Efficient separation and comprehensive recovery method based on copper-nickel complex platinum-palladium concentrate

Abstract

The invention relates to an efficient separation and comprehensive recovery method based on copper-nickel complex platinum-palladium concentrate. The efficient separation and comprehensive recovery method comprises the following steps: (1) smelting and enriching; (2) carrying out two-stage cascade blowing enrichment; (3) stage selective crystallization; (4) separating mixed ore; (5) selective separation of alloyed gold; (6) separating copper-nickel bulk concentrate; and (7) mixing the nickel sulfate crystal obtained in the step (5) and the nickel oxide obtained in the step (6), feeding the mixture into a reverberatory furnace, smelting to obtain a nickel anode plate with crude nickel as a main component, and then feeding the nickel anode plate into an electrolysis process for electrolysis to obtain cathode nickel, anode mud and silver-cobalt slag products. According to the method, valuable metals such as platinum, palladium, gold, silver, copper, nickel and cobalt in the complex platinum-palladium concentrate can be efficiently separated and recovered, the separation effect is good, the comprehensive recovery rate is high, comprehensive recycling of resources is fully achieved, the problems that an existing recovery process is poor in recovery effect and high in cost are solved, and good economic benefits and social benefits are achieved.

Description

Technical Field

[0001] The present invention belongs to the technical field of precious metal metallurgy, and particularly relates to an efficient separation and comprehensive recovery method based on copper-nickel complex platinum-palladium concentrate. Background Art

[0002] This type of platinum-palladium concentrate is rich in valuable metals such as platinum, palladium, gold, silver, copper, and nickel. Traditionally, roasting leaching or all-wet leaching is used, resulting in a large amount of slag, complex leaching solution components, difficult waste liquid treatment, or difficult recovery of poor copper, nickel, and cobalt during the complex process of recovering precious metals, or difficult recovery of the main component precious metals with a simple process for recovering poor copper, nickel, and cobalt, and high production costs, leading to ineffective comprehensive recovery of resources. Therefore, exploring a scientific and reasonable recovery process has become an urgent technical problem in the industry.

[0003] A method for separating and recovering nickel, copper, cobalt, and precious metals such as gold, silver, platinum, and palladium from platinum-palladium concentrate disclosed in CN117051252A. This patent first obtains matte through primary smelting and primary converting, then conducts high-temperature and high-pressure oxidative acid leaching using high-temperature and normal pressure and autoclave, and then separates and recovers valuable components through fractional crystallization. Although better recovery indexes can be obtained for platinum and palladium, high-temperature and high-pressure leaching results in complex leaching solution components, and there is serious interference between different components during the fractional crystallization process, easily forming mixed crystals, and poor recovery indexes for low-content metals. Secondly, this method uses high-pressure equipment to oxidize sulfides at high temperatures, requiring high design and material requirements for process equipment, higher capital investment, and corresponding improvement in operation and maintenance levels. In addition, this method completely returns the primary converting slag to the smelting process for re-treatment, which is bound to increase the smelting charge and lead to increased smelting costs.

[0004] Based on the existing defects and deficiencies, the present invention aims to provide an efficient separation and comprehensive recovery method based on copper-nickel complex platinum-palladium concentrate with scientific and reasonable process layout, simple equipment, and low production costs. Summary of the Invention

[0005] The present invention provides an efficient separation and comprehensive recovery method based on copper-nickel complex platinum-palladium concentrate, mainly targeting complex platinum-palladium concentrate with a copper: nickel ratio of 1:1.1 - 2.5. A combined process of smelting enrichment, two-stage cascade converting enrichment, stage selective crystallization, crushing-grinding - weak magnetic field pulsed high-gradient magnetic separation, roasting - sulfuric acid selective leaching combined with atmospheric pressure separation, and classified refining of separation products is successively adopted to efficiently recover valuable metals such as platinum, palladium, gold, silver, copper, nickel, and cobalt.

[0006] The specific technical solution is as follows: The present invention provides an efficient separation and comprehensive recovery method based on copper-nickel complex platinum-palladium concentrate, including the following steps: (1) Smelting enrichment: Mix the platinum-palladium concentrate to be treated with a slag former and pulverized coal for smelting to obtain low-grade matte, slag, and sulfur dioxide flue gas; (2) Two-stage cascade smelting enrichment: The low-grade matte obtained in step (1) is charged with silica and lump coal for primary converter smelting. When the primary smelting is approaching the end point, it is promptly transferred to another converter for short-time secondary smelting to obtain a high-grade matte liquid with an iron content of about 10% and a sulfur content of 22% - 25% at the bottom, iron oxide slag (for recycling), and sulfur dioxide flue gas. By combining smelting with two-stage smelting, the iron content and sulfur content in the matte liquid can be increased, enabling more iron to exist in the form of native gold, which is beneficial to improving the subsequent native gold yield, taking into account the noble metal enrichment ratio and recovery rate, reducing the content of dispersed noble metals in the slag, fully ensuring the recovery of valuable components in the matte liquid, and enhancing the differentiation efficiency. At the same time, the formation of iron sulfides and overly fine mineral crystals (less than 0.01 grains) is avoided, providing favorable conditions for the next separation step. It can also significantly reduce the amount of iron oxide slag returned for smelting and lower the smelting production cost.

[0007] (3) Stage selective crystallization: The high-grade matte liquid obtained in step (2) is promptly transported to a stage temperature-controlled cooling furnace for selective crystallization to obtain a differentiated structure large-particle mineral-bound matte ore (including native gold, cuprous sulfide, and nickel sulfide ore). By combining stage cooling with smelting and two-stage smelting, through the synergistic effect, a combination of high-content platinum-palladium native gold with relatively thick single crystals and simple structure and doped extremely low copper sulfide crystals and nickel sulfide crystals is formed, effectively avoiding or reducing the formation of mixed-melting crystals and overly fine mineral crystals (less than 0.01 grains). This combination is easily broken along its grain boundaries, which is very beneficial for the separation between different chemical phases and creates favorable conditions for subsequent mechanical and chemical separation.

[0008] (4) Separation of mixed ore: The differentiated structure matte ore obtained in step (3) is transported to a grinding - weak magnetic field pulsed high-gradient magnetic separation (improved type) separation system for separation. First, the differentiated structure matte ore is finely ground, and after fine grinding, it undergoes weak magnetic field pulsed high-gradient magnetic separation to obtain native gold (platinum-palladium-gold-copper-nickel-iron) and copper-nickel mixed concentrate. Among them, the improvement points of the weak magnetic field pulsed high-gradient magnetic separator mainly lie in that the diameter and spacing of the magnetic medium rods in the upper half of the magnetic box are 3 - 5 mm, the diameter and spacing of the magnetic medium rods in the lower half are 4 - 8 mm, and a grid sieve with a sieve hole of 1 mm is installed on the first layer of magnetic medium rods in the upper part.

[0009] This step applies weak magnetic field pulsed high-gradient magnetic separation to the separation of native gold and sulfide ore, greatly improving the separation efficiency, avoiding the technical defect that the flaky coarse native gold generated during the grinding process is prone to clogging the magnetic medium box, and simplifying the process flow.

[0010] (5)Selective separation of crude gold: The crude gold obtained in step (4) is subjected to atmospheric pressure heating acid leaching - filtration - precipitation - filtration - crystallization. Sulfuric acid is added to the crude gold, and after heating acid leaching, filtration is carried out to obtain crude gold (platinum, palladium, gold, copper); continue to inflate, then add nickel carbonate to remove iron, and filter - wash to remove iron to obtain iron oxide slag; the washed mother liquor is then subjected to nickel sulfate crystallization to obtain nickel sulfate crystals and nickel sulfate crystallization mother liquor (for recycling); (6)Separation of copper - nickel mixed concentrate: The copper - nickel mixed concentrate obtained in step (4) is roasted, and the roasted slag is then subjected to acid leaching, filtration, and crystallization to obtain nickel oxide, copper sulfate crystals, and crystallization mother liquor (for recycling); (7)Mix the nickel sulfate crystals obtained in step (5) with the nickel oxide obtained in step (6) and send them to a reverberatory furnace for smelting to obtain nickel anode plates mainly composed of crude nickel. Then, the nickel anode plates are sent to the electrolysis process for electrolysis to obtain cathode nickel, anode slime, and silver - cobalt slag products.

[0011] Further, the slag - forming agent in step (1) is selected from one or two of limestone, silicon dioxide, and iron(III) oxide.

[0012] Furthermore, the dosages of the slag - forming agents are as follows: limestone is 7% - 12% of the weight of the platinum - palladium concentrate, silicon dioxide is 20% - 25% of the weight of the platinum - palladium concentrate, and iron(III) oxide is 7% - 15% of the weight of the platinum - palladium concentrate.

[0013] Furthermore, the iron(III) oxide can be selected from converter iron oxide slag or crystallization iron - removing slag.

[0014] Further, the smelting temperature in step (1) is 1300 - 1450 °C Further, in step (2), the blowing temperature of the first converter is 1200 - 1300 °C, and the dosage of silicon dioxide is 15% - 30% of the weight of the smelting product; the blowing temperature of the second converter is 1100 - 1200 °C.

[0015] Further, the specific cooling process in step (3) is as follows: the slow cooling time from 1100 °C to 580 °C is 1 - 2 days, the slow cooling time from 580 °C to 520 °C is 4 - 5 days, the slow cooling time from 520 °C to 370 °C is 1 - 2 days, and then natural cooling.

[0016] Further, in step (4), the fine grinding particle size is such that - 0.074 mm accounts for 50% - 65%; the magnetic separation concentration is 20% - 25%; the magnetic induction intensity of the weak - magnetic - field pulsed high - gradient magnetic separator is 100 - 200 mT, and the pulse intensity is 100 - 180 times / min.

[0017] Furthermore, the diameters and spacings of the magnetic medium rods in the upper half of the mass box of the weak magnetic field pulse high-gradient magnetic separator are 3 - 5 mm, and those in the lower half are 4 - 8 mm. A grid sieve with a sieve hole of 1 mm is installed on the first layer of magnetic medium rods in the upper part.

[0018] Further, in step (5), the pH of the heating acid leaching is 1 - 2, the leaching temperature is 65 - 85 °C, the liquid-solid ratio is 1:1 - 4:1, and the leaching time is 2 - 3 hours; for iron removal, nickel carbonate is slowly used to adjust the pH to 3.0 - 4.7.

[0019] Further, in step (6), the roasting temperature is 750 - 850 °C; the liquid-solid ratio of the acid leaching is 2:1 - 5:1, the pH is 1 - 3, and the leaching temperature is 65 - 85 °C.

[0020] Further, a small amount of elemental copper contained in the crude gold (platinum-palladium-gold-copper) in step (5) is recycled in the subsequent traditional precious metal smelting process.

[0021] Further, the sulfur dioxide-containing flue gas generated in steps (1) and (2) is dust-removed by gravity, cyclone dust collectors, electrostatic precipitators and bag filters, and the flue gas with a sulfur dioxide concentration of more than 8.0% and dust are obtained. The flue gas can be transported to a sulfuric acid plant to produce sulfuric acid, and the dust can be made into pellets and returned to step (1) for reuse.

[0022] The beneficial effects of the present invention: The present invention adopts a combined process of smelting enrichment, two-stage cascade blowing enrichment and impurity removal, stage selective crystallization, weak magnetic field pulse high-gradient magnetic separation, roasting-sulfuric acid selective leaching combination for atmospheric pressure separation and classification and refining of separated products for complex platinum-palladium concentrate based on copper-nickel. It can efficiently separate and recover valuable metals such as platinum, palladium, gold, silver, copper, nickel and cobalt in the complex platinum-palladium concentrate, with good separation effect and high comprehensive recovery rate, fully realizing the comprehensive recovery and utilization of resources, solving the problems of poor recovery effect and high cost of the existing recovery process, and having good economic and social benefits.

[0023] The main technical effects of the present invention are as follows: (1) The present invention skillfully combines smelting and two-stage cascade blowing. When the first blowing is approaching the end, it is timely transferred to another converter and continuously blown for a short time at 1100 - 1200 °C for secondary blowing, and a bottom matte liquid with an iron content of about 10% and a sulfur content of 22% - 25% can be obtained, which can effectively increase the iron and sulfur contents in the matte liquid, make more iron exist in the form of crude gold, be conducive to improving the subsequent crude gold yield, taking into account the enrichment ratio and recovery rate of precious metals, reducing the content of dispersed precious metals in the slag, fully ensuring the recovery of valuable components in the matte liquid, and enhancing the differentiation efficiency. At the same time, it avoids the formation of iron-containing sulfides and overly fine mineral crystals (less than 0.01 grain size), provides favorable conditions for the next separation, and can also greatly reduce the amount of iron oxide slag returned to smelting, reducing the smelting production cost.

[0024] (2) The present invention combines stage cooling with smelting and two-stage cascade blowing. Through the synergistic effect, a large-grained mineral aggregate with a differentiated structure is formed, where the single crystals are relatively thick and the structure is simple. This effectively avoids or reduces the formation of mixed-melting crystals and overly fine mineral crystals (less than 0.01 grain size), facilitating fracture along the grain boundaries during the crushing and grinding processes, and providing sufficient separation conditions for subsequent processes.

[0025] (3) The present invention ingeniously applies weak magnetic field pulsed high-gradient magnetic separation to the separation of native gold and sulfide ores, significantly improving the separation efficiency, avoiding the technical defect that the flaky coarse-grained native gold generated during the grinding process is prone to clogging the magnetic medium box, and simplifying the process flow.

[0026] (4) For the differentiated structure mixed ore, the present invention ingeniously combines grinding - pulsed high-gradient magnetic separation with roasting - sulfuric acid selective leaching for separation. Under the condition of appropriate grinding fineness, it improves the recovery rate of precious metals and the separation effect of associated valuable components, simplifies the process flow, and effectively avoids the defects such as high requirements for equipment and materials, high energy consumption, high reagent consumption, high requirements for operation level and infrastructure, and high difficulty in industrial production control when using high-temperature and high-pressure oxidative leaching. Description of the Drawings

[0027] Figure 1 is a flow chart of an efficient separation and comprehensive recovery method for a copper-nickel complex platinum-palladium concentrate according to the present invention. Detailed Embodiments

[0028] In order to make the technical problems and technical solutions solved by the present invention clearer, the present invention will be further described in detail below with reference to the drawings and embodiments. It should be understood that the specific embodiments described herein are only used to explain the present invention and are not used to limit the present invention. Embodiment 1

[0029] Raw material 1#: A copper-nickel complex platinum-palladium concentrate, with the main elements and their contents being: platinum 26.68 g / t, palladium 36.89 g / t, gold 5 g / t, silver 15 g / t, copper 2.34%, nickel 2.50%, cobalt 0.70%, iron 18.26%, sulfur 14.62%, and magnesium 11.29%.

[0030] As Figure 1 shown, the present invention is implemented on this complex platinum-palladium concentrate, and the specific steps are as follows: (1)Smelting enrichment: The platinum-palladium concentrate to be processed is mixed with limestone, silica, and iron(III) oxide or preferably with converter iron oxide slag (select one or two according to the material properties) and pulverized coal for smelting to obtain low-grade matte, slag, and sulfur dioxide flue gas. During this process, the smelting temperature is 1400 °C, and the dosages of the smelting slag formers are as follows: limestone is 8% of the weight of the platinum-palladium concentrate, silica is 20% of the weight of the platinum-palladium concentrate, and the dosage of iron(III) oxide (or preferably converter iron oxide slag and iron-removing slag by crystallization) is 7.5% of the weight of the platinum-palladium concentrate.

[0031] (2)Two-stage cascade smelting enrichment: The low-grade matte obtained in step (1) is charged with silica and lump coal for primary converter smelting. When the primary smelting is approaching the end point, it is promptly transferred to another converter for a short secondary smelting to obtain a bottom-settling high-grade matte liquid with an iron content of about 10% and a sulfur content of 22% - 25%, iron oxide slag (for recycling), and sulfur dioxide flue gas. During this process, the temperature of the primary converter smelting is 1280 °C, and the dosage of silica for the primary converter smelting is 23% of the weight of the smelting product; the temperature of the secondary converter smelting is 1180 °C.

[0032] (3)Stage selective crystallization: The bottom-settling high-grade matte liquid obtained in step (2) is promptly transported to a stage temperature-controlled cooling furnace for selective crystallization to obtain a large-particle mineral-bound matte ore with a differentiated structure (including native gold, cuprous sulfide, and nickel sulfide ore). The specific process of stage temperature-controlled cooling is as follows: the slow cooling time from 1180 °C to 580 °C is 1 day, the slow cooling time from 580 °C to 520 °C is 4 days, the slow cooling time from 520 °C to 370 °C is 1 day, and then natural cooling is carried out.

[0033] (4)Separation of mixed ore: The differentiated structure matte ore obtained in step (3) is transported to a grinding - weak magnetic field pulsed high-gradient magnetic separation (improved type) separation system for separation. First, the differentiated structure matte ore is finely ground, and after fine grinding, it is separated by weak magnetic field pulsed high-gradient magnetic separation to obtain native gold (platinum-palladium-gold-copper-nickel-iron) and copper-nickel mixed concentrate. During this process, the fineness of fine grinding is -0.074 mm accounting for 50% - 65%, the magnetic separation concentration is 20% - 25%, the magnetic induction intensity of the weak magnetic field pulsed high-gradient magnetic separator (improved type) is 100 - 200 mT, the diameter of the magnetic medium and the working gap are 3 - 5 mm, and the pulse intensity is 100 - 180 times / min; (5)Selective separation of crude gold: The crude gold obtained in step (4) is subjected to atmospheric pressure heating acid leaching - filtration - precipitation - filtration - crystallization. Sulfuric acid is added to the crude gold, and after heating acid leaching, filtration is carried out to obtain crude gold (platinum - palladium - gold - copper); continue to inflate, then add nickel carbonate to remove iron, and filter - wash to remove iron to obtain iron oxide sediment; the washed mother liquor is then subjected to nickel sulfate crystallization to obtain nickel sulfate crystals and nickel sulfate crystallization mother liquor (for recycling). In this process, the pH of heating acid leaching is 1 - 2, the leaching temperature is 65 - 85 °C, the liquid - solid ratio is 1:1 - 4:1, and the leaching time is 2 - 3 hours; nickel carbonate is used to slowly adjust the pH to 3.0 - 4.7 for iron removal.

[0034] (6)Separation of copper - nickel mixed concentrate: The copper - nickel mixed concentrate obtained in step (4) is roasted, and the roasted slag is then subjected to acid leaching, filtration, and crystallization to obtain nickel oxide, copper sulfate crystals, and crystallization mother liquor (for recycling). In this process, the roasting temperature is 750 - 850 °C, the liquid - solid ratio is 2:1 - 5:1, the pH of acid leaching is 1 - 3, and the leaching temperature is 65 - 85 °C.

[0035] (7)Mix the nickel sulfate crystals obtained in step (5) with the nickel oxide obtained in step (6) and send them to a reverberatory furnace for smelting to obtain nickel anode plates mainly composed of crude nickel. Then, the nickel anode plates are sent to the electrolysis process for electrolysis to obtain cathode nickel, anode slime, and silver - cobalt slag products.

[0036] In addition, a small amount of elemental copper contained in the crude gold (platinum - palladium - gold - copper) in step (5) is recycled in the subsequent traditional precious metal smelting process. After the sulfur dioxide - containing flue gases generated in steps (1) and (2) are combined, after dust removal by gravity, cyclone dust collectors, electrostatic precipitators, and bag filters, flue gas with a sulfur dioxide concentration of more than 8.0% and soot are obtained. Then, the flue gas is transported to a sulfuric acid plant to produce sulfuric acid, and the soot can be made into pellets and returned to step (1) for reuse.

[0037] The test results obtained are as follows: the comprehensive recovery rate of platinum and palladium is 96.17%, the comprehensive recovery rate of gold and silver is 94.88%, the recovery rate of copper is 90.67%, the recovery rate of nickel is 90.25%, and the recovery rate of cobalt is 70.45%. Example 2

[0038] Raw material 2#: A copper - nickel complex platinum - palladium concentrate, with the main elements and contents being: platinum 37.45 g / t, palladium 47.48 g / t, gold 7 g / t, silver 17 g / t, copper 2.08%, nickel 2.40%, cobalt 0.50%, iron 16.75%, sulfur 12.67%, magnesium 8.27%.

[0039] As Figure 1 shown, implementing the present invention on this complex platinum - palladium concentrate, the specific steps are as follows: (1)Smelting enrichment: The platinum-palladium concentrate to be treated is mixed with limestone, silica, and iron(III) oxide or preferably converter iron oxide slag (select one or two according to the material properties) and pulverized coal for smelting to obtain low-grade matte, slag, and sulfur dioxide flue gas. During this process, the smelting temperature is 1380 °C, and the dosages of the smelting slag formers are as follows: limestone is 8% of the weight of the platinum-palladium concentrate, silica is 21% of the weight of the platinum-palladium concentrate, and iron(III) oxide (or preferably converter iron oxide slag and iron-removing slag by crystallization) is 8% of the weight of the platinum-palladium concentrate.

[0040] (2)Two-stage cascade blowing enrichment: The low-grade matte obtained in step (1) is charged with silica and lump coal for primary converter blowing. When the primary blowing is approaching the end point, it is promptly transferred to another converter for a short secondary blowing to obtain a bottom-settling high-grade matte liquid with an iron content of about 10% and a sulfur content of 22% - 25%, iron oxide slag (for recycling), and sulfur dioxide flue gas. During this process, the primary converter blowing temperature is 1280 °C, and the dosage of silica for primary converter blowing is 22% of the weight of the smelting product; the secondary converter blowing temperature is 1150 °C.

[0041] (3)Stage selective crystallization: The bottom-settling high-grade matte liquid obtained in step (2) is promptly transported to a stage temperature-controlled cooling furnace for selective crystallization to obtain a large-particle mineral-bound matte ore with a differentiated structure (including alloyed gold, cuprous sulfide, and nickel sulfide ore). The specific process of stage temperature-controlled cooling is as follows: the slow cooling time from 1150 °C to 580 °C is 1 day, the slow cooling time from 580 °C to 520 °C is 4 days, the slow cooling time from 520 °C to 370 °C is 1 day, and then natural cooling can be carried out.

[0042] (4)Separation of mixed ore: The differentiated-structure matte ore obtained in step (3) is transported to a grinding - weak magnetic field pulsed high-gradient magnetic separation (improved type) separation system for separation. First, the differentiated-structure matte ore is finely ground. After fine grinding, it is separated by weak magnetic field pulsed high-gradient magnetic separation to obtain alloyed gold (platinum-palladium-gold-copper-nickel-iron) and copper-nickel mixed concentrate. During this process, the fineness of fine grinding is that -0.074 mm accounts for 50% - 65%, the magnetic separation concentration is 20% - 25%, the magnetic induction intensity of the weak magnetic field pulsed high-gradient magnetic separator (improved type) is 100 - 200 mT, the diameter of the magnetic medium and the working gap are 3 - 5 mm, and the pulse intensity is 100 - 180 times / min; (5)Selective separation of crude gold: The crude gold obtained in step (4) is subjected to atmospheric pressure heating acid leaching - filtration - precipitation - filtration - crystallization. Sulfuric acid is added to the crude gold, and after heating acid leaching, filtration is carried out to obtain crude gold (platinum - palladium - gold - copper); continue to inflate, then add nickel carbonate to remove iron, filter - wash to remove iron, and obtain iron oxide sediment; the washed mother liquor is then subjected to nickel sulfate crystallization to obtain nickel sulfate crystals and nickel sulfate crystallization mother liquor (for recycling). During this process, the pH of the heating acid leaching is 1 - 2, the leaching temperature is 65 - 85 °C, the liquid - solid ratio is 1:1 - 4:1, and the leaching time is 2 - 3 hours; for iron removal, nickel carbonate is slowly used to adjust the pH to 3.0 - 4.7.

[0043] (6)Separation of copper - nickel mixed concentrate: The copper - nickel mixed concentrate obtained in step (4) is roasted, and the roasted slag is then subjected to acid leaching, filtration, and crystallization to obtain nickel oxide, copper sulfate crystals, and crystallization mother liquor (for recycling). During this process, the roasting temperature is 750 - 850 °C, the liquid - solid ratio is 2:1 - 5:1, the pH of acid leaching is 1 - 3, and the leaching temperature is 65 - 85 °C.

[0044] (7)Mix the nickel sulfate crystals obtained in step (5) with the nickel oxide obtained in step (6) and send them to a reverberatory furnace for smelting to obtain nickel anode plates with the main component being crude nickel. Then, the nickel anode plates are sent to the electrolysis process for electrolysis to obtain cathode nickel, anode slime, and silver - cobalt slag products.

[0045] In addition, a small amount of elemental copper contained in the crude gold (platinum - palladium - gold - copper) in step (5) is recycled in the subsequent traditional precious metal smelting process. After the sulfur dioxide - containing flue gases generated in steps (1) and (2) are combined, after dust removal by gravity, cyclone dust collectors, electrostatic precipitators, and bag filters, flue gas with a sulfur dioxide concentration of more than 8.0% and soot are obtained. Then, the flue gas is transported to a sulfuric acid plant to produce sulfuric acid, and the soot can be made into pellets and returned to step (1) for reuse.

[0046] The test results obtained are as follows: The comprehensive recovery rate of platinum and palladium is 96.48%, the comprehensive recovery rate of gold and silver is 94.79%, the recovery rate of copper is 89.00%, the recovery rate of nickel is 88.75%, and the recovery rate of cobalt is 68.87%. Example 3

[0047] Raw material 3#: A copper - nickel - based complex platinum - palladium concentrate, with the main elements and their contents being: platinum 30.65 g / t, palladium 80.92 g / t, gold 6 g / t, silver 16 g / t, copper 2.25%, nickel 2.15%, cobalt 0.74%, iron 16.84%, sulfur 15.15%, magnesium 10.66%.

[0048] As Figure 1 shown, the present invention is implemented for this complex platinum - palladium concentrate, and the specific steps are as follows: (1)Smelting enrichment: The platinum-palladium concentrate to be treated is mixed with limestone, silica, ferric oxide or preferably the blown iron oxide slag (select one or two according to the material properties) and pulverized coal for smelting to obtain low-grade matte, slag and sulfur dioxide flue gas. During this process, the smelting temperature is 1360 °C, and the dosages of the smelting slag formers are as follows: limestone is 7.5% of the weight of the platinum-palladium concentrate, silica is 20% of the weight of the platinum-palladium concentrate, and ferric oxide (or preferably the blown iron oxide slag and the crystallized iron-removing slag) is 8% of the weight of the platinum-palladium concentrate.

[0049] (2)Two-stage cascade blowing enrichment: The low-grade matte obtained in step (1) is charged with silica and lump coal for primary converter blowing. When the primary blowing is approaching the end point, it is promptly transferred to another converter for short-time secondary blowing to obtain a bottom-settling high-grade matte liquid with an iron content of about 10% and a sulfur content of 22% - 25%, iron oxide slag (recycled) and sulfur dioxide flue gas. During this process, the temperature of the primary converter blowing is 1280 °C, and the dosage of silica for the primary converter blowing is 22% of the weight of the smelting product; the temperature of the secondary converter blowing is 1200 °C.

[0050] (3)Stage selective crystallization: The bottom-settling high-grade matte liquid obtained in step (2) is promptly transported to a stage temperature-controlled cooling furnace for selective crystallization to obtain a large-particle mineral-bound matte ore with a differentiated structure (including alloyed gold, cuprous sulfide and nickel sulfide ore). The specific process of stage temperature-controlled cooling is as follows: the slow cooling time from 1200 °C to 580 °C is 1 day, the slow cooling time from 580 °C to 520 °C is 4 days, the slow cooling time from 520 °C to 370 °C is 1 day, and then natural cooling can be carried out.

[0051] (4)Separation of mixed ore: The differentiated structure matte ore obtained in step (3) is transported to a grinding - weak magnetic field pulsed high-gradient magnetic separation (improved type) separation system for separation. First, the differentiated structure matte ore is finely ground, and after fine grinding, it is separated by weak magnetic field pulsed high-gradient magnetic separation to obtain alloyed gold (platinum-palladium-gold-copper-nickel-iron) and copper-nickel mixed concentrate. During this process, the fineness of fine grinding is that -0.074 mm accounts for 50% - 65%, the magnetic separation concentration is 20% - 25%, the magnetic induction intensity of the weak magnetic field pulsed high-gradient magnetic separator (improved type) is 100 - 200 mT, the diameter of the magnetic medium and the working gap are 3 - 5 mm, and the pulse intensity is 100 - 180 times / min; (5)Selective separation of crude gold: The crude gold obtained in step (4) is subjected to atmospheric pressure heating acid leaching - filtration - precipitation - filtration - crystallization. Sulfuric acid is added to the crude gold, and after heating acid leaching, filtration is carried out to obtain crude gold (platinum - palladium - gold - copper); continue to inflate, then add nickel carbonate to remove iron, and filter - wash to remove iron to obtain iron oxide sediment; the washed mother liquor is then subjected to nickel sulfate crystallization to obtain nickel sulfate crystals and nickel sulfate crystallization mother liquor (for recycling). During this process, the pH of heating acid leaching is 1 - 2, the leaching temperature is 65 - 85 °C, the liquid - solid ratio is 1:1 - 4:1, and the leaching time is 2 - 3 hours; iron removal is carried out by slowly adjusting the pH to 3.0 - 4.7 with nickel carbonate.

[0052] (6)Separation of copper - nickel mixed concentrate: The copper - nickel mixed concentrate obtained in step (4) is roasted, and the roasted slag is then subjected to acid leaching, filtration, and crystallization to obtain nickel oxide, copper sulfate crystals, and crystallization mother liquor (for recycling). During this process, the roasting temperature is 750 - 850 °C, the liquid - solid ratio is 2:1 - 5:1, the pH of acid leaching is 1 - 3, and the leaching temperature is 65 - 85 °C.

[0053] (7)Mix the nickel sulfate crystals obtained in step (5) with the nickel oxide obtained in step (6) and send them to a reverberatory furnace for smelting to obtain nickel anode plates with the main component being crude nickel. Then, the nickel anode plates are sent to the electrolysis process for electrolysis to obtain cathode nickel, anode slime, and silver - cobalt slag products.

[0054] In addition, a small amount of elemental copper contained in the crude gold (platinum - palladium - gold - copper) in step (5) is recycled in the subsequent traditional precious metal smelting process. After the sulfur dioxide - containing flue gases generated in steps (1) and (2) are combined, after dust removal by gravity, cyclone dust collectors, electrostatic precipitators, and bag filters, flue gas with a sulfur dioxide concentration of more than 8.0% and dust are obtained. Then, the flue gas is transported to a sulfuric acid plant to produce sulfuric acid, and the dust can be made into pellets and returned to step (1) for reuse.

[0055] The test results obtained are as follows: The comprehensive recovery rate of platinum and palladium is 97.53%, the comprehensive recovery rate of gold and silver is 95.29%, the recovery rate of copper is 90.00%, the recovery rate of nickel is 89.00%, and the recovery rate of cobalt is 69.08%.

[0056] To sum up, when using the method described in the present invention to process such complex platinum - palladium concentrates, where the ratio of copper to nickel is 1:1.1 - 2.5, the comprehensive recovery rate of platinum and palladium is above 96%, the comprehensive recovery rate of gold and silver is about 95%, the recovery rate of copper is about 90%, the recovery rate of nickel is about 89%, and the recovery rate of cobalt is about 69%. It can be seen that this method can efficiently separate and recover valuable metals in complex platinum - palladium concentrates, fully realizing the comprehensive recovery and utilization of resources, and having good economic and social benefits.

[0057] The present invention has been described in detail through specific and preferred embodiments. However, those skilled in the art should understand that the present invention is not limited to the above-described embodiments. Any modifications, equivalent substitutions, etc. made within the spirit and principles of the present invention shall be included within the protection scope of the present invention.

Claims

1. A highly efficient separation and comprehensive recovery method based on copper-nickel complex platinum-palladium concentrate, characterized in that: The steps include: (1) Smelting and enrichment: The platinum-palladium concentrate to be processed is mixed with a slag-making agent and coal powder for smelting to obtain low-grade matte, slag and sulfur dioxide flue gas; (2) Two-stage step blowing and enrichment: The low-grade matte obtained in step (1) is mixed with silicon dioxide and lump coal for a converter blowing. When the first blowing is close to the end, it is promptly transferred to another converter for a short-term secondary blowing to obtain high-grade matte liquid, iron oxide slag and sulfur dioxide flue gas at the bottom. (3) Stage selective crystallization: The high-grade matte solution obtained in step (2) is promptly transported to a stage temperature-controlled cooling furnace for selective crystallization to obtain a matte ore with large-particle minerals in a differentiated structure, wherein the matte ore in a differentiated structure contains gold, cuprous sulfide and nickel sulfide; (4) Mixed ore separation: The matte ore with differentiated structure obtained in step (3) is transported to a crushing and grinding-weak magnetic field pulse high gradient magnetic separation system for separation. The matte ore with differentiated structure is first finely ground, and then subjected to weak magnetic field pulse high gradient magnetic separation to obtain a gold mixture, which contains platinum, palladium, gold, copper, nickel and iron, as well as a copper-nickel mixed concentrate; (5) Selective separation of gold alloy: subjecting the gold alloy obtained in step (4) to acid leaching at normal pressure, filtration, precipitation, filtration and crystallization, adding sulfuric acid to the gold alloy, heating the acid leaching and filtering to obtain gold alloy, wherein the gold alloy contains platinum, palladium, gold and copper; continuing to aerate, adding nickel carbonate to remove iron, filtering and washing to remove iron, and obtaining iron oxide precipitate; subjecting the mother liquor after washing to nickel sulfate crystallization, obtaining nickel sulfate crystals and nickel sulfate crystallization mother liquor; (6) Separation of copper-nickel mixed concentrate: roasting the copper-nickel mixed concentrate obtained in step (4), and then acid leaching, filtering and crystallizing the roasted slag to obtain nickel oxide, copper sulfate crystals and crystallization mother liquor; (7) The nickel sulfate crystals obtained in step (5) and the nickel oxide obtained in step (6) are mixed and sent to a reverberatory furnace for smelting to obtain a nickel anode plate whose main component is crude nickel, and then the nickel anode plate is sent to an electrolysis process for electrolysis to obtain cathode nickel, anode mud and silver-cobalt slag products.

2. The method for efficient separation and comprehensive recovery of copper-nickel complex platinum-palladium concentrate according to claim 1, characterized in that: The slag-forming agent in step (1) is selected from one or two of limestone, silicon dioxide and ferric oxide, and the amounts of the slag-forming agents are: limestone is 7% to 12% by weight of the platinum-palladium concentrate, silicon dioxide is 20% to 25% by weight of the platinum-palladium concentrate, and ferric oxide is 7% to 15% by weight of the platinum-palladium concentrate.

3. A highly efficient separation and comprehensive recovery method based on copper-nickel complex platinum-palladium concentrate according to claim 2, characterized in that: The ferric oxide can be made from blowing iron oxide slag or crystallization iron removal slag.

4. The method for efficient separation and comprehensive recovery of copper-nickel complex platinum-palladium concentrate according to claim 1, characterized in that: The smelting temperature in step (1) is 1300-1450°C.

5. The method for efficient separation and comprehensive recovery of copper-nickel complex platinum-palladium concentrate according to claim 1, characterized in that: In step (2), the blowing temperature of the primary converter is 1200-1300°C, and the amount of silicon dioxide used is 15%-30% of the weight of the smelted product; the blowing temperature of the secondary converter is 1100-1200°C.

6. The method for efficient separation and comprehensive recovery of copper-nickel complex platinum-palladium concentrate according to claim 1, characterized in that: The cooling process in step (3) is as follows: the slow cooling time from 1100°C to 580°C is 1 to 2 days, the slow cooling time from 580°C to 520°C is 4 to 5 days, the slow cooling time from 520°C to 370°C is 1 to 2 days, and then natural cooling is performed.

7. The method for efficient separation and comprehensive recovery of copper-nickel complex platinum-palladium concentrate according to claim 1, characterized in that: In step (4), the fine grinding particle size is -0.074 mm, accounting for 50% to 65%; the magnetic separation concentration is 20% to 25%; the magnetic induction intensity of the weak magnetic field pulse high gradient magnetic separator is 100 to 200 mT, and the pulse intensity is 100 to 180 times / min.

8. The method for efficient separation and comprehensive recovery of copper-nickel complex platinum-palladium concentrate according to claim 7, characterized in that: The diameter and spacing of the magnetic medium rods in the upper half of the mass box of the weak magnetic field pulse high gradient magnetic separator are 3-5 mm, and the diameter and spacing of the magnetic medium rods in the lower half are 4-8 mm, and a grid screen with a sieve hole of 1 mm is installed on the upper first layer of magnetic medium rods.

9. The method for efficient separation and comprehensive recovery of copper-nickel complex platinum-palladium concentrate according to claim 1, characterized in that: In step (5), the pH of the heated acid leaching is 1 to 2, the leaching temperature is 65 to 85° C., the liquid-to-solid ratio is 1:1 to 4:1, and the leaching time is 2 to 3 hours; and nickel carbonate is used to slowly adjust the pH to 3.0 to 4.7 for iron removal.

10. The method for efficient separation and comprehensive recovery of copper-nickel complex platinum-palladium concentrate according to claim 1, characterized in that: In step (6), the roasting temperature is 750-850°C; the liquid-to-solid ratio of the acid leaching is 2:1-5:1, the pH is 1-3, and the leaching temperature is 65-85°C.

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