%, and the product purity can reach over 99.95%.

[0055] Furthermore, in step (4), the specific steps for extracting rhodium, ruthenium, and iridium from aqua regia residue are as follows: Mix aqua regia residue with sodium hydroxide and sodium peroxide at a mass ratio of 1:(1.5~2):(0.5~1) and melt at 550~700℃ for 1~2 hours. After cooling, leach the melt with water at a solid-liquid ratio of 1:5~1:10, filter, and obtain the water leachate and water leachate residue.

[0056] Add concentrated sulfuric acid or concentrated hydrochloric acid to the aqueous extract to adjust the pH to 1-3, heat to 80-95℃, and then pass chlorine gas through for oxidative distillation. The distilled ruthenium tetroxide is absorbed by hydrochloric acid solution and reduced to recover the ruthenium product.

[0057] The water-leached residue is dissolved in concentrated hydrochloric acid to obtain a rhodium-iridium-containing solution. Ammonium chloride is added to the rhodium-iridium-containing solution to precipitate the residue. The precipitate is then calcined and reduced to obtain rhodium-iridium-enriched gold powder or rhodium / iridium products.

[0058] Furthermore, after the rhodium and iridium solution is precipitated with ammonium chloride and calcined and reduced, a rhodium-iridium enriched gold powder with a total rhodium and iridium mass percentage of 75% to 85% is obtained.

[0059] The main components of aqua regia residue are insoluble metallic rhodium (Rh), ruthenium (Ru), and iridium (Ir), as well as small amounts of impurities such as silicon dioxide and lead sulfate. After drying, the aqua regia residue is reacted with sodium hydroxide (NaOH) and sodium peroxide (NaOH). Mix the ingredients thoroughly at a mass ratio of 1:(1.5~2):(0.5~1), place the mixture in a high-alumina crucible, and melt it in a muffle furnace at 550~700℃ (preferably 600~650℃) for 1~2 hours. Under this high-temperature, strongly oxidizing alkaline fusion system, the sparingly soluble Rh, Ru, and Ir lattices are destroyed, transforming into soluble oxyacid salts or high-valence oxide complex ions (such as water-soluble sodium ruthenium oxide). , and acid-soluble rhodium / iridium oxides).

[0060] After the melt cools, deionized water is added at a solid-liquid ratio of 1:5 to 1:10, and the mixture is soaked at 60 to 80°C for 1 to 2 hours, followed by filtration. Ruthenium enters the leaching solution in the form of water-soluble high-valent ruthenium salts. Chlorine gas is then introduced into the leaching solution to oxidize and distill off ruthenium tetroxide. High-purity ruthenium powder is obtained by absorption and reduction with hydrochloric acid; the water leaching residue is mainly composed of rhodium and iridium hydroxides / oxides, which are dissolved by heating with concentrated hydrochloric acid or dilute aqua regia to obtain a chloride solution containing rhodium and iridium. Rhodium powder and iridium powder are then obtained by conventional ammonium chloride precipitation, extraction or displacement calcination reduction (or directly reduced to obtain rhodium-ruthenium-iridium enriched gold powder with a total platinum group metal mass percentage of 75%~85%).

[0061] Benefically, the alkaline fusion activation open circuit method overcomes the stubborn problem that platinum group metals such as rhodium, ruthenium, and iridium are extremely chemically inert and cannot be dissolved by traditional acid methods, and achieves efficient and complete recovery of all components of platinum group metals (Pt, Pd, Rh, Ru, Ir) in tin-containing matte.

[0062] Specific embodiments are provided below. These embodiments are intended to enable those skilled in the art to more fully understand the present invention, but do not limit the present invention in any way.

[0063] In the following examples and comparative examples, the elemental analysis, phase identification, and performance index determination all followed the following standards: 1. Grade / content of platinum group metals (PGMs: Pt, Pd, Rh, Ru, Ir) and gold (Au): determined by inductively coupled plasma atomic emission spectrometry (ICP-OES).

[0064] 2. Residual amounts of impurity elements (Sn, Fe, Cu): Quantitative analysis was performed using flame atomic absorption spectrometry (AAS) and ICP-OES.

[0065] 3. Endpoint potential of reduction / precipitation reaction: The entire process is monitored online using a high-precision redox potentiometer (ORP meter with saturated calomel electrode).

[0066] 4. PGMs comprehensive recovery rate: Calculated based on the ratio of the total mass of PGMs in the initial raw materials to the total mass of PGMs in the final extracted high-purity precious metal products (and enriched gold powder).

[0067] 5. Complete dissolution time of aqua regia: determined by a combination of visual inspection and ICP-OES. During the dissolution process at 90℃ with stirring, a small amount of suspension was taken every 0.5 hours, filtered through a 0.22μm filter membrane, and the concentrations of Pt and Pd in ​​the filtrate were determined by ICP-OES. The endpoint of complete dissolution of aqua regia was recorded when the concentration change rate of two consecutive measurements was within ±1%, and no visible black solid residue remained at the bottom of the container.

[0068] Example 1

[0069] Weigh the tin-containing copper matte raw material (main component) produced in the electric furnace workshop. The eutectic mixture contains 32.5% Cu, 28.1% Fe, 3.8% Sn, 42 g / t Pt, 28 g / t Pd, 8 g / t Rh, 5 g / t Ru, and 3 g / t Ir (total PGMs mass 17.20 g) in 200 kg. The raw materials are ground in a mill and pulverizer, and sieved to obtain copper matte powder with a particle size ≤150 μm (100 mesh).

[0070] The copper matte powder was added to an acid leaching tank, and 20% dilute hydrochloric acid was added at a solid-liquid ratio of 1:6. The mixture was heated to 80°C, and 22 kg of sodium chlorate was slowly added in four batches with stirring. The leaching was maintained at a constant temperature of 80°C with stirring for 6 hours. After the reaction was completed, the mixture was filtered, and the selective leaching rate of Sn in the raw material was measured to be 93.5%. The filtrate was discharged into a base metal recovery system, and the filter residue was washed and dried to obtain 37.0 kg of primary precious metal slag, with a residual Sn content of 0.35 kg.

[0071] The dried primary precious metal slag was mixed with 2.22 kg of quartz sand, 1.48 kg of sodium carbonate, and 2.96 kg of crude copper powder, and placed in a medium-frequency induction furnace. The mixture was heated to 1500℃ and held for 1.5 hours for smelting. During the smelting process, impurities formed slag and floated to the surface, while precious metals settled and accumulated at the bottom. After natural cooling, slag and gold were separated to obtain 3.24 kg of smelted alloy blocks. ICP-OES analysis showed that the total PGMs content in the alloy blocks was 0.50% (containing approximately 16.20 g of PGMs, with a smelting recovery rate of 94.2%).

[0072] The obtained alloy block was mechanically crushed and ground a second time to 100 mesh. Industrial hydrochloric acid with a mass concentration of 15% was added at a solid-liquid ratio of 1:5, and the mixture was stirred and leached at 65°C for 3 hours. The residue was filtered and washed until neutral, yielding 0.87 kg of refined precious metal slag. The total PGMs mass percentage was measured to be 1.83% (containing approximately 15.92 g of PGMs).

[0073] Take 0.80 kg of the above-mentioned precious metal slag and add freshly prepared aqua regia (concentrated hydrochloric acid to concentrated nitric acid in a volume ratio of 3:1) at a solid-liquid ratio of 1:5. Stir and dissolve at 90°C for 6 hours. Filter to obtain 4.0 L of aqua regia solution and 0.12 kg of aqua regia residue.

[0074] The aqua regia solution was heated to remove nitrates until no brown gas was emitted, and the free hydrochloric acid concentration was controlled at 2.0 mol / L. 30 g of anhydrous sodium sulfite was added in portions at 45 °C, and the reaction was stopped when the potential stabilized at 720 mV. The mixture was filtered to obtain sponge gold powder. Hydrochloric acid was added to the filtrate to adjust the free acid concentration to 4.0 mol / L. 240 mL of saturated ammonium chloride solution was added at 70 °C, and the mixture was stirred for 2 hours. After cooling and settling, the filtrate was filtered to obtain a yellow ammonium chloroplatinate precipitate. Concentrated nitric acid was added to the filtrate to adjust the free acid concentration to 2.0 mol / L, and 40 g of ammonium chloride was added at 80 °C. The precipitate was filtered to obtain a red ammonium chloropalladate precipitate.

[0075] 0.12 kg of aqua regia residue was thoroughly mixed with 0.21 kg of sodium hydroxide and 0.09 kg of sodium peroxide, and melted in a muffle furnace at 620 °C for 1.5 hours. After cooling, deionized water at 80 °C was added at a solid-liquid ratio of 1:8, and the mixture was leached for 1.5 hours, followed by filtration. The leaching solution was then subjected to chlorine gas oxidation distillation to absorb and recover ruthenium, yielding 0.80 g of pure ruthenium powder. The leaching residue was dissolved in concentrated hydrochloric acid upon heating, and after precipitation with ammonium chloride and calcination reduction, 1.45 g of rhodium-iridium enriched gold powder (Rh+Ir purity 80.2%) was obtained. The overall recovery rate of PGMs in the entire process was 91.2%.

[0076] Example 2

[0077] Weigh 200 kg of tin-containing copper matte raw material (total mass of PGMs is 17.20 g) from the same batch as in Example 1, and crush and grind it to 100 mesh.

[0078] Add 20% dilute hydrochloric acid at a solid-liquid ratio of 1:6, heat to 80℃, and add 22 kg of sodium chlorate in 4 batches. Leach at a constant temperature with stirring for 6 hours. After filtration and washing, 36.4 kg of primary precious metal slag was obtained. Analysis showed that the residual Sn content in the primary precious metal slag was 0.25 kg.

[0079] After drying the primary precious metal slag, 1.82 kg of quartz sand, 1.09 kg of sodium carbonate, and acid-leached activated self-reinforcing Cu-Sn microalloying scavengers (2.91 kg of copper powder and 0.36 kg of borax) were added. The raw materials were thoroughly dry-mixed and then placed in a medium-frequency induction furnace. A graphite plate was used to maintain a weak reducing atmosphere, and the furnace was smelted at 1350℃ for 1.5 hours. After cooling, the slag and gold were separated, yielding 3.02 kg of smelted alloy blocks. Analysis showed that the total PGMs content in the alloy blocks was 0.54% (approximately 16.31 g of PGMs, with a smelting recovery rate of 94.8%), with only 0.8% Sn and 0.5% Fe residue.

[0080] The alloy block was crushed and ground to 100 mesh, and then leached at 65°C for 3 hours with 15% hydrochloric acid at a solid-liquid ratio of 1:5. After filtration and washing, 0.82 kg of precious metal slag was obtained, and the total PGMs content was measured to be 1.95% (approximately 16.00 g of PGMs).

[0081] Take 0.80 kg of the above-mentioned precious metal residue (approximately 15.61 g of PGMs), add aqua regia (hydrochloric acid to nitric acid volume ratio 3.5:1) at a solid-liquid ratio of 1:4, and stir at 85°C for 4 hours to achieve complete dissolution. Filter to obtain 3.2 L of aqua regia solution and 0.10 kg of aqua regia residue.

[0082] After removing nitrates from the aqua regia solution, the free acid concentration was controlled at 2.0 mol / L. Anhydrous sodium sulfite was added in batches at 45℃ (endpoint potential 715mV) for efficient precipitation and gold extraction. The free acid concentration of the filtrate was adjusted to 4.5 mol / L, and ammonium chloride was added at 70℃ to precipitate and extract platinum. The free acid concentration of the filtrate was adjusted to 2.0 mol / L, and nitric acid was added. Ammonium chloride was added at 80℃ to precipitate and extract palladium.

[0083] 0.10 kg of aquatic residue was melted with 0.175 kg of sodium hydroxide and 0.075 kg of sodium peroxide at 620 °C for 1.5 hours. After water leaching and phase separation, concentrated sulfuric acid was added to the water leaching solution to adjust the pH to 1.5, and the solution was heated to 85 °C. Chlorine gas was introduced for oxidation, distillation, absorption, and reduction to obtain 0.88 g of pure ruthenium powder. The water leaching residue was acid-dissolved to extract rhodium and iridium, yielding 1.85 g of rhodium-iridium enriched gold powder (Rh+Ir purity 81.5%). The overall recovery rate of PGMs in the entire process was 98.6%.

[0084] Example 3

[0085] Weigh 200 kg of tin-containing copper matte raw material (total mass of PGMs is 17.20 g) from the same batch as in Example 1, and grind it to 100 mesh.

[0086] Oxidative acid leaching was performed under the same conditions as in Example 2, yielding 36.6 kg of primary precious metal slag. The residual Sn content in the slag was determined to be 0.08 kg. Adding 1.83 kg of copper oxide powder yields a total of 1.46 kg of pure Cu. Based on the target mass ratio K = 8.1, the required total Sn mass is 1.46 / 8.1 ≈ 0.18 kg. Therefore, the calculated additional tin powder amount is... =0.18-0.08=0.10 kg.

[0087] 1.10 kg of quartz sand and 0.73 kg of sodium carbonate were added according to the slag weight. 1.83 kg of copper oxide powder, 0.18 kg of boric acid, and 0.10 kg of tin powder were added as a collector. The mixture was smelted in a medium-frequency furnace at 1250℃ for 2.0 hours. After cooling, 2.76 kg of alloy blocks were obtained. The total PGMs content was measured to be 0.58% (approximately 16.00 g of PGMs), with 1.1% Sn residue and 0.8% Fe residue.

[0088] The conditions for subsequent secondary acid leaching for impurity removal, aqua regia dissolution, and refining extraction were the same as in Example 2.

[0089] Example 4

[0090] Weigh 200 kg of tin-containing copper matte raw material (total mass of PGMs is 17.20 g) from the same batch as in Example 1, and grind it to 100 mesh.

[0091] Oxidative acid leaching was carried out under the same acid leaching conditions as in Example 2, yielding 36.2 kg of primary precious metal slag.

[0092] Add 2.90 kg of quartz sand and 2.17 kg of sodium carbonate by weight of slag, and add 5.43 kg of copper powder and fluorite powder as a collector. 0.54 kg. Melted in a medium-frequency furnace at 1450℃ for 1.0 hour. After cooling, 4.50 kg of alloy block was obtained. The total PGMs content was measured to be 0.36% (approximately 16.20 g of PGMs), with Sn residue of 0.6% and Fe residue of 0.3%.

[0093] The subsequent secondary acid leaching for impurity removal, aqua regia dissolution, and refining extraction processes were all conducted under the same conditions as in Example 2, and the final overall PGMs recovery rate was 97.5%.

[0094] Comparative Example 1 Weigh 200 kg of tin-containing copper matte raw material from the same batch as in Example 1 and grind it to 100 mesh.

[0095] Oxidative acid leaching was carried out under the same acid leaching conditions as in Example 1 (20% dilute hydrochloric acid, solid-liquid ratio 1:6, leaching with 11% sodium chlorate in batches at 80°C for 6 hours) to obtain 37.0 kg of primary precious metal slag.

[0096] The same slag-forming and scavenging agents as in Example 1 were added: 2.22 kg of quartz sand, 1.48 kg of sodium carbonate, and 2.96 kg of crude copper powder. The mixture was heated to 1350°C and held for 1.5 hours in a medium-frequency induction furnace. After cooling, the slag-metal interface became blurred, and only 2.15 kg of crude alloy blocks separated. ICP-OES analysis showed that the total PGMs content in the alloy blocks was 0.23% (PGMs content was 4.95 g), and the alloy contained 3.8% Sn and 2.9% Fe residues.

[0097] Subsequently, a second acid leaching to remove impurities and aqua regia dissolution and refining extraction were carried out under the conditions of Example 1. The aqua regia dissolution was slow and incomplete, and the final comprehensive recovery rate of PGMs was 58.3%.

[0098] Comparative Example 2 Weigh 200 kg of tin-containing copper matte raw material from the same batch as in Example 1 and grind it to 100 mesh.

[0099] Without adding dilute hydrochloric acid and sodium chlorate, only deionized water was added at a solid-liquid ratio of 1:6, and the mixture was stirred and washed at 80°C for 6 hours. After filtration and drying, 192 kg of washing residue was obtained.

[0100] Quartz sand, sodium carbonate, and Cu-Sn microalloying scavenger (15.36 kg copper powder and 1.92 kg borax) were added in the same proportions as in Example 2. The mixture was smelted in a medium-frequency furnace at 1350°C for 1.5 hours. The slag was viscous, and slag-metal separation was difficult after cooling, yielding 11.6 kg of alloy blocks. The total PGMs content was measured to be 0.03% (PGMs content was 3.48 g).

[0101] The obtained alloy block was mechanically crushed and ground to 100 mesh. 15% industrial hydrochloric acid was added at a solid-liquid ratio of 1:5, and the mixture was stirred and leached at 65°C for 3 hours. After filtration and washing, 10.8 kg of secondary acid leaching residue was obtained.

[0102] Take 10.0 kg of the above-mentioned secondary acid leaching residue, add aqua regia (concentrated hydrochloric acid to concentrated nitric acid volume ratio 3.5:1) at a solid-liquid ratio of 1:4, and stir to dissolve for 4 hours at 85℃. Filter to obtain aqua regia solution. After subsequent gold, platinum, and palladium precipitation, the final comprehensive recovery rate of PGMs was 32.4%.

[0103] Comparative Example 3 Weigh 200 kg of tin-containing copper matte raw material from the same batch as in Example 1 and grind it to 100 mesh.

[0104] Oxidative acid leaching was carried out under the same acid leaching conditions as in Example 2, yielding 36.4 kg of precious metal slag.

[0105] 1.82 kg of quartz sand and 1.09 kg of sodium carbonate were added by weight of slag, and 2.91 kg of crude copper powder was added as a collector. No boron source or calcium fluoride was added. The alloy was smelted in a medium-frequency furnace at 1350℃ for 1.5 hours. After cooling, 3.70 kg of alloy blocks were obtained. The total PGMs content was determined to be 0.40% (PGMs content was 14.80 g), and the alloy contained 4.5% Sn and 2.3% Fe.

[0106] The alloy blocks were mechanically crushed and ground to 100 mesh. Industrial hydrochloric acid with a mass concentration of 15% was added at a solid-liquid ratio of 1:5, and the mixture was stirred and leached at 65°C for 3 hours. After filtration and washing, 1.12 kg of secondary acid leaching residue was obtained. The Sn residue in the secondary acid leaching residue still reached 3.8%.

[0107] Take 1.0 kg of the above-mentioned secondary acid leaching residue, add aqua regia (concentrated hydrochloric acid to concentrated nitric acid, volume ratio 3.5:1) at a solid-liquid ratio of 1:4, and stir to dissolve for 4 hours at 85°C. During the dissolution process, the residue dissolution rate is 65.2% after 4 hours of reaction.

[0108] The obtained aqua regia solution was extracted in stages. The purity of the gold powder precipitate was 96.2%, the purity of the platinum / palladium was 98.85%, and the final comprehensive recovery rate of PGMs was 82.1%.

[0109] The performance test indicators of the above embodiments and comparative examples are shown in Table 1 below:

[0110] This article uses specific examples to illustrate the principles and implementation methods of the present invention. The descriptions of the above embodiments are only for the purpose of helping to understand the method and core ideas of the present invention. The above are merely preferred embodiments of the present invention, but the scope of protection of the present invention is not limited thereto. Any equivalent substitutions or modifications made by those skilled in the art within the technical scope disclosed in the present invention, based on the technical solution and inventive concept of the present invention, should be within the scope of protection of the present invention.

Claims

1. A method for extracting platinum group metals from copper matte, characterized in that, Includes the following steps: S1. Crush the tin-containing copper matte raw material, add dilute hydrochloric acid and sodium chlorate at a solid-liquid ratio of 1:5 to 1:8, stir and leach at 60 to 90°C, and filter to obtain primary precious metal slag. S2. The primary precious metal slag is dried, slag-forming agent and scavenging agent are added, and it is smelted in a medium frequency furnace at 1250~1550℃. After the impurities are slagified and floated to the surface, the bottom is cooled to obtain an alloy block containing platinum group metals. S3. The alloy block is crushed, leached with dilute hydrochloric acid, filtered and washed to obtain precious metal slag; S4. Dissolve the precious metal slag in aqua regia, filter to obtain aqua regia solution and aqua regia slag; add a reducing agent to the aqua regia solution to extract gold, platinum and palladium in stages; perform alkaline fusion activation on the aqua regia slag to extract rhodium, ruthenium and iridium.

2. The method according to claim 1, characterized in that, In step S1, the mass concentration of the dilute hydrochloric acid is 15%~25%; the sodium chlorate is added in 3~5 batches, and the amount added is 8%~15% of the mass of the tin-containing copper matte raw material; the leaching time is 4~8 hours.

3. The method according to claim 1, characterized in that, In step S2, based on the mass of the primary precious metal slag, the collector includes: 5%~15% of added copper source, 0.5%~2.0% of boron source or 0.3%~1.5% of calcium fluoride; before smelting, the residual tin content in the primary precious metal slag is detected, and the tin source is controlled or supplemented according to the mass ratio of total Sn to Cu in the added copper source of 1:8~1:

15.

4. The method according to claim 3, characterized in that, In step S2, the external copper source is copper powder or copper oxide; the boron source is borax or boric acid; the melting temperature is 1250~1450℃, and the melting time is 1~2 hours.

5. The method according to claim 1, characterized in that, In step S2, based on the mass of the primary precious metal slag, the slag-forming agent includes 3% to 8% quartz sand and 2% to 6% sodium carbonate; the collecting agent is copper or lead.

6. The method according to claim 1, characterized in that, In step S3, the alloy block is crushed to 100 mesh and then leached for 2-4 hours at 50-80°C using hydrochloric acid with a mass concentration of 10%-20% at a solid-liquid ratio of 1:4-1:

6.

7. The method according to claim 3 or 4, characterized in that, In step S4, the aqua regia is a mixture of concentrated hydrochloric acid and concentrated nitric acid in a volume ratio of 3:1 to 4:1; the aqua regia is added according to the solid-liquid ratio of the precious metal slag to the aqua regia of 1:3 to 1:6, and dissolved at 80 to 95°C for 3 to 6 hours.

8. The method according to claim 1, characterized in that, In step S4, the specific steps for the fractional extraction of gold, platinum, and palladium from the aqua regia are as follows: Gold extraction: Heat the aqua regia solution to remove nitrates, control the concentration of free hydrochloric acid to 1~3 mol / L, add anhydrous sodium sulfite or sodium bisulfite in batches at 40~50℃ at 1.2~1.5 times the theoretical amount of gold, control the endpoint potential to 700~750 mV, and filter to obtain gold powder. Platinum extraction: Maintain the free hydrochloric acid concentration in the filtrate at 3~6 mol / L, heat to 50~80℃, add ammonium chloride and react for 1~3 hours, let stand to precipitate to obtain ammonium chloroplatinate; Palladium extraction: Add nitric acid to the filtrate after platinum extraction to adjust the free acid concentration to 1-3 mol / L, and add ammonium chloride at 60-90℃ to precipitate and extract palladium.

9. The method according to claim 1, characterized in that, In step S4, the specific steps for extracting rhodium, ruthenium, and iridium from the aqua regia residue are as follows: The aqua regia residue is mixed with sodium hydroxide and sodium peroxide at a mass ratio of 1:(1.5~2):(0.5~1) and melted at 550~700℃ for 1~2 hours; after the melt is cooled, it is leached with water at a solid-liquid ratio of 1:5~1:10, filtered, and the water leaching solution and water leaching residue are obtained. Add concentrated sulfuric acid or concentrated hydrochloric acid to the aqueous extract to adjust the pH to 1-3, heat to 80-95℃, and then pass chlorine gas through for oxidative distillation. The distilled ruthenium tetroxide is absorbed by hydrochloric acid solution and reduced to recover the ruthenium product. The water-leached residue is dissolved in concentrated hydrochloric acid to obtain a rhodium-iridium-containing solution. Ammonium chloride is added to the rhodium-iridium-containing solution to precipitate the residue. The precipitate is then calcined and reduced to obtain rhodium-iridium-enriched gold powder or rhodium / iridium products.

10. The method according to claim 9, characterized in that, The rhodium and iridium-containing solution is precipitated with ammonium chloride and then calcined and reduced to obtain rhodium-iridium enriched gold powder, wherein the total mass percentage of rhodium and iridium is 75%~85%.

Images