A method for efficiently recovering valuable components from complex low-grade platinum-palladium oxygen-sulfur mixed ore

Abstract

The present invention relates to a method for efficiently recovering valuable components of a complex low-grade platinum-palladium oxygen-sulfur mixed ore, comprising the following steps: (1) pre-treating the low-grade platinum-palladium oxygen-sulfur mixed ore to be treated; (2) primary enrichment of useful components; (3) secondary enrichment by pyrometallurgy; (4) differentiation ore-making to obtain differentiated structure matte ore; (5) separation of differentiated ore to obtain gold, silver-containing copper sulfide concentrate and nickel-cobalt sulfide concentrate; (6) separation of gold; and (7) separate refining of byproducts copper and nickel. The present invention has a significant effect on the efficient recovery of valuable components of a complex low-grade platinum-palladium oxygen-sulfur mixed ore, and the recovery rates of valuable components such as platinum, palladium, gold, silver, copper, and nickel are all above 60%, with good recovery efficiency, stable process and strong adaptability, and is of great significance for the comprehensive utilization of low-grade platinum-palladium mineral resources with poor copper and nickel.

Description

Technical Field

[0001] The invention belongs to the technical field of precious metal metallurgy, and specifically relates to a method for efficiently recovering valuable components of a complex low-grade platinum-palladium oxygen-sulfur mixed ore. Background Art

[0002] This type of complex low-grade platinum-palladium ore has a complex structure and composition. The gangue minerals are mostly serpentine, pyroxene and other minerals that are easy to mud and have good floatability, which makes sorting difficult. Moreover, this type of ore often undergoes certain shallow alteration in different spaces. While the sulfide ore is shallowly altered, the surrounding rock also undergoes alteration, the gangue minerals undergo talcization, chlorite and silicification, and a large amount of water-soluble salts are precipitated, which also brings more difficulties to process control. At present, the sulfide ore under research and development is first enriched by the traditional flotation process, and then the traditional roasting leaching or full wet leaching method is adopted. However, due to the good floatability and easy mudification of the gangue minerals, the grade and recovery rate of the flotation products obtained are low, resulting in a large amount of slag in the metallurgical process, high production costs, and low comprehensive recovery rates; or the oxide ore is directly leached by full wet metallurgy, but the comprehensive recovery rate is low, the production cost is high, and it causes environmental pollution and poor safety controllability. In general, this type of mine is still in a stagnant state, and the resources cannot be effectively developed and utilized.

[0003] In short, the current R&D process either has a complex process for recovering precious metals while ignoring the recovery of lean copper and nickel, or has a simple process for recovering lean copper and nickel but the recovery of the main precious metals is difficult and the production cost is high; the traditional pyrometallurgical process ignores the recovery of precious metals, and resources cannot be effectively recovered, resulting in a huge waste of resources. Therefore, exploring scientific and reasonable recovery processes has become a technical problem that needs to be solved urgently in this industry.

[0004] Based on the existing defects and deficiencies, the present invention provides a method for efficiently recovering valuable components of complex low-grade platinum-palladium ores with scientific and reasonable process, appropriate production cost and strong adaptability, so as to promote the effective comprehensive utilization of resources and improve the economic and social benefits of the enterprise. Summary of the invention

[0005] In view of the above problems, the present invention provides a method for efficiently recovering valuable components in complex low-grade platinum-palladium oxygen-sulfur mixed ore, which adopts a comprehensive resource recovery method of multi-stage quality improvement in the classification stage, efficient separation, separate refining, and closed-loop circulation treatment.

[0006] The specific technical solution is: a method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore, comprising the following steps:

[0007] (1) Pretreatment: The low-grade platinum-palladium sulfide ore and oxide ore to be treated are mixed in a ratio of 1:1-2:1, and then transported to a two-stage ball mill and hydrocyclone classification system for low-concentration grinding and low-concentration classification to obtain a slurry with a concentration of 18%-20%, of which -0.074 mm accounts for 80%-90% of the slurry;

[0008] (2) Primary enrichment of useful components: The slurry after classification in step (1) is transported to a flotation system for flotation to obtain a platinum-palladium concentrate and flotation tailings, and then the flotation tailings are transported to a pulse high-gradient magnetic separation system for magnetic separation to obtain a low-grade platinum-palladium-iron concentrate and magnetic separation tailings;

[0009] (3) Secondary enrichment by pyrometallurgy: The platinum-palladium concentrate and the low-grade platinum-palladium-iron concentrate obtained in step (2) are dried and then mixed with lime and coal powder in a ratio of 4:1-5:1 for slag smelting. After removing the scum, the product is mixed with silicon dioxide and lump coal for blowing. After completion, a bottom matte solution (with an iron content of 3.2%-3.4% and a sulfur content of 20%-25%) and iron silicate are obtained;

[0010] (4) Differentiation ore production: The precipitated matte solution obtained in step (3) is promptly transported to a temperature-controlled insulation cooling furnace for staged insulation cooling to obtain a matte ore with a differentiated structure;

[0011] (5) Separation of differentiated ores: The differentiated matte ore obtained in step (4) is transported to a fine grinding-pulse high gradient magnetic separation-preferential copper flotation separation system for low-concentration separation to obtain gold concentrate (containing 15% to 20% copper, 50% to 55% nickel, and 1400 to 1500 g / t of palladium, platinum and gold), silver-containing copper sulfide concentrate (containing ≧70% copper, ≦5% nickel, and ≧500 g / t silver), and nickel-cobalt sulfide concentrate (containing ≧70% nickel, ≦0.5% copper);

[0012] (6) Separation of gold alloy: subjecting the gold alloy (platinum, palladium, gold, copper, nickel, and iron) obtained in step (5) to acid leaching at normal pressure, filtration, precipitation, filtration, and crystallization to obtain platinum, palladium, gold, and copper alloy, iron hydroxide, nickel sulfate crystals, and nickel sulfate crystal mother liquor; wherein the specific process is: acid dissolution followed by filtration to obtain platinum, palladium, gold, and copper alloy, and then adding nickel carbonate to remove iron and filtering, and then subjecting the mother liquor to nickel sulfate crystallization to obtain nickel sulfate crystals and nickel sulfate crystal mother liquor (which can be recycled);

[0013] (7) Refining the by-products copper and nickel separately: the silver-containing copper sulfide concentrate obtained in step (5) is subjected to smelting, blowing and refining to obtain copper anode plates, which are then sent to the copper electrolysis process for electrolysis to obtain copper cathode plates and silver-containing anode mud; the nickel-cobalt sulfide concentrate obtained in step (5) is quenched with high-temperature water, dissolved with concentrated hydrochloric acid, filtered and crystallized to obtain nickel chloride crystals and crystallization mother liquor, and then the nickel chloride crystals are mixed with the nickel sulfate crystals obtained in step (6) and electrolyzed to obtain cathode nickel, anode mud and cobalt slag products.

[0014] Furthermore, in step (1), 1000 g / t of sodium carbonate and 150 g / t of sodium hexametaphosphate are added to a first-stage ball mill, and the grinding concentrations in both stages are 50% to 55%, and the graded concentration is 30% to 35%.

[0015] Furthermore, in step (2), 300 g / t of sodium hexametaphosphate as an inhibitor, 100 g / t of butyl xanthate + butyl ammonium black powder as a collector, and 40 g / t of POC as a frother are added to the flotation system.

[0016] Furthermore, the magnetic induction intensity of the pulse high gradient magnetic separation system in step (2) is 900-1200 mT, and the pulse intensity is 200-300 times / min.

[0017] Furthermore, in step (3), the smelting temperature is 1300-1350°C, the blowing temperature is 1180-1300°C, the amount of smelting lime is 20%-25% of the platinum-palladium concentrate, and the amount of blowing silica is 25%-30% of the smelting product.

[0018] Furthermore, in step (3), the iron content of the precipitated matte solution is controlled to be 3.2%-3.4%, and the sulfur content is controlled to be 20%-25%.

[0019] Furthermore, the specific operation of the heat preservation and cooling stage in step (4) is as follows: the slow cooling time from 1180°C to 580°C is 2 to 3 days, the slow cooling time from 580°C to 520°C is 6 to 7 days, and the slow cooling time from 520°C to 370°C is 2 to 3 days.

[0020] Furthermore, in step (5), the particle size of the finely ground ore is -0.043 mm, accounting for 75% to 80%; the magnetic separation concentration is 18% to 20%; the magnetic induction intensity of the pulse high gradient magnetic separation is 200 to 300 mT, and the pulse intensity is 200 to 300 times / min; the preferential copper flotation separation system is added with lime to adjust the slurry pH to 12.2 to 12.6, and a highly selective collector short-chain dithiosulfate 50 g / t and a foaming agent isobutyl carbinol 30 g / t are added.

[0021] Furthermore, the leaching conditions of the acid leaching in step (6) are controlled as follows: the leaching pH is 1-2, the leaching temperature is 75-85° C., the liquid-solid ratio is 2:1-4:1, and the leaching time is 2-3 hours; and when removing iron, nickel carbonate is used to adjust the pH to 2.7-3.5.

[0022] Furthermore, a small amount of elemental copper in the gold alloy (platinum, palladium, gold and copper) in step (6) is separated and recovered during the traditional precious metal smelting process.

[0023] Furthermore, the sulfur dioxide-containing flue gas generated in steps (2) to (7) is dedusted by an electrostatic precipitator and then mixed to obtain a mixed flue gas with a sulfur dioxide concentration of more than 8.0%, which is then transported to an acid plant for recycling in the production of sulfuric acid.

[0024] Furthermore, if the intermediate products obtained in steps (3) to (7) have a high content of mineral elements, they can be returned to step (3) for recycling.

[0025] Beneficial effects of the present invention: The present invention has a significant effect on the efficient recovery of valuable components of complex low-grade platinum-palladium oxygen-sulfur mixed ore, and the recovery rates of valuable components such as platinum, palladium, gold, silver, copper, and nickel are all above 60%, with good recovery efficiency, stable process and strong adaptability, which is of great significance for the comprehensive utilization of low-grade platinum-palladium mineral resources with poor copper and nickel. The beneficial effects of each process are as follows:

[0026] (1) The present invention first adopts a low-concentration coarse grinding-fine grinding-classification closed-loop circulation system to separate materials with relatively uniform particle size for a flotation to avoid over-grinding of easy-to-grind and easy-to-float impurity ores that affect flotation indicators; then, high-grade platinum-palladium concentrate and low-grade platinum-palladium-iron concentrate are separated by low-concentration flotation-pulse high-gradient magnetic separation to ensure that the concentrate is fully recovered. In this process, since this type of ore is a mixture of sulfide ore and oxide ore in proportion, it is beneficial to improve and stabilize the flotation grade and recovery rate of the platinum-palladium concentrate, and can also provide the required additives (iron oxide) for subsequent smelting, reduce the production cost of the pyrometallurgical enrichment process, and ensure the full utilization of low-grade oxide ore. Subsequently, the high-grade platinum-palladium concentrate and the low-grade platinum-palladium-iron concentrate are pyrometallurgically enriched in a ratio of 4:1 to 5:1 to further improve the grade of the valuable components.

[0027] (2) The present invention cools the bottom matte liquid obtained by blowing through the temperature control stage to form a mineral combination with a differentiated structure, such as large sulfide single crystal particle size and appropriate gold yield, which can effectively avoid the formation of mixed crystals and create favorable conditions for subsequent separation. Then, the fine grinding-pulse high gradient magnetic separation-preferential copper flotation process is used for the differentiated ore to effectively improve the separation effect and recovery rate of platinum, palladium, copper and nickel.

[0028] (3) For the separated gold and silver-containing copper sulfide concentrates and nickel-cobalt sulfide concentrates, a large amount of waste is discarded by a combination of physical and chemical methods according to their material characteristics, so that the valuable components can be efficiently enriched. Then, a process of re-separation and separate refining is adopted, which greatly reduces the processing volume in the high-energy consumption stage and reduces the production cost. At the same time, a higher comprehensive recovery rate is obtained, so that the low-grade platinum and palladium mineral resources with poor copper and nickel are effectively utilized. BRIEF DESCRIPTION OF THE DRAWINGS

[0029] Figure 1 is a flow chart of the method of the present invention; DETAILED DESCRIPTION

[0030] In order to make the technical problems and technical solutions solved by the present invention more clearly understood, the present invention is further described in detail below in conjunction with the accompanying 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. Example 1

[0031] Raw material 1#: A complex low-grade platinum-palladium oxygen-sulfur mixed ore, in which the mixing ratio of sulfide ore and oxide ore is 1:1. The main elements and contents of the sulfide ore are: 1.00g / t platinum, 1.71g / t palladium, 0.02g / t gold, 0.12% copper, 0.16% nickel, 0.018% cobalt, 4g / t silver, and 8.19% iron; the main elements and contents of the oxide ore are: 0.61g / t platinum, 0.86g / t palladium, 0.01g / t gold, 0.06% copper, 0.14% nickel, 0.016% cobalt, 5g / t silver, and 10.42% iron.

[0032] like Figure 1 As shown, the present invention is used to recover the complex low-grade platinum-palladium oxygen-sulfur mixed ore, and the specific steps are as follows:

[0033] (1) Pretreatment: Raw material 1# is transported to a two-stage ball mill and hydrocyclone classification system for low-concentration grinding and low-concentration classification to obtain a slurry with a concentration of 18% to 20%, of which -0.074 mm accounts for 80% to 90% of the slurry; during this process, 1000 g / t of sodium carbonate and 150 g / t of sodium hexametaphosphate are added to a first-stage ball mill, and the two-stage grinding concentration is controlled at 50% to 55%, and the classification concentration is 30% to 35%.

[0034] (2) Primary enrichment of useful components: The ore pulp after classification in step (1) is transported to a flotation system for flotation. During the flotation process, 300 g / t of sodium hexametaphosphate inhibitor, 100 g / t of butyl xanthate + butyl ammonium black powder collector, and 40 g / t of POC frother are added to the flotation system to obtain platinum-palladium concentrate and flotation tailings by flotation. The flotation tailings are then transported to a pulse high gradient magnetic separation system for magnetic separation. The magnetic induction intensity of the magnetic separator is 900-1200 mT and the pulse intensity is 200-300 times / min. Low-grade platinum-palladium iron concentrate and magnetic separation tailings are obtained by magnetic separation.

[0035] (3) Secondary enrichment by pyrometallurgy: The platinum-palladium concentrate and low-grade platinum-palladium-iron concentrate obtained in step (2) are dried and then mixed with lime and coal powder in a ratio of 5:1 for slag smelting. After removing the scum, the product is mixed with silicon dioxide and lump coal for blowing. After completion, a bottom matte liquid (with an iron content of 3.2% to 3.4% and a sulfur content of 20% to 25%) and iron silicate are obtained. In this process, the smelting temperature is 1300 to 1350°C, the blowing temperature is 1180 to 1300°C, the amount of smelting lime used is 20% to 25% of the platinum-palladium concentrate, and the amount of blowing silicon dioxide used is 25% to 30% of the smelting product.

[0036] (4) Differentiation ore production: The precipitated matte liquid obtained in step (3) is promptly transported to a temperature-controlled insulation cooling furnace for staged insulation cooling to obtain a differentiated structure matte ore; wherein the specific operation of the insulation cooling stage is to cool the temperature from 1180°C to 580°C for 3 days, to cool the temperature from 580°C to 520°C for 7 days, and to cool the temperature from 520°C to 370°C for 3 days.

[0037] (5) Separation of differentiated ores: The differentiated structure matte ore obtained in step (4) is transported to a fine grinding-pulse high gradient magnetic separation-preferential copper flotation separation system for low-concentration separation to obtain gold concentrate (containing 15% to 20% copper, 50% to 55% nickel, and 1400 to 1500 g / t of palladium, platinum and gold), silver-containing copper sulfide concentrate (containing ≧70% copper, ≦5% nickel, and ≧500 g / t silver), and nickel-cobalt sulfide concentrate (containing ≧70% nickel, ≦0.5% copper). During this process, the particle size of finely ground ore is -0.043mm, accounting for 75%~80%; the magnetic separation concentration is 18%~20%; the magnetic induction intensity of pulse high gradient magnetic separation is 200~300mT, and the pulse intensity is 200~300 times / min; the preferential copper flotation separation system adds lime to adjust the slurry pH to 12.2~12.6, and adds 50g / t of highly selective collector short-chain dithiosulfate and 30g / t of foaming agent isobutyl methanol.

[0038] (6) Separation of gold alloy: The gold alloy obtained in step (5) is subjected to acid leaching at normal pressure, filtration, precipitation, filtration and crystallization to obtain platinum-palladium-gold-copper gold alloy, iron hydroxide, nickel sulfate crystals and nickel sulfate crystal mother liquor; the specific process is as follows: acid dissolution followed by filtration to obtain platinum-palladium-gold-copper gold alloy, nickel carbonate is added to remove iron, and after filtration, nickel sulfate crystallization is performed on the mother liquor to obtain nickel sulfate crystals and nickel sulfate crystal mother liquor (which can be recycled); in this process, the leaching conditions are controlled as follows: leaching pH is 1-2, leaching temperature is 75-85°C, liquid-to-solid ratio is 4:1, and leaching time is 3 hours; when removing iron, the pH is adjusted to 2.7-3.5 using nickel carbonate.

[0039] (7) Refining the by-products copper and nickel separately: the silver-containing copper sulfide concentrate obtained in step (5) is subjected to smelting, blowing and refining to obtain copper anode plates, which are then sent to the copper electrolysis process for electrolysis to obtain copper cathode plates and silver-containing anode mud; the nickel-cobalt sulfide concentrate obtained in step (5) is quenched with high-temperature water, dissolved with concentrated hydrochloric acid, filtered and crystallized to obtain nickel chloride crystals and crystallization mother liquor, and then the nickel chloride crystals are mixed with the nickel sulfate crystals obtained in step (6) and electrolyzed to obtain cathode nickel, anode mud and cobalt slag products.

[0040] The present invention is implemented on the complex low-grade platinum-palladium oxygen-sulfur mixed ore, and the test results obtained are: platinum recovery rate is 80.87%, palladium recovery rate is 82.52%, gold recovery rate is 66.54%, silver recovery rate is 64.85%, copper recovery rate is 77.44%, and nickel recovery rate is 70.72%. Example 2

[0041] Raw material 2#: A complex low-grade platinum-palladium oxygen-sulfur mixed ore, in which the mixing ratio of sulfide ore to oxide ore is 1.5:1. The main elements and contents of the sulfide ore are: 1.00g / t platinum, 1.71g / t palladium, 0.02g / t gold, 0.12% copper, 0.16% nickel, 0.018% cobalt, 4g / t silver, and 8.19% iron; the main elements and contents of the oxide ore are: 0.61g / t platinum, 0.86g / t palladium, 0.01g / t gold, 0.06% copper, 0.14% nickel, 0.016% cobalt, 5g / t silver, and 10.42% iron.

[0042] like Figure 1 As shown, the present invention is used to recover the complex low-grade platinum-palladium oxygen-sulfur mixed ore, and the specific steps are as follows:

[0043] (1) Pretreatment: Raw material 2# is transported to a two-stage ball mill and hydrocyclone classification system for low-concentration grinding and low-concentration classification to obtain a slurry with a concentration of 18% to 20%, of which -0.074 mm accounts for 80% to 90% of the slurry; during this process, 1000 g / t of sodium carbonate and 150 g / t of sodium hexametaphosphate are added to a first-stage ball mill, and the two-stage grinding concentration is controlled at 50% to 55%, and the classification concentration is 30% to 35%.

[0044] (2) Primary enrichment of useful components: The ore pulp after classification in step (1) is transported to a flotation system for flotation. During the flotation process, 300 g / t of sodium hexametaphosphate inhibitor, 100 g / t of butyl xanthate + butyl ammonium black powder collector, and 40 g / t of POC frother are added to the flotation system to obtain platinum-palladium concentrate and flotation tailings by flotation. The flotation tailings are then transported to a pulse high gradient magnetic separation system for magnetic separation. The magnetic induction intensity of the magnetic separator is 900-1200 mT and the pulse intensity is 200-300 times / min. Low-grade platinum-palladium iron concentrate and magnetic separation tailings are obtained by magnetic separation.

[0045] (3) Secondary enrichment by pyrometallurgy: The platinum-palladium concentrate and low-grade platinum-palladium-iron concentrate obtained in step (2) are dried and then mixed with lime and coal powder in a ratio of 4:1 for slag smelting. After removing the scum, the product is mixed with silicon dioxide and lump coal for blowing. After completion, a bottom matte liquid (with an iron content controlled at 3.2% to 3.4% and a sulfur content of 20% to 25%) and iron silicate are obtained. In this process, the smelting temperature is 1300 to 1350°C, the blowing temperature is 1180 to 1300°C, the amount of smelting lime used is 20% to 25% of the platinum-palladium concentrate, and the amount of blowing silicon dioxide used is 25% to 30% of the smelting product.

[0046] (4) Differentiation ore production: The precipitated matte liquid obtained in step (3) is promptly transported to a temperature-controlled insulation cooling furnace for staged insulation cooling to obtain a differentiated structure matte ore; wherein the specific operation of the insulation cooling stage is to cool the temperature from 1180°C to 580°C for 2 days, to cool the temperature from 580°C to 520°C for 6 days, and to cool the temperature from 520°C to 370°C for 2 days.

[0047] (5) Separation of differentiated ores: The differentiated structure matte ore obtained in step (4) is transported to a fine grinding-pulse high gradient magnetic separation-preferential copper flotation separation system for low-concentration separation to obtain gold concentrate (containing 15% to 20% copper, 50% to 55% nickel, and 1400 to 1500 g / t of palladium, platinum and gold), silver-containing copper sulfide concentrate (containing ≧70% copper, ≦5% nickel, and ≧500 g / t silver), and nickel-cobalt sulfide concentrate (containing ≧70% nickel, ≦0.5% copper). During this process, the particle size of finely ground ore is -0.043mm, accounting for 75%~80%; the magnetic separation concentration is 18%~20%; the magnetic induction intensity of pulse high gradient magnetic separation is 200~300mT, and the pulse intensity is 200~300 times / min; the preferential copper flotation separation system adds lime to adjust the slurry pH to 12.2~12.6, and adds 50g / t of highly selective collector short-chain dithiosulfate and 30g / t of foaming agent isobutyl methanol.

[0048] (6) Separation of gold alloy: The gold alloy obtained in step (5) is subjected to acid leaching at normal pressure, filtration, precipitation, filtration and crystallization to obtain platinum-palladium-gold-copper gold alloy, iron hydroxide, nickel sulfate crystals and nickel sulfate crystal mother liquor; the specific process is as follows: acid dissolution followed by filtration to obtain platinum-palladium-gold-copper gold alloy, nickel carbonate is added to remove iron, and after filtration, nickel sulfate crystallization is performed on the mother liquor to obtain nickel sulfate crystals and nickel sulfate crystal mother liquor (which can be recycled); in this process, the leaching conditions are controlled as follows: leaching pH is 1-2, leaching temperature is 75-85°C, liquid-to-solid ratio is 3:1, and leaching time is 2 hours; when removing iron, the pH is adjusted to 2.7-3.5 using nickel carbonate.

[0049] (7) Refining the by-products copper and nickel separately: the silver-containing copper sulfide concentrate obtained in step (5) is subjected to smelting, blowing and refining to obtain copper anode plates, which are then sent to the copper electrolysis process for electrolysis to obtain copper cathode plates and silver-containing anode mud; the nickel-cobalt sulfide concentrate obtained in step (5) is quenched with high-temperature water, dissolved with concentrated hydrochloric acid, filtered and crystallized to obtain nickel chloride crystals and crystallization mother liquor, and then the nickel chloride crystals are mixed with the nickel sulfate crystals obtained in step (6) and electrolyzed to obtain cathode nickel, anode mud and cobalt slag products.

[0050] The present invention was implemented on the complex low-grade platinum-palladium oxygen-sulfur mixed ore, and the test results obtained were: platinum recovery rate was 81.35%, palladium recovery rate was 82.98%, gold recovery rate was 66.92%, silver recovery rate was 65.63%, copper recovery rate was 78.57%, and nickel recovery rate was 71.09%. Example 3

[0051] Raw material 3#: A complex low-grade platinum-palladium oxygen-sulfur mixed ore, in which the mixing ratio of sulfide ore to oxide ore is 2:1. The main elements and contents of the sulfide ore are: 1.00g / t platinum, 1.71g / t palladium, 0.02g / t gold, 0.12% copper, 0.16% nickel, 0.018% cobalt, 4g / t silver, and 8.19% iron; the main elements and contents of the oxide ore are: 0.61g / t platinum, 0.86g / t palladium, 0.01g / t gold, 0.06% copper, 0.14% nickel, 0.016% cobalt, 5g / t silver, and 10.42% iron.

[0052] like Figure 1 As shown, the present invention is used to recover the complex low-grade platinum-palladium oxygen-sulfur mixed ore, and the specific steps are as follows:

[0053] (1) Pretreatment: Raw material 2# is transported to a two-stage ball mill and hydrocyclone classification system for low-concentration grinding and low-concentration classification to obtain a slurry with a concentration of 18% to 20%, of which -0.074 mm accounts for 80% to 90% of the slurry; during this process, 1000 g / t of sodium carbonate and 150 g / t of sodium hexametaphosphate are added to a first-stage ball mill, and the two-stage grinding concentration is controlled at 50% to 55%, and the classification concentration is 30% to 35%.

[0054] (2) Primary enrichment of useful components: The ore pulp after classification in step (1) is transported to a flotation system for flotation. During the flotation process, 300 g / t of sodium hexametaphosphate inhibitor, 100 g / t of butyl xanthate + butyl ammonium black powder collector, and 40 g / t of POC frother are added to the flotation system to obtain platinum-palladium concentrate and flotation tailings by flotation. The flotation tailings are then transported to a pulse high gradient magnetic separation system for magnetic separation. The magnetic induction intensity of the magnetic separator is 900-1200 mT and the pulse intensity is 200-300 times / min. Low-grade platinum-palladium iron concentrate and magnetic separation tailings are obtained by magnetic separation.

[0055] (3) Secondary enrichment by pyrometallurgy: The platinum-palladium concentrate and low-grade platinum-palladium-iron concentrate obtained in step (2) are dried and then mixed with lime and coal powder in a ratio of 4.5:1 for slag smelting. After removing the scum, the product is mixed with silicon dioxide and lump coal for blowing. After completion, a bottom matte liquid (with an iron content controlled at 3.2% to 3.4% and a sulfur content of 20% to 25%) and iron silicate are obtained. In this process, the smelting temperature is 1300 to 1350°C, the blowing temperature is 1180 to 1300°C, the amount of smelting lime used is 20% to 25% of the platinum-palladium concentrate, and the amount of blowing silicon dioxide used is 25% to 30% of the smelting product.

[0056] (4) Differentiation ore production: The precipitated matte liquid obtained in step (3) is promptly transported to a temperature-controlled insulation cooling furnace for staged insulation cooling to obtain a differentiated structure matte ore; wherein the specific operation of the insulation cooling stage is to cool the temperature from 1180°C to 580°C for 3 days, to cool the temperature from 580°C to 520°C for 6 days, and to cool the temperature from 520°C to 370°C for 3 days.

[0057] (5) Separation of differentiated ores: The differentiated structure matte ore obtained in step (4) is transported to a fine grinding-pulse high gradient magnetic separation-preferential copper flotation separation system for low-concentration separation to obtain gold concentrate (containing 15% to 20% copper, 50% to 55% nickel, and 1400 to 1500 g / t of palladium, platinum and gold), silver-containing copper sulfide concentrate (containing ≧70% copper, ≦5% nickel, and ≧500 g / t silver), and nickel-cobalt sulfide concentrate (containing ≧70% nickel, ≦0.5% copper). During this process, the particle size of finely ground ore is -0.043mm, accounting for 75%~80%; the magnetic separation concentration is 18%~20%; the magnetic induction intensity of pulse high gradient magnetic separation is 200~300mT, and the pulse intensity is 200~300 times / min; the preferential copper flotation separation system adds lime to adjust the slurry pH to 12.2~12.6, and adds 50g / t of highly selective collector short-chain dithiosulfate and 30g / t of foaming agent isobutyl methanol.

[0058] (6) Separation of gold alloy: The gold alloy obtained in step (5) is subjected to acid leaching at normal pressure, filtration, precipitation, filtration and crystallization to obtain platinum-palladium-gold-copper gold alloy, iron hydroxide, nickel sulfate crystals and nickel sulfate crystal mother liquor; the specific process is as follows: acid dissolution followed by filtration to obtain platinum-palladium-gold-copper gold alloy, nickel carbonate is added to remove iron and filtered, and then nickel sulfate crystallization is performed on the mother liquor to obtain nickel sulfate crystals and nickel sulfate crystal mother liquor (which can be recycled); in this process, the leaching conditions are controlled as follows: leaching pH is 1-2, leaching temperature is 75-85°C, liquid-to-solid ratio is 2:1, and leaching time is 2 hours; when removing iron, the pH is adjusted to 2.7-3.5 using nickel carbonate.

[0059] (7) Refining the by-products copper and nickel separately: the silver-containing copper sulfide concentrate obtained in step (5) is subjected to smelting, blowing and refining to obtain copper anode plates, which are then sent to the copper electrolysis process for electrolysis to obtain copper cathode plates and silver-containing anode mud; the nickel-cobalt sulfide concentrate obtained in step (5) is quenched with high-temperature water, dissolved with concentrated hydrochloric acid, filtered and crystallized to obtain nickel chloride crystals and crystallization mother liquor, and then the nickel chloride crystals are mixed with the nickel sulfate crystals obtained in step (6) and electrolyzed to obtain cathode nickel, anode mud and cobalt slag products.

[0060] The present invention was implemented on the complex low-grade platinum-palladium oxygen-sulfur mixed ore, and the test results obtained were: platinum recovery rate was 79.59%, palladium recovery rate was 80.88%, gold recovery rate was 63.94%, silver recovery rate was 63.55%, copper recovery rate was 76.86%, and nickel recovery rate was 69.76%.

[0061] In summary, adopting the method described in the present invention to process complex low-grade platinum-palladium-oxygen-sulfur mixed ores can effectively recover valuable metals in such ores, with the platinum recovery rate reaching about 80%, the palladium recovery rate reaching about 81%, the gold recovery rate reaching about 64%, the silver recovery rate reaching about 65%, the copper recovery rate reaching about 77%, and the nickel recovery rate reaching about 70%. The recovery efficiency is good, the process is stable and highly adaptable, which is of great significance for the comprehensive utilization of poor copper-nickel low-grade platinum-palladium mineral resources.

[0062] The present invention has been described in detail above 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 principle of the present invention shall be included within the protection scope of the present invention.

Claims

1. A method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore, characterized in that: The steps include: (1) Pretreatment: The low-grade platinum-palladium sulfide ore and oxide ore to be treated are mixed in a ratio of 1:1-2:1, and then transported to a two-stage ball mill and hydrocyclone classification system for low-concentration grinding and low-concentration classification to obtain a slurry with a concentration of 18%-20%, of which -0.074 mm accounts for 80%-90% of the slurry; (2) Primary enrichment of useful components: The slurry after classification in step (1) is transported to a flotation system for flotation to obtain a platinum-palladium concentrate and flotation tailings, and then the flotation tailings are transported to a pulse high-gradient magnetic separation system for magnetic separation to obtain a low-grade platinum-palladium-iron concentrate and magnetic separation tailings; (3) Secondary enrichment by pyrometallurgy: The platinum-palladium concentrate and the low-grade platinum-palladium-iron concentrate obtained in step (2) are dried, mixed with lime and coal powder in a ratio of 4:1 to 5:1 for slag smelting, and the product is then mixed with silicon dioxide and lump coal for blowing after removing the scum. After completion, a bottom matte liquid and iron silicate are obtained; (4) Differentiation ore production: The precipitated matte solution obtained in step (3) is promptly transported to a temperature-controlled insulation cooling furnace for staged insulation cooling to obtain a matte ore with a differentiated structure; (5) Separation of differentiated ores: The differentiated matte ore obtained in step (4) is transported to a fine grinding-pulse high gradient magnetic separation-preferential copper flotation separation system for low-concentration separation to obtain a gold-containing, silver-containing copper sulfide concentrate and a nickel-cobalt sulfide concentrate; (6) separation of gold alloy: subjecting the gold alloy obtained in step (5) to acid leaching at normal pressure, filtration, precipitation, filtration and crystallization to obtain platinum-palladium-gold-copper gold alloy, iron hydroxide, nickel sulfate crystals and nickel sulfate crystallization mother liquor; the specific process is: acid dissolution followed by filtration to obtain platinum-palladium-gold-copper gold alloy, adding nickel carbonate to remove iron and filtering, and then subjecting the mother liquor to nickel sulfate crystallization to obtain nickel sulfate crystals and nickel sulfate crystallization mother liquor; (7) Refining the by-products copper and nickel separately: the silver-containing copper sulfide concentrate obtained in step (5) is subjected to smelting, blowing and refining to obtain copper anode plates, which are then sent to the copper electrolysis process for electrolysis to obtain copper cathode plates and silver-containing anode mud; the nickel-cobalt sulfide concentrate obtained in step (5) is quenched with high-temperature water, dissolved with concentrated hydrochloric acid, filtered and crystallized to obtain nickel chloride crystals and crystallization mother liquor, and then the nickel chloride crystals are mixed with the nickel sulfate crystals obtained in step (6) and electrolyzed to obtain cathode nickel, anode mud and cobalt slag products.

2. The method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore according to claim 1, characterized in that: Step (1) Add 1000 g / t of sodium carbonate and 150 g / t of sodium hexametaphosphate into a first-stage ball mill. The grinding concentrations in both stages are 50% to 55%, and the graded concentration is 30% to 35%.

3. The method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore according to claim 1, characterized in that: Step (2) Add 300 g / t of sodium hexametaphosphate as an inhibitor, 100 g / t of butyl xanthate + butyl ammonium black powder as a collector, and 40 g / t of POC as a frother into the flotation system.

4. The method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore according to claim 1, characterized in that: The magnetic induction intensity of the pulse high gradient magnetic separation system in step (2) is 900-1200 mT, and the pulse intensity is 200-300 times / min.

5. The method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore according to claim 1, characterized in that: In step (3), the smelting temperature is 1300-1350°C, the blowing temperature is 1180-1300°C, the amount of smelting lime is 20%-25% of the platinum-palladium concentrate, and the amount of blowing silicon dioxide is 25%-30% of the smelting product.

6. The method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore according to claim 1, characterized in that: In step (3), the iron content of the precipitated matte solution is controlled to be 3.2%-3.4%, and the sulfur content is controlled to be 20%-25%.

7. The method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore according to claim 1, characterized in that: The specific operation of the stage insulation cooling in step (4) is as follows: the slow cooling time from 1180°C to 580°C is 2 to 3 days, the slow cooling time from 580°C to 520°C is 6 to 7 days, and the slow cooling time from 520°C to 370°C is 2 to 3 days.

8. The method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore according to claim 1, characterized in that: In step (5), the particle size of the finely ground ore is -0.043 mm, accounting for 75% to 80%; the magnetic separation concentration is 18% to 20%; the magnetic induction intensity of the pulse high gradient magnetic separation is 200 to 300 mT, and the pulse intensity is 200 to 300 times / min; the preferential copper flotation separation system is added with lime to adjust the slurry pH to 12.2 to 12.6, and a highly selective collector short-chain dithiosulfate 50 g / t and a foaming agent isobutyl carbinol 30 g / t are added.

9. The method for efficiently recovering valuable components from a complex low-grade platinum-palladium oxygen-sulfur mixed ore according to claim 1, characterized in that: The leaching conditions of the acid leaching in step (6) are controlled as follows: the leaching pH is 1-2, the leaching temperature is 75-85° C., the liquid-to-solid ratio is 2:1-4:1, and the leaching time is 2-3 hours; when removing iron, nickel carbonate is used to adjust the pH to 2.7-3.5.

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