CN115418486B - Method for jointly recovering cobalt and manganese in zinc purification slag by acid leaching-precipitation flotation method - Google Patents

Abstract

The invention discloses a method for jointly recovering cobalt and manganese in zinc purification slag by an acid leaching-precipitation flotation method, which comprises the following steps: leaching the zinc purification slag by sulfuric acid to obtain zinc-rich leaching solution and leaching slag; leaching the leaching slag through acidic reduction to obtain cobalt-manganese leaching solution; sequentially precipitating and flocculating the cobalt-manganese leaching solution, then adding a surfactant to carry out froth flotation, and collecting cobalt-rich froth and manganese-rich solution. The method has the advantages of short process flow, low operation difficulty, easy control of technical conditions, environmental friendliness, realization of efficient separation and recovery of cobalt and manganese in the purified cobalt slag of the zinc hydrometallurgy, remarkable improvement of the comprehensive utilization added value of the purified slag, strong adaptability to a main system of the zinc hydrometallurgy, and good industrial application prospect.

Description

A method for combined acid leaching and sedimentation flotation to recover cobalt and manganese in zinc purification residue

Technical field

The invention relates to a method for processing zinc hydrometallurgy purification slag, specifically to a method for jointly recovering cobalt and manganese from zinc purification slag by an acid leaching-precipitation flotation method, and belongs to the technical field of comprehensive recycling and utilization of non-ferrous metal smelting.

Background technique

Cobalt and its compounds are widely used in key fields such as new energy materials and special performance alloys. With the rapid development of smart devices and new energy vehicle industries, the demand for cobalt will be increasing. my country's cobalt resources are mainly associated with associated minerals, and their reserves account for only 1.1% of the world's cobalt resource reserves. China is highly dependent on the import of cobalt resources. The recycling of cobalt-containing secondary resources can alleviate the shortage of cobalt resources in my country, while improving resource utilization efficiency and reducing the environmental impact caused by traditional stacking or incineration. In the zinc hydrometallurgy process, in order to eliminate the adverse effects of cobalt and other impurity metal ions in the solution on the subsequent electroplating steps, the leachate needs to be purified. Commonly used cobalt removal processes include zinc powder replacement, α-nitroso-β- Cobalt removal with naphthol and cobalt removal with xanthate will also produce a large amount of purification residue containing valuable metals such as zinc, cobalt, and manganese. The zinc in the purified slag can be recycled to the hydro-zinc smelting system after treatment, and other valuable metals can also be recycled after separation and enrichment, which has a high recovery value.

At present, the recovery of zinc hydrometallurgy purification residue is mainly through acid leaching or alkali leaching, dissolving all or part of the valuable metals into the solution, and performing separation and enrichment in the solution. Commonly used separation methods include hydroxide precipitation, sulfide precipitation, oxidation precipitation, solvent extraction, etc. However, since the precipitates are easily entrained with each other during the implementation of various precipitation methods, it is difficult to perform deep separation. Although the solvent extraction method has a better separation effect Good, but the cost of using the extraction agent is higher. For purified cobalt slag that contains many types of valuable metals, a single separation method usually cannot enrich and recover all valuable metals, and a combined process needs to be used to gradually separate and enrich it.

Contents of the invention

Based on the defects of poor selectivity of valuable metals, low separation efficiency and incomplete separation existing in the existing hydrometallurgical zinc purification slag recovery technology, the purpose of the present invention is to provide a combined acid leaching-precipitation flotation method to recover zinc. Method for purifying cobalt and manganese in slag. This method uses a combined process of selective acid leaching-reducing acid leaching-sedimentation flotation to obtain cobalt-rich foam products and manganese-enriched solutions. The process is short, the treatment cost is low, and it can achieve green and high-value utilization of purified cobalt slag.

In order to achieve the above object, the present invention provides a method for jointly recovering cobalt and manganese in zinc purification residue by acid leaching and precipitation flotation, which includes the following steps:

1) Leach the zinc smelting purification residue with sulfuric acid to obtain leachate and leaching residue;

2) Leaching the leaching residue through acidic reduction to obtain cobalt and manganese leaching liquid;

3) Precipitate and flocculate the cobalt-manganese leachate in sequence, then add surfactant to perform foam flotation, and collect cobalt-rich foam and manganese-rich solution.

The present invention can gradually enrich and recover the three valuable metals of cobalt, manganese and zinc in the purified residue through the combined process of selective acid leaching-reductive acid leaching-precipitation flotation. It is simple to operate and has strong controllability. Among them, each The reaction principle in the steps is as follows:

Sulfuric acid leaching: Zn+H ₂ SO ₄ =ZnSO ₄ +H ₂

ZnO+H ₂ SO ₄ =ZnSO ₄ +H ₂ O

CoO+H ₂ SO ₄ =CoSO ₄ +H ₂ O

Reductive acid leaching: MnO ₂ +H ₂ SO ₄ +H ₂ O ₂ =MnSO ₄ +2H ₂ O+O ₂ ↑

Co ₂ O ₃ +2H ₂ SO ₄ +H ₂ O ₂ =2CoSO ₄ +3H ₂ O+O ₂ ↑

Sedimentation flotation: CoSO ₄ +Na ₂ S＝CoS+Na ₂ SO ₄

Co ³⁺ +3C ₁₀ H ₆ ONOH＝CO(C ₁₀ H ₆ ONO) ₃ ↓+3H ⁺

As a preferred solution, in step 1), the leaching conditions are: sulfuric acid concentration is 2 to 4 mol/L, liquid-to-solid ratio is 10 to 20 mL: 1 g, temperature is 20 to 40°C, and time is 10 to 30 minutes. The leaching process uses mechanical stirring with a rotation speed of 200 to 400 rpm to assist leaching. The cobalt-containing particle suspension is transferred to a microbubble flotation column for aeration flotation.

During the leaching reaction process, it is necessary to control the sulfuric acid concentration and liquid-to-solid ratio within appropriate ranges. If the sulfuric acid concentration or liquid-to-solid ratio is too low, the leaching rate of zinc will be low and selective leaching cannot be achieved; if the sulfuric acid concentration is too high, it will cause Waste of reagents, and too high liquid-to-solid ratio will reduce the degree of zinc enrichment. Reaction temperature and time also affect the leaching efficiency. Too high temperature or too long reaction time will increase the leaching rate of cobalt and reduce the cobalt in the leaching residue.

As a preferred solution, the acidic reduction leaching uses a leaching agent containing sulfuric acid and a reducing agent. The reducing agent is at least one of hydrogen peroxide, sucrose and citric acid. Using hydrogen peroxide, sucrose, citric acid, etc. as reducing agents can reduce Co ³⁺ to Co ²⁺ and Mn ⁴⁺ to Mn ²⁺ , thereby dissolving cobalt and manganese in the solution to facilitate subsequent cobalt and manganese separation.

As a preferred solution, in step 2), the conditions for the acidic reduction leaching are: the sulfuric acid concentration in the leaching agent is 0.6~1.4mol/L, the reducing agent concentration is 0.8~1.2mol/L, and the liquid-to-solid ratio is 8～12mL: 1g.

As a preferred solution, in step 2), the acidic reduction leaching conditions are: temperature is 20-40°C, and time is 10-30 minutes. When the leaching residue is leached with a reducing acid solution, mechanical stirring with a rotation speed of 200 to 400 rpm is used to assist the leaching.

During the reduction leaching reaction process, controlling the sulfuric acid concentration, reducing agent concentration (added amount) and liquid-to-solid ratio within appropriate ranges can improve the selective leaching efficiency. Too low sulfuric acid concentration, reducing agent concentration (added amount) or liquid-to-solid ratio will result in low cobalt and manganese leaching rate; too high sulfuric acid concentration or reducing agent concentration (added amount) will lead to waste of reagents and increase production costs. , and too high a liquid-to-solid ratio will reduce the enrichment of cobalt and manganese.

As a preferred solution, the precipitating agent is agent A, which contains α-nitroso-β-naphthol and Na ₂ S, and the mole ratio of α-nitroso-β-naphthol to Na ₂ S The ratio is 1~3:1. The dosage of the precipitant is 0.8 to 1.2 times the molar amount of cobalt ions in the solution.

The above precipitant components can be used to selectively precipitate cobalt ions in the solution. In this selective precipitation reaction process, the amount of precipitant needs to be controlled within an appropriate range. Too little amount of precipitant will lead to incomplete cobalt ion precipitation reaction, and too much amount of precipitant will cause excess precipitant to react with manganese to form precipitation. , unable to achieve selective precipitation.

As a preferred solution, the flocculant is polyacrylamide. The concentration of the flocculant in the solution is 10-100 mg/L. During the flocculation process, magnetic stirring with a rotation speed of 200 to 400 rpm is used to assist leaching.

The use of flocculants can aggregate cobalt precipitated particles, and controlling the flocculants in a suitable range can improve the efficiency of subsequent flotation separation. If the amount of flocculant added is too small (the concentration is too low), the sediment particles will not be able to aggregate, which will adversely affect the subsequent flotation recovery rate; if the amount of flocculant added is too much (the concentration is too high), it will cause waste of reagents and increase production. cost.

As a preferred solution, the surfactant is CTAB. The concentration of the surfactant in the solution is 10-50 mg/L.

Adding surfactants can enhance foam stability and improve particle floatability. Regulating the concentration of surfactant in the solution within a suitable range is beneficial to the efficient separation of cobalt and manganese. Adding too little surfactant (concentration is too low) will result in weak floatability of the particles and affect the flotation recovery rate; while surfactant Adding too much (too high a concentration) will cause surfactant residues in the solution and affect the separation efficiency.

As a preferred option, in step 3), the precipitation reaction conditions are: the pH of the solution system is 2-4. The temperature is 20~40℃ and the time is 10~30min.

During the selective precipitation reaction process, controlling the pH of the solution system within a suitable range is conducive to the efficient separation and recovery of cobalt and manganese. Too high a pH value of the solution system will cause hydrolysis and precipitation of cobalt ions and manganese ions, and the pH value of the solution system If it is too low, it will have an adverse effect on the subsequent sedimentation and flotation reaction.

Compared with the prior art, the present invention has the following beneficial effects:

(1) In this method, zinc is first selectively leached using sulfuric acid, so that more than 95% of the zinc is leached into the leachate, and the leachate can be returned to the main wet zinc smelting system to recover zinc; secondly, cobalt-rich manganese leachate is obtained through acid reduction leaching, and the leachate is added A specific precipitant selectively precipitates cobalt ions, and by adding flocculants and surfactants, the cobalt-containing precipitated particles and manganese-rich solution are completely separated after foam flotation, gradually achieving efficient separation and high value of zinc, cobalt, and manganese in the purification slag. Recycle.

(2) This method has a short process flow, low operating difficulty, easy control of technical conditions, environmental friendliness, strong adaptability to the main hydrometallurgical zinc smelting system, and good industrial application prospects.

Description of drawings

Figure 1 is a schematic diagram of the process flow of the present invention.

Detailed ways

The present invention will be further described below with reference to the examples and drawings, but the present invention is not limited in any way. Any transformation or replacement based on the teachings of the present invention falls within the protection scope of the present invention.

Unless otherwise specified, various raw materials, reagents, instruments and equipment used in the present invention can be purchased in the market or prepared by existing methods.

The main chemical components of the purification slag used in the examples are shown in Table 1:

Table 1 Main chemical components of purification slag

Example 1

This embodiment provides a method for jointly recovering cobalt and manganese in zinc purification residue by acid leaching and precipitation flotation, which includes the following steps:

(1) Selective acid leaching: Crush, grind, and sieve the purified residue until the mass proportion of particles with a particle size less than 0.074mm accounts for more than 90%. The obtained samples are used for acid leaching experiments; configure a sulfuric acid leaching agent with a concentration of 4mol/L , mixed with the treated purified residue at a liquid-to-solid ratio of 20 mL: 1 g, heated to 30°C, leached for 10 minutes under mechanical stirring at 300 rpm, and solid-liquid separation was performed through vacuum filtration to obtain leachate and leaching residue. The leaching rates of Co, Mn, and Zn metals are 9.3%, 0.1%, and 96.6% respectively.

(2) Reductive acid leaching: Wash the leaching residue obtained in step (1) until it is neutral and then dry it. Prepare a sulfuric acid leaching agent with a concentration of 1 mol/L, mix it with the leaching residue at a liquid-to-solid ratio of 10 mL: 1 g, and heat to 30°C, add the reducing agent H ₂ O ₂ to the solution to a concentration of 1 mol/L, leaching for 20 minutes under mechanical stirring at 300 rpm, and perform solid-liquid separation through vacuum filtration. The leaching rates of Co and Mn are 99.3% respectively. , 98.7%, to obtain a cobalt-manganese solution.

(3) Precipitation reaction: Adjust the cobalt and manganese solution obtained in step (2) to pH 3 with NaOH solution, add agent A that is 1.2 times the theoretical reaction amount with cobalt, and the molar ratio of agent A is 1:1 Composed of α-nitroso-β-naphthol and Na ₂ S; heated to 30°C and reacted for 10 minutes under magnetic stirring at 300 rpm. The precipitation rates of Co and Mn were 96.3% and 5.4% respectively. Then add the flocculant polyacrylamide so that the concentration in the solution is 50 mg/L. After stirring evenly, add the surfactant CTAB so that the concentration in the solution is 50 mg/L and stir evenly.

(4) Foam flotation: Transfer the solution system obtained in step (3) to a microbubble flotation column for air-filled flotation to collect cobalt-rich foam products and manganese-rich solutions. The flotation recovery rates of Co and Mn are respectively 92.5%, 2.7%.

Example 2

(1) Selective acid leaching: Crush, grind, and screen the purified residue until the mass proportion of particles with a particle size less than 0.074 mm accounts for more than 90%. The obtained samples are used for acid leaching experiments; configure a sulfuric acid leach agent with a concentration of 2 mol/L , mixed with the treated purified residue at a liquid-to-solid ratio of 10 mL: 1 g, heated to 40°C, leached for 20 minutes under mechanical stirring at 300 rpm, and solid-liquid separation was performed through vacuum filtration to obtain leachate and leaching residue. The leaching rates of Co, Mn and Zn are 8.9%, 0.1% and 93.4% respectively.

(2) Reductive acid leaching: Wash the leaching residue obtained in step (1) until it is neutral and then dry it. Prepare a sulfuric acid leaching agent with a concentration of 0.8 mol/L, mix it with the leaching residue at a liquid-to-solid ratio of 8 mL: 1 g, and heat to 40°C, add the reducing agent H ₂ O ₂ so that the concentration in the solution is 0.8 mol/L, leaching for 30 minutes under mechanical stirring at 300 rpm, and perform solid-liquid separation through vacuum filtration. The leaching rates of Co and Mn are respectively 98.6%, 98.1%, and a cobalt-manganese solution was obtained.

(3) Precipitation reaction: Adjust the cobalt-manganese solution obtained in step (2) to pH 2.5 with NaOH solution, add agent A that is 1 times the theoretical reaction amount, and the agent A is composed of α- Nitroso-β-naphthol is composed of Na ₂ S; heated to 40°C and reacted for 20 minutes under magnetic stirring at 300 rpm. The precipitation rates of Co and Mn were 95.2% and 3.5% respectively. Then add the flocculant polyacrylamide so that the concentration in the solution is 30 mg/L. After stirring evenly, add the surfactant CTAB so that the concentration in the solution is 30 mg/L and stir evenly.

(4) Foam flotation: Transfer the solution system obtained in step (3) to a microbubble flotation column for air-filled flotation to collect cobalt-rich foam products and manganese-rich solutions. The flotation recovery rates of Co and Mn are respectively 92.3%, 2.4%.

Example 3

(1) Selective acid leaching: Crush, grind, and screen the purified residue until the mass proportion of particles with a particle size less than 0.074 mm accounts for more than 90%. The obtained samples are used for acid leaching experiments; configure a sulfuric acid leach agent with a concentration of 2 mol/L , mixed with the treated purified residue at a liquid-to-solid ratio of 15 mL: 1 g, heated to 35°C, leached for 15 minutes under mechanical stirring at 300 rpm, and solid-liquid separation was performed through vacuum filtration to obtain leachate and leaching residue. The leaching rates of Co, Mn and Zn are 9.1%, 0.1% and 95.8% respectively.

(2) Reductive acid leaching: Wash the leaching residue obtained in step (1) until it is neutral and then dry it. Prepare a sulfuric acid leaching agent with a concentration of 1.2 mol/L, mix it with the leaching residue at a liquid-to-solid ratio of 12 mL: 1 g, and heat to 35°C, add the reducing agent H ₂ O ₂ so that the concentration in the solution is 1.2 mol/L, leaching for 25 minutes under mechanical stirring at 300 rpm, and perform solid-liquid separation through vacuum filtration. The leaching rates of Co and Mn are respectively 99.5%, 99.1% to obtain cobalt manganese solution.

(3) Precipitation reaction: Adjust the cobalt and manganese solution obtained in step (2) to pH 3.5 with NaOH solution, add agent A with a theoretical reaction amount of 0.8 times, and the agent A is composed of α- at a molar ratio of 2:1 Nitroso-β-naphthol is composed of Na ₂ S; heated to 35°C and reacted for 15 minutes under magnetic stirring at 300 rpm. The precipitation rates of Co and Mn were 94.6% and 3.2% respectively. Then add the flocculant polyacrylamide so that the concentration in the solution is 60 mg/L. After stirring evenly, add the surfactant CTAB so that the concentration in the solution is 40 mg/L and stir evenly.

(4) Foam flotation: Transfer the solution system obtained in step (3) to a microbubble flotation column for air-filled flotation to collect cobalt-rich foam products and manganese-rich solutions. The flotation recovery rates of Co and Mn are respectively 91.7%, 2.1%.

Comparative example 1

(1) Selective acid leaching: Crush, grind, and sieve the purified residue until the mass proportion of particles with a particle size less than 0.074mm accounts for more than 90%. The obtained samples are used for acid leaching experiments; configure a sulfuric acid leaching agent with a concentration of 4mol/L , mixed with the treated purified residue at a liquid-to-solid ratio of 20 mL: 1 g, heated to 30°C, leached for 10 minutes under mechanical stirring at 300 rpm, and solid-liquid separation was performed through vacuum filtration to obtain leachate and leaching residue. The leaching rates of Co, Mn, and Zn metals are 9.3%, 0.1%, and 96.6% respectively.

(2) Reductive acid leaching: Wash the leaching residue obtained in step (1) until it is neutral and then dry it. Prepare a sulfuric acid leaching agent with a concentration of 1 mol/L, mix it with the leaching residue at a liquid-to-solid ratio of 10 mL: 1 g, and heat to 30°C, add the reducing agent H ₂ O ₂ to the solution to a concentration of 0.1 mol/L, leaching for 20 minutes under mechanical stirring at 300 rpm, and perform solid-liquid separation through vacuum filtration. The leaching rates of Co and Mn are 67.3 respectively. %, 51.7%, and a cobalt-manganese solution was obtained.

(3) Precipitation reaction: Adjust the cobalt-manganese solution obtained in step (2) to pH 3 with NaOH solution, add agent A with a theoretical reaction amount of 1.2 times, and the agent A is composed of α- at a molar ratio of 1:1 Nitroso-β-naphthol is composed of Na ₂ S; heated to 30°C and reacted for 10 minutes under magnetic stirring at 300 rpm. The precipitation rates of Co and Mn were 97.2% and 7.3% respectively. Then add the flocculant polyacrylamide so that the concentration in the solution is 50 mg/L. After stirring evenly, add the surfactant CTAB so that the concentration in the solution is 50 mg/L and stir evenly.

(4) Foam flotation: Transfer the solution system obtained in step (3) to a microbubble flotation column for air-filled flotation to collect cobalt-rich foam products and manganese-rich solutions. The flotation recovery rates of Co and Mn are respectively 90.2%, 4.3%.

Comparative example 2

(1) Selective acid leaching: Crush, grind, and screen the purified residue until the mass proportion of particles with a particle size less than 0.074mm accounts for more than 90%. The obtained samples are used for acid leaching experiments; configure a sulfuric acid leaching agent with a concentration of 1 mol/L , mixed with the treated purified residue at a liquid-to-solid ratio of 10 mL: 1 g, heated to 40°C, leached for 20 minutes under mechanical stirring at 300 rpm, and solid-liquid separation was performed through vacuum filtration to obtain leachate and leaching residue. The leaching rates of Co, Mn and Zn are 8.9%, 0.1% and 93.4% respectively.

(2) Reductive acid leaching: Wash the leaching residue obtained in step (1) until it is neutral and then dry it. Prepare a sulfuric acid leaching agent with a concentration of 0.8 mol/L, mix it with the leaching residue at a liquid-to-solid ratio of 8 mL: 1 g, and heat to 40°C, add the reducing agent H ₂ O ₂ so that the concentration in the solution is 0.8 mol/L, leaching for 30 minutes under mechanical stirring at 300 rpm, and perform solid-liquid separation through vacuum filtration. The leaching rates of Co and Mn are respectively 98.6%, 98.1%, and a cobalt-manganese solution was obtained.

(3) Precipitation reaction: Adjust the cobalt-manganese solution obtained in step (2) to pH 2.5 with NaOH solution, add agent A 0.5 times the theoretical reaction amount, and agent A is composed of α-substituted phosphate with a molar ratio of 3:1. Nitro-β-naphthol is composed of Na ₂ S; heated to 40°C and reacted for 20 minutes under magnetic stirring at 300 rpm. The precipitation rates of Co and Mn were 47.3% and 0.6% respectively. Then add the flocculant polyacrylamide so that the concentration in the solution is 30 mg/L. After stirring evenly, add the surfactant CTAB so that the concentration in the solution is 30 mg/L and stir evenly.

(4) Foam flotation: Transfer the solution system obtained in step (3) to a microbubble flotation column for air-filled flotation to collect cobalt-rich foam products and manganese-rich solutions. The flotation recovery rates of Co and Mn are respectively 93.4%, 2.4%.

Comparative example 3

(1) Selective acid leaching: Crush, grind, and screen the purified residue until the mass proportion of particles with a particle size less than 0.074 mm accounts for more than 90%. The obtained samples are used for acid leaching experiments; configure a sulfuric acid leach agent with a concentration of 2 mol/L , mixed with the treated purified residue at a liquid-to-solid ratio of 15 mL: 1 g, heated to 35°C, leached for 15 minutes under mechanical stirring at 300 rpm, and solid-liquid separation was performed through vacuum filtration to obtain leachate and leaching residue. The leaching rates of Co, Mn and Zn are 9.1%, 0.1% and 95.8% respectively.

(2) Reductive acid leaching: Wash the leaching residue obtained in step (1) until it is neutral and then dry it. Prepare a sulfuric acid leaching agent with a concentration of 1.2 mol/L, mix it with the leaching residue at a liquid-to-solid ratio of 12 mL: 1 g, and heat to 35°C, add the reducing agent H ₂ O ₂ so that the concentration in the solution is 1.2 mol/L, leaching for 25 minutes under mechanical stirring at 300 rpm, and perform solid-liquid separation through vacuum filtration. The leaching rates of Co and Mn are respectively 99.5%, 99.1% to obtain cobalt manganese solution.

(3) Precipitation reaction: Adjust the cobalt and manganese solution obtained in step (2) to pH 3.5 with NaOH solution, add agent A with a theoretical reaction amount of 0.8 times, and agent A is composed of α-substituted with a molar ratio of 2:1. Composed of nitro-β-naphthol and Na ₂ S; heated to 35°C and reacted for 15 minutes under magnetic stirring at 300 rpm. The precipitation rates of Co and Mn were 94.6% and 3.2% respectively. Then add the flocculant polyacrylamide so that the concentration in the solution is 1 mg/L. After stirring evenly, add the surfactant CTAB so that the concentration in the solution is 40 mg/L and stir evenly.

(4) Foam flotation: Transfer the solution system obtained in step (3) to a microbubble flotation column for air-filled flotation to collect cobalt-rich foam products and manganese-rich solutions. The flotation recovery rates of Co and Mn are respectively 84.2%, 1.1%.

The embodiments of the present invention have been described in detail above in conjunction with the examples, but the present invention is not limited to the described embodiments. For those skilled in the art, various changes, modifications, substitutions and modifications can be made to these embodiments without departing from the principle and spirit of the invention, and they still fall within the protection scope of the invention.

Claims (4)

1. A method for jointly recovering cobalt and manganese from zinc purification residue by acid leaching and precipitation flotation, which is characterized by: including the following steps: 1) Leach the zinc smelting purification residue with sulfuric acid to obtain leachate and leaching residue; 2) Leaching the leaching residue through acidic reduction to obtain cobalt and manganese leaching liquid; 3) Precipitate and flocculate the cobalt-manganese leachate in sequence, then add surfactant to perform foam flotation, and collect cobalt-rich foam and manganese-rich solution; In step 1), the leaching conditions are: sulfuric acid concentration is 2~4mol/L, liquid-to-solid ratio is 10~20mL:1g, temperature is 20~40°C, and time is 10~30min; In step 2), the acidic reduction leaching uses a leaching agent containing sulfuric acid and a reducing agent; the conditions for the acidic reduction leaching are: the sulfuric acid concentration in the leaching agent is 0.6~1.4mol/L, and the reducing agent concentration is 0.8~ 1.2mol/L, liquid-solid ratio is 8~12mL:1g; temperature is 20~40℃, time is 10~30min; The precipitating agent is agent A, which contains α-nitroso-β-naphthol and Na ₂ S. The molar ratio of α-nitroso-β-naphthol to Na ₂ S is 1 to 3:1; The amount of precipitant used is 0.8~1.2 times the molar amount of cobalt ions in the solution; The flocculant is polyacrylamide; the concentration of the flocculant in the solution is 10~100mg/L. 2. A method for jointly recovering cobalt and manganese in zinc purification residue by acid leaching and precipitation flotation according to claim 1, characterized in that: the reducing agent is at least one of hydrogen peroxide, sucrose, and citric acid. 3. A method for jointly recovering cobalt and manganese in zinc purification residue by acid leaching and sedimentation flotation according to claim 1, characterized by: The surfactant is CTAB; The concentration of the surfactant in the solution is 10~50mg/L. 4. A method for jointly recovering cobalt and manganese in zinc purification slag by an acid leaching-precipitation flotation method according to claim 1, characterized in that: in step 3), the reaction conditions of the precipitation are: solution system The pH is 2~4; the temperature is 20~40℃, and the time is 10~30min.