Method for recovering valuable metals in nickel soot

Abstract

The invention discloses a method for recovering valuable metals in nickel soot, which comprises the following steps: realizing high-efficiency leaching of nickel, copper, cobalt and zinc through reduction-oxidation combined pressure leaching, and respectively recovering copper sulfate, zinc sulfate and nickel sulfate cobalt solution by adopting a multi-stage extraction separation technology. Separation and recovery of valuable metals in the nickel soot are achieved, impurities such as harmful elements arsenic and lead in the nickel soot are removed, and the quality of nickel products is effectively guaranteed. The method solves the problems of impurity enrichment and low valuable metal leaching rate in the traditional process, the open circuit rate of harmful elements such as lead and iron is greater than or equal to 95%, and the nickel-cobalt recovery rate is greater than or equal to 90%.

Description

Technical Field

[0001] The present invention belongs to the technical field of nonferrous metal metallurgy, and in particular relates to a method for recovering valuable metals (nickel, copper, cobalt, zinc) in nickel ash produced by a nickel smelting pyrometallurgical process. Background Art

[0002] In the pyrometallurgical production process of nickel-copper sulfide ore, the copper-nickel concentrate obtained by beneficiation enters the pyrometallurgical smelting system to obtain high-nickel matte, and nickel ash is produced at the same time. The nickel ash will continue to return to the pyrometallurgical system, resulting in the inability of impurity elements such as arsenic, lead, and zinc that affect the quality of electrolytic nickel to open the circuit, and the harmful impurity elements continue to enrich, which ultimately affects the quality of electrolytic nickel products.

[0003] The pyrometallurgical smelting of nickel-copper sulfide ores primarily involves oxidizing the sulfur in the raw materials into sulfur dioxide (to produce sulfuric acid) while simultaneously adding elements like iron to the slag in an oxidized state (silicates), thereby enriching valuable metals like copper and nickel. The pyrometallurgical smelting of nickel-copper sulfide ores primarily occurs in an oxidizing atmosphere, so the metallic elements in the ash are partially present in an oxidized state. Furthermore, the ash also contains a significant amount of sulfur-containing materials, either due to the sulfide nature of the raw materials fed into the furnace or other factors. Consequently, the ash produced by pyrometallurgical smelting of copper-nickel sulfide ores has a complex chemical and phase composition, making its processing more challenging.

[0004] Due to the complex elemental composition of nickel ash, sulfuric acid leaching processes result in low leaching rates for valuable metals such as nickel, copper, and cobalt, making recovery difficult and separating the valuable metals from impurity elements challenging. To address these issues, the present invention proposes a combined oxidation-reduction method for treating nickel ash, which improves the leaching rates of nickel, copper, and cobalt from the ash, enabling the reuse of valuable metals and the removal of impurity elements. Summary of the Invention

[0005] In order to solve the above technical problems, the present invention provides a method for recovering valuable metals in nickel ash.

[0006] A method for recovering valuable metals in nickel ash comprises the following steps: (1) Reduction pressure leaching: nickel ash is slurried with raffinate and sulfuric acid solution and then added to a pressure autoclave. The temperature is raised and sulfur dioxide gas is introduced to react with stirring. The amount of sulfuric acid used is 0.85-1.0 times the theoretical amount of leached valuable metals, the liquid-solid ratio is 4-10:1, the reaction temperature is 150-200°C, the sulfur dioxide gas partial pressure is 0.1-0.5 MPa, the reaction time is 2-6 hours, and the introduction of sulfur dioxide gas is stopped after the reaction is completed; (2) Oxidative pressure leaching: oxygen is introduced into the reduction leaching reaction slurry obtained in step (1) to continue the reaction in a pressure autoclave, the reaction temperature is 150-200°C, the oxygen partial pressure is controlled to be 0.1-0.3 MPa, the reaction time is 2-6 hours, and the pH value at the reaction end point is 1.0-2.0. After the reaction is completed, solid-liquid separation is performed to obtain a leachate and lead-iron slag; (3) Copper extraction with lix984: The leachate obtained in step (2) is subjected to organic copper extraction with lix984, with an extraction phase O / A ratio of 1.5:1-3:1, two-stage countercurrent extraction, two-stage countercurrent washing, washing with a sulfuric acid solution having a pH of 1.5-2.5, and two-stage back extraction with a 150-220 g / L sulfuric acid solution to obtain a copper sulfate solution containing 45-55 g / L of copper. The copper raffinate is used in the neutralization process; (4) neutralizing the copper raffinate obtained in step (3) with sodium carbonate, controlling the neutralization endpoint pH to 2.5-4.0, and returning the neutralized slag to the reduction pressure leaching process; (5) The pH of the neutralized solution obtained in step (4) is adjusted to 2.0-2.5, zinc is extracted with P204, the extraction phase O / A ratio is 1:5-3:1, 4-6 stage countercurrent extraction is performed, 0.05-0.1 mol / L sulfuric acid solution is used for washing, 4-5 stage countercurrent washing is performed, the washing phase O / A ratio is 8:1-10:1, and P204 is loaded for back extraction using 0.35-0.4 mol / L sulfuric acid solution for four stage countercurrent stripping to obtain a zinc sulfate solution; (6) After adjusting the pH of the zinc raffinate obtained in step (5) to 4.0-5.0, nickel and cobalt are extracted using HBL116 extractant, the extraction phase O / A ratio is 1.5:1-5:1, 4-6 stage countercurrent extraction is performed, and the zinc raffinate is washed with a sulfuric acid solution with a pH of 1.5-2.5, 4-5 stage countercurrent washing, the washing phase O / A ratio is 8:1-10:1, and HBL116 is loaded for back extraction using a 0.5-0.75 mol / L sulfuric acid solution for 4-6 stages of countercurrent stripping to obtain a nickel and cobalt sulfate solution.

[0007] The chemical composition of nickel fly ash includes 4%-9% nickel, 4%-8% copper, 0.1%-0.3% cobalt, 10%-30% iron, 0.2%-7% zinc, 0.3%-10% lead, and 0.2%-0.4% arsenic. This fly ash, derived from the pyrometallurgical treatment process of nickel-copper sulfide ore, has a complex composition and phase structure. Based on its production principles and phase formation, a two-stage pressure leaching process (combined reduction-oxidation) is employed to extract the valuable metals from the nickel fly ash and recover them into the corresponding products. This represents a comprehensive method for recycling and treating nickel fly ash.

[0008] The present invention discloses a combined oxidation-reduction method for treating nickel fly ash produced by pyrometallurgical smelting of nickel-copper sulfide ores. This method utilizes a combined reduction pressure leaching and oxidation pressure leaching process to leach and recover valuable metals from the raw material. The combined leaching process selectively extracts copper, zinc, nickel, and cobalt from the fly ash, ultimately trapping iron and lead in the leached residue. Furthermore, extraction and separation techniques are employed to recover copper as a copper sulfate solution, zinc as a zinc sulfate solution, and nickel and cobalt as a nickel-cobalt sulfate solution.

[0009] Compared with the prior art, the present invention has the following beneficial effects: 1. Environmental benefits: efficient removal and resource utilization of harmful elements 1. Directional separation of impurity elements Through combined reduction-oxidation leaching, lead and iron are enriched in the slag as stable compounds (PbSO₄, Fe₂O₃, and Fe(OH)SO₄), achieving a lead open-circuit rate of ≥98% (the lead slag in Example 1 contained 0.36% Pb). Arsenic is removed through iron salt co-precipitation (As residue <0.01%), solving the problem of impurity cyclic enrichment in traditional processes. While conventional reprocessing can achieve lead enrichment of 10%-15%, the lead-iron slag produced by this invention can be directly used as a lead smelting raw material or safely landfilled (in compliance with GB 5085.3-2007 hazardous waste standards).

[0010] 2. Wastewater reduction: Recycling of raffinate (e.g., returning it to the pulping process) and returning the neutralized residue to the leaching process can reduce wastewater emissions throughout the entire process and mitigate the risk of heavy metal pollution.

[0011] 2. Economic Benefits: High Recovery Rate of Valuable Metals and Low Cost 1. Metal recovery rate significantly improved Nickel / Cobalt: SO2 pressure reduction in the reduction stage enhances the dissolution of metal oxides (especially high-valent states). O2 pressure oxidation leaching in the oxidation stage promotes the dissolution of metal sulfides (Ni3S2 + H2SO4 + 1 / 2O2 = NiSO4 + 2NiS + H2O; MeS + 2O2 = MeSO4 (Me = Ni, Co)). Leaching rates for nickel and copper are ≥95%, and for cobalt are ≥90%. High-temperature leaching increases chemical reaction kinetics and promotes material dissolution. A copper / zinc / nickel / cobalt step extraction system uses the lix984 to selectively extract copper, while P204 extracts zinc at a pH of 2.0-2.5. The nickel and cobalt are retained in the raffinate, which is then co-extracted using the HBL116, achieving recovery rates of ≥99% for copper, ≥99% for zinc, and ≥99% for nickel and cobalt.

[0012] 2. Reduced operating costs Energy consumption optimization: The two-stage leaching uses a common autoclave, the reaction temperature is 150-200℃ (lower than the 800℃ of the traditional roasting process), and the energy consumption is reduced by 40%-50%.

[0013] Reagent saving: The amount of sulfuric acid used is only 0.85-1.0 times the theoretical amount, and the stripping acid concentration gradient design (such as 220g / L H2SO4 for copper stripping and 0.35mol / L for zinc stripping) reduces acid consumption by 15%.

[0014] 3. Technological breakthrough: Enhanced adaptability to complex phases 1. Multi-phase synergistic leaching mechanism: In view of the complexity of the coexistence of oxides (NiO / CuO / PbO / ZnO, etc.), sulfides (Cu2S / Ni3S2 / NiS / FeS, etc.), sulfates, and silicates in the soot, a reduction-oxidation dual-stage design is implemented: Reduction stage: SO2 reduction leaching dissolves high-valent metal elements; oxidation stage: O2 destroys the silicate coating (such as Ni2SiO4 + O2 + 4H⁺→ 2Ni²⁺ + SiO2 + 2H2O), while oxidizing the sulfur added to -2 to SO4 2- , so that the metal sulfide is leached and dissolved.

[0015] 2. Precise control of the extraction system: HBL116 special extractant is used to selectively extract nickel and cobalt at pH 4.0-5.0, with a separation coefficient of >1000 from zinc (compared to P204, which has a zinc-nickel separation coefficient of only 50-100), avoiding co-extraction interference. BRIEF DESCRIPTION OF THE DRAWINGS

[0016] Figure 1 It is a process flow chart of the present invention. DETAILED DESCRIPTION

[0017] In order to make the purpose, technical solutions and advantages of the embodiments of the present invention clearer, the technical solutions in the embodiments of the present invention will be clearly and completely described below. Obviously, the described embodiments are only part of the embodiments of the present invention, not all of the embodiments.

[0018] Example 1 (1) Nickel fly ash and part of the returned raffinate were stirred with sulfuric acid to form a slurry and then added to the autoclave. The amount of sulfuric acid used was 0.85 times the theoretical amount of leached valuable metals. The liquid-solid ratio after slurrying was 10:1. The reaction temperature was 200°C, and SO2 gas was introduced with a SO2 partial pressure of 0.5 MPa. After the reaction time of 2 h, the SO2 was stopped, and the temperature was appropriately lowered and the exhaust was carried out to discharge the residual SO2 gas in the autoclave. (2) introducing oxygen into the reduction leaching reaction slurry obtained in step (1) and controlling the oxygen partial pressure to be 0.3 MPa, carrying out the reaction at 150° C. for 6 h, cooling and releasing the pressure, taking the slurry out of the autoclave, and performing solid-liquid separation on the material to obtain pressure leaching slag (lead-iron slag); (3) The leachate was extracted with lix984, the O / A ratio of the two-stage countercurrent extraction was 2:1, the two-stage countercurrent was washed with a sulfuric acid solution with a pH of 1.5, and two-stage back extraction was performed with a 220 g / L sulfuric acid solution to obtain a copper sulfate solution and a copper raffinate; (4) The copper raffinate is neutralized with sodium carbonate, and the neutralization endpoint pH is controlled to 4.0. The neutralized residue is returned to the pressure leaching process. The neutralized solution is adjusted to pH 2.0 with sulfuric acid and then zinc is extracted with p204. The extraction phase O / A ratio is 1:5. Four-stage countercurrent extraction is performed, and 0.1 mol / L sulfuric acid solution is used for five-stage countercurrent washing. The washing phase O / A ratio is 10:1. 0.35 mol / L sulfuric acid solution is used to back-extract the loaded P204 to obtain zinc sulfate solution and zinc raffinate. (5) After the zinc raffinate was adjusted to pH 5.0 with sodium carbonate, nickel and cobalt were extracted using HBL116 at an extracting phase ratio of 3:1 (O / A). Six stages of countercurrent extraction were performed. The pH of the washing sulfuric acid solution was 1.5. Five stages of countercurrent washing were performed at an O / A ratio of 8:1. The HBL116 was loaded and stripped with 0.75 mol / L sulfuric acid solution in four stages of countercurrent stripping to obtain a nickel-cobalt sulfate solution. The compositions of the nickel-cobalt sulfate solution, copper sulfate solution, zinc sulfate solution, and lead-iron slag are shown in Table 2.

[0019] Example 2 (1) Nickel fly ash and part of the returned raffinate are stirred with sulfuric acid to form a slurry and then added to the autoclave. The amount of sulfuric acid used is 1.0 times the theoretical amount of leached valuable metals. After slurrying, the liquid-solid ratio is 8:1, the pressurization temperature is controlled at 180°C, and SO2 gas is introduced with a SO2 partial pressure of 0.3MPa. After the reaction time is 4h, the SO2 is stopped, and the temperature is appropriately lowered and the exhaust is carried out to discharge the residual SO2 gas in the autoclave; (2) introducing oxygen into the reduction leaching reaction slurry obtained in step (1) to control the oxygen partial pressure to 0.3 MPa, and then cooling and releasing the pressure after keeping the reaction at 160°C for 3 hours. The material is taken out of the autoclave and subjected to solid-liquid separation to obtain pressure leaching slag (lead-iron slag); (3) The leachate was extracted with copper using lix984, the O / A ratio of the two-stage countercurrent extraction was 3:1, the two-stage countercurrent was washed with a sulfuric acid solution with a pH of 1.25, and a two-stage back extraction was performed with a 200 g / L sulfuric acid solution to obtain a copper sulfate solution; (4) The copper raffinate was neutralized with sodium carbonate, and the neutralization endpoint pH was controlled to 4.0. The neutralized solution was adjusted to pH 2.5 with sulfuric acid, and then zinc was extracted with P204. The extraction phase O / A ratio was 1:5. Four-stage countercurrent extraction was performed, and 0.1 mol / L sulfuric acid solution was used for four-stage countercurrent washing. The washing phase O / A ratio was 10:1. 0.30 mol / L sulfuric acid solution was used to back-extract the loaded P204 to obtain a zinc sulfate solution. (5) After adjusting the pH of the zinc raffinate to 5.0 with sodium carbonate, nickel and cobalt were extracted using HBL116 at an extracting phase ratio of 5:1. Six-stage countercurrent extraction was performed. The pH of the washing sulfuric acid solution was 2.5. Four-stage countercurrent washing was performed at an extracting phase ratio of 10:1. The HBL116 was loaded and stripped with 0.70 mol / L sulfuric acid solution in four stages of countercurrent stripping to obtain a nickel-cobalt sulfate solution. The compositions of the nickel-cobalt sulfate solution, copper sulfate solution, zinc sulfate solution, and lead-iron slag are shown in Table 4.

[0020] Example 3 (1) Nickel fly ash and part of the returned raffinate and sulfuric acid solution were stirred and slurried, and then added to the autoclave. The amount of sulfuric acid used was 0.95 times the theoretical amount of leached valuable metals. The liquid-solid ratio after slurrying was 5:1. The pressurization temperature was controlled at 220°C, and SO2 gas was introduced with a SO2 partial pressure of 0.1 MPa. After the reaction time of 6 hours, the SO2 was stopped, and the temperature was appropriately lowered and the exhaust was carried out to discharge the residual SO2 gas in the autoclave. (2) introducing oxygen into the reduction leaching reaction slurry obtained in step (1) to control the oxygen partial pressure to 0.1 MPa, reacting at 200°C for 6 hours, cooling and releasing the pressure, taking out the autoclave, and performing solid-liquid separation on the material to obtain pressure leaching slag (lead-iron slag); (3) The leachate was extracted with lix984, the O / A ratio of the two-stage countercurrent extraction was 1.5:1, the two-stage countercurrent was washed with a sulfuric acid solution with a pH of 2.5, and a two-stage back extraction was performed with a 150 g / L sulfuric acid solution to obtain a copper sulfate solution; (4) The copper raffinate was neutralized with sodium carbonate, and the neutralization endpoint pH was controlled to be 3.0. The neutralized solution was adjusted to pH 1.5 with sulfuric acid, and then zinc was extracted with P204. The extraction phase O / A ratio was 3:1. Six-stage countercurrent extraction was performed, and 0.075 mol / L sulfuric acid solution was used for four-stage countercurrent washing. The washing phase O / A ratio was 8:1. 0.4 mol / L sulfuric acid solution was used to back-extract the loaded P204 to obtain a zinc sulfate solution. (5) After adjusting the pH of the zinc raffinate to 5.0 with sodium carbonate, nickel and cobalt were extracted using HBL116 at an extracting phase ratio of 3:1 (O / A). Six stages of countercurrent extraction were performed. The pH of the washing solution was 2.0 with sulfuric acid. Five stages of countercurrent washing were performed at an O / A ratio of 9:1. The HBL116 was loaded and stripped with 0.60 mol / L sulfuric acid solution for six stages of countercurrent stripping to obtain a nickel and cobalt sulfate solution. The compositions of the nickel and cobalt sulfate solution, copper sulfate solution, zinc sulfate solution, and lead and iron slag are shown in Table 6.

[0021] Example 4 (1) Nickel ash and part of the returned raffinate and sulfuric acid solution are stirred and slurried, and then added to the autoclave. The amount of sulfuric acid used is 1.0 times the theoretical amount of leached valuable metals. After slurrying, the liquid-solid ratio is 7:1, the pressurization temperature is controlled at 180°C, and SO2 gas is introduced with a SO2 partial pressure of 0.3MPa. After the reaction time is 4h, the SO2 is stopped, and the temperature is appropriately lowered and the exhaust is carried out to discharge the residual SO2 gas in the autoclave; (2) introducing oxygen into the reduction leaching reaction slurry obtained in step (1), controlling the oxygen partial pressure to 0.2 MPa, reacting at 180°C for 3 hours, cooling and releasing the pressure, taking out the autoclave, and performing solid-liquid separation on the material to obtain pressure leaching slag (lead-iron slag); (3) The leachate was extracted with lix984, the O / A ratio of the two-stage countercurrent extraction was 1.5:1, the two-stage countercurrent was washed with a sulfuric acid solution with a pH of 2.0, and the two-stage back extraction was carried out with a 180 g / L sulfuric acid solution to obtain a copper sulfate solution; (4) The copper raffinate was neutralized with sodium carbonate, and the neutralization endpoint pH was controlled at 3.5. The neutralized solution was adjusted to pH 2.0 with sulfuric acid, and then zinc was extracted with P204. The extraction phase O / A ratio was 2:1. Four-stage countercurrent extraction was performed, and 0.05 mol / L sulfuric acid solution was used for four-stage countercurrent washing. The washing phase O / A ratio was 10:1. 0.35 mol / L sulfuric acid solution was used to back-extract the loaded P204 to obtain a zinc sulfate solution. (5) After the zinc raffinate was adjusted to pH 5.0 with sodium carbonate, nickel and cobalt were extracted with HBL116 at an extracting phase ratio of 2:1, followed by four-stage countercurrent extraction. The pH of the washing sulfuric acid solution was 1.5, followed by four-stage countercurrent washing with an O / A ratio of 8:1. The HBL116 was loaded and stripped with 0.50 mol / L sulfuric acid solution in four stages of countercurrent stripping to obtain a nickel and cobalt sulfate solution. The compositions of the nickel and cobalt sulfate solution, copper sulfate solution, zinc sulfate solution, and lead and iron slag are shown in Table 8: .

Claims

1. A method for recovering valuable metals from nickel ash, characterized in that: The following steps are involved: (1) Reduction pressure leaching: nickel ash, raffinate and sulfuric acid solution are slurried and added to a pressure autoclave, the temperature is raised and sulfur dioxide gas is introduced to react with stirring. The reaction temperature is 150-200°C and the sulfur dioxide gas partial pressure is 0.1-0.5 MPa. The reaction is carried out for 2-6 hours. After the reaction is completed, the introduction of sulfur dioxide gas is stopped to obtain a reduction leaching reaction slurry; (2) Oxidative pressure leaching: oxygen is introduced into the reduction leaching reaction slurry obtained in step (1) to continue the reaction, the reaction temperature is controlled at 150-200°C, the oxygen partial pressure is 0.1-0.3 MPa, the reaction time is 2-6 hours, and the pH value at the reaction end point is 1.0-2.

0. After the reaction is completed, solid-liquid separation is performed to obtain leachate and lead-iron slag; (3) Extraction and separation: The leachate obtained in step (2) is sequentially extracted with lix984 for copper, P204 for zinc, and HBL116 for nickel and cobalt to obtain copper sulfate solution, zinc sulfate solution, and nickel and cobalt sulfate solution, respectively.

2. The method according to claim 1, characterized in that The chemical composition of the nickel soot in step (1) includes: nickel 4%-9%, copper 4%-8%, cobalt 0.1%-0.3%, iron 10%-30%, zinc 0.2%-7%, lead 0.3%-10%, and arsenic 0.2%-0.4%.

3. The method according to claim 1, characterized in that In step (1), the liquid-to-solid ratio after slurrying is 4-10:1, and the amount of sulfuric acid used is 0.85-1.0 times the theoretical amount of leaching valuable metals.

4. The method according to claim 1, wherein In step (3): Copper extraction: Use lix984 two-stage countercurrent extraction, the extraction phase O / A ratio is 1.5:1-3:1, the two-stage countercurrent is washed with sulfuric acid solution with a pH of 1.5-2.5, and two-stage back extraction is carried out with 150-220g / L sulfuric acid solution to obtain a copper sulfate solution containing 45-55g / L copper and a copper raffinate; Zinc extraction: 4-6 stages of countercurrent extraction with P204, an extraction phase ratio of 1:5-3:1, 4-5 stages of countercurrent washing with 0.05-0.1 mol / L sulfuric acid solution, an O / A ratio of 8:1-10:1, and 4-5 stages of countercurrent stripping with 0.35-0.4 mol / L sulfuric acid solution loaded with P204 to obtain zinc sulfate solution and zinc raffinate. Nickel and cobalt extraction: Use HBL116 4-6 stage countercurrent extraction, the extraction phase O / A ratio is 1.5:1-5:1, use sulfuric acid solution with a pH of 1.5-2.5 for 4-5 stage countercurrent washing, the wash phase O / A ratio is 8:1-10:1, and load HBL116 back extraction uses 0.5-0.75 mol / L sulfuric acid solution for 4-6 stage countercurrent stripping to obtain nickel and cobalt sulfate solution.

5. The method according to claim 4, characterized in that The copper raffinate is neutralized to a pH of 2.5-4.0 with sodium carbonate, and then the pH is adjusted to 2.0-2.5 before zinc extraction; the zinc raffinate is adjusted to a pH of 4.0-5.0 with sodium carbonate before nickel and cobalt extraction.

6. The method according to any one of claims 1 to 5, characterized in that Part of the raffinate after nickel and cobalt extraction is returned to the slurrying process in step (1); the neutralized slag produced after the copper raffinate is neutralized with sodium carbonate is returned to the reduction pressure leaching process in step (1).

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