Method for recovering valuable metals from smelting electric dust

Abstract

The invention belongs to the technical field of metallurgy, and particularly relates to a method for recovering valuable metals from smelting electric dust. The method comprises the following steps: S1, adding sulfuric acid, ferric sulfate and hydrogen peroxide into the smelting electric dust, and after reaction, carrying out solid-liquid separation to obtain an acidic leaching solution and acidic leaching residues; s2, sodium hydroxide is added into the acid leaching residues obtained in the step S1, the pH value of the acid leaching residues is adjusted to 13.5 or above, and alkaline leaching liquid and alkaline leaching residues are obtained after reaction; s3, sulfuric acid is added into the alkaline leachate obtained in the step S2, the pH value of the alkaline leachate is adjusted to 10.0-12.5, and neutralization liquid and lead slag are obtained after reaction; and S4, roasting the lead slag obtained in the step S3 to obtain lead oxide. According to the method, ferric sulfate is added in the acid leaching process of the smelting electric dust, leaching of arsenic can be effectively inhibited, and efficient separation and recovery of valuable metal can be effectively achieved by subsequently changing the acid-base environment of the reaction.

Description

Technical Field

[0001] The invention belongs to the technical field of metallurgy, and in particular relates to a method for recovering valuable metals by smelting electrostatic dust. Background Art

[0002] The recovery of valuable elements from copper smelting dust mainly adopts pyrometallurgical, hydrometallurgical or hydrometallurgical-pyrometallurgical combined processes. Among them, the pyrometallurgical treatment process mainly returns the dust to smelting, but the impurities such as arsenic and lead in the dust are enriched, which has a serious impact on the quality of the product and equipment corrosion; the hydrometallurgical-pyrometallurgical combined process usually starts with wet leaching, and the leached slag is then treated by pyrometallurgical treatment. Secondary dust is generated during the pyrometallurgical treatment process and needs to be treated again, resulting in low recovery rate of valuable elements and high treatment cost; the wet treatment process mainly adopts acidic atmospheric pressure leaching, most of the copper, zinc and part of the arsenic in the dust are leached, and the leaching solution recovers copper, arsenic and zinc respectively, but due to the partial leaching of arsenic, the leaching solution and the leaching slag contain arsenic, which is very easy to cause secondary arsenic pollution during the treatment of the leaching solution and leaching slag.

[0003] Chinese invention patent application CN 112359213 A discloses a method for the coordinated disposal of hazardous wastes from copper smelting and the comprehensive recovery of valuable metals. The method comprises the following steps: firstly, the copper smelting dust is subjected to two-stage countercurrent leaching at normal pressure and pressure to obtain a normal pressure leaching solution and lead, silver and bismuth slag, and the normal pressure leaching solution is subjected to electrolytic pre-copper removal to obtain electrolytic copper and a pre-copper removal solution; then, arsenic sulfide slag is added to the pre-copper removal solution to replace copper precipitation to obtain a copper sulfide concentrate and a copper precipitation solution; SO2-containing flue gas is introduced into the copper precipitation solution to reduce arsenic precipitation to obtain arsenic trioxide and a arsenic precipitation solution, and 75% of the arsenic precipitation solution is returned to the copper smelting dust leaching process, and 25% is opened to the sulfide arsenic removal process, resulting in a low arsenic recovery rate, a high arsenic content in the slag, and a high subsequent recovery cost. Summary of the invention

[0004] In view of the deficiencies of the above-mentioned prior art, the present invention provides a method for recovering valuable metals from smelting electrostatic ash, which can effectively recover the valuable metals in the smelting electrostatic ash and has a high arsenic recovery rate.

[0005] The specific technical solutions are as follows: The method for recovering valuable metals from smelting electrostatic dust comprises the following steps: S1: adding sulfuric acid, ferric sulfate and hydrogen peroxide to the smelting dust ash, and after the reaction, the solid and liquid are separated to obtain acidic leaching solution and acidic leaching slag; S2: adding sodium hydroxide to the acidic leaching residue obtained in step S1 to adjust the pH value to above 13.5, and obtaining alkaline leaching solution and alkaline leaching residue after reaction; S3: adding sulfuric acid to the alkaline leaching solution obtained in step S2 to adjust the pH value thereof to 10.0-12.5, and obtaining a neutralized solution and lead slag after reaction; S4: roasting the lead slag obtained in step S3 to obtain lead oxide.

[0006] Wherein, the smelting electric dust ash can come from: smoke collected from the copper pyrometallurgical smelting section.

[0007] Specifically, the main component contents of the smelting electric dust ash are: Au: 5g / t~20g / t, Ag: 50g / t~200g / t, Cu: 5wt%~10wt%, Pb: 25wt%~35wt%, Zn: 5wt%~10wt%, Fe: 1wt%~3wt%, As: 5wt%~10wt%.

[0008] The reaction mechanism of the present invention is as follows: The present invention retains lead, arsenic, gold and silver in smelting electric dust ash in the form of lead sulfate, ferric arsenate and other precipitations in acidic leaching residues by acid leaching, while copper, zinc and iron are retained in the form of metal salts in acidic leaching solution and can be separated through subsequent steps such as sulfidation and neutralization; the acidic leaching residue converts lead sulfate and ferric arsenate into soluble salts sodium plumbite and sodium arsenate in a reaction environment with a pH value of ≥13.5, thereby realizing efficient separation of lead sulfate and ferric arsenate from gold and silver; and subsequently controlling the pH value of the leaching solution to be 10.0-12.5 so that the sodium plumbite is hydrolyzed to generate lead hydroxide, thereby realizing separation and recovery of lead and arsenic.

[0009] In step S1, adding ferric sulfate to the smelting dust ash can effectively inhibit the leaching of arsenic, wherein the obtained acidic leaching solution mainly contains copper, zinc and iron, and the acidic leaching slag mainly contains valuable metals such as lead, arsenic, gold and silver. The main reaction is: PbO+H2SO4→PbSO4↓+H2O; ZnO+H2SO4→ZnSO4+H2O; CuO+H2SO4→CuSO4+H2O; Fe2O3+3H2SO4→Fe2(SO4)3+3H2O; As2O3+H2O+2H2O2→2H3AsO4; 2H3AsO4+Fe2(SO4)3→2FeAsO4↓+3H2SO4.

[0010] Furthermore, in step S1, the concentration of sulfuric acid is 100-200 g / L, the liquid-solid mass ratio of sulfuric acid to smelting electrostatic ash is (2-6):1, the molar ratio of hydrogen peroxide to the total amount of arsenic in the smelting electrostatic ash is (0.5-1.5):1, and the molar ratio of iron sulfate to the total amount of arsenic in the smelting electrostatic ash is (0.8-1.2):1.

[0011] Among them, the main components of the acid leaching residue (dry basis) obtained in step S1 are: As content 5wt%~15wt%, Pb content 35wt%~50wt%; As content in the acid leaching solution is ≤2.0mg / L.

[0012] Further, in step S1, the reaction conditions are: reaction temperature is 40-80° C., and reaction time is 1-4 h.

[0013] In step S2, the acidic leaching residue is in a strong alkaline condition (pH ≥ 13.5), and the arsenic-containing substances such as lead sulfate and ferric arsenate in the residue react with sodium hydroxide to generate sodium lead sulfate and sodium arsenate, which are dissolved in the alkaline leaching solution, while the gold and silver-containing substances in the acidic leaching residue continue to remain in the leaching residue and are separated from the solid to obtain the alkaline leaching residue; preferably, the reaction conditions are: reacting at room temperature for 1 to 4 hours. The main reactions are: PbSO4+4NaOH→Na2PbO2+Na2SO4+2H2O; FeAsO4+3NaOH→Na3AsO4+Fe(OH)3. In step S3, the alkaline leaching solution is adjusted to a pH value of 10.0-12.5 by adding acid, and the sodium lead sulfate therein undergoes a hydrolysis reaction to obtain a lead hydroxide precipitate, i.e., lead slag; preferably, the reaction conditions are: reacting at room temperature for 1-4 hours. The main reactions are: Na2PbO2+2H2O→Pb(OH)2↓+2NaOH.

[0014] In step S4, the lead slag is roasted to obtain lead oxide, and the roasting conditions are: the roasting temperature is 300-400° C., and the roasting time is 2-4 hours.

[0015] Furthermore, the alkaline leaching residue mainly contains valuable metals gold and silver, which can be recovered through a precious metal smelting system.

[0016] Furthermore, the neutralized liquid in step S3 is returned to the multiple-effect evaporation system to recover the sodium arsenate.

[0017] Furthermore, the method also includes step S5: recovering copper and zinc by subjecting the acidic leaching solution obtained in step S1 to sulfidation precipitation, neutralization and iron removal, and neutralization precipitation steps in sequence.

[0018] The sulfide precipitation step is specifically as follows: sodium sulfide is added to the acidic leaching solution obtained in step S1, and after the reaction, the solid and liquid are separated to obtain copper sulfide and sulfide liquid. The main reaction is: CuSO4+Na2S→CuS↓+Na2SO4.

[0019] Preferably, the molar ratio of the total amount of copper in the acidic leaching solution to sodium sulfide is 1:(1-1.1).

[0020] The neutralization and iron removal step is specifically as follows: adding zinc oxide to the sulfided liquid to adjust the pH of the solution to 3.5-4.0, and after the reaction, obtaining iron slag and iron-removed liquid. The main reaction is: ZnO+H2SO4→ZnSO4+H2O; Fe2(SO4)3+3H2O+3ZnO→2Fe(OH)3↓+3ZnSO4. Furthermore, the copper sulfide and iron slag are returned to the copper smelting system to recover the copper.

[0021] The neutralization precipitation step is specifically as follows: sodium hydroxide is added to the deironing solution to adjust the pH of the solution to 7.0-8.0, and zinc slag and neutralized solution are obtained after the reaction. The main reaction is: 2NaOH+H2SO4→Na2SO4+2H2O; ZnSO4+2NaOH→Zn(OH)2↓+Na2SO4.

[0022] Furthermore, the zinc slag obtained in step S5 is returned to the zinc electrowinning system to recover zinc, and the neutralized liquid is returned to the multiple-effect evaporation system to recover sodium sulfate.

[0023] The beneficial effects of the present invention are as follows: The present invention adds ferric sulfate during the acid leaching process of smelting electrostatic dust ash, which can effectively inhibit the leaching of arsenic, has the advantages of high metal recovery rate, low processing cost, etc., and at the same time avoids the high content of arsenic in the leaching solution, which causes secondary pollution in the subsequent recycling process; on the other hand, the present invention can effectively separate valuable metals by controlling the acidity and alkalinity of the reaction environment, has a simple process and a short process flow, can meet the needs of smoke resource utilization and harmless treatment, and has significant economic and environmental benefits. BRIEF DESCRIPTION OF THE DRAWINGS

[0024] Figure 1 The present invention is a process flow chart for recovering valuable metals from electrostatic ash in a specific implementation manner. DETAILED DESCRIPTION

[0025] The principles and features of the present invention are described below in conjunction with examples, which are only used to explain the present invention and are not used to limit the scope of the present invention. The experimental methods used in the following examples are conventional methods unless otherwise specified. The materials, reagents, etc. used in the following examples, unless otherwise specified, can all be obtained from commercial sources. Example 1

[0026] A method for recovering valuable metals from smelting electrostatic dust, wherein: The smelting electric dust ash comes from the smoke collected from the copper pyrometallurgical smelting section, and its main components are: Au: 5.02g / t, Ag: 50.42g / t, Cu: 9.89wt%, Pb: 25.41wt%, Zn: 5.21wt%, Fe: 2.87wt%, As: 9.89wt%.

[0027] The method steps are as follows: S1 Place the smelting electric dust ash in a reaction tank, slowly add a sulfuric acid solution with a concentration of 100g / L, a hydrogen peroxide solution with a concentration of 30wt% and ferric sulfate thereto, the reaction temperature is 40°C, the reaction time is 4h, and after solid-liquid separation, an acidic leaching solution and an acidic leaching slag are obtained; wherein, the liquid-solid mass ratio of sulfuric acid to smelting electric dust ash is 2:1, the molar ratio of hydrogen peroxide to the total amount of arsenic in the smelting electric dust ash is 0.5:1, and the molar ratio of ferric sulfate to the total amount of arsenic in the smelting electric dust ash is 0.8:1.

[0028] S2: At room temperature, the acidic leaching solution in step S1 is placed in a reaction tank, and then sodium hydroxide is added to the reaction tank to adjust the pH value to 13.5. After reacting for 1 hour, the solid-liquid separation is performed to obtain alkaline leaching residue and alkaline leaching solution; the alkaline leaching residue is returned to the precious metal smelting system to recover gold and silver ingots.

[0029] S3: at room temperature, placing the alkaline leaching solution obtained in step S2 in a reaction tank, then adding sulfuric acid to the reaction tank to adjust the pH value to 10.0, reacting for 1 hour, performing solid-liquid separation to obtain lead slag and neutralization solution; the neutralization solution is returned to the multiple-effect evaporation system to recover sodium arsenate.

[0030] S4: The lead slag obtained in step S3 is placed in an electric furnace and roasted to obtain lead oxide. The roasting temperature is 400° C. and the roasting time is 2 hours.

[0031] S5: placing the acidic leaching solution in step S1 in a reaction tank at room temperature, then adding sodium sulfide into the reaction tank, and obtaining copper sulfide and sulfided liquid by solid-liquid separation after the reaction; wherein the molar ratio of the total amount of copper in the acidic leaching solution to the sodium sulfide is 1:1; The sulfided liquid is placed in a reaction tank, and then zinc oxide is added to the reaction tank to adjust the pH of the solution to 3.5 for neutralization and iron removal. After the reaction, solid-liquid separation is performed to obtain iron slag and iron-removed liquid; the iron slag and copper sulfide are returned to the copper smelting system for recovery to obtain cathode copper; The iron-removed liquid is placed in a reaction tank, and then sodium hydroxide is added into the reaction tank to adjust the solution pH to 7.0, and zinc is removed by neutralization precipitation. After the reaction, zinc slag and neutralized liquid are obtained through solid-liquid separation; the zinc slag is returned to the zinc electrowinning system to be recovered to obtain zinc ingots, and the neutralized liquid is recovered through a multi-effect evaporation system to obtain sodium sulfate. Example 2

[0032] A method for recovering valuable metals from smelting electrostatic dust, wherein: The smelting electric dust ash comes from the smoke collected from the copper pyrometallurgical smelting section, and its main components are: Au: 19.58g / t, Ag: 199.21g / t, Cu: 5.41wt%, Pb: 34.86wt%, Zn: 5.32wt%, Fe: 1.26wt%, As: 5.31wt%.

[0033] The method steps are as follows: S1 Place the smelting electric dust ash in a reaction tank, slowly add a sulfuric acid solution with a concentration of 150g / L, a hydrogen peroxide solution with a concentration of 30wt% and ferric sulfate thereto, the reaction temperature is 80°C, the reaction time is 1h, and after solid-liquid separation, an acidic leaching solution and an acidic leaching slag are obtained; wherein, the liquid-solid mass ratio of sulfuric acid to smelting electric dust ash is 6:1, the molar ratio of hydrogen peroxide to the total amount of arsenic in the smelting electric dust ash is 1:1, and the molar ratio of ferric sulfate to the total amount of arsenic in the smelting electric dust ash is 1.2:1.

[0034] S2: At room temperature, the acidic leaching solution in step S1 is placed in a reaction tank, and then sodium hydroxide is added to the reaction tank to adjust the pH value to 14. After reacting for 4 hours, the solid-liquid separation is performed to obtain alkaline leaching residue and alkaline leaching solution; the alkaline leaching residue is returned to the precious metal smelting system to recover gold and silver ingots.

[0035] S3: at room temperature, the alkaline leaching solution obtained in step S2 is placed in a reaction tank, and then sulfuric acid is added to the reaction tank to adjust the pH value to 12.5. After reacting for 4 hours, the solid-liquid separation is performed to obtain lead slag and neutralization solution; the neutralization solution is returned to the multiple-effect evaporation system to recover sodium arsenate.

[0036] S4: The lead slag obtained in step S3 is placed in an electric furnace and roasted to obtain lead oxide. The roasting temperature is 300° C. and the roasting time is 4 hours.

[0037] S5. At room temperature, the acidic leaching solution in step S1 is placed in a reaction tank, and then sodium sulfide is added into the reaction tank. After the reaction, copper sulfide and sulfide solution are obtained by solid-liquid separation; wherein the molar ratio of the total amount of copper in the acidic leaching solution to the sodium sulfide is 1:1.1; The sulfided liquid is placed in a reaction tank, and then zinc oxide is added to the reaction tank to adjust the pH of the solution to 4.0 for neutralization and iron removal. After the reaction, iron slag and iron-removed liquid are obtained through solid-liquid separation; the iron slag and copper sulfide are returned to the copper smelting system for recovery to obtain cathode copper; The liquid after iron removal is placed in a reaction tank, and then sodium hydroxide is added into the reaction tank to adjust the pH value of the solution to 8.0, and zinc is removed by neutralization precipitation. After the reaction, zinc slag and neutralized liquid are obtained through solid-liquid separation; the zinc slag is returned to the zinc electrowinning system to be recovered to obtain zinc ingots, and the neutralized liquid is recovered through a multi-effect evaporation system to obtain sodium sulfate. Example 3

[0038] A method for recovering valuable metals from smelting electrostatic dust, wherein: The smelting electric dust ash comes from the smoke collected from the copper pyrometallurgical smelting section, and its main components are: Au: 10.25g / t, Ag: 121.24g / t, Cu: 6.58wt%, Pb: 31.21wt%, Zn: 7.82wt%, Fe: 2.31wt%, As: 8.64wt%.

[0039] The method steps are as follows: S1 Place the smelting electric dust ash in a reaction tank, slowly add a sulfuric acid solution with a concentration of 200g / L, a hydrogen peroxide solution with a concentration of 30wt% and ferric sulfate thereto, the reaction temperature is 60°C, the reaction time is 3h, and acidic leaching solution and acidic leaching slag are obtained after solid-liquid separation; wherein, the liquid-solid mass ratio of sulfuric acid to smelting electric dust ash is 4:1, the molar ratio of hydrogen peroxide to the total amount of arsenic in the smelting electric dust ash is 1.5:1, and the molar ratio of ferric sulfate to the total amount of arsenic in the smelting electric dust ash is 1:1.

[0040] S2: At room temperature, the acidic leaching solution in step S1 is placed in a reaction tank, and then sodium hydroxide is added to the reaction tank to adjust the pH value to 13.7. After reacting for 3 hours, the solid-liquid separation is performed to obtain alkaline leaching residue and alkaline leaching solution; the alkaline leaching residue is returned to the precious metal smelting system to recover gold and silver ingots.

[0041] S3: at room temperature, placing the alkaline leaching solution obtained in step S2 in a reaction tank, then adding sulfuric acid to the reaction tank to adjust the pH value to 11.0, reacting for 2 hours, performing solid-liquid separation to obtain lead slag and neutralization solution; the neutralization solution is returned to the multiple-effect evaporation system to recover sodium arsenate.

[0042] S4 The lead slag obtained in step S3 is placed in an electric furnace and roasted to obtain lead oxide. The roasting temperature is 350° C. and the roasting time is 3 hours.

[0043] S5: placing the acidic leaching solution in step S1 in a reaction tank at room temperature, then adding sodium sulfide into the reaction tank, and obtaining copper sulfide and sulfided liquid by solid-liquid separation after the reaction; wherein the molar ratio of the total amount of copper in the acidic leaching solution to the sodium sulfide is 1:1.05; The sulfided liquid is placed in a reaction tank, and then zinc oxide is added to the reaction tank to adjust the pH of the solution to 3.7 for neutralization and iron removal. After the reaction, solid-liquid separation is performed to obtain iron slag and iron-removed liquid; the iron slag and copper sulfide are returned to the copper smelting system for recovery to obtain cathode copper; The iron-removed liquid is placed in a reaction tank, and then sodium hydroxide is added to the reaction tank to adjust the solution pH to 7.5, and zinc is removed by neutralization precipitation. After the reaction, zinc slag and neutralized liquid are obtained through solid-liquid separation; the zinc slag is returned to the zinc electrowinning system to be recovered to obtain zinc ingots, and the neutralized liquid is recovered through a multi-effect evaporation system to obtain sodium sulfate. Comparative Example 1

[0044] The specific steps are shown in Example 1, except that no iron sulfate is added during the acid leaching process of step S1. test

[0045] The Pb and As contents (wt%) in the acidic leaching residues (dry basis) obtained in step S1 of Examples 1 to 3 and Comparative Example 1 and the As content (mg / L) in the acidic leaching solution were detected, and the results are shown in Table 1; the recovery rates of each element in Examples 1 to 3 and Comparative Example 1 were calculated, and the results are shown in Table 2.

[0046] The lead in acid leaching slag is detected according to the Na2EDTA titration method in "Chemical Analysis Methods for Copper Smelting Dust Part 2: Determination of Lead Content by Flame Atomic Absorption Spectrometry and Na2EDTA Titration Method" (YS / T 1512.2-2021); the arsenic in acid leaching slag and acid leaching liquid is detected according to "Chemical Analysis Methods for Copper Smelting Dust Part 5: Determination of Arsenic Content by Ammonium Ferrous Sulfate Titration Method" (YS / T 1512.5-2021).

[0047] Table 1 Pb and As contents in acid leaching residue (dry basis) and As content in acid leaching solution content Example 1 Example 2 Example 3 Comparative Example 1 Pb (wt%) 35.24 49.65 42.35 42.14 As (wt%) 14.87 5.41 10.21 10.04 As (mg / L) 0.58 0.95 0.35 7802.52

[0048] Table 2 Recovery rate of each element Total recovery rate Example 1 Example 2 Example 3 Comparative Example 1 Pb (%) 99.21 99.23 99.18 99.08 As(%) 99.12 99.01 99.21 67.25 Au(%) 98.24 98.15 98.02 98.04 Ag(%) 97.65 97.26 97.38 97.14 Cu (%) 99.52 99.61 99.58 92.35 Zn(%) 99.21 99.15 99.18 91.25 The above description is only a preferred embodiment of the present invention and is not intended to limit the present invention. Any modifications, equivalent substitutions, improvements, etc. made within the spirit and principle of the present invention should be included in the protection scope of the present invention.

Claims

1. A method for recovering valuable metals from electrolytic dust ash, characterized in that: The steps include: S1: adding sulfuric acid, ferric sulfate and hydrogen peroxide to the smelting dust ash, and after the reaction, the solid and liquid are separated to obtain acidic leaching solution and acidic leaching slag; S2: adding sodium hydroxide to the acidic leaching residue obtained in step S1 to adjust the pH value to above 13.5, and obtaining alkaline leaching solution and alkaline leaching residue after reaction; S3: adding sulfuric acid to the alkaline leaching solution obtained in step S2 to adjust the pH value thereof to 10.0-12.5, and obtaining a neutralized solution and lead slag after reaction; S4: roasting the lead slag obtained in step S3 to obtain lead oxide.

2. The method according to claim 1, characterized in that In step S1, the concentration of sulfuric acid is 100-200 g / L, the liquid-solid mass ratio of sulfuric acid to smelting electrostatic ash is (2-6):1, the molar ratio of hydrogen peroxide to the total amount of arsenic in the smelting electrostatic ash is (0.5-1.5):1, and the molar ratio of ferric sulfate to the total amount of arsenic in the smelting electrostatic ash is (0.8-1.2):

1.

3. The method according to claim 1, characterized in that In step S1, the reaction conditions are: reaction temperature is 40-80° C., and reaction time is 1-4 h.

4. The method according to claim 1, characterized in that: The alkaline leaching residue obtained in step S2 is returned to the precious metal smelting system to recover gold and silver; the neutralized liquid obtained in step S3 is returned to the multiple-effect evaporation system to recover sodium arsenate.

5. The method according to claim 1, characterized in that The method further comprises step S5: subjecting the acidic leaching solution obtained in step S1 to sulfidation precipitation, neutralization and iron removal, and neutralization and precipitation steps in sequence to recover copper and zinc.

6. The method according to claim 5, characterized in that The sulfide precipitation step is: adding sodium sulfide to the acidic leaching solution obtained in step S1, and after the reaction, separating the solid and the liquid to obtain copper sulfide and the sulfide liquid.

7. The method according to claim 6, characterized in that The neutralization and iron removal step is: adding zinc oxide to the sulfided liquid to adjust the pH of the solution to 3.5-4.0, and after the reaction, obtaining iron slag and iron-removed liquid.

8. The method according to claim 7, characterized in that It also includes returning copper sulfide and iron slag to the copper smelting system to recover copper.

9. The method according to claim 7, characterized in that: The neutralization precipitation step is: adding sodium hydroxide to the iron removal solution to adjust the pH of the solution to 7.0-8.0, and obtaining zinc slag and neutralized solution after reaction.

10. The method according to claim 9, characterized in that It also includes returning the zinc slag to the zinc electrowinning system to recover zinc, and returning the neutralized liquid to the multiple-effect evaporation system to recover sodium sulfate.

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