Recovery method of precipitation slag produced in copper anode electrolytic refining process

Abstract

The invention discloses a method for recycling precipitation residues produced in the copper anode electrolytic refining process, and relates to the technical field of copper anode electrolytic refining. According to the method, the precipitation slag containing arsenic, tin, nickel, antimony and copper is treated, the physically entrained electrolyte is removed through water washing, most of copper, tin, antimony and the like are separated through a simple and low-cost flotation process, the obtained flotation copper concentrate is subjected to alkali liquor leaching to remove arsenic, antimony and tin, and the flotation effect is good. The leaching liquid is subjected to oxidative hydrolysis to remove antimony, and then calcium stannate slag and calcium arsenate slag are obtained through fractional precipitation of lime. The recycling method provided by the invention is low in cost, and resource utilization of arsenic, tin, antimony, nickel and copper can be realized.

Description

Technical Field

[0001] This invention relates to the field of copper anode electrolytic refining technology, and more specifically, to a method for recovering precipitated residue produced in the copper anode electrolytic refining process. Background Technology

[0002] The precipitate residue produced during the electrolytic refining of high-arsenic, high-tin, and high-nickel copper anodes is rich in copper, nickel, tin, arsenic, antimony, and lead, as well as trace amounts of precious metals. It is primarily used for wet oxygen pressure leaching or strong oxidative leaching of copper and nickel. The remaining residue is then subjected to strong reduction sulfidation smelting to recover valuable metals such as tin and antimony. However, oxygen pressure leaching requires expensive pressure vessel equipment, while direct strong reduction sulfidation smelting requires batching due to the high copper content, diluting the enrichment ratio of precious metals. Furthermore, strong reduction sulfidation smelting requires a large amount of sulfidizing agent, and the resulting copper product (copper matte produced by strong reduction sulfidation smelting) has a very low grade, significantly increasing the cost of strong reduction sulfidation smelting and subsequent deep processing.

[0003] There is an urgent need to develop a recycling process for the precipitate residue produced in the electrolytic refining process of high-arsenic, high-tin, and high-nickel copper anodes, so as to achieve low-cost recycling and utilization of resources.

[0004] In view of this, the present invention is proposed. Summary of the Invention

[0005] The purpose of this invention is to provide a method for recovering precipitated residue produced in the copper anode electrolytic refining process, aiming to provide a simple and easy-to-implement process to achieve the separation of arsenic, tin, antimony, nickel and copper.

[0006] This invention is implemented as follows: In a first aspect, the present invention provides a method for recovering precipitated residue produced in a copper anode electrolytic refining process, comprising: providing precipitated residue containing arsenic, tin, nickel, antimony and copper; The precipitate was washed with water, and the solid-liquid separation was performed to obtain the washing residue. The washing residue is ground and floated to obtain flotation copper concentrate and flotation tailings; The flotation copper concentrate and alkaline solution are mixed and leached. After solid-liquid separation, leaching residue and leaching solution are obtained. The leaching residue is then used in the copper-nickel recovery process. The leachate was subjected to oxidative hydrolysis to remove antimony, and after solid-liquid separation, antimony salt products and antimony precipitate filtrate were obtained. The antimony filtrate was treated by stepwise lime precipitation, and calcium stannate slag and calcium arsenate slag were recovered in sequence.

[0007] In an optional implementation, a countercurrent washing method is used for washing, and the amount of fresh washing water is controlled to be 2-3 times the weight of the sediment, with a washing stage of 3 or more.

[0008] In an optional implementation, the washing slag is ground to a particle size of -325 mesh with a mass ratio of more than 80%, and the slurry is adjusted to a pH of 7-10.

[0009] In an optional embodiment, butyl xanthate and pine oil are added during the flotation process. The amount of butyl xanthate added is (80-150) g / t dry basis material, and the amount of pine oil added is (15-25) g / t dry basis material. And / or, after flotation, the total mass fraction of copper and nickel in the flotation copper concentrate is ≥55%, and the distribution ratio of copper in the flotation copper concentrate is ≥80%, wherein the flotation copper concentrate is used for alkaline removal of tin, antimony and arsenic before copper and nickel are recovered. And / or, after flotation, the total mass fraction of copper and nickel in the flotation tailings is ≤5%, and the distribution ratio of tin and antimony in the flotation tailings is ≥85%; the flotation tailings are used for strong reduction sulfidation smelting to recover tin and antimony, copper and nickel are recovered in the form of matte, and tin and antimony are recovered in the flue dust.

[0010] In an optional embodiment, the alkaline solution is an aqueous solution of sodium hydroxide with a concentration of 0.75 mol / L to 1.50 mol / L; And / or, the amount of alkali solution used is 1.3 to 1.5 times the theoretical amount needed to dissolve tin, antimony, and arsenic; And / or, the leaching temperature is 80℃-90℃, the leaching time is 30min-120min, and the liquid-to-solid ratio is 3-10L / kg.

[0011] In an optional embodiment, the oxidant used in the process of descaling antimony by oxidative hydrolysis is selected from hydrogen peroxide, air, or oxygen.

[0012] In an optional embodiment, when hydrogen peroxide is used as the oxidant, the amount of hydrogen peroxide is controlled to be 2-3 times the theoretical amount. The reaction is carried out at room temperature without additional heating or heat preservation, and the reaction time is 60-120 minutes. The theoretical amount of hydrogen peroxide is calculated based on 1 mol of antimony equaling 2 mol of hydrogen peroxide.

[0013] And / or, when the oxidant is air or oxygen, the reaction temperature is 60℃-90℃ and the reaction time is 90min-180min.

[0014] In an optional embodiment, the process of recovering calcium stannate slag includes: mixing lime and antimony precipitation filtrate, and controlling the molar ratio of effective calcium oxide to tin to be (1.1-1.5):1; Preferably, the lime is added in at least two separate additions, with each addition not exceeding 50% of the total mass. The interval between each addition is 20-30 minutes, and the reaction temperature is controlled at 80-90°C. After the last addition, the reaction continues for 60-120 minutes.

[0015] In an optional embodiment, the process of recovering calcium arsenate slag includes: mixing the liquid to be treated obtained after recovering calcium stannate slag with lime, controlling the molar ratio of effective calcium oxide to arsenic to be (2.0-2.5):1, the reaction is carried out at room temperature without additional heating or heat preservation, and the reaction time is 60min-120min. And / or, after recovering the calcium arsenate residue, perform solid-liquid separation, and return the resulting filtrate to the alkaline leaching stage after adding alkali.

[0016] In an optional embodiment, the precipitate contains, by mass fraction (dry basis), 10%~20% arsenic, 15%~25% tin, 2%~5% nickel, 7%~15% antimony, and 15%~25% copper; And / or, the mass percentage of sulfates, elements, alloy phases and sulfides in the copper and nickel phases of the precipitate is >70%.

[0017] This invention offers the following advantages: It treats precipitated slag containing arsenic, tin, nickel, antimony, and copper. First, the slag is washed with water to remove physically entrained electrolyte. A simple, low-cost flotation process separates most of the copper from tin and antimony. The resulting copper concentrate is then leached with alkaline solution to remove arsenic, antimony, and tin. After antimony removal through oxidation and hydrolysis, the leachate is further precipitated with lime to obtain calcium stannate slag and calcium arsenate slag. The recovery method provided by this invention is low-cost and enables the resource utilization of arsenic, tin, antimony, nickel, and copper. Attached Figure Description

[0018] To more clearly illustrate the technical solutions of the embodiments of the present invention, the accompanying drawings used in the embodiments will be briefly introduced below. It should be understood that the following drawings only show some embodiments of the present invention and should not be regarded as a limitation on the scope. For those skilled in the art, other related drawings can be obtained based on these drawings without creative effort.

[0019] Figure 1 The present invention provides a process flow diagram of a method for recovering precipitated residue produced during the electrolytic refining process of copper anodes. Detailed Implementation

[0020] To make the objectives, 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. Where specific conditions are not specified in the embodiments, conventional conditions or conditions recommended by the manufacturer shall apply. Reagents or instruments whose manufacturers are not specified are all conventional products that can be purchased commercially.

[0021] High-arsenic, high-tin, and high-nickel copper anode electrolytic refining is a process of further purifying crude copper obtained through pyrometallurgical processes to remove impurities (such as gold, silver, arsenic, antimony, tin, bismuth, nickel, lead, selenium, and tellurium) to obtain high-purity cathode copper suitable for high-end applications in electrical and electronic fields. The removed impurities are enriched in the precipitate residue. The recovery method provided by this invention is for treating the precipitate residue produced during the copper anode electrolytic refining process to recover arsenic, tin, nickel, antimony, and copper separately.

[0022] This invention provides a method for recovering precipitated residue produced during the electrolytic refining process of copper anodes, such as... Figure 1 As shown, the steps are as follows: S1, providing sediment. The precipitate residue produced during the electrolytic refining process of copper anode is collected. The precipitate residue contains arsenic, tin, nickel, antimony and copper. By mass fraction (dry basis), the precipitate residue contains 10%~20% arsenic, 15%~25% tin, 2%~5% nickel, 7%~15% antimony and 15%~25% copper.

[0023] Specifically, the mass percentage of sulfates, elements, alloy phases and sulfides in the copper and nickel phases of the precipitate is >70%, meaning that most of the copper and nickel phases exist in the form of copper sulfate (nickel sulfate), copper-nickel alloys and cuprous sulfide (nickel sulfide).

[0024] S2, Water Washing The precipitate is washed with water, and the solid-liquid separation yields the washed residue. Simple water washing removes the physically entrained electrolyte, and the filtrate obtained from the solid-liquid separation is returned to the electrolysis system for reuse, without increasing the cost of metal recovery. Especially during high-nickel anode plate electrolysis, the nickel removal rate of this precipitate can reach 50%-60%, significantly shortening the nickel recovery process path.

[0025] In some embodiments, countercurrent washing (or "countercurrent leaching") is used for washing, with the amount of fresh washing water controlled to be 2-3 times the weight of the precipitate, and the number of washing stages being greater than or equal to 3, in order to fully remove the entrained electrolytes.

[0026] S3, Grinding The washing residue obtained in step S2 is ground, and after grinding, it is classified and screened. The material that meets the particle size requirements (i.e., the material on the screen) enters step S4.

[0027] In some embodiments, after grinding the washing slag, the proportion of particles with a mass ratio of -325 mesh is greater than 80%, and the larger particles after screening can be further ground. The resulting slurry is adjusted to a pH of 7-10, such as 7, 8, 9, or 10.

[0028] S4, Flotation The slurry obtained after grinding is subjected to flotation, yielding flotation copper concentrate and flotation tailings. The flotation tailings are used for strong reduction sulfidation smelting to recover tin and antimony, while the flotation copper concentrate proceeds to the next stage. The working principle and equipment of flotation can be found in existing technologies and will not be elaborated upon here.

[0029] Specifically, the lower copper grade of the tailings obtained by flotation can meet the standard for direct feeding into the furnace in the strong reduction sulfidation smelting process. In addition, the small amount of copper also serves the purpose of enriching precious metals (using matte produced in the strong reduction sulfidation smelting process to dissolve and enrich precious metals). At the same time, because the copper grade of the material is reduced, the amount of sulfiding agent used in the strong reduction sulfidation smelting process can be greatly reduced. Furthermore, the copper grade of flotation refining is much higher than that of matte obtained from strong reduction sulfidation smelting, which is more conducive to deep processing and reduces the overall production cost.

[0030] In some embodiments, butyl xanthate and pine oil are added during the flotation process, with butyl xanthate acting as a collector and pine oil acting as a frother. The amount of butyl xanthate added is (80-150) g / t dry basis material, and the amount of pine oil added is (15-25) g / t dry basis material.

[0031] After flotation, the total mass fraction of copper and nickel in the flotation copper concentrate is ≥55%, with copper accounting for ≥80% of the concentrate, and tin and antimony each accounting for ≤15%. The flotation copper concentrate is used for alkaline removal of tin, antimony, and arsenic before copper and nickel recovery. After flotation, the total mass fraction of copper and nickel in the flotation tailings is ≤5%, with tin and antimony each accounting for ≥85%. The flotation tailings are used for strong reducing sulfidation smelting to recover tin and antimony, while copper and nickel are recovered as matte, and tin and antimony are recovered in the flue dust. Flotation effectively separates copper and nickel from tin and antimony, allowing the flotation copper concentrate to proceed to subsequent separation stages. This simplest and lowest-cost flotation process achieves the separation of most copper from tin and antimony, with the unit processing cost of flotation typically being 10-15% of the cost of pyrometallurgical processing.

[0032] Specifically, "copper-nickel total grade" refers to the total content of copper and nickel, two valuable metal elements, in the material, expressed as a percentage by mass.

[0033] Specifically, the "distribution ratio" in the flotation process refers to the ratio of the mass of valuable metals distributed in the flotation concentrate or flotation tailings to their total mass (the sum of the masses of metals in the flotation concentrate and flotation tailings).

[0034] S5, alkaline leaching The copper concentrate from flotation is mixed with alkaline solution for leaching. After solid-liquid separation, leaching residue and leachate are obtained. The leaching residue is then used in the copper-nickel recovery process. Although only small amounts of arsenic, tin, and antimony are distributed in the copper concentrate during flotation, this still leads to the dispersion of antimony and tin resources and increases the impurity removal costs in the copper pyrometallurgical refining process. Alkaline leaching not only removes and separates arsenic, tin, and antimony for recovery, but also significantly reduces the processing cost of alkaline leaching through calcification precipitation regeneration of the alkaline solution.

[0035] In some embodiments, the alkaline solution used is an aqueous sodium hydroxide solution with a concentration of 0.75 mol / L to 1.50 mol / L, specifically 0.75 mol / L, 0.80 mol / L, 0.90 mol / L, 1.00 mol / L, 1.10 mol / L, 1.20 mol / L, 1.30 mol / L, 1.40 mol / L, 1.50 mol / L, etc. The amount of alkaline solution used is 1.3 to 1.5 times the theoretical amount required to dissolve tin, antimony, and arsenic, such as 1.3, 1.4, or 1.5 times. The calculation method for the amount of sodium hydroxide used is as follows: n(NaOH) = (1.3~1.5) × {3n(As) + n(Sb) + 2n(Sn)} (Note: n represents the molar amount, the same below).

[0036] Furthermore, the leaching temperature is 80℃-90℃, such as 80℃, 83℃, 85℃, 88℃, 90℃, etc.; the leaching time is 30min-120min, such as 30min, 50min, 80min, 100min, 120min, etc.; the liquid-to-solid ratio is 3-10 L / kg, such as 3L / kg, 4L / kg, 5L / kg, 6L / kg, 7L / kg, 8L / kg, 9L / kg, 10L / kg, etc.

[0037] S6, Oxidative hydrolysis The leachate is subjected to oxidative hydrolysis to remove antimony, and after solid-liquid separation, antimony salt products (such as basic sodium antimonate) and antimony precipitate filtrate are obtained.

[0038] In some embodiments, the oxidant used in the descaling process of antimony via oxidative hydrolysis is selected from hydrogen peroxide, air, or oxygen, and the oxidant can be any one or more of the above. When hydrogen peroxide is selected as the oxidant, the amount of hydrogen peroxide is controlled to be 2-3 times the theoretical amount required for the complete reaction of antimony, so as to ensure that the antimony reacts fully. The reaction is carried out at room temperature without additional heating or heat preservation, specifically at room temperature; the reaction time is 60 min-120 min, such as 60 min, 70 min, 80 min, 90 min, 100 min, 110 min, 120 min, etc.

[0039] In some embodiments, when the oxidant used in the descaling process of antimony by oxidative hydrolysis is air or oxygen, the reaction temperature is 60℃-90℃, such as 60℃, 70℃, 80℃, 90℃, etc.; the reaction time is 90min-180min, such as 90min, 100min, 110min, 130min, 140min, 150min, 160min, 170min, 180min, etc.

[0040] S7, lime stepwise precipitation The antimony filtrate was treated by stepwise lime precipitation to recover calcium stannate slag and calcium arsenate slag sequentially. Taking advantage of the significantly lower solubility product constant (Ksp) of calcium stannate compared to calcium arsenate, the calcium stannate slag and calcium arsenate slag were precipitated stepwise.

[0041] In some embodiments, the process of recovering calcium stannate slag includes: mixing lime and antimony precipitation filtrate, controlling the molar ratio of effective calcium oxide to tin to be (1.1-1.5):1, such as 1.1:1, 1.2:1, 1.3:1, 1.4:1, 1.5:1, etc., and the lime can be quicklime or hydrated lime. The lime is added at least twice (or three times), and regardless of the number of additions, the amount added each time should not exceed 50% of the total mass. If added in three batches, each addition can be 20%, 40%, 40%, or 40%, 40%, 20%, etc. The interval between each addition is 20-30 minutes, such as 20 minutes, 25 minutes, 30 minutes, etc. The reaction temperature is controlled at 80℃-90℃, such as 80℃, 83℃, 85℃, 88℃, 90℃, etc. After the final addition, continue the reaction for 60-120 minutes, such as 60, 70, 80, 90, 100, 110, or 120 minutes. After the reaction, the tin precipitation rate should be ≥95%, and the arsenic precipitation rate ≤1%.

[0042] In some embodiments, the process of recovering calcium arsenate slag includes: mixing the liquid to be treated obtained after recovering calcium stannate slag with lime, controlling the molar ratio of effective calcium oxide to arsenic to be (2.0-2.5):1, such as 2.0:1, 2.1:1, 2.2:1, 2.3:1, 2.4:1, 2.5:1, etc., and the lime can be quicklime or hydrated lime. The reaction is carried out at room temperature without additional heating or heat preservation; the reaction time is 60min-120min, such as 60min, 70min, 80min, 90min, 100min, 110min, 120min, etc. After the reaction, the precipitation rate of arsenic is detected to be ≥98%. After recovering the calcium arsenate slag, solid-liquid separation is performed by filtration, and the obtained filtrate is a sodium hydroxide solution, which is directly returned to the front-end alkaline leaching after alkali replenishment.

[0043] The features and performance of the present invention will be further described in detail below with reference to embodiments.

[0044] Example 1 This embodiment provides a method for recovering precipitated residue produced during the electrolytic refining process of copper anodes, such as... Figure 1 As shown, the steps are as follows: (1) Provide sediment The precipitate residue produced during the electrolytic refining process of copper anodes was collected. By mass fraction, the precipitate residue contained 15.4% arsenic, 18.1% tin, 2.7% nickel, 8.2% antimony, and 21.5% copper. The copper and nickel phases in the precipitate residue accounted for 75.3% by mass of sulfates, elemental phases, alloy phases, and sulfides. In other words, most of the copper and nickel phases existed in the form of copper sulfate (nickel sulfate), copper-nickel alloys, and cuprous sulfide (nickel sulfide).

[0045] (2) Washing with water The precipitated residue was washed countercurrently, filtered, and the resulting washed residue was returned to the electrolysis system. The volume of fresh wash water was controlled to be three times the weight of the precipitated residue, and the washing process consisted of three stages. Testing showed that the nickel removal rate reached 61.8%.

[0046] (3) Grinding The washing residue obtained in step (2) is ground and then screened. The material that meets the particle size requirements (i.e., the oversize material) enters step (4). After grinding the washing residue, the mass ratio of particles with a mesh size of -325 is greater than 80%. The pH of the resulting slurry is adjusted to 8.

[0047] (4) Flotation The slurry obtained after grinding is subjected to flotation, and flotation yields flotation copper concentrate and flotation tailings. The flotation tailings are used for strong reduction sulfidation smelting to recover tin and antimony, while the flotation copper concentrate proceeds to step (5).

[0048] Butyl xanthate and pine oil are added during the flotation process. The dosage of butyl xanthate is 120 g / t dry basis material, and the dosage of pine oil is 20 g / t dry basis material.

[0049] After testing, the total mass fraction of copper and nickel in the flotation copper concentrate was 56.5%, with copper accounting for 85.9%, tin for 8.8%, and antimony for 7.5%. After flotation, the total grade of copper and nickel in the flotation tailings was 4.5%, with tin accounting for 91.2% and antimony for 92.5%. These materials can be directly used for strong reducing sulfidation smelting to recover tin and antimony and enrich precious metals.

[0050] (5) Alkaline leaching The copper concentrate from flotation is mixed with alkaline solution for leaching. After filtration, leaching residue and leaching solution are obtained. The leaching residue is then used in the copper and nickel recovery process.

[0051] The alkaline solution was a 1.0 mol / L sodium hydroxide aqueous solution, and the amount of sodium hydroxide used was 1.4 times the theoretical amount required for dissolving tin, antimony, and arsenic. The leaching temperature was 85℃, and the leaching time was 80 min. The leaching rate of arsenic was 85.6%, and the leaching rate of tin was 38.1%.

[0052] (6) Oxidative hydrolysis The leachate was subjected to oxidative hydrolysis to remove antimony, and after filtration, sodium antimonyate and antimony precipitate filtrate were obtained. Hydrogen peroxide was used as the oxidant in the oxidative hydrolysis process, and the amount of hydrogen peroxide was controlled to be 2.5 times the theoretical amount required for complete antimony reaction. The reaction temperature was room temperature, and the reaction time was 100 min.

[0053] (7) Stepwise precipitation of lime The antimony filtrate was treated by stepwise lime precipitation, and calcium stannate slag and calcium arsenate slag were recovered in sequence.

[0054] The process for recovering calcium stannate slag includes: mixing quicklime and antimony precipitation filtrate, controlling the molar ratio of effective calcium oxide to tin to be 1.3:1. The quicklime is added in three portions, with the first addition accounting for 40% of the total mass. The remaining amount is added in two subsequent equal portions, 25 minutes apart, while maintaining the reaction temperature at 85°C. After the final addition, the reaction continues for 100 minutes. Post-reaction analysis shows that the tin precipitation rate is 97.5%, and the arsenic precipitation rate is 0.72%.

[0055] The process of recovering calcium arsenate slag includes: mixing the liquid to be treated obtained after recovering calcium stannate slag with quicklime, controlling the molar ratio of effective calcium oxide to arsenic to be 2.3:1, controlling the reaction temperature to be room temperature, and the reaction time to be 100 min. Post-reaction testing showed that the arsenic precipitation rate was 98.7%. After recovering the calcium arsenate slag, solid-liquid separation was performed by filtration, and the resulting filtrate was a sodium hydroxide solution, which was directly returned to the upstream alkaline leaching stage for reuse.

[0056] Tests showed that the tin content in the calcium stannate in this embodiment was 33.5%, which can be used as a synthetic tin concentrate for tin recovery; the arsenic content in the calcium arsenate was 20.7%, with trace amounts of other valuable metals, which need to be outsourced for harmless treatment.

[0057] Example 2 This embodiment provides a method for recovering precipitated residue produced during the electrolytic refining process of copper anodes, such as... Figure 1 As shown, the steps are as follows: (1) Provide sediment Same as Example 1 (2) Washing with water The precipitated residue was washed countercurrently, filtered, and the resulting washed residue was returned to the electrolysis system. The volume of fresh wash water was controlled to be three times the weight of the precipitated residue, and the washing process consisted of three stages. Testing showed that the nickel removal rate reached 62.1%.

[0058] (3) Grinding The washing residue obtained in step (2) is ground and then screened. The material that meets the particle size requirements (i.e., the oversize material) enters step (4). After grinding the washing residue, the mass ratio of particles with a mesh size of -325 is greater than 80%, and the pH of the resulting slurry is adjusted to 7.

[0059] (4) Flotation The slurry obtained after grinding is subjected to flotation, and flotation yields flotation copper concentrate and flotation tailings. The flotation tailings are used for strong reduction sulfidation smelting to recover tin and antimony, while the flotation copper concentrate proceeds to step (5).

[0060] Butyl xanthate and pine oil are added during the flotation process. The dosage of butyl xanthate is 80 g / t dry basis material, and the dosage of pine oil is 15 g / t dry basis material.

[0061] After testing, the total mass fraction of copper and nickel in the flotation copper concentrate was 55.7%, with copper accounting for 82.4%, tin for 6.9%, and antimony for 6.2%. After flotation, the total grade of copper and nickel in the flotation tailings was 4.6%, with tin accounting for 93.1% and antimony for 93.8%. These materials can be directly used for strong reducing sulfidation smelting to recover tin and antimony and enrich precious metals.

[0062] (5) Alkaline leaching The copper concentrate from flotation is mixed with alkaline solution for leaching. After filtration, leaching residue and leaching solution are obtained. The leaching residue is then used in the copper and nickel recovery process.

[0063] The alkaline solution was a 0.75 mol / L sodium hydroxide aqueous solution, and the amount of sodium hydroxide used was 1.3 times the theoretical amount. The leaching temperature was 80℃, the leaching time was 120 min, and the leaching rate of arsenic was 88.2% and the leaching rate of tin was 40.3%.

[0064] (6) Oxidative hydrolysis The leachate was subjected to oxidative hydrolysis to remove antimony, and after filtration, sodium antimonyate and antimony-precipitated filtrate were obtained. Air was used as the oxidant in the oxidative hydrolysis process, the reaction temperature was 60℃, and the reaction time was 180 min.

[0065] (7) Stepwise precipitation of lime The antimony filtrate was treated by stepwise lime precipitation, and calcium stannate slag and calcium arsenate slag were recovered in sequence.

[0066] The process for recovering calcium stannate slag includes: mixing quicklime and antimony precipitation filtrate, controlling the molar ratio of effective calcium oxide to tin to be 1.1:1. The quicklime is added in three portions, with the first addition accounting for 50% of the total mass. The remaining amount is added in two subsequent equal portions, 20 minutes apart, while maintaining the reaction temperature at 80°C. After the final addition, the reaction continues for 120 minutes. Post-reaction analysis shows a tin precipitation rate of 96.1% and an arsenic precipitation rate of 0.38%.

[0067] The process of recovering calcium arsenate slag includes: mixing the liquid to be treated obtained after recovering calcium stannate slag with quicklime, controlling the molar ratio of effective calcium oxide to arsenic to be 2.0:1, the reaction temperature to be room temperature, and the reaction time to be 120 min. Post-reaction testing showed that the arsenic precipitation rate was 98.2%. After recovering the calcium arsenate slag, solid-liquid separation was performed by filtration, and the resulting filtrate was a sodium hydroxide solution, which was directly returned to the upstream alkaline leaching stage for reuse.

[0068] Example 3 This embodiment provides a method for recovering precipitated residue produced during the electrolytic refining process of copper anodes, such as... Figure 1 As shown, the steps are as follows: (1) Provide sediment Same as Example 1 (2) Washing with water The precipitated residue was washed countercurrently, filtered, and the resulting washed residue was returned to the electrolysis system. The volume of fresh wash water was controlled to be three times the weight of the precipitated residue, and the washing process consisted of three stages. Testing showed that the nickel removal rate reached 62.9%.

[0069] (3) Grinding The washing residue obtained in step (2) is ground and then screened. The material that meets the particle size requirements (i.e., the oversize material) enters step (4). After grinding the washing residue, the mass ratio of particles with a mesh size of -325 is greater than 80%. The pH value of the resulting slurry is adjusted to 10.

[0070] (4) Flotation The slurry obtained after grinding is subjected to flotation, and flotation yields flotation copper concentrate and flotation tailings. The flotation tailings are used for strong reduction sulfidation smelting to recover tin and antimony, while the flotation copper concentrate proceeds to step (5).

[0071] Butyl xanthate and pine oil are added during the flotation process. The dosage of butyl xanthate is 150 g / t dry basis material, and the dosage of pine oil is 25 g / t dry basis material.

[0072] After testing, the total mass fraction of copper and nickel in the flotation copper concentrate was 56.2%, with copper accounting for 83.1%, tin for 7.4%, and antimony for 6.8%. In the flotation tailings, the total grade of copper and nickel was 3.8%, with tin accounting for 92.6% and antimony for 93.2%. These materials can be directly used for strong reducing sulfidation smelting to recover tin and antimony and enrich precious metals.

[0073] (5) Alkaline leaching The copper concentrate from flotation is mixed with alkaline solution for leaching. After filtration, leaching residue and leaching solution are obtained. The leaching residue is then used in the copper and nickel recovery process.

[0074] The alkaline solution was a 1.5 mol / L sodium hydroxide aqueous solution, with the amount of sodium hydroxide used being 1.5 times the theoretical amount. The leaching temperature was 90℃, and the leaching time was 30 min. The leaching rate of arsenic was 87.6%, and the leaching rate of tin was 35.8%.

[0075] (6) Oxidative hydrolysis The leachate was subjected to oxidative hydrolysis to remove antimony, and after filtration, sodium antimony basicate and antimony-precipitated filtrate were obtained. Oxygen was used as the oxidant in the oxidative hydrolysis process, the reaction temperature was 90℃, and the reaction time was 90 min.

[0076] (7) Stepwise precipitation of lime The antimony filtrate was treated by stepwise lime precipitation, and calcium stannate slag and calcium arsenate slag were recovered in sequence.

[0077] The process for recovering calcium stannate slag includes: mixing quicklime and antimony precipitation filtrate, controlling the molar ratio of effective calcium oxide to tin to be 1.5:1. The quicklime is added in three portions, with the first addition accounting for 45% of the total mass. The remaining amount is added in two subsequent equal portions, 30 minutes apart, while maintaining the reaction temperature at 90℃. After the final addition, the reaction continues for 60 minutes. Post-reaction analysis shows a tin precipitation rate of 98.9% and an arsenic precipitation rate of 0.92%.

[0078] The process of recovering calcium arsenate slag includes: mixing the liquid to be treated obtained after recovering calcium stannate slag with quicklime, controlling the molar ratio of effective calcium oxide to arsenic at 2.5:1, controlling the reaction temperature at 35℃, and the reaction time at 60 min. Post-reaction testing showed that the arsenic precipitation rate was 99.1%. After recovering the calcium arsenate slag, solid-liquid separation was performed by filtration, and the resulting filtrate was a sodium hydroxide solution, which was directly returned to the upstream alkaline leaching stage for reuse.

[0079] Comparative Example 1 The only difference from Example 1 is that lime is added all at once when recovering calcium stannate slag in step (7).

[0080] Tests showed that the precipitation rate of tin in this comparative example was 92.6%, while the precipitation rate of arsenic was 5.72%, indicating a significant increase in the precipitation rate of arsenic.

[0081] Comparative Example 2 The only difference from Example 1 is that when recovering calcium stannate slag in step (7), the amount of lime added for the first time is 70%.

[0082] Tests showed that the precipitation rate of tin in this comparative example was 94.2%, while the precipitation rate of arsenic was 1.42%, indicating that the precipitation rate of arsenic was relatively high.

[0083] Comparative Example 3 The only difference from Example 1 is that the leaching temperature in step (5) is room temperature.

[0084] Tests showed that the leaching rate of arsenic in this comparative sample was 43.6%, and the leaching rate of tin was 29.7%, both of which were relatively low.

[0085] Comparative Example 4 The only difference from Example 1 is that in step (7), the molar ratio of effective calcium oxide to tin is controlled to be 2:1.

[0086] Tests showed that the precipitation rate of tin in this comparative example was 99.7%, and the precipitation rate of arsenic was 37.3%. Tin precipitation was very thorough, but tin and arsenic were not separated, and the arsenic content in the tin slag was 8.7%.

[0087] The above description is merely a preferred embodiment of the present invention and is not intended to limit the invention. Various modifications and variations can be made to the present invention by those skilled in the art. Any modifications, equivalent substitutions, improvements, etc., made within the spirit and principles of the present invention should be included within the scope of protection of the present invention.

Claims

1. A method for recovering precipitated residue produced in the copper anode electrolytic refining process, characterized in that, include: A precipitate is provided, the precipitate containing arsenic, tin, nickel, antimony and copper; The precipitate was washed with water, and the solid-liquid separation was performed to obtain the washing residue. The washing residue is ground and flotated to obtain flotation copper concentrate and flotation tailings; The flotation copper concentrate and alkaline solution are mixed and leached. After solid-liquid separation, leaching residue and leaching solution are obtained. The leaching residue is then used in the copper-nickel recovery process. The leachate was subjected to oxidative hydrolysis to remove antimony, and after solid-liquid separation, antimony salt products and antimony precipitate filtrate were obtained. The antimony filtrate was treated by stepwise lime precipitation to recover calcium stannate slag and calcium arsenate slag in sequence.

2. The recycling method according to claim 1, characterized in that, The washing process is carried out using a counter-current water washing method, with the amount of fresh washing water controlled to be 2 to 3 times the weight of the sediment, and the number of washing stages is greater than or equal to 3.

3. The recycling method according to claim 1, characterized in that, After grinding the washing residue, the mass ratio of particles with a mesh size of -325 is greater than 80%, and the pH of the slurry is adjusted to 7-10.

4. The recycling method according to claim 1, characterized in that, Butyl xanthate and pine oil are added during the flotation process. The amount of butyl xanthate added is (80-150) g / t dry basis material, and the amount of pine oil added is (15-25) g / t dry basis material. And / or, after flotation, the total mass fraction of copper and nickel in the flotation copper concentrate is ≥55%, the distribution ratio of copper in the flotation copper concentrate is ≥80%, and the flotation copper concentrate is used for alkaline removal of tin, antimony and arsenic before copper and nickel are recovered. And / or, after flotation, the total mass fraction of copper and nickel in the flotation tailings is ≤5%, and the distribution ratio of tin and antimony in the flotation tailings is ≥85%; the flotation tailings are used for strong reduction sulfidation smelting to recover tin and antimony, copper and nickel are recovered in the form of matte, and tin and antimony are recovered in the flue dust.

5. The recycling method according to claim 1, characterized in that, The alkaline solution is an aqueous solution of sodium hydroxide with a concentration of 0.75 mol / L to 1.50 mol / L; And / or, the amount of alkaline solution used is 1.3 to 1.5 times the theoretical amount required for dissolving tin, antimony, and arsenic; And / or, the leaching temperature is 80℃-90℃, the leaching time is 30min-120min, and the liquid-to-solid ratio is 3-10L / kg.

6. The recycling method according to claim 1, characterized in that, The oxidant used in the process of descaling antimony by oxidative hydrolysis is selected from hydrogen peroxide, air, or oxygen.

7. The recycling method according to claim 6, characterized in that, When the oxidant is hydrogen peroxide, the amount of hydrogen peroxide used should be controlled to be 2-3 times the theoretical amount, and the reaction time should be 60 min-120 min. And / or, when the oxidant is air or oxygen, the reaction temperature is 60℃-90℃ and the reaction time is 90min-180min.

8. The recycling method according to claim 1, characterized in that, The process of recovering calcium stannate slag includes: mixing lime and the antimony precipitation filtrate, and controlling the molar ratio of effective calcium oxide to tin to be (1.1-1.5):1; Preferably, the lime is added in at least two separate additions, with each addition not exceeding 50% of the total mass. The interval between each addition is 20-30 minutes, and the reaction temperature is controlled at 80℃-90℃. After the last addition, the reaction continues for 60-120 minutes.

9. The recycling method according to claim 1, characterized in that, The process of recovering calcium arsenate slag includes: mixing the liquid to be treated obtained after recovering calcium stannate slag with lime, controlling the molar ratio of effective calcium oxide to arsenic to be (2.0-2.5):1, and controlling the reaction time to be 60min-120min; And / or, after recovering the calcium arsenate residue, perform solid-liquid separation, and return the resulting filtrate to the alkaline leaching stage after adding alkali.

10. The recycling method according to claim 1, characterized in that, The precipitate contains 10%–20% arsenic, 15%–25% tin, 2%–5% nickel, 7%–15% antimony, and 15%–25% copper by mass fraction. And / or, the mass percentage of sulfates, elements, alloy phases and sulfides in the copper and nickel phases of the precipitate is >70%.

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