Method for Recycling Valuable Metals from Smelting Electric Dust Ash

Abstract

The present invention belongs to the field of metallurgical technology, and particularly relates to a method for recovering valuable metals from smelting electric dust ash. The method of the present invention comprises the following steps: S1 adding sulfuric acid, ferric sulfate and hydrogen peroxide to the smelting electric dust ash, and after reaction, performing solid-liquid separation to obtain an acidic leaching solution and an acidic leaching residue; S2 adding sodium hydroxide to the acidic leaching residue obtained in step S1, adjusting its pH value to above 13.5, and after reaction, obtaining an alkaline leaching solution and an alkaline leaching residue; S3 adding sulfuric acid to the alkaline leaching solution obtained in step S2, adjusting its pH value to 10.0-12.5, and after reaction, obtaining a neutralization solution and lead slag; S4 roasting the lead slag obtained in step S3 to obtain lead oxide. In the acid leaching process of the smelting electric dust ash of the present invention, adding ferric sulfate can effectively inhibit the leaching of arsenic, and subsequent changing the acid-base environment of the reaction can effectively realize the efficient separation and recovery of valuable metals.

Description

Technical Field

[0001] The present invention belongs to the technical field of metallurgy, and particularly relates to a method for recovering valuable metals from smelting electric dust ash. Background Art

[0002] To recover valuable elements from copper smelting soot, mainly pyrometallurgical, hydrometallurgical or combined hydrometallurgical-pyrometallurgical processes are adopted. Among them, the pyrometallurgical treatment process mainly returns the soot to smelting. However, impurities such as arsenic and lead in the soot are enriched, which seriously affects the quality of the product and the corrosion of equipment; the combined hydrometallurgical-pyrometallurgical process usually first conducts hydrometallurgical leaching, and the leaching residue is then treated by pyrometallurgy. Secondary soot 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 hydrometallurgical treatment process mainly adopts acidic atmospheric leaching. Most of the copper, zinc and part of the arsenic in the soot are leached. The leaching solution recovers copper, arsenic and zinc respectively. However, due to partial leaching of arsenic, both the leaching solution and the leaching residue contain arsenic, and secondary pollution of arsenic is extremely likely to occur during the treatment of the leaching solution and the leaching residue.

[0003] Chinese Patent Application CN 112359213 A discloses a method for co-disposing copper smelting hazardous waste and comprehensively recovering valuable metals. First, copper smelting soot is subjected to two-stage countercurrent leaching under atmospheric pressure and pressure to obtain an atmospheric leaching solution and a lead-silver-bismuth slag. The atmospheric leaching solution is pre-desilvered by electrowinning to obtain electrolytic copper and a solution after pre-desilvering; then, arsenic sulfide slag is added to the solution after pre-desilvering for replacement copper precipitation to obtain copper sulfide concentrate and a solution after copper precipitation; sulfur dioxide-containing 2 flue gas is introduced into the solution after copper precipitation for reduction arsenic precipitation to obtain arsenic trioxide and a solution after arsenic precipitation. 75% of the solution after arsenic precipitation is returned to the copper smelting soot leaching process, and 25% is discharged to the arsenic removal by sulfidation process, resulting in low arsenic recovery rate, high arsenic content in the slag, and high subsequent recovery cost. Summary of the Invention

[0004] Aiming at the deficiencies of the above-mentioned existing technologies, the present invention provides a method for recovering valuable metals from smelting electric dust ash, which can effectively recover valuable metals in the smelting electric dust ash and has a high arsenic recovery rate.

[0005] The specific technical solution is as follows:

[0006] A method for recovering valuable metals from smelting electric dust ash includes the following steps:

[0007] S1 Sulfuric acid, ferric sulfate and hydrogen peroxide are added to the smelting electric dust ash. After the reaction, solid-liquid separation is carried out to obtain an acidic leaching solution and an acidic leaching residue;

[0008] S2 Sodium hydroxide is added to the acidic leaching residue obtained in step S1 to adjust its pH value to above 13.5. After the reaction, an alkaline leaching solution and an alkaline leaching residue are obtained;

[0009] In S3, sulfuric acid is added to the alkaline leaching solution obtained in step S2 to adjust its pH value to 10.0 - 12.5. After the reaction, a neutralized solution and lead slag are obtained;

[0010] In S4, the lead slag obtained in step S3 is calcined to obtain lead oxide.

[0011] Among them, the smelting electric dust ash can come from: the soot collected from the copper pyrometallurgy smelting section.

[0012] Specifically, the main component contents in the smelting electric dust ash are: Au: 5 g / t - 20 g / t, Ag: 50 g / t - 200 g / t, Cu: 5 wt% - 10 wt%, Pb: 25 wt% - 35 wt%, Zn: 5 wt% - 10 wt%, Fe: 1 wt% - 3 wt%, As: 5 wt% - 10 wt%.

[0013] The reaction mechanism of the present invention is as follows:

[0014] In the present invention, lead, arsenic, gold, and silver in the smelting electric dust ash are retained in the acid leaching residue in the form of precipitates such as lead sulfate and iron arsenate by means of acid leaching, while copper, zinc, and iron are retained in the acid leaching solution in the form of metal salts and can be separated through subsequent steps such as sulfidation and neutralization; in a reaction environment with pH ≥ 13.5, the acid leaching residue converts lead sulfate and iron arsenate into soluble salts sodium plumbite and sodium arsenate, thereby achieving efficient separation from gold and silver; subsequently, by controlling the pH value of the leaching solution to 10.0 - 12.5, sodium plumbite hydrolyzes to form lead hydroxide to achieve the separation and recovery of lead and arsenic.

[0015] In step S1, adding ferric sulfate to the smelting electric dust ash can effectively inhibit the leaching of arsenic. The obtained acid leaching solution mainly contains copper, zinc, and iron, and the acid leaching residue mainly contains valuable metals lead, arsenic, gold, and silver. The main reactions are:

[0016] PbO + H 2 SO 4 →PbSO 4 ↓ + H 2 O;

[0017] ZnO + H 2 SO 4 →ZnSO 4 + H 2 O;

[0018] CuO + H 2 SO 4 →CuSO 4 + H 2 O;

[0019] Fe 2 O 3+3H 2 SO 4 →Fe 2 (SO 4 ) 3 +3H 2 O;

[0020] As 2 O 3 +H 2 O+2H 2 O 2 →2H 3 AsO 4 ;

[0021] 2H 3 AsO 4 +Fe 2 (SO 4 ) 3 →2FeAsO 4 ↓+3H 2 SO 4 。

[0022] Furthermore, in step S1, the sulfuric acid concentration is 100 - 200 g / L, the liquid-solid mass ratio of sulfuric acid to smelting electric dust ash is (2 - 6):1, the molar ratio of hydrogen peroxide to the total amount of arsenic in the smelting electric dust ash is (0.5 - 1.5):1, and the molar ratio of ferric sulfate to the total amount of arsenic in the smelting electric dust ash is (0.8 - 1.2):1.

[0023] Among them, the main component proportion of the acidic leaching residue (dry basis) obtained in step S1 is: the As content is 5 wt% - 15 wt%, and the Pb content is 35 wt% - 50 wt%; the As content in the acidic leaching solution is ≤ 2.0 mg / L.

[0024] Furthermore, in step S1, the reaction conditions are: the reaction temperature is 40 - 80 °C, and the reaction time is 1 - 4 h.

[0025] In step S2, under strong alkaline conditions (pH ≥ 13.5), arsenic-containing substances such as lead sulfate and ferric arsenate in the slag react with sodium hydroxide to form sodium plumbite and sodium arsenate and dissolve in the alkaline leaching solution, while the gold and silver substances in the acidic leaching residue continue to remain in the leaching residue, and the alkaline leaching residue is obtained through solid-liquid separation; preferably, the reaction conditions are: reacting at room temperature for 1 - 4 h. The main reactions are:

[0026] PbSO 4 +4NaOH→Na 2 PbO 2 +Na 2 SO 4 +2H 2 O;

[0027] FeAsO 4 + 3NaOH → Na 3 AsO 4 + Fe(OH) 3 。

[0028] In step S3, the pH value of the alkaline leaching solution is adjusted to 10.0 - 12.5 by adding acid, and sodium plumbite therein undergoes a hydrolysis reaction to obtain lead hydroxide precipitate, i.e., lead slag; preferably, the reaction conditions are: reacting at room temperature for 1 - 4 h. The main reaction is:

[0029] Na 2 PbO 2 + 2H 2 O → Pb(OH) 2 ↓ + 2NaOH.

[0030] 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 h.

[0031] Furthermore, the alkaline leaching residue mainly contains valuable metals such as gold and silver, and gold and silver can be recovered through the precious metal smelting system.

[0032] Furthermore, the neutralization solution in step S3 is returned to the multi-effect evaporation system to recover sodium arsenate.

[0033] Furthermore, the method further includes step S5: the acidic leaching solution obtained in step S1 is successively subjected to sulfide precipitation, neutralization for iron removal, and neutralization precipitation steps to recover copper and zinc.

[0034] Among them, the sulfide precipitation step is specifically: sodium sulfide is added to the acidic leaching solution obtained in step S1, and after the reaction, solid-liquid separation is carried out to obtain copper sulfide and the sulfided solution. The main reaction is:

[0035] CuSO 4 + Na 2 S → CuS↓ + Na 2 SO 4 。

[0036] Preferably, the molar ratio of the total amount of copper in the acidic leaching solution to sodium sulfide is 1:(1 - 1.1).

[0037] Among them, the neutralization for iron removal step is specifically: zinc oxide is added to the sulfided solution to adjust the solution pH to 3.5 - 4.0, and after the reaction, iron slag and the iron-removed solution are obtained. The main reaction is:

[0038] ZnO + H 2 SO 4 → ZnSO 4 + H 2 O;

[0039] Fe 2 (SO 4 ) 3 +3H 2 O+3ZnO→2Fe(OH) 3 ↓+3ZnSO 4 。

[0040] Furthermore, copper sulfide and iron slag are returned to the copper smelting system for copper recovery.

[0041] Among them, the neutralization precipitation step is specifically: adding sodium hydroxide to the solution after iron removal to adjust the solution pH to 7.0 - 8.0, and obtaining zinc slag and neutralized solution after reaction. The main reactions are:

[0042] 2NaOH + H 2 SO 4 →Na 2 SO 4 +2H 2 O;

[0043] ZnSO 4 +2NaOH → Zn(OH) 2 ↓ + Na 2 SO 4 。

[0044] Furthermore, the zinc slag obtained in step S5 is returned to the zinc electrowinning system for zinc recovery, and the neutralized solution is returned to the multiple-effect evaporation system for sodium sulfate recovery.

[0045] The beneficial effects of the present invention are as follows:

[0046] In the acid leaching process of smelting electric dust ash, ferric sulfate is added in the present invention, which can effectively inhibit the leaching of arsenic, and has the advantages of high metal recovery rate and low treatment cost. At the same time, it avoids the high content of arsenic element in the leaching solution, causing 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, with a simple process and a short process flow, which can meet the requirements of resource utilization and harmless treatment of dust, and has remarkable economic and environmental benefits. BRIEF DESCRIPTION OF THE DRAWINGS

[0047] Figure 1 It is a process flow chart for recovering valuable metals from smelting electric dust ash in the specific embodiment. SPECIFIC EMBODIMENT

[0048] The principles and features of the present invention will be described below in conjunction with examples. The examples are only used to explain the present invention and are not intended to limit the scope of the present invention. Unless otherwise specified, the experimental methods used in the following examples are all conventional methods. Unless otherwise specified, the materials, reagents, etc. used in the following examples can all be obtained from commercial sources. Example 1

[0049] A method for recovering valuable metals from smelting electric dust ash, wherein,

[0050] The smelting electric dust ash is from the soot collected in the copper pyrometallurgy smelting section, and the main component contents are: Au: 5.02 g / t, Ag: 50.42 g / t, Cu: 9.89 wt%, Pb: 25.41 wt%, Zn: 5.21 wt%, Fe: 2.87 wt%, As: 9.89 wt%.

[0051] The method steps are as follows:

[0052] S1 Place the smelting electric dust ash in a reaction tank, slowly add a sulfuric acid solution with a concentration of 100 g / L, a hydrogen peroxide solution with a concentration of 30 wt%, and ferric sulfate thereto. The reaction temperature is 40 °C and the reaction time is 4 h. After solid-liquid separation, an acidic leaching solution and an acidic leaching residue are obtained; wherein, the liquid-solid mass ratio of sulfuric acid to the 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.

[0053] S2 At room temperature, place the acidic leaching solution in step S1 in a reaction tank, then add sodium hydroxide to the reaction tank to adjust the pH value to 13.5. After reacting for 1 h, solid-liquid separation is carried out to obtain an alkaline leaching residue and an alkaline leaching solution; the alkaline leaching residue is returned to the precious metal smelting system to recover gold ingots and silver ingots.

[0054] S3 At room temperature, place the alkaline leaching solution obtained in step S2 in a reaction tank, then add sulfuric acid to the reaction tank to adjust the pH value to 10.0. After reacting for 1 h, solid-liquid separation is carried out to obtain a lead slag and a neutralization solution; the neutralization solution is returned to the multi-effect evaporation system to recover sodium arsenate.

[0055] S4 Place the lead slag obtained in step S3 in an electric furnace for roasting to obtain lead oxide. The roasting temperature is 400 °C and the roasting time is 2 h.

[0056] S5 At room temperature, place the acidic leaching solution in step S1 in a reaction tank, then add sodium sulfide into the reaction tank. After reaction, solid-liquid separation is carried out to obtain copper sulfide and a post-sulfidation solution; wherein, the molar ratio of the total amount of copper in the acidic leaching solution to sodium sulfide is 1:1;

[0057] Place the post - vulcanization solution in a reaction tank, then add zinc oxide to the reaction tank to adjust the pH of the solution to 3.5 for neutralizing and removing iron. After the reaction, solid - liquid separation is carried out to obtain iron slag and the iron - removed solution; return the iron slag and copper sulfide to the copper smelting system to recover cathode copper;

[0058] Place the iron - removed solution in a reaction tank, then add sodium hydroxide to the reaction tank to adjust the pH of the solution to 7.0 for neutralizing and precipitating zinc. After the reaction, solid - liquid separation is carried out to obtain zinc slag and the neutralized solution; return the zinc slag to the zinc electrowinning system to recover zinc ingots, and the neutralized solution is recovered to obtain sodium sulfate through a multi - effect evaporation system. Example 2

[0059] A method for recovering valuable metals from smelting electric dust, wherein,

[0060] The smelting electric dust comes from the soot collected in the copper pyrometallurgy smelting section, and the main component contents are: Au: 19.58 g / t, Ag: 199.21 g / t, Cu: 5.41 wt%, Pb: 34.86 wt%, Zn: 5.32 wt%, Fe: 1.26 wt%, As: 5.31 wt%.

[0061] The method steps are as follows:

[0062] S1 Place the smelting electric dust in a reaction tank, slowly add a sulfuric acid solution with a concentration of 150 g / L, a hydrogen peroxide solution with a concentration of 30 wt% and ferric sulfate to it. The reaction temperature is 80 °C and the reaction time is 1 h. After solid - liquid separation, an acidic leaching solution and an acidic leaching residue are obtained; among them, the liquid - solid mass ratio of sulfuric acid to smelting electric dust is 6:1, the molar ratio of hydrogen peroxide to the total amount of arsenic in the smelting electric dust is 1:1, and the molar ratio of ferric sulfate to the total amount of arsenic in the smelting electric dust is 1.2:1.

[0063] S2 At room temperature, place the acidic leaching solution in step S1 in a reaction tank, then add sodium hydroxide to the reaction tank to adjust the pH value to 14. After reacting for 4 h, solid - liquid separation is carried out to obtain an alkaline leaching residue and an alkaline leaching solution; return the alkaline leaching residue to the precious metal smelting system to recover gold ingots and silver ingots.

[0064] S3 At room temperature, place the alkaline leaching solution obtained in step S2 in a reaction tank, then add sulfuric acid to the reaction tank to adjust the pH value to 12.5. After reacting for 4 h, solid - liquid separation is carried out to obtain lead slag and a neutralized solution; return the neutralized solution to the multi - effect evaporation system to recover sodium arsenate.

[0065] S4 Roast the lead slag obtained in step S3 in an electric furnace to obtain lead oxide. The roasting temperature is 300 °C and the roasting time is 4 h.

[0066] At room temperature, place the acidic leaching solution in step S1 into a reaction tank, then add sodium sulfide into the reaction tank. After the reaction, copper sulfide and post-sulfidation solution are obtained through solid-liquid separation; wherein, the molar ratio of the total amount of copper in the acidic leaching solution to sodium sulfide is 1:1.1.

[0067] Place the post-sulfidation solution in a reaction tank, then add zinc oxide into the reaction tank to adjust the pH value of the solution to 4.0 for neutralization and iron removal. After the reaction, iron slag and post-iron-removal solution are obtained through solid-liquid separation; return the iron slag and copper sulfide to the copper smelting system for recycling to obtain cathode copper.

[0068] Place the post-iron-removal solution in a reaction tank, then add sodium hydroxide into the reaction tank to adjust the pH value of the solution to 8.0 for neutralization precipitation and zinc removal. After the reaction, zinc slag and post-neutralization solution are obtained through solid-liquid separation; return the zinc slag to the zinc electrowinning system for recycling to obtain zinc ingots, and the post-neutralization solution is recycled through a multiple-effect evaporation system to obtain sodium sulfate. Example 3

[0069] A method for recovering valuable metals from smelting electric dust, wherein

[0070] The smelting electric dust comes from the soot collected in the copper pyrometallurgy smelting section, and the main component contents are: Au: 10.25 g / t, Ag: 121.24 g / t, Cu: 6.58 wt%, Pb: 31.21 wt%, Zn: 7.82 wt%, Fe: 2.31 wt%, As: 8.64 wt%.

[0071] The method steps are as follows:

[0072] S1 Place the smelting electric dust in a reaction tank, slowly add a sulfuric acid solution with a concentration of 200 g / L, a hydrogen peroxide solution with a concentration of 30 wt% and ferric sulfate into it. The reaction temperature is 60 °C and the reaction time is 3 h. After solid-liquid separation, an acidic leaching solution and acidic leaching residue are obtained; wherein, the liquid-solid mass ratio of sulfuric acid to smelting electric dust is 4:1, the molar ratio of hydrogen peroxide to the total amount of arsenic in the smelting electric dust is 1.5:1, and the molar ratio of ferric sulfate to the total amount of arsenic in the smelting electric dust is 1:1.

[0073] S2 At room temperature, place the acidic leaching solution in step S1 into a reaction tank, then add sodium hydroxide into the reaction tank to adjust the pH value to 13.7. After reacting for 3 h, solid-liquid separation is carried out to obtain an alkaline leaching residue and an alkaline leaching solution; return the alkaline leaching residue to the precious metal smelting system for recycling to obtain gold ingots and silver ingots.

[0074] S3 At room temperature, place the alkaline leaching solution obtained in step S2 into a reaction tank, then add sulfuric acid into the reaction tank to adjust the pH value to 11.0. After reacting for 2 h, solid-liquid separation is carried out to obtain lead slag and a neutralization solution; return the neutralization solution to the multiple-effect evaporation system for recycling sodium arsenate.

[0075] S4 Roast the lead slag obtained in step S3 in an electric furnace to obtain lead oxide. The roasting temperature is 350 °C and the roasting time is 3 h.

[0076] S5 At room temperature, place the acidic leaching solution in step S1 into a reaction tank, and then add sodium sulfide into the reaction tank. After the reaction, carry out solid-liquid separation to obtain copper sulfide and sulfided solution; among them, the molar ratio of the total copper in the acidic leaching solution to sodium sulfide is 1:1.05;

[0077] Place the sulfided solution in a reaction tank, and then add zinc oxide into the reaction tank to adjust the pH of the solution to 3.7 for neutralization and iron removal. After the reaction, carry out solid-liquid separation to obtain iron slag and iron-removed solution; return the iron slag and copper sulfide to the copper smelting system to recover cathode copper;

[0078] Place the iron-removed solution in a reaction tank, and then add sodium hydroxide into the reaction tank to adjust the pH of the solution to 7.5 for neutralization precipitation and zinc removal. After the reaction, carry out solid-liquid separation to obtain zinc slag and neutralized solution; return the zinc slag to the zinc electrowinning system to recover zinc ingots, and recover sodium sulfate from the neutralized solution through a multi-effect evaporation system. Comparative Example 1

[0079] For the specific steps, refer to Example 1, the difference is that: ferric sulfate is not added during the acid leaching process in step S1.

[0080] Test

[0081] Detect the Pb and As contents (wt%) in the acidic leaching residue (dry basis) and the As content (mg / L) in the acidic leaching solution obtained in step S1 of Examples 1 to 3 and Comparative Example 1. The results are shown in Table 1; calculate the recovery rates of each element in Examples 1 to 3 and Comparative Example 1. The results are shown in Table 2.

[0082] For the lead in the acidic leaching residue, it is detected according to the Na 2 EDTA titration method in "Methods for Chemical Analysis of Copper Smelting Fume - Part 2: Determination of Lead Content - Flame Atomic Absorption Spectrometry and Na 2 EDTA Titration Method" (YS / T 1512.2 - 2021); for the arsenic in the acidic leaching residue and acidic leaching solution, it is detected according to "Methods for Chemical Analysis of Copper Smelting Fume - Part 5: Determination of Arsenic Content - Ferrous Ammonium Sulfate Titration Method" (YS / T 1512.5 - 2021).

[0083] Table 1 Pb and As contents in the acidic leaching residue (dry basis) and As content in the acidic leaching solution

[0084] 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

[0085] Table 2 Recovery rates of each element

[0086] 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

[0087] The above are only the preferred embodiments of the present invention and are not intended to limit the present invention. Any modifications, equivalent replacements, improvements, etc. made within the spirit and principles of the present invention shall be included within 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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