Method for recovering valuable metals from copper anode slime

Abstract

The invention belongs to the technical field of metallurgy, and relates to a method for recovering valuable metals from copper anode slime, which comprises the following steps: S1, adding sulfuric acid and copper powder into copper anode slime sulfating roasting slag, adjusting the pH value to be less than 1.0, and after reaction, carrying out solid-liquid separation to obtain acid leaching liquid and acid leaching slag; s2, a sodium hydroxide solution is added into the acid leaching residues obtained in the step S1, the pH value of the acid leaching residues is adjusted to 13.0 or above, and alkaline leaching liquid and alkaline leaching residues are obtained after reaction; s3, the alkaline leaching residues obtained in the step S2 are subjected to acid pickling, and washing liquid and washing residues are obtained; s4, hydrochloric acid, hydrogen peroxide and ferric chloride are added into the washing residues obtained in the step S3, the pH value is adjusted to be 0.5 or below, and leaching liquid and leaching residues are obtained after reaction; and S5, the leaching residues obtained in the step S4 are sequentially returned to a silver recovery system, a gold recovery system and a platinum refining system to recover silver, gold and platinum. Valuable metals in the copper anode slime can be recovered, and the recovery rate is high.

Description

Technical Field

[0001] The invention belongs to the technical field of metallurgy, and in particular relates to a method for recovering valuable metals from copper anode mud. Background Art

[0002] In the process of electrolytic refining of copper, the copper anode plate serves as the anode and the stainless steel electrode plate serves as the cathode. When current passes through this system, the copper and other metal impurities on the anode that are more active than copper will release electrons, convert into cations and dissolve in the electrolyte. Metal cations that are more difficult to reduce than copper ions (such as zinc, iron, nickel, etc.) will not precipitate on the cathode, but continue to exist in the electrolyte; at the same time, those impurities with lower metal activity than copper, such as silver and gold, are not easy to dissolve on the anode because of their relatively weak ability to lose electrons. Therefore, when the copper and other metals in the anode lose electrons and dissolve, these inactive metal impurities are deposited in solid form at the bottom of the electrolytic cell, forming the so-called anode mud. Copper anode mud is mainly composed of metals and compounds in the anode crude metal that are insoluble in the electrolyte. Its specific composition is significantly affected by the anode composition and electrolysis technical conditions. In the copper anode mud, precious metals such as gold, silver, selenium, tellurium, platinum, palladium, etc. are contained. The presence of these metals gives the copper anode mud a very high recovery value.

[0003] The treatment technologies widely used in copper anode mud at home and abroad are mainly divided into three categories: pyrometallurgical process, hydrometallurgical process and combined metallurgical process. In the traditional pyrometallurgical process, the direct recovery rate of gold and silver is about 97%. In contrast, the hydrometallurgical process can increase the recovery rate to more than 99%; in the pyrometallurgical process, precious metals are often dispersed in various intermediate products, such as primary slag, secondary slag, copper-bismuth slag, tellurium slag, etc., which makes the recovery process complicated and inefficient. In particular, platinum group metals such as platinum and palladium are more dispersed in the pyrometallurgical process, and the direct recovery rate is much lower than that of gold and silver. They need to go through multiple steps of sorting and extraction, which will undoubtedly further increase the processing cost. In addition, the pyrometallurgical process is also accompanied by serious environmental pollution, high energy consumption, and long production cycle.

[0004] The chemical composition of copper anode mud is complex and diverse. The wet processing technology puts forward specific requirements for the stability of the chemical composition of the raw materials in the process of extracting and separating valuable metals. If the chemical composition of the raw materials changes significantly, it will interfere with the normal operation of the processing process, resulting in a significant decrease in the metal recovery efficiency, and it is difficult to achieve the expected good recovery effect. At present, the wet process generally uses the method of gold reduction and then liquid precipitation of platinum and palladium for separation. The common techniques for platinum and palladium separation include precipitation and extraction. The precipitation method mainly achieves separation by adding specific chemical reagents to precipitate platinum or palladium. The extraction method relies on the selective extraction ability of the extractant for platinum and palladium to separate platinum and palladium from the solution. However, these methods still face some challenges in practical applications, such as low separation efficiency, long process, low recovery rate and environmental pollution. Summary of the invention

[0005] In view of the deficiencies in the above-mentioned prior art, the present invention provides a method for recovering valuable metals from copper anode mud, which can effectively recover the valuable metals in the copper anode mud with a high recovery rate and a simple recovery process.

[0006] The specific technical solutions are as follows: The method for recovering valuable metals from copper anode mud comprises the following steps: S1: adding sulfuric acid and copper powder to the copper anode mud sulfate roasting slag, adjusting the pH value to below 1.0, and after the reaction, separating the solid and liquid to obtain acidic leaching solution and acidic leaching slag; S2: adding sodium hydroxide solution to the acidic leached residue obtained in step S1 to adjust the pH value to above 13.0, and obtaining alkaline leaching solution and alkaline leaching residue after reaction; S3: acid-washing the alkaline leaching residue obtained in step S2 to obtain a washing liquid and washing residue; S4: adding hydrochloric acid, hydrogen peroxide and ferric chloride to the washing residue obtained in step S3, adjusting the pH value to below 0.5, and obtaining a leaching solution and leaching residue after reaction; S5 The leaching residue obtained in step S4 is returned to the silver recovery system, the gold recovery system and the platinum refining system in sequence to recover silver, gold and platinum.

[0007] Among them, the main components of copper anode mud are: Au content ≤1.0wt%, Ag content ≤30wt%, Cu content ≤30wt%, Te content ≤10wt%, Se content ≤10wt%, Pt content ≤50g / t, and Pd content ≤500g / t.

[0008] The reaction mechanism of the present invention is as follows: The invention adds sulfuric acid and copper powder to the copper anode mud sulfate roasting slag, and copper reacts with tellurium oxide therein to generate copper telluride precipitation which is retained in the acid leaching slag. Subsequently, the copper telluride is converted into sodium tellurite and separated from the alkaline leaching slag containing elements such as copper, silver, gold, palladium and platinum through alkaline leaching. During the acid washing process of the alkaline leaching slag, the copper-containing substances therein react with the acid solution and are transferred to the washing liquid, and the washing slag selectively leaches palladium in the presence of hydrochloric acid, hydrogen peroxide and ferric chloride, and the gold, silver and platinum in the leaching slag are sequentially returned to the recovery and refining system to be separated and recovered.

[0009] In step S1, copper anode mud is sulphated and roasted with concentrated sulfuric acid at 800-850°C to obtain copper anode mud sulphated roasted slag and flue gas. The flue gas is returned to the selenium recovery system to recover selenium dioxide, and the sulphated roasted slag is used to recover valuable metals. The main reactions are: Cu2Se+6H2SO4=2CuSO4+SeO2↑+4SO2↑+6H2O; Ag2Se+4H2SO4=Ag2SO4+SeO2↑+3SO2↑+4H2O; Cu2Te+6H2SO4=2CuSO4+TeO2+4SO2↑+6H2O; Ag2Te+4H2SO4=Ag2SO4+TeO2+3SO2↑+4H2O; Se+2H2SO4=SeO2↑+2SO2↑+2H2O; Cu+2H2SO4=CuSO4+SO2↑+2H2O.

[0010] Preferably, the solid-liquid mass ratio of copper anode mud to concentrated sulfuric acid is (1-4):1, and the roasting time is 3-6 hours.

[0011] In step S1, sulfuric acid and copper powder are added to the copper anode mud sulfate roasting slag, and the copper reacts with the tellurium oxide in the roasting slag to form copper telluride precipitation which is retained in the acidic leaching slag, thereby avoiding the leaching of tellurium under acidic conditions. The main reactions are: TeO2+4Cu+2H2SO4=Cu2Te+2CuSO4+2H2O.

[0012] Preferably, the liquid-solid mass ratio of sulfuric acid to copper anode mud sulfated roasted slag is (3-6):1, and the molar ratio of copper powder to the total amount of tellurium in the copper anode mud sulfated roasted slag is (1.5-2.0):1.

[0013] Preferably, in step S1, the reaction temperature is 60-80° C., and the reaction time is 2-4 h.

[0014] In step S2, copper telluride in the acidic leaching residue reacts with sodium hydroxide to generate sodium tellurite which is dissolved in the alkaline leaching solution. The main reaction is: 2Cu2Te+4NaOH+4O2=4CuO+2Na2TeO3+2H2O.

[0015] Preferably, the liquid-solid mass ratio of the sodium hydroxide solution to the acidic leaching residue is (3-6):1.

[0016] Preferably, in step S2, the reaction temperature is 60-80° C., and the reaction time is 2-6 h.

[0017] Further, the alkaline leaching solution obtained in step S2 is returned to the tellurium electrowinning system to recover tellurium, and the electrowinning current density is 120~200A / m 2 .

[0018] In step S3, preferably, the alkaline leached residue is exposed to air and pickled with sulfuric acid (concentration of 1.5-10.0 wt%) at 60-90° C. for 3-6 hours to remove copper oxide and copper element, the main reaction being: CuO+H2SO4=CuSO4+H2O; 2Cu+O2+2H2SO4=2CuSO4+2H2O.

[0019] Furthermore, the acidic leaching solution obtained in step S1 and the washing solution obtained in step S3 are returned to the copper recovery system to recover copper.

[0020] In step S4, hydrochloric acid, hydrogen peroxide and ferric chloride are added to the washing residue obtained in step S3 to selectively leach palladium. Ferric chloride can be used as an inhibitor to inhibit the leaching of platinum, while gold is relatively stable and will not be leached. The main reaction is: Pd + 4HCl + H2O2 = H2PdCl4 + 2H2O; 2Ag+2HCl+H2O2=2AgCl+2H2O.

[0021] Without the addition of ferric chloride, a large amount of Pt will dissolve in the leaching solution, and the following reaction will occur: Pt+6HCl+4H2O2=H2PtCl6+O2↑+6H2O.

[0022] Preferably, the mass ratio of hydrogen peroxide to washing residue is (0.03-0.10):1, the mass ratio of ferric chloride to washing residue is (0.01-0.05):1, and the liquid-solid mass ratio of hydrochloric acid to washing residue is (3-6):1.

[0023] Preferably, in step S4, the reaction temperature is 5-25° C., and the reaction time is 2-4 h.

[0024] Preferably, the concentration of hydrogen peroxide is 10-30wt%.

[0025] Furthermore, the leaching solution obtained in step S4 is recovered to obtain palladium through a replacement reaction.

[0026] In step S5, since the leached residue contains valuable metal elements such as silver, gold and platinum, it can be returned to the silver recovery system to recover silver, and black gold powder can be obtained at the same time. It can be sent to the gold recovery system to obtain gold and platinum-rich liquid, and the platinum-rich liquid can be returned to the platinum refining system to recover platinum.

[0027] The beneficial effects of the present invention are as follows: The present invention adds sulfuric acid and copper powder to the sulfated roasted slag of copper anode mud to prevent tellurium from leaching under acidic conditions, and has a high tellurium recovery rate. Tellurium ingots with a grade of more than 99.90% can be recovered, and the copper recovery process in the acidic leaching solution is simplified, and copper can be recovered without impurity removal. The present invention utilizes hydrochloric acid, hydrogen peroxide and ferric chloride to selectively leach palladium, avoiding the leaching of platinum, making the subsequent recovery and separation process simple and the recovery rate high, and sponge palladium with a grade of more than 99.90% can be recovered. BRIEF DESCRIPTION OF THE DRAWINGS

[0028] Figure 1 The figure is a process flow chart for recovering valuable metals from copper anode mud in a specific implementation manner. DETAILED DESCRIPTION

[0029] 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

[0030] A method for recovering valuable metals from copper anode mud, wherein: The main components of copper anode mud are: Au content is 0.95wt%, Ag content is 28.54wt%, Cu content is 12.34wt%, Te content is 2.34wt%, Se content is 3.25wt%, Pt content is 46.57g / t, and Pd content is 473.82g / t.

[0031] The steps are as follows: S1 Copper anode mud is mixed with concentrated sulfuric acid (concentration of 98wt%) in a solid-liquid mass ratio of 1:1, and sulfate roasting is carried out at 800℃ for 3h to obtain copper anode mud sulfate roasting slag and flue gas, and the flue gas is returned to the selenium recovery system to recover selenium dioxide; sulfuric acid and copper powder are added to the copper anode mud sulfate roasting slag, the pH value is adjusted to 1.0, and the reaction is carried out at 60℃ for 2h. After the reaction, the solid and liquid are separated to obtain acidic leaching solution and acidic leaching slag, and the acidic leaching solution is returned to the copper recovery system to recover cathode copper; wherein, the liquid-solid mass ratio of sulfuric acid to copper anode mud sulfate roasting slag is 3:1, and the molar ratio of copper powder to the total amount of tellurium in the copper anode mud sulfate roasting slag is 1.5:1.

[0032] S2: Add sodium hydroxide solution to the acidic leached residue obtained in step S1, adjust its pH value to 13.0, react at 60°C for 6 hours, and obtain alkaline leaching solution and alkaline leaching residue through solid-liquid separation. The alkaline leaching solution is returned to the tellurium electrowinning system at an electrowinning current density of 120A / m 2 Under the conditions of , tellurium powder is recovered, and the tellurium powder is washed with deionized water, dried, and melted to obtain tellurium ingots; wherein, the liquid-solid mass ratio of the sodium hydroxide solution to the acidic leaching residue is 6:1.

[0033] S3: The alkaline leaching residue obtained in step S2 is exposed to air and pickled with 1.5 wt % sulfuric acid at 90° C. for 6 h to obtain a washing liquid and washing residue, and the washing liquid is returned to the copper recovery system to recover cathode copper.

[0034] S4: adding hydrochloric acid, hydrogen peroxide (with a concentration of 30 wt%) and ferric chloride to the washing residue obtained in step S3, adjusting the pH value to 0.5, carrying out chlorination selective leaching reaction at 25°C for 2 hours, obtaining a leachate and a leaching residue through solid-liquid separation, using iron powder to undergo a replacement reaction in the leachate to obtain a crude palladium powder, and washing the crude palladium powder with dilute sulfuric acid, washing with clean water, and drying to obtain a sponge palladium; wherein the mass ratio of hydrogen peroxide to the washing residue is 0.03:1, the mass ratio of ferric chloride to the washing residue is 0.01:1, and the liquid-solid mass ratio of hydrochloric acid to the washing residue is 3:1.

[0035] S5: The leached residue obtained in step S4 is returned to the silver recovery system to obtain silver ingots and black gold powder, the black gold powder is sent to the gold recovery system to obtain gold ingots and platinum-rich liquid, and the platinum-rich liquid is returned to the platinum refining system to recover sponge platinum. Example 2

[0036] A method for recovering valuable metals from copper anode mud, wherein: The main components of copper anode mud are: Au content is 0.23wt%, Ag content is 11.25wt%, Cu content is 29.21wt%, Te content is 9.31wt%, Se content is 9.01wt%, Pt content is 24.31g / t, and Pd content is 263.21g / t.

[0037] The steps are as follows: S1 Copper anode mud is mixed with concentrated sulfuric acid (concentration of 98wt%) at a solid-liquid mass ratio of 4:1, and sulfate roasting is carried out at 830℃ for 6h to obtain copper anode mud sulfate roasting slag and flue gas, and the flue gas is returned to the selenium recovery system to recover selenium dioxide; sulfuric acid and copper powder are added to the copper anode mud sulfate roasting slag, the pH value is adjusted to 0.5, and the reaction is carried out at 80℃ for 3h. After the reaction, the solid and liquid are separated to obtain acidic leaching solution and acidic leaching slag, and the acidic leaching solution is returned to the copper recovery system to recover cathode copper; wherein, the liquid-solid mass ratio of sulfuric acid to copper anode mud sulfate roasting slag is 6:1, and the molar ratio of copper powder to the total amount of tellurium in the copper anode mud sulfate roasting slag is 1.7:1.

[0038] S2: Add sodium hydroxide solution to the acidic leached residue obtained in step S1, adjust its pH value to 14.0, react at 80°C for 2h, obtain alkaline leaching solution and alkaline leaching residue through solid-liquid separation, and return the alkaline leaching solution to the tellurium electrowinning system at an electrowinning current density of 150A / m 2 Under the conditions of , tellurium powder is recovered, and the tellurium powder is washed with deionized water, dried, and melted to obtain tellurium ingots; wherein, the liquid-solid mass ratio of the sodium hydroxide solution to the acidic leaching residue is 3:1.

[0039] S3: The alkaline leaching residue obtained in step S2 is exposed to air and pickled with 10 wt % sulfuric acid at 60° C. for 3 h to obtain a washing liquid and washing residue, and the washing liquid is returned to the copper recovery system to recover cathode copper.

[0040] S4: adding hydrochloric acid, hydrogen peroxide (with a concentration of 30 wt%) and ferric chloride to the washing residue obtained in step S3, adjusting the pH value to 0.3, carrying out chlorination selective leaching reaction at 15°C for 4 hours, obtaining a leachate and a leaching residue through solid-liquid separation, using iron powder to undergo a replacement reaction in the leachate to obtain a crude palladium powder, and washing the crude palladium powder with dilute sulfuric acid, washing with clean water, and drying to obtain a sponge palladium; wherein the mass ratio of hydrogen peroxide to the washing residue is 0.05:1, the mass ratio of ferric chloride to the washing residue is 0.05:1, and the liquid-solid mass ratio of hydrochloric acid to the washing residue is 6:1.

[0041] S5: The leached residue obtained in step S4 is returned to the silver recovery system to obtain silver ingots and black gold powder, the black gold powder is sent to the gold recovery system to obtain gold ingots and platinum-rich liquid, and the platinum-rich liquid is returned to the platinum refining system to recover sponge platinum. Example 3

[0042] A method for recovering valuable metals from copper anode mud, wherein: The main components of copper anode mud are: Au content is 5.36wt%, Ag content is 16.25wt%, Cu content is 19.24wt%, Te content is 4.18wt%, Se content is 6.21wt%, Pt content is 27.14g / t, and Pd content is 284.17g / t.

[0043] The steps are as follows: S1 Copper anode mud is mixed with concentrated sulfuric acid (concentration of 98wt%) in a solid-liquid mass ratio of 3:1, and sulfated roasted at 850℃ for 5h to obtain copper anode mud sulfated roasted slag and flue gas, and the flue gas is returned to the selenium recovery system to recover selenium dioxide; sulfuric acid and copper powder are added to the copper anode mud sulfated roasted slag, the pH value is adjusted to 0.1, and the reaction is carried out at 70℃ for 4h. After the reaction, the solid and liquid are separated to obtain acidic leaching solution and acidic leaching slag, and the acidic leaching solution is returned to the copper recovery system to recover cathode copper; wherein, the liquid-solid mass ratio of sulfuric acid to copper anode mud sulfated roasted slag is 4:1, and the molar ratio of copper powder to the total amount of tellurium in the copper anode mud sulfated roasted slag is 2.0:1.

[0044] S2: Add sodium hydroxide solution to the acidic leached residue obtained in step S1, adjust its pH value to 13.5, react at 70°C for 5h, obtain alkaline leaching solution and alkaline leaching residue through solid-liquid separation, and return the alkaline leaching solution to the tellurium electrowinning system at an electrowinning current density of 200A / m 2 Under the conditions of , tellurium powder is recovered, and the tellurium powder is washed with deionized water, dried, and melted to obtain tellurium ingots; wherein, the liquid-solid mass ratio of the sodium hydroxide solution to the acidic leaching residue is 4:1.

[0045] S3: The alkaline leaching residue obtained in step S2 is exposed to air and pickled with 5.0 wt% sulfuric acid at 80°C for 4 hours to obtain a washing liquid and washing residue, and the washing liquid is returned to the copper recovery system to recover cathode copper.

[0046] S4: adding hydrochloric acid, hydrogen peroxide (with a concentration of 30 wt%) and ferric chloride to the washing residue obtained in step S3, adjusting the pH value to 0.1, carrying out chlorination selective leaching reaction at 5°C for 3 hours, obtaining a leachate and a leaching residue through solid-liquid separation, using iron powder to undergo a replacement reaction in the leachate to obtain a crude palladium powder, and washing the crude palladium powder with dilute sulfuric acid, washing with clean water, and drying to obtain a sponge palladium; wherein the mass ratio of hydrogen peroxide to the washing residue is 0.10:1, the mass ratio of ferric chloride to the washing residue is 0.03:1, and the liquid-solid mass ratio of hydrochloric acid to the washing residue is 5:1.

[0047] S5: The leached residue obtained in step S4 is returned to the silver recovery system to obtain silver ingots and black gold powder, the black gold powder is sent to the gold recovery system to obtain gold ingots and platinum-rich liquid, and the platinum-rich liquid is returned to the platinum refining system to recover sponge platinum. Comparative Example 1

[0048] The specific steps are as in Example 1, except that: copper powder is not added in step S1. Comparative Example 2

[0049] The specific steps are as in Example 1, except that: ferric chloride is not added in step S4. Comparative Example 3

[0050] The specific steps are as in Example 1, except that: copper powder is not added in step S1, and ferric chloride is not added in step S4. test

[0051] The Te content in the acidic leaching solution obtained in step S1 of Examples 1 to 3 and Comparative Examples 1 to 3 and the Pt content in the leaching solution obtained in step S4 were detected, and the results are shown in Table 1; the grades of the tellurium ingots obtained in step S2 of Examples 1 to 3 and Comparative Examples 1 to 3 and the sponge palladium obtained in step S4 were detected, and the results are shown in Table 2; the recovery rate of each element in Examples 1 to 3 and Comparative Examples 1 to 3 was calculated, and the results are shown in Table 3.

[0052] Tellurium in the acidic leaching solution was detected according to the "Chemical Analysis Methods for Copper Anode Mud - Determination of Arsenic, Bismuth, Iron, Nickel, Lead, Antimony, Selenium and Tellurium Content - Inductively Coupled Plasma Atomic Emission Spectrometry" (GB / T 23607-2009), and platinum in the leaching solution was detected according to the "Chemical Analysis Methods for Precious Metal Alloy Electroplating Wastewater - Part 1: Determination of Gold, Silver, Platinum, Palladium and Iridium Content - Inductively Coupled Plasma Atomic Emission Spectrometry" (GB / T The tellurium ingot is tested in accordance with the Chemical Analysis Methods for Tellurium Part 12: Determination of Bismuth, Aluminum, Lead, Iron, Selenium, Copper, Magnesium, Sodium and Arsenic by Inductively Coupled Plasma Atomic Emission Spectrometry (YS / T 227.12-2011), and the sponge palladium is tested in accordance with the Chemical Analysis Methods for Sponge Palladium - Determination of Magnesium, Aluminum, Silicon, Cr, Manganese, Iron, Nickel, Copper, Zinc, Ruthenium, Rhodium, Silver, Tin, Iridium, Platinum, Gold, Lead and Bismuth by Inductively Coupled Plasma Mass Spectrometry (YS / T 1119-2016).

[0053] Table 1 Te content in acidic leaching solution and Pt content in leaching solution content Example 1 Example 2 Example 3 Comparative Example 1 Comparative Example 2 Comparative Example 3 Te (mg / L) 3.24 2.58 4.32 1253.56 3.25 1289.04 Pt (mg / L) 2.12 3.12 2.58 1.31 125.31 118.94

[0054] Table 2 Grades of tellurium ingot and palladium sponge grade Example 1 Example 2 Example 3 Comparative Example 1 Comparative Example 2 Comparative Example 3 Tellurium ingot (wt%) 99.92 99.91 99.94 99.94 99.91 99.93 Palladium sponge (wt%) 99.92 99.91 99.93 99.92 94.35 93.87

[0055] Table 3 Recovery rate of each element Total recovery rate Example 1 Example 2 Example 3 Comparative Example 1 Comparative Example 2 Comparative Example 3 Au(%) 99.21 99.18 99.14 99.08 99.11 99.06 Ag(%) 98.62 98.74 98.57 98.54 98.62 98.54 Pt(%) 95.68 96.24 95.87 95.18 90.24 90.45 Pd (%) 96.24 95.83 96.17 95.62 95.37 95.18 Te (%) 99.58 99.62 99.54 90.12 99.51 91.54 According to Tables 1 to 3, the present invention can recover valuable metals in copper anode mud to obtain a higher recovery rate and a high product quality. Among them, since comparative examples 1 and 3 do not add copper powder during the recovery of copper anode mud, the Te content in the acidic leaching solution is greatly increased compared with Examples 1 to 3, which can reach more than 1250 mg / L, resulting in a Te recovery rate much lower than that of Examples 1 to 3, with a maximum of only 91.54%; similarly, since comparative examples 2 and 3 do not add ferric chloride during the recovery of copper anode mud, the Pt content in the leaching solution is greatly increased compared with Examples 1 to 3, which can reach more than 118 mg / L, resulting in a decrease in the grade of sponge palladium, with a maximum of only 94.35wt%, and the recovery rate of Pt is also greatly affected, with a recovery rate of about 90%.

[0056] 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 copper anode mud, characterized in that: The steps include: S1: adding sulfuric acid and copper powder to the copper anode mud sulfate roasting slag, adjusting the pH value to below 1.0, and after the reaction, separating the solid and liquid to obtain acidic leaching solution and acidic leaching slag; S2: adding sodium hydroxide solution to the acidic leached residue obtained in step S1 to adjust the pH value to above 13.0, and obtaining alkaline leaching solution and alkaline leaching residue after reaction; S3: acid-washing the alkaline leaching residue obtained in step S2 to obtain a washing liquid and washing residue; S4: adding hydrochloric acid, hydrogen peroxide and ferric chloride to the washing residue obtained in step S3, adjusting the pH value to below 0.5, and obtaining a leaching solution and leaching residue after reaction; S5: recovering silver, gold and platinum in sequence from the leaching residue obtained in step S4.

2. The method according to claim 1, characterized in that In step S1, the liquid-solid mass ratio of sulfuric acid to copper anode mud sulfated roasted slag is (3-6):1, and the molar ratio of copper powder to the total amount of tellurium in the copper anode mud sulfated roasted slag is (1.5-2.0):

1.

3. The method according to claim 1, characterized in that In step S1, the reaction temperature is 60-80°C and the reaction time is 2-4h.

4. The method according to claim 1, characterized in that: In step S2, the liquid-solid mass ratio of the sodium hydroxide solution to the acidic leaching residue is (3-6):1, the reaction temperature is 60-80°C, and the reaction time is 2-6h.

5. The method according to claim 1, characterized in that In step S4, the mass ratio of hydrogen peroxide to washing residue is (0.03-0.10):1, the mass ratio of ferric chloride to washing residue is (0.01-0.05):1, and the liquid-solid mass ratio of hydrochloric acid to washing residue is (3-6):

1.

6. The method according to claim 1, characterized in that In step S4, the reaction temperature is 5-25° C. and the reaction time is 2-4 h.

7. The method according to claim 1, characterized in that In step S1, copper anode mud and concentrated sulfuric acid are subjected to sulfation roasting to obtain copper anode mud sulfation roasting slag and flue gas, and the flue gas is returned to the selenium recovery system to recover selenium dioxide.

8. The method according to claim 1, characterized in that The acidic leaching solution obtained in step S1 and the washing solution obtained in step S3 are returned to the copper recovery system to recover copper.

9. The method according to claim 1, characterized in that: The alkaline leaching solution obtained in step S2 is returned to the tellurium electrowinning system to recover tellurium.

10. The method according to claim 1, characterized in that The leaching solution obtained in step S4 is recovered to obtain palladium through a replacement reaction.

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