A method for comprehensive recovery of valuable metals from copper soot

Abstract

The application provides a method for comprehensively recovering valuable metals from copper soot, which comprises the following steps: firstly, obtaining arsenic, lead and zinc enrichment liquid and copper, antimony and bismuth enrichment residue through oxidative alkali leaching separation; then, obtaining arsenate and lead-zinc enrichment residue through low-temperature co-precipitation and water washing to remove arsenic from the arsenic, lead and zinc enrichment liquid; finally, obtaining sponge antimony, sponge copper and bismuth oxychloride products through coordination alkali leaching, acid leaching and electrodeposition, and coordination acid leaching, respectively. Compared with the prior art, the method provided by the application not only has simple operation and high metal recovery rate, but also realizes efficient separation and comprehensive recovery of multiple metals, realizes closed circuit of process water, greatly reduces wastewater discharge and reagent consumption, is a more green, economic and efficient resource utilization approach, and has industrial application value.

Description

Technical Field

[0001] This invention belongs to the field of metallurgical separation technology, specifically relating to a method for the comprehensive recovery of valuable metals from copper flue dust. Background Technology

[0002] During copper pyrometallurgical processes, a large amount of copper smelting ash is generated due to the volatilization, oxidation, and airflow carryover of metals at high temperatures. This ash accounts for approximately 2-50 wt% of the raw materials, with large copper smelting enterprises producing thousands or even tens of thousands of tons of copper-containing ash annually. The composition, phases, and properties of copper smelting ash vary depending on the furnace structure, smelting process, and raw materials. Copper ash is not only rich in major valuable metals such as copper, zinc, and lead, but also contains associated rare and dispersed metals such as arsenic, antimony, bismuth, and tin. Efficiently recovering and utilizing these valuable metals has significant strategic and industrial value.

[0003] Currently, the treatment and extraction of valuable metals from copper smelting flue dust are mainly divided into two categories: pyrometallurgical enrichment and hydrometallurgical extraction, but both have limitations to varying degrees. The core of pyrometallurgical enrichment lies in using high-temperature reduction or oxidation to concentrate metals such as lead and bismuth in the crude metal phase, while volatile elements such as zinc and indium enter the secondary flue dust. Although this method has a large processing capacity and a relatively short process, it has significant inherent drawbacks: under high-temperature conditions, highly toxic elements such as arsenic and cadmium are easily volatilized, leading to pollution diffusion and high environmental risks; valuable metals tend to disperse, especially rare metals, resulting in low recovery rates; energy consumption is high and pretreatment is required; the products are mostly crude metals or intermediate alloys with low added value, requiring a long subsequent refining process. In contrast, hydrometallurgical extraction processes have significant advantages in terms of metal recovery rate, environmental friendliness, and mild operating conditions, and have become the mainstream direction of technological development. However, existing hydrometallurgical technologies mostly focus on the selective extraction of single or a few metals, such as using a sulfuric acid system to recover zinc and copper, or separating arsenic and antimony through alkaline leaching.

[0004] While the aforementioned "segmented" or "selective" processes can solve the problem of localized recycling, they are difficult to achieve the synergistic and comprehensive utilization of multiple metals in flue ash throughout the entire process. Some valuable metals are still lost in the intermediate slag phase or waste liquid, thus limiting the overall resource utilization efficiency and economic viability.

[0005] In view of the shortcomings of the existing processes, there is an urgent need in the field to develop a wet process that can comprehensively recover the main valuable metals in copper flue dust, and has a simple process and outstanding environmental benefits. Summary of the Invention

[0006] In view of the current situation where existing copper smelting flue dust treatment processes are generally lengthy, incomplete separation of valuable metals, high consumption of auxiliary materials, complex wastewater treatment, and difficult to achieve efficient and synergistic recovery of multiple metals such as copper, lead, zinc, arsenic, antimony, and bismuth, the purpose of this invention is to provide a method for comprehensively recovering valuable metals from copper flue dust. The method has a simple operation process, can reduce overall costs, and significantly improve the level of comprehensive resource utilization.

[0007] To achieve this objective, the present invention adopts the following technical solution: This invention provides a method for the comprehensive recovery of valuable metals from copper flue dust, the method comprising the following steps: (1) Mix copper ash, first alkaline solution and oxidant for oxidative alkaline leaching to obtain leachate and leaching residue; (2) The leachate obtained in step (1) is subjected to low-temperature precipitation to obtain precipitate residue and precipitate solution; then the precipitate residue is circulated and washed to obtain lead-zinc slag and wash water. When As in the wash water reaches the critical concentration, it is cooled and crystallized to obtain arsenate and crystallized solution. (3) Mix the leaching residue, the second alkaline solution and the precipitant to perform complexation alkaline leaching, and obtain alkaline leaching residue and alkaline leaching solution; (4) Mix the reducing agent and the alkaline leaching solution obtained in step (3) to reduce and precipitate, and obtain filtrate and sponge antimony; The mixed sulfuric acid and the alkaline leaching residue obtained in step (3) were subjected to acid leaching treatment to obtain acid leaching residue and copper-rich leaching solution; (5) Mix sulfuric acid, complexing agent and acid leaching residue obtained in step (4) for complexation acid leaching to obtain acid leaching solution, and then adjust the acid and base to obtain bismuth oxychloride and precipitated solution; The copper-rich leachate obtained in step (4) is subjected to electrowinning to recover cathode copper and waste acid; Steps (2) and (3) are not in any particular order.

[0008] Based on a systematic analysis of the occurrence forms and acid-base chemical properties of various metals in copper flue dust, this invention innovatively proposes an integrated process route of "acid-base leaching synergy and stepwise separation and recovery." By cleverly utilizing the differences in redox behavior between copper, lead, zinc, and arsenic, antimony, and bismuth in acidic and alkaline media, it achieves efficient separation and comprehensive extraction of multiple metals. The process provided by this invention is not only simple to operate and has a high metal recovery rate, but also enables closed-loop recycling of process water, significantly reducing wastewater discharge and reagent consumption. It represents a green, economical, and efficient resource recovery approach. As a preferred technical solution of the present invention, the elements in the copper soot in step (1) include calcium, iron, aluminum, copper, nickel, cobalt, zinc, lead, arsenic, antimony and bismuth.

[0009] Preferably, in step (1), the first alkaline solution comprises a sodium hydroxide solution or a potassium hydroxide solution.

[0010] Preferably, the concentration of the first alkaline solution is 100~400g / L, for example, it can be 100g / L, 200g / L, 300g / L or 400g / L, but is not limited to the listed values. Other values ​​not listed within the range are also applicable.

[0011] Preferably, the oxidant in step (1) includes any one or a combination of at least two of hydrogen peroxide, oxygen, air, ozone or sodium persulfate. Typical but non-limiting combinations include: a combination of hydrogen peroxide and sodium persulfate, a combination of oxygen, air and ozone, or a combination of hydrogen peroxide, oxygen, air, ozone and sodium persulfate.

[0012] Preferably, the mixing amount of the oxidant is 1.0 to 2.0 times the total molar amount of arsenic, antimony and copper in the copper flue ash, for example, it can be 1.0 times, 1.2 times, 1.4 times, 1.6 times, 1.8 times or 2.0 times, etc., but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0013] Preferably, the solid-liquid ratio of the oxidative alkali leaching in step (1) is 1g:(2~10)mL, for example, it can be 1g:2mL, 1g:4mL, 1g:6mL, 1g:8mL or 1g:10mL, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0014] Preferably, the temperature of the oxidative alkali leaching in step (1) is 50~200℃, for example, it can be 50℃, 80℃, 110℃, 140℃, 170℃ or 200℃, etc., but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0015] Preferably, the oxidative alkali leaching in step (1) is accompanied by stirring, and the stirring time is 0.5~5h, for example, it can be 0.5h, 1h, 2h, 3h, 4h or 5h, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0016] It is worth noting that the oxidative leaching described in this invention specifically involves: first mixing copper ash and alkaline solution, heating to the temperature required for oxidative leaching, and then adding an oxidant to carry out the leaching process.

[0017] In addition, the purpose of the oxidative alkaline leaching in step (1) of this invention is: (1) to oxidize and decompose the sulfides (Cu2S, As2S3, Sb2S3, PbS, ZnS) in copper flue dust, and at the same time oxidize As(III) and Sb(III) to As(V) and Sb(V); (2) to dissolve and separate As(V), Pb, and Zn by utilizing the solubility of Pb and Zn in strong alkali. Taking hydrogen peroxide as an oxidant as an example, the reactions included in the oxidative alkaline leaching process are as follows: As a preferred embodiment of the present invention, the precipitant in steps (2) and (3) is independently a sulfur-containing compound.

[0018] Preferably, the sulfur-containing compound includes any one or a combination of at least two of sodium sulfide, ammonium sulfide, hydrogen sulfide, sodium hydrosulfide, ammonium hydrosulfide, thiourea, or thioacetamide. Typical but non-limiting combinations include: a combination of sodium sulfide and sodium hydrosulfide, a combination of hydrogen sulfide and ammonium sulfide, a combination of ammonium hydrosulfide, hydrogen sulfide, and ammonium sulfide, a combination of thiourea and thioacetamide, a combination of sodium sulfide, hydrogen sulfide, and sodium hydrosulfide, or a combination of sodium sulfide, ammonium sulfide, ammonium hydrosulfide, hydrogen sulfide, sodium hydrosulfide, thiourea, and thioacetamide.

[0019] Preferably, the molar ratio of S content in the precipitant to the total amount of Pb and Zn in the leachate is (0.8~1.1):1, for example, it can be 0.8:1, 0.9:1, 1:1 or 1.1:1, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0020] Preferably, the temperature of the low-temperature precipitation in step (2) is 5~25℃, for example, it can be 5℃, 10℃, 15℃, 20℃ or 25℃, etc., but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0021] Preferably, the low-temperature precipitation in step (2) is accompanied by stirring, and the stirring time is 0.5~4h, for example, it can be 0.5h, 1h, 2h, 3h or 4h, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0022] Preferably, the precipitate obtained in step (2) can be reused for the low-temperature precipitation.

[0023] It is worth noting that the chemical reaction that occurs during the low-temperature precipitation process in step (2) of this invention is as follows: ZnO2 2- +S 2- +2H₂O→ZnS+4OH⁻ - ; HPbO2- +S 2- +H₂O→PbS+3OH - ; AsO4 3- +3Na + +12H2O→Na3AsO4∙12H2O.

[0024] As a preferred technical solution of the present invention, the washing in step (2) includes water washing.

[0025] Preferably, the solid-liquid ratio of the water wash is 1g:(1~10)mL, for example, it can be 1g:1mL, 1g:2mL, 1g:4mL, 1g:6mL, 1g:8mL or 1g:10mL, etc., but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0026] Preferably, the water washing temperature is 25~50℃, for example, it can be 25℃, 30℃, 35℃, 40℃, 45℃ or 50℃, etc., but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0027] Preferably, the water washing is accompanied by stirring, and the stirring time is 0.5 to 5 hours, for example, it can be 0.5 hours, 1 hour, 2 hours, 3 hours, 4 hours or 5 hours, but it is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0028] Preferably, the critical concentration in step (2) is 50~150 g / L, for example, it can be 50 g / L, 70 g / L, 90 g / L, 110 g / L, 130 g / L or 150 g / L, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0029] Preferably, the cooling crystallization temperature in step (2) is 5~25℃, for example, it can be 5℃, 10℃, 15℃, 20℃ or 25℃, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0030] Preferably, the cooling crystallization in step (2) is accompanied by stirring, and the stirring time is 0.5~4h, for example, it can be 0.5h, 1h, 2h, 3h or 4h, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0031] Preferably, the crystallized liquid obtained in step (2) can be reused for cyclic washing.

[0032] As a preferred technical solution of the present invention, the second alkaline solution in step (3) includes a sodium hydroxide solution or a potassium hydroxide solution.

[0033] Preferably, the concentration of the second alkaline solution in step (3) is 5~25 g / L, for example, it can be 5 g / L, 10 g / L, 15 g / L, 20 g / L or 25 g / L, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0034] Preferably, the molar ratio of S element in the precipitant to Sb element in the leaching residue in step (3) is (4.0~5.0):1, for example, it can be 4.0:1, 4.2:1, 4.4:1, 4.6:1, 4.8:1 or 5.0:1, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0035] Preferably, the solid-liquid ratio of the complexing alkali leaching in step (3) is 1g:(1~10)mL, for example, it can be 1g:1mL, 1g:2mL, 1g:4mL, 1g:6mL, 1g:8mL or 1g:10mL, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0036] Preferably, the temperature of the complexing alkali leaching in step (3) is 50~100℃, for example, it can be 50℃, 60℃, 70℃, 80℃, 90℃ or 100℃, etc., but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0037] Preferably, the complexing alkali leaching in step (3) is accompanied by stirring, and the stirring time is 0.5~4h, for example, it can be 0.5h, 1h, 2h, 3h or 4h, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0038] It is worth noting that the chemical reaction occurring in the complexation alkali leaching process described in step (3) of this invention is as follows: NaSb(OH)6+4S 2- →SbS4 3– +Na + +6OH - .

[0039] As a preferred technical solution of the present invention, the reducing agent in step (4) includes any one or a combination of at least two of zinc powder, aluminum powder, hydrazine hydrate, glycerol, sodium borohydride, starch or glucose. Typical but non-limiting combinations include: a combination of aluminum powder and zinc powder, a combination of hydrazine hydrate, glycerol and sodium borohydride, a combination of starch and glucose, or a combination of zinc powder, aluminum powder, hydrazine hydrate, glycerol, sodium borohydride, starch and glucose.

[0040] Preferably, the amount of reducing agent used in step (4) is 0.5 to 1.0 times the theoretical molar amount of antimony required in the reducing alkaline leaching solution. For example, it can be 0.5 times, 0.6 times, 0.7 times, 0.8 times, 0.9 times or 1.0 times, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0041] Preferably, the temperature for reducing precipitation in step (4) is 25~200℃, for example, it can be 25℃, 50℃, 80℃, 110℃, 140℃, 170℃ or 200℃, etc., but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0042] Preferably, the reduction precipitation immersion in step (4) is accompanied by stirring, and the stirring time is 0.5~5h, for example, it can be 0.5h, 1h, 2h, 3h, 4h or 5h, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0043] It is worth noting that the purpose of the reduction precipitation in step (4) of this invention is to utilize the reducing agent to oxidize the Sb(V) element in the alkaline leaching solution, thereby achieving the recovery of antimony element from the alkaline leaching solution. Taking aluminum powder as an example: 3SbS4 3- +20OH - +5Al→3Sb+12S 2- +5AlO2 - +10H2O.

[0044] Preferably, the filtrate from step (4) can be reused for complexing alkaline leaching.

[0045] As a preferred technical solution of the present invention, the concentration of sulfuric acid in steps (4) and (5) is 50~200g / L, for example, it can be 50g / L, 80g / L, 110g / L, 140g / L, 170g / L or 200g / L, etc., but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0046] Preferably, the solid-liquid ratio of the acid leaching treatment in step (4) is 1g:(2~10)mL, for example, it can be 1g:2mL, 1g:4mL, 1g:6mL, 1g:8mL or 1g:10mL, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0047] Preferably, the acid leaching temperature in step (4) is 25~100℃, for example, it can be 25℃, 40℃, 60℃, 80℃ or 100℃, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0048] Preferably, the acid leaching process in step (4) is accompanied by stirring, and the stirring time is 0.5 to 5 hours, for example, it can be 0.5 hours, 1 hour, 2 hours, 3 hours, 4 hours or 5 hours, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0049] It is worth noting that the chemical reaction that occurs during the acid leaching process in step (4) of this invention is as follows: CuO + 2H+ + →Cu 2+ +H2O; Fe2O3+6H + →2Fe 3+ +3H2O; CaO + 2H+ + +SO4 2- →CaSO4 + H2O; NiO + 2H + →Ni 2+ +H2O; CoO+2H + →Co 2+ +H2O.

[0050] As a preferred technical solution of the present invention, the complexing agent in step (5) includes any one or a combination of at least two of sodium chloride, potassium chloride, hydrochloric acid, calcium chloride, barium chloride, strontium chloride or lead chloride. Typical but non-limiting combinations include: a combination of sodium chloride and potassium chloride, a combination of calcium chloride, barium chloride, strontium chloride and lead chloride, a combination of sodium chloride, potassium chloride, hydrochloric acid and calcium chloride, or a combination of sodium chloride, potassium chloride, hydrochloric acid, calcium chloride, barium chloride, strontium chloride and lead chloride.

[0051] Preferably, the molar ratio of Cl element in the complexing agent to Bi element in the acid leaching residue in step (5) is (4.0~6.0):1, for example, it can be 4.0:1, 4.4:1, 4.8:1, 5.2:1, 5.6:1 or 6.0:1, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0052] Preferably, the solid-liquid ratio of the complexing acid leaching in step (5) is 1g:(2~10)mL, for example, it can be 1g:2mL, 1g:4mL, 1g:6mL, 1g:8mL or 1g:10mL, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0053] Preferably, the temperature of the complexing acid leaching in step (5) is 25~100℃, for example, it can be 25℃, 40℃, 60℃, 80℃ or 100℃, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0054] Preferably, the complexing acid leaching in step (5) is accompanied by stirring, and the stirring time is 0.5~5h, for example, it can be 0.5h, 1h, 2h, 3h, 4h or 5h, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0055] It is worth noting that the chemical reaction occurring in the complexation acid leaching process described in step (5) of this invention is as follows: Bi₂O₃ + 6H₂ + +8Cl - =2BiCl4 - +3H2O.

[0056] As a preferred technical solution of the present invention, the neutralizing agent used for acid-base adjustment in step (5) includes any one or at least two combinations of sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, ammonia, or sodium hydroxide. Typical but non-limiting combinations include: a combination of sodium carbonate and sodium bicarbonate, a combination of ammonium carbonate and ammonium bicarbonate, a combination of ammonia and sodium hydroxide, a combination of sodium carbonate, ammonium carbonate, sodium bicarbonate, and ammonium bicarbonate, or a combination of sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, ammonia, and sodium hydroxide.

[0057] Preferably, the endpoint pH value of the acid-base adjustment in step (5) is 2 to 3, for example, it can be 2, 2.2, 2.4, 2.6, 2.8 or 3, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0058] Preferably, the temperature for acid-base adjustment in step (5) is 25~50℃, for example, it can be 25℃, 30℃, 35℃, 40℃, 45℃ or 50℃, etc., but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0059] Preferably, the acid-base adjustment in step (5) is accompanied by stirring, and the stirring time is 0.5~5h, for example, it can be 0.5h, 1h, 2h, 3h, 4h or 5h, but is not limited to the listed values. Other unlisted values ​​within the range are also applicable.

[0060] Preferably, the precipitate solution obtained in step (5) can be reused for complexing acid leaching.

[0061] Preferably, the waste acid in step (5) can be reused for the acid leaching treatment in step (4).

[0062] The numerical range described in this invention includes not only the point values ​​listed above, but also any point values ​​within the numerical ranges not listed above. Due to space limitations and for the sake of brevity, this invention will not exhaustively list all the specific point values ​​included in the range.

[0063] Compared with the prior art, the present invention has the following beneficial effects: The method provided by this invention is not only simple to operate and has a high metal recovery rate, but also achieves efficient separation and comprehensive recovery of multiple metals and realizes closed-loop circulation of process water, which greatly reduces wastewater discharge and reagent consumption. It is a green, economical and efficient resource utilization approach with industrial application value. Attached Figure Description

[0064] Figure 1 This is a process flow diagram of the method for comprehensively recovering valuable metals from copper flue dust provided in Embodiment 1 of the present invention. Detailed Implementation

[0065] The technical solution of the present invention will be further illustrated below through specific embodiments. Those skilled in the art should understand that the embodiments described are merely illustrative of the present invention and should not be construed as limiting the invention in any way.

[0066] The main chemical composition of the copper fume treated in the following examples and comparative examples is shown in the table below: Example 1 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust, such as... Figure 1 As shown, the method includes the following steps: (1) Mix 100g of copper ash, sodium hydroxide solution and oxidant for oxidative alkaline leaching to obtain leachate and leaching residue; The concentration of the sodium hydroxide solution is 300 g / L; the oxidant is hydrogen peroxide; the mixing amount of the oxidant is 1.2 times the total molar amount of arsenic, antimony and copper in the copper flue ash; The solid-liquid ratio of the oxidative alkali leaching was 1g:5mL, the temperature was 80℃, and the stirring time was 2h. (2) The leachate obtained in step (1) is subjected to low-temperature precipitation to obtain precipitate residue and precipitate solution; then the precipitate residue is washed with circulating water to obtain lead-zinc slag and wash water. When the As in the wash water reaches the critical concentration (120g / L), it is cooled and crystallized to obtain sodium arsenate with a purity of 94.2% and crystallized solution. The precipitated liquid is returned to step (1) for cyclic leaching; The precipitant is sodium sulfide; the molar ratio of S content in the precipitant to the total amount of Pb and Zn in the leachate is 0.9:1; the low-temperature precipitation temperature is 15℃, and the stirring time is 1h; The solid-liquid ratio of the water washing was 1g:5mL, the temperature was 40℃, and the stirring time was 2h. The cooling crystallization temperature is 10°C, and the stirring time is 1 hour; the crystallized liquid can be reused for circulating washing. (3) Mix the leaching residue, sodium hydroxide solution and precipitant to perform complexation alkali leaching to obtain alkali leaching residue and alkali leaching solution; The concentration of the sodium hydroxide solution is 15 g / L; the precipitant is sodium sulfide; the molar ratio of sulfur (S) in the precipitant to sludge (Sb) is 4.5:1. The solid-liquid ratio of the complexed alkaline leaching was 1g:6mL, the temperature was 60℃, and the stirring time was 3h. (4) Mix the reducing agent and the alkaline leaching solution obtained in step (3) to reduce and precipitate, and obtain filtrate and sponge antimony with a purity of 99.2%; The mixed sulfuric acid and the alkaline leaching residue obtained in step (3) were subjected to acid leaching treatment to obtain acid leaching residue and copper-rich leaching solution; The reducing agent is zinc powder; the amount of the reducing agent used is 1.0 times the theoretical molar amount of antimony required to reduce the alkaline leaching solution; The reduction precipitation temperature is 80℃, and the stirring time is 15h; the filtrate can be reused for complexing alkali leaching. The concentration of the sulfuric acid is 100 g / L; the solid-liquid ratio of the acid leaching treatment is 1 g: 3 mL, the temperature is 80 °C, and the stirring time is 1 h; (5) Mix sulfuric acid, complexing agent and acid leaching residue obtained in step (4) for complexation acid leaching to obtain acid leaching solution, and then adjust the acid and base to obtain bismuth oxychloride and precipitated solution; The copper-rich leachate obtained in step (4) is subjected to electrowinning to recover cathode copper and waste acid; The concentration of the sulfuric acid is 150 g / L; the complexing agent is sodium chloride; the molar ratio of Cl element in the complexing agent to Bi element in the acid leaching residue is 5.0:1; The solid-liquid ratio of the complex acid leaching was 1g:4mL, the temperature was 80℃, and the stirring time was 2h. The neutralizing agent used for the acid-base adjustment is sodium bicarbonate; the endpoint pH value of the acid-base adjustment is 3, the temperature is 40℃, and the stirring time is 1 hour. The precipitated liquid can be reused for complexing acid leaching; the waste acid can be reused for the acid leaching treatment in step (4).

[0067] Example 2 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust, the method comprising the following steps: (1) Mix 100g of copper ash, sodium hydroxide solution and oxidant for oxidative alkaline leaching to obtain leachate and leaching residue; The concentration of the sodium hydroxide solution is 200 g / L; the oxidant is air; the mixing amount of the oxidant is 1.5 times the total molar amount of arsenic, antimony and copper in the copper flue dust; The solid-liquid ratio of the oxidative alkali leaching was 1g:3mL, the temperature was 50℃, and the stirring time was 3h. (2) The leachate obtained in step (1) is subjected to low-temperature precipitation to obtain precipitate residue and precipitate solution; then the precipitate residue is washed with circulating water to obtain lead-zinc slag and wash water. When As in the wash water reaches the critical concentration (100g / L), it is cooled and crystallized to obtain sodium arsenate and crystallized solution. The precipitant is sodium hydrosulfide; the molar ratio of S content in the precipitant to the total amount of Pb and Zn in the leachate is 1.1:1; The low-temperature precipitation was carried out at a temperature of 10°C for 1.5 hours. The solid-liquid ratio of the water washing was 1g:4mL, the temperature was 50℃, and the stirring time was 3h. The cooling crystallization temperature is 10°C, and the stirring time is 2 hours; the crystallized liquid can be reused for circulating washing. (3) Mix the leaching residue, sodium hydroxide solution and precipitant to perform complexation alkali leaching to obtain alkali leaching residue and alkali leaching solution; The concentration of the sodium hydroxide solution is 10 g / L; the precipitant is sodium sulfide; the molar ratio of sulfur (S) in the precipitant to sludge (Sb) is 4.0:1. The solid-liquid ratio of the complexed alkaline leaching was 1g:4mL, the temperature was 70℃, and the stirring time was 1.5h. (4) Mix the reducing agent and the alkaline leaching solution obtained in step (3) to reduce and precipitate, and obtain filtrate and sponge antimony; The mixed sulfuric acid and the alkaline leaching residue obtained in step (3) were subjected to acid leaching treatment to obtain acid leaching residue and copper-rich leaching solution; The reducing agent is aluminum powder; the amount of the reducing agent used is 0.6 times the theoretical molar amount of antimony required to reduce the alkaline leaching solution; The reduction precipitation temperature is 100℃, and the stirring time is 2 hours; the filtrate can be reused for complexing alkali leaching. The concentration of the sulfuric acid is 120 g / L; the solid-liquid ratio of the acid leaching treatment is 1 g: 3 mL, the temperature is 60 °C, and the stirring time is 2 h; (5) Mix sulfuric acid, complexing agent and acid leaching residue obtained in step (4) for complexation acid leaching to obtain acid leaching solution, and then adjust the acid and base to obtain bismuth oxychloride and precipitated solution; The copper-rich leachate obtained in step (4) is subjected to electrowinning to recover cathode copper and waste acid; The concentration of the sulfuric acid is 160 g / L; the complexing agent is calcium chloride; the molar ratio of Cl element in the complexing agent to Bi element in the acid leaching residue is 4.6:1; The solid-liquid ratio of the complexing acid leaching was 1g:4mL, the temperature was 70℃, and the stirring time was 2.5h. The neutralizing agent used for the acid-base adjustment is sodium carbonate; the endpoint pH value of the acid-base adjustment is 2.2, the temperature is 38℃, and the stirring time is 1.2h. The precipitated liquid can be reused for complexing acid leaching; the waste acid can be reused for the acid leaching treatment in step (4).

[0068] Example 3 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust, the method comprising the following steps: (1) Mix 100g of copper ash, sodium hydroxide solution and oxidant for oxidative alkaline leaching to obtain the leached product. The concentration of the sodium hydroxide solution is 400 g / L; the oxidant is ozone; the mixing amount of the oxidant is 1.8 times the total molar amount of arsenic, antimony and copper in the copper flue ash; The solid-liquid ratio of the oxidative alkali leaching was 1g:6mL, the temperature was 120℃, and the stirring time was 1h. (2) The leachate obtained in step (1) is subjected to low-temperature precipitation to obtain precipitate residue and precipitate solution; then the precipitate residue is washed with circulating water to obtain lead-zinc slag and wash water. When the As concentration in the wash water reaches the critical concentration (80g / L), it is cooled and crystallized to obtain arsenate and crystallized solution. The precipitant is thioacetamide; the molar ratio of S content in the precipitant to the total amount of Pb and Zn in the leachate is 0.9:1; The low-temperature precipitation was carried out at a temperature of 8°C for 1 hour. The solid-liquid ratio of the water washing process was 1g:6mL, the temperature was 25℃, and the stirring time was 0.8h. The cooling crystallization temperature is 5°C, and the stirring time is 0.5 h; the crystallized liquid can be reused for circulating washing. (3) Mix the leaching residue, sodium hydroxide solution and precipitant to perform complexation alkali leaching to obtain alkali leaching residue and alkali leaching solution; The concentration of the sodium hydroxide solution is 8 g / L; the precipitant is sodium hydrosulfide; the molar ratio of sulfur (S) in the precipitant to sludge (Sb) is 5.0:1. The solid-liquid ratio of the complexed alkaline leaching was 1g:5mL, the temperature was 50℃, and the stirring time was 3h. (4) Mix the reducing agent and the alkaline leaching solution obtained in step (3) to reduce and precipitate, and obtain filtrate and sponge antimony; The mixed sulfuric acid and the alkaline leaching residue obtained in step (3) were subjected to acid leaching treatment to obtain acid leaching residue and copper-rich leaching solution; The reducing agent is sodium borohydride; the amount of the reducing agent used is 0.5 times the theoretical molar amount of antimony required to reduce the alkaline leaching solution; The reduction precipitation temperature is 60℃, and the stirring time is 1.5h; the filtrate can be reused for complexing alkali leaching; The concentration of the sulfuric acid is 150 g / L; the solid-liquid ratio of the acid leaching treatment is 1 g: 8 mL, the temperature is 50 °C, and the stirring time is 1.2 h; (5) Mix sulfuric acid, complexing agent and acid leaching residue obtained in step (4) for complexation acid leaching to obtain acid leaching solution, and then adjust the acid and base to obtain bismuth oxychloride and precipitated solution; The copper-rich leachate obtained in step (4) is subjected to electrowinning to recover cathode copper and waste acid; The concentration of the sulfuric acid is 120 g / L; the complexing agent is potassium chloride; the molar ratio of Cl element in the complexing agent to Bi element in the acid leaching residue is 5.0:1; The solid-liquid ratio of the complex acid leaching was 1g:4mL, the temperature was 80℃, and the stirring time was 2h. The neutralizing agent used for acid-base adjustment is ammonia; the endpoint pH value of the acid-base adjustment is 2.5, the temperature is 40℃, and the stirring time is 1 hour. The precipitated liquid can be reused for complexing acid leaching; the waste acid can be reused for the acid leaching treatment in step (4).

[0069] Example 4 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust, the method comprising the following steps: (1) Mix 100g of copper ash, sodium hydroxide solution and oxidant for oxidative alkaline leaching to obtain leachate and leaching residue; The concentration of the sodium hydroxide solution is 250 g / L; the oxidant is oxygen; the mixing amount of the oxidant is 1.2 times the total molar amount of arsenic, antimony and copper in the copper flue ash; The solid-liquid ratio of the oxidative alkali leaching was 1g:4mL, the temperature was 100℃, and the stirring time was 2h. (2) The leachate obtained in step (1) is subjected to low-temperature precipitation to obtain precipitate residue and precipitate solution; then the precipitate residue is washed with circulating water to obtain lead-zinc slag and wash water. When As in the wash water reaches the critical concentration (100g / L), it is cooled and crystallized to obtain sodium arsenate and crystallized solution. The precipitant is hydrogen sulfide; the molar ratio of S content in the precipitant to the total amount of Pb and Zn in the leachate is 0.8:1; the low-temperature precipitation temperature is 15℃, and the stirring time is 2h; The solid-liquid ratio of the water washing was 1g:4mL, the temperature was 40℃, and the stirring time was 25h. The cooling crystallization temperature is 10°C, and the stirring time is 1.5 hours; the crystallized liquid can be reused for circulating washing. (3) Mix the leaching residue, sodium hydroxide solution and precipitant to perform complexation alkali leaching to obtain alkali leaching residue and alkali leaching solution; The concentration of the sodium hydroxide solution is 20 g / L; the precipitant is sodium sulfide; the molar ratio of sulfur (S) in the precipitant to sludge (Sb) is 4.5:1. The solid-liquid ratio of the complexed alkaline leaching was 1g:4mL, the temperature was 60℃, and the stirring time was 2.5h. (4) Mix the reducing agent and the alkaline leaching solution obtained in step (3) to reduce and precipitate, and obtain filtrate and sponge antimony; The mixed sulfuric acid and the alkaline leaching residue obtained in step (3) were subjected to acid leaching treatment to obtain acid leaching residue and copper-rich leaching solution; The reducing agent is glucose; the amount of the reducing agent used is 0.8 times the theoretical molar amount of antimony required to reduce the alkaline leaching solution; The reduction precipitation temperature is 150℃, and the stirring time is 1.5h; the filtrate can be reused for complexing alkali leaching. The concentration of the sulfuric acid is 120 g / L; the solid-liquid ratio of the acid leaching treatment is 1 g: 5 mL, the temperature is 60 °C, and the stirring time is 2 h; (5) Mix sulfuric acid, complexing agent and acid leaching residue obtained in step (4) for complexation acid leaching to obtain acid leaching solution, and then adjust the acid and base to obtain bismuth oxychloride and precipitated solution; The copper-rich leachate obtained in step (4) is subjected to electrowinning to recover cathode copper and waste acid; The concentration of the sulfuric acid is 150 g / L; the complexing agent is hydrochloric acid; the molar ratio of Cl element in the complexing agent to Bi element in the acid leaching residue is 4.0:1; The solid-liquid ratio of the complex acid leaching was 1g:5mL, the temperature was 80℃, and the stirring time was 2h. The neutralizing agent used for acid-base adjustment is sodium hydroxide; the endpoint pH value of the acid-base adjustment is 2.0, the temperature is 40℃, and the stirring time is 1 hour. The precipitated liquid can be reused for complexing acid leaching; the waste acid can be reused for the acid leaching treatment in step (4).

[0070] Example 5 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust. The only difference between this method and Embodiment 1 is that: In this embodiment, the temperature of the oxidative alkali leaching in step (1) is adjusted to 30°C.

[0071] Example 6 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust. The only difference between this method and Embodiment 1 is that: In this embodiment, the molar ratio of S content in the precipitant to the total amount of Pb and Zn in the leachate in step (2) is adjusted to 0.5:1.

[0072] Example 7 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust. The only difference between this method and Embodiment 1 is that: In this embodiment, the temperature of the low-temperature precipitation in step (2) is adjusted to 30°C.

[0073] Example 8 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust. The only difference between this method and Embodiment 1 is that: In this embodiment, the concentration of the sodium hydroxide solution used in the complexation alkali leaching in step (3) is adjusted to 100 g / L.

[0074] Example 9 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust. The only difference between this method and Embodiment 1 is that: In this embodiment, the temperature of the complexation alkali immersion in step (3) is adjusted to 30°C.

[0075] Example 10 This embodiment provides a method for the comprehensive recovery of valuable metals from copper flue dust. The only difference between this method and Embodiment 1 is that: In this embodiment, the temperature of the complexation alkali leaching in step (3) is adjusted to 120°C.

[0076] Comparative Example 1 This comparative example provides a method for the comprehensive recovery of valuable metals from copper flue dust. The only difference between this method and Example 1 is that: In this comparative example, the subsequent treatment process of the leachate obtained in step (1) is adjusted as follows: the leachate is cooled and crystallized to obtain mother liquor and sodium arsenate crystals; then the mother liquor and sodium sulfide are mixed and subjected to precipitation reaction to obtain purified liquid and precipitate residue. The precipitation reaction was carried out at a temperature of 50°C.

[0077] Comparative Example 2 This comparative example provides a method for the comprehensive recovery of valuable metals from copper flue dust. The only difference between this method and Example 1 is that: In this comparative example, the complexation alkali leaching in step (3) is changed to acid leaching; the leaching solution used for acid leaching is sulfuric acid.

[0078] Application examples The purity / content of the products obtained by the recovery methods provided in the above embodiments and comparative examples were tested, and the results are shown in Table 1: Table 1 According to Table 1, the following points can be observed: (1) Comprehensive analysis of Examples 1-4 shows that the method provided by the present invention can simultaneously recover high-purity lead-zinc slag, sodium arsenate, sponge antimony, bismuth oxychloride and copper sulfate solution by treating copper flue dust. Moreover, the metal recovery efficiency is high and the closed-loop circulation of process water greatly reduces wastewater discharge and reagent consumption, which has industrial application value. (2) Comprehensive analysis of Examples 1 and 5 shows that if the temperature used for the oxidative leaching is too low, the copper conversion will be incomplete, and some copper sulfides will be difficult to leach out during the acid leaching process in step (4), while some copper will leach out during the complexation leaching process of bismuth in step (5), affecting the purity of the bismuth product; in addition, the oxidation of Sb(III) and As(III) is incomplete, reducing their recovery rate. A comprehensive analysis of Examples 1 and 6 shows that if the amount of precipitant used in the low-temperature precipitation process described in step (2) is too low, it will lead to incomplete precipitation of Pb and Zn, excessively high residual concentration of alkali solution, and reduced leaching rate of Pb and Zn during the circulation process. A comprehensive analysis of Examples 1 and 7 shows that if the temperature of the low-temperature precipitation is too high, the arsenic crystallization rate will be too low, but it will not affect the purity of the product. A comprehensive analysis of Examples 1 and 8 shows that if the concentration of the sodium hydroxide solution used in the complexation leaching is too high, it will lead to incomplete Sb leaching. In addition, the complexation leaching and bismuth leaching will occur simultaneously in step (5), and the hydrolysis precipitation of bismuth will also cause precipitation, affecting the purity of BiOCl. A comprehensive analysis of Examples 1 and 9-10 shows that if the temperature of the complexing alkaline leaching is too low, the antimony leaching rate will be too low, and the unleached antimony will be partially leached out during the acid leaching of copper in step (4) and the complexing acid leaching of bismuth in step (5), and will further affect the copper electrowinning and bismuth deposition process as an impurity. If the temperature is too high, the antimony leaching rate will not change much, and further increasing the temperature will only increase the energy cost. (3) Compared with Example 1, the method provided in Comparative Example 1 first cools and crystallizes the leaching solution of oxidized alkaline leaching, and then carries out a precipitation reaction (which is exactly the opposite of the recovery order provided in Example 1). This will result in a lower arsenic concentration in the initial leaching solution, and direct cooling and crystallization may result in poor recovery effect. Compared to Example 1, Comparative Example 2 directly acid-leaching the leaching residue obtained from oxidative alkaline leaching will make it difficult for antimony to leach out, and it will enter the residue along with bismuth. In step (5), complexation leaching and bismuth leaching are carried out simultaneously, and bismuth precipitation occurs during the hydrolysis precipitation process, which affects the purity of BiOCl.

[0079] In summary, the method provided by this invention is not only simple to operate and has a high metal recovery rate, but also achieves efficient separation and comprehensive recovery of multiple metals, and realizes closed-loop circulation of process water, which greatly reduces wastewater discharge and reagent consumption. It is a green, economical, and efficient resource utilization approach with industrial application value.

[0080] The applicant declares that the above description is only a specific embodiment of the present invention, but the protection scope of the present invention is not limited thereto. Those skilled in the art should understand that any changes or substitutions that can be easily conceived by those skilled in the art within the technical scope disclosed in the present invention fall within the protection and disclosure scope of the present invention.

Claims

1. A method for comprehensively recovering valuable metals from copper flue dust, characterized in that, The method includes the following steps: (1) Mix copper ash, first alkaline solution and oxidant for oxidative alkaline leaching to obtain leachate and leaching residue; (2) The leachate obtained in step (1) is subjected to low-temperature precipitation to obtain precipitate residue and precipitate solution; then the precipitate residue is circulated and washed to obtain lead-zinc slag and wash water. When As in the wash water reaches the critical concentration, it is cooled and crystallized to obtain arsenate and crystallized solution. (3) Mix the leaching residue, the second alkaline solution and the precipitant to perform complexation alkaline leaching, and obtain alkaline leaching residue and alkaline leaching solution; (4) Mix the reducing agent and the alkaline leaching solution obtained in step (3) to reduce and precipitate, and obtain filtrate and sponge antimony; The mixed sulfuric acid and the alkaline leaching residue obtained in step (3) were subjected to acid leaching treatment to obtain acid leaching residue and copper-rich leaching solution; (5) Mix sulfuric acid, complexing agent and acid leaching residue obtained in step (4) for complexation acid leaching to obtain acid leaching solution, and then adjust the acid and base to obtain bismuth oxychloride and precipitated solution; The copper-rich leachate obtained in step (4) is subjected to electrowinning to recover cathode copper and waste acid; Steps (2) and (3) are not in any particular order.

2. The method for comprehensively recovering valuable metals from copper flue dust according to claim 1, characterized in that, The elements in the copper soot in step (1) include calcium, iron, aluminum, copper, nickel, cobalt, zinc, lead, arsenic, antimony and bismuth; Preferably, in step (1), the first alkaline solution comprises a sodium hydroxide solution or a potassium hydroxide solution; Preferably, the concentration of the first alkaline solution is 100~400g / L; Preferably, the oxidant in step (1) includes any one or a combination of at least two of hydrogen peroxide, oxygen, air, ozone or sodium persulfate; Preferably, the mixing amount of the oxidant is 1.0 to 2.0 times the total molar amount of arsenic, antimony and copper in the copper flue ash.

3. The method for comprehensively recovering valuable metals from copper flue dust according to claim 1 or 2, characterized in that, The solid-liquid ratio of the oxidative alkali leaching in step (1) is 1g:(2~10)mL; Preferably, the temperature of the oxidative alkali leaching in step (1) is 50~200℃; Preferably, the oxidative alkali leaching in step (1) is accompanied by stirring, and the stirring time is 0.5~5h.

4. The method for comprehensively recovering valuable metals from copper flue dust according to any one of claims 1-3, characterized in that, The precipitants mentioned in steps (2) and (3) are each independently sulfur-containing compounds; Preferably, the sulfur-containing compound includes any one or a combination of at least two of sodium sulfide, ammonium sulfide, hydrogen sulfide, sodium hydrosulfide, ammonium hydrosulfide, thiourea, or thioacetamide; Preferably, the molar ratio of S content in the precipitant to the total amount of Pb and Zn in the leachate is (0.8~1.1):1; Preferably, the temperature of the low-temperature precipitation in step (2) is 5~25℃; Preferably, the low-temperature precipitation in step (2) is accompanied by stirring, and the stirring time is 0.5~4h; Preferably, the precipitate obtained in step (2) can be reused for the low-temperature precipitation.

5. The method for comprehensively recovering valuable metals from copper flue dust according to any one of claims 1-4, characterized in that, The washing described in step (2) includes washing with water; Preferably, the solid-liquid ratio of the water washing is 1g:(1~10)mL; Preferably, the temperature of the water wash is 25~50℃; Preferably, the water washing is accompanied by stirring, and the stirring time is 0.5~5 hours; Preferably, the critical concentration in step (2) is 50~150 g / L; Preferably, the cooling crystallization temperature in step (2) is 5~25℃; Preferably, the cooling crystallization in step (2) is accompanied by stirring, and the stirring time is 0.5~4h; Preferably, the crystallized liquid obtained in step (2) can be reused for cyclic washing.

6. The method for comprehensively recovering valuable metals from copper flue dust according to any one of claims 1-5, characterized in that, Step (3) The second alkaline solution includes; Preferably, in step (3), the concentration of the second alkaline solution is 5~25 g / L; Preferably, the molar ratio of S element in the precipitant to Sb element in the leaching residue in step (3) is (4.0~5.0):1; Preferably, the solid-liquid ratio of the complexing alkali leaching in step (3) is 1g:(1~10)mL; Preferably, the temperature of the complexing alkali leaching in step (3) is 50~100℃; Preferably, the complexing alkali immersion in step (3) is accompanied by stirring, and the stirring time is 0.5~4h.

7. The method for comprehensively recovering valuable metals from copper flue dust according to any one of claims 1-6, characterized in that, The reducing agent in step (4) includes any one or a combination of at least two of the following: zinc powder, aluminum powder, hydrazine hydrate, glycerol, sodium borohydride, starch, or glucose; Preferably, the amount of reducing agent used in step (4) is 0.5 to 1.0 times the theoretical molar amount of antimony required to reduce the alkaline leaching solution; Preferably, the temperature for reducing the precipitate in step (4) is 25~200℃; Preferably, the reduction precipitate soaking in step (4) is accompanied by stirring, and the stirring time is 0.5~5h; Preferably, the filtrate from step (4) can be reused for complexing alkaline leaching.

8. The method for comprehensively recovering valuable metals from copper flue dust according to any one of claims 1-7, characterized in that, The concentration of sulfuric acid in steps (4) and (5) is 50~200 g / L; Preferably, the solid-liquid ratio of the acid leaching treatment in step (4) is 1g:(2~10)mL; Preferably, the acid leaching temperature in step (4) is 25~100℃; Preferably, the acid leaching process in step (4) is accompanied by stirring, and the stirring time is 0.5~5h.

9. The method for comprehensively recovering valuable metals from copper flue dust according to any one of claims 1-8, characterized in that, The complexing agent in step (5) includes any one or a combination of at least two of sodium chloride, hydrochloric acid, potassium chloride, calcium chloride, barium chloride, strontium chloride, or lead chloride; Preferably, the molar ratio of Cl element in the complexing agent to Bi element in the acid leaching residue in step (5) is (4.0~6.0):1; Preferably, the solid-liquid ratio of the complexing acid leaching in step (5) is 1g:(2~10)mL; Preferably, the temperature of the complexing acid leaching in step (5) is 25~100℃; Preferably, the complexing acid leaching in step (5) is accompanied by stirring, and the stirring time is 0.5~5h.

10. The method for comprehensively recovering valuable metals from copper flue dust according to any one of claims 1-9, characterized in that, The neutralizing agent used for acid-base adjustment in step (5) includes any one or a combination of at least two of sodium carbonate, ammonium carbonate, sodium bicarbonate, ammonium bicarbonate, ammonia, or sodium hydroxide. Preferably, the endpoint pH value of the acid-base adjustment in step (5) is 2~3; Preferably, the temperature for acid-base adjustment in step (5) is 25~50℃. Preferably, the acid-base adjustment in step (5) is accompanied by stirring, and the stirring time is 0.5~5h; Preferably, the precipitate solution obtained in step (5) can be reused for complexing acid leaching; Preferably, the waste acid in step (5) can be reused for the acid leaching treatment in step (4).

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