A method for preparing battery-grade iron phosphate using sulfuric acid residue and recovering gold, copper and cobalt

Abstract

The invention discloses a method for preparing battery-grade iron phosphate using sulfuric acid slag and recovering gold, copper and cobalt. The method comprises the following steps: leaching sulfuric acid slag with phosphoric acid solution and elemental iron to obtain ferrous dihydrogen phosphate leaching solution and leaching slag; adding an oxidant to the ferrous dihydrogen phosphate leaching solution and adjusting the pH to perform a precipitation reaction to obtain a dihydrate iron phosphate precipitate and a phosphoric acid solution; the dihydrate iron phosphate precipitate is calcined to obtain iron phosphate; and the leaching slag is used to recover gold, copper and cobalt. The method has a high iron recovery rate, and the obtained iron phosphate has a high purity. At the same time, gold, copper and cobalt in sulfuric acid slag can be effectively recovered. The process is simple, the cost is low, and the method is suitable for industrial production.

Description

Technical Field

[0001] The present invention relates to a method for preparing ferric phosphate, and in particular to a method for preparing battery-grade ferric phosphate by utilizing sulfuric acid slag and recovering gold, copper and cobalt, belonging to the field of new energy battery materials. Background Art

[0002] Sulfate slag is a byproduct of pyrite-based acid production. Its primary component is iron oxide. Every ton of sulfuric acid produced generates approximately 0.5 tons of slag. my country produces approximately 20 million tons of slag annually, the vast majority of which is used as a steelmaking raw material. The iron grade of slag generally ranges from 55% to 60%. In addition to iron, it also contains impurities such as silicon, calcium, magnesium, and non-ferrous metals. When used as an ironmaking raw material, these impurities can affect the smelting process. Consequently, the application of slag in the steel industry is severely limited, hindering its full resource utilization.

[0003] With the rapid development of the new energy industry, the production capacity of lithium iron phosphate materials has expanded rapidly. Among them, iron phosphate, as a precursor of lithium iron phosphate materials, has also been attracting much attention. The main methods for synthesizing battery-grade iron phosphate include liquid phase precipitation, hydrothermal synthesis, sol-gel, air oxidation and controlled crystallization, among which liquid phase precipitation is the most widely used. In the liquid phase precipitation method, cheap iron source is an important link in controlling its production cost. Sulfuric acid slag is a cheap iron-containing secondary resource with large output and wide distribution range. If sulphuric acid slag can be used as an iron source for iron phosphate production, it can not only reduce the production cost of iron phosphate, but also increase the value of sulphuric acid slag. Previous studies have used sulphuric acid slag as an iron source for direct phosphoric acid leaching to prepare iron phosphate products. However, in this process, the silicon component impurities in sulphuric acid slag cannot be dissolved in phosphoric acid and will precipitate together with iron phosphate dihydrate, affecting the purity of iron phosphate. Hydrofluoric acid is required for subsequent impurity removal, which is a complicated process.

[0004] In addition, the current recovery of gold, copper and cobalt using sulfuric acid slag requires first removing iron and then separating the gold, copper and cobalt, which is a cumbersome process. The present invention can prepare iron phosphate while effectively separating gold, copper and cobalt, simplifying the recovery steps and improving the separation recovery rate of each element.

[0005] Therefore, developing a simple method for preparing battery-grade iron phosphate using sulfuric acid slag as an iron source and recovering gold, copper and cobalt is of great significance to the sustainable development of my country's chemical and new energy industries. Summary of the Invention

[0006] To address the shortcomings of the prior art, the present invention aims to provide a method for preparing battery-grade iron phosphate from sulfate slag and recovering gold, copper, and cobalt. This method improves the recovery rate of iron from sulfate slag, yielding high-purity iron phosphate that can be directly used in battery production. It also effectively recovers gold, copper, and cobalt. The method is simple, low-cost, and suitable for industrial production.

[0007] To achieve the above technical objectives, the present invention provides a method for preparing battery-grade ferric phosphate and recovering gold, copper and cobalt using sulfuric acid slag. The method comprises the following steps: leaching the sulfuric acid slag with a phosphoric acid solution and elemental iron to obtain a ferrous dihydrogen phosphate leachate and a leaching residue; adding an oxidant to the ferrous dihydrogen phosphate leachate and adjusting the pH to carry out a precipitation reaction to obtain a ferric phosphate dihydrate precipitate and a phosphoric acid solution; calcining the ferric phosphate dihydrate precipitate to obtain ferric phosphate; and using the leaching residue to recover gold, copper and cobalt.

[0008] The present invention cleverly utilizes the reducing property of elemental iron to reduce trivalent iron in the sulfuric acid slag to divalent iron during the phosphoric acid leaching process, and reacts with phosphoric acid to generate ferrous dihydrogen phosphate, so that all iron components in the raw material are converted into soluble ferrous dihydrogen phosphate salts, while silicon, calcium impurities, copper, cobalt and other components remain insoluble. On the other hand, the entry of elemental iron into the system can reduce ionic gold, copper and cobalt elements in the solution to elemental gold, copper and cobalt, thereby reducing the loss rate of gold, copper and cobalt during the leaching process, improving the gold, copper and cobalt content and recovery rate in the leaching residue, and also enabling the iron component in the sulfuric acid slag to be leached with high selectivity to obtain a high-purity ferrous dihydrogen phosphate solution. At the same time, gold, copper and cobalt in the leaching residue are enriched and can be directly used for separating and recovering gold, copper and cobalt.

[0009] As a preferred solution, the molar ratio of Fe2O3 to phosphoric acid and elemental iron in the sulfuric acid slag is 1:6-10:1.5-3.

[0010] As a preferred solution, during the leaching process, the concentration of the phosphoric acid solution is 3-8 mol / L, the liquid-solid ratio is 6-15 ml / g, the temperature is 30-100° C., more preferably 60-100° C., and the time is 2-6 h.

[0011] As a preferred solution, the oxidant includes at least one of hydrogen peroxide, air, and oxygen.

[0012] As a preferred solution, when the oxidant is hydrogen peroxide, the molar ratio of H2O2 to ferrous dihydrogen phosphate in the hydrogen peroxide is 1:2 to 2.5.

[0013] As a preferred solution, when the oxidant is air or oxygen, the gas flow rate is 50 to 200 mL / min.

[0014] As a preferred embodiment, the precipitation reaction process is as follows: stirring the reaction at a temperature of 60-80° C. for 1-3 hours. The precipitation reaction is completed by stirring the reaction at a temperature of 60-80° C. until the iron content in the solution does not exceed 0.4 g / L.

[0015] As a preferred solution, after the precipitation reaction is completed, the iron content in the solution is ≤0.4 g / L.

[0016] As a preferred solution, the calcination conditions are: temperature of 500-700° C., and time of 60-180 min.

[0017] As a preferred solution, the process for recovering gold, copper and cobalt is as follows: leaching the leaching residue with sulfuric acid and then performing solid-liquid separation to obtain a copper-cobalt leachate and a gold-containing leaching residue; mixing the copper-cobalt leachate with a sodium thiosulfate solution and adjusting the pH to perform a copper precipitation reaction, and then performing solid-liquid separation to obtain copper slag and a cobalt-containing solution; and leaching the gold-containing leaching residue with a gold leaching agent to obtain a gold-containing leachate.

[0018] As a preferred solution, during the copper and cobalt recovery process, the mass concentration of the sodium thiosulfate solution is 8-10%; the amount of sodium thiosulfate in the sodium thiosulfate solution is 2-3 times the molar amount of the copper element; the pH is adjusted to 4.0-6.0, and the pH regulator is sodium carbonate; the temperature is 80-100° C., and the time is 3-4 hours.

[0019] As a preferred solution, the cobalt-containing solution uses a sodium carbonate solution with a mass fraction of 8-10% as a precipitant, adjusts the solution pH to 8.0-9.0, reacts at room temperature with stirring for 3-4 hours, and then separates the solid and liquid to obtain cobalt slag.

[0020] As a preferred solution, the gold leaching agent includes at least one of cyanide, thiosulfate, and thiourea.

[0021] As a preferred solution, the molar dosage of the gold leaching agent is determined according to the gold content in the sulfuric acid slag, and the dosage is 1.2 to 1.5 times the theoretical dosage.

[0022] As a preferred solution, the leaching temperature of the gold-containing leaching residue is 20-70° C., and the leaching time is 6-10 hours.

[0023] As a preferred solution, in the desorption solution, the mass concentration of sodium chloride is 1-5%, and the mass concentration of sodium hydroxide is 1-2%.

[0024] As a preferred solution, the gold mud is dried and then melted to obtain elemental gold.

[0025] Compared with the prior art, the present invention has the following beneficial effects:

[0026] (1) The present invention uses sulfuric acid slag as the iron source and prepares a high-value-added battery-grade iron phosphate product through a phosphoric acid reduction leaching-oxidation precipitation-roasting dehydration process. Compared with the traditional iron phosphate production process, the present invention has the advantages of a wide range of raw material sources and low cost, thereby realizing the high-value-added utilization of industrial by-products.

[0027] (2) The present invention adds an iron reducing agent during the phosphoric acid leaching process, cleverly utilizing the reducing property of elemental iron to reduce the trivalent iron ions generated during the acid leaching process to divalent iron ions, thereby causing them to react with phosphate radicals to generate soluble ferrous dihydrogen phosphate. Impurity components in the sulfate residue do not dissolve into the leaching residue, thereby achieving phase transformation of iron and phosphorus and impurity removal in one step.

[0028] (3) While obtaining the ferric phosphate product, the present invention can also obtain the leaching residue obtained in the phosphoric acid reduction leaching process, which can be directly used to separate and recover gold, copper and cobalt. The gold, copper and cobalt grades in the leaching residue are greatly improved compared with the sulfate residue raw material. In the subsequent recovery process, the reagent consumption can be reduced, and the gold, copper and cobalt ion concentration of the system can be increased, which greatly simplifies the process of recovering gold, copper and cobalt using sulfate residue, and significantly improves resource utilization. BRIEF DESCRIPTION OF THE DRAWINGS

[0029] In order to more clearly illustrate the embodiments of the present invention or the technical solutions in the prior art, the following briefly introduces the drawings required for use in the embodiments or the description of the prior art. Obviously, the drawings described below are merely embodiments of the present invention. For ordinary technicians in this field, other drawings can be obtained based on the provided drawings without paying any creative work.

[0030] Figure 1 This is a process flow chart of the present invention for preparing battery-grade iron phosphate and recovering gold, copper and cobalt using sulfuric acid slag.

[0031] Figure 2 This is the XRD pattern of the leached residue obtained after reduction leaching of the sulfuric acid residue in Example 1.

[0032] Figure 3 This is the XRD spectrum of the ferric phosphate dihydrate prepared in Example 1.

[0033] Figure 4 This is the SEM image of the ferric phosphate dihydrate prepared in Example 2.

[0034] Figure 5 This is the XRD diagram of the leached residue obtained after reduction leaching of the sulfuric acid residue in Comparative Example 1. DETAILED DESCRIPTION

[0035] The following will clearly and completely describe the technical solutions of the present invention in conjunction with the embodiments of the present invention and the accompanying drawings. Obviously, the embodiments described are only part of the embodiments of the present invention, not all of the embodiments. Based on the embodiments of the present invention, all other embodiments obtained by ordinary technicians in this field without making creative efforts are within the scope of protection of the present invention.

[0036] Example 1

[0037] This embodiment uses sulfuric acid slag from a sulfuric acid plant with an iron grade of 60.87%. The other main components are shown in Table 1.

[0038] Table 1 Main chemical components of sulfate slag / %

[0039]

[0040] Phosphoric acid reduction leaching test was carried out using this as raw material. The molar ratio of Fe2O3, phosphoric acid, and reducing agent in the sulfate slag was controlled to be 1:8:1.6. The reducing agent was elemental iron powder, the phosphoric acid concentration was 5 mol / L, the liquid-solid ratio was 8 ml / g, and the leaching temperature was controlled at 70°C. After 4 hours of leaching, the iron components in the sulfate slag were basically converted into soluble ferrous dihydrogen phosphate. The main components of the leached slag were siliceous gangue and insoluble metals such as gold, copper, and cobalt (( Figure 2 The leached residue can be used for subsequent extraction of valuable metals. The chemical composition of the leached residue is shown in Table 2.

[0041] Table 2 Chemical composition analysis of reduction leaching residue / %

[0042]

[0043] After solid-liquid separation, a slightly excess amount of 25% mass concentration hydrogen peroxide was added to the ferrous dihydrogen phosphate solution, wherein the molar ratio of Fe2O3 to H2O2 was 1:1.7; the reaction temperature of the hydrogen peroxide oxidation step was controlled at 60°C and stirred for 3 hours until the iron ion content of the supernatant was less than 0.4g / L. After solid-liquid separation, the obtained ferric phosphate dihydrate product (( Figure 3 The product was calcined at 550°C for 2 hours to obtain an iron phosphate product with an iron recovery rate of 76.63%. The chemical composition of the iron phosphate product is shown in Table 3.

[0044] Table 3 All element analysis of iron phosphate / %

[0045]

[0046] The leaching residue obtained during the phosphoric acid leaching process is used to recover gold, copper and cobalt:

[0047] The leaching residue was mixed with 1 mol / L sulfuric acid at a solid-liquid mass ratio of 1:3 and stirred for 3 hours to obtain a copper-cobalt leachate and a gold-containing leaching residue. Sodium thiosulfate with a mass concentration of 10% was added to the copper-cobalt leachate, and the amount was twice the copper ion concentration. Sodium carbonate was used as a pH adjuster to adjust the pH value of the mixed solution to 5. The stirring temperature was 80°C and the stirring time was 3 hours. Finally, copper precipitate residue and cobalt-rich solution were obtained, and the copper recovery rate was 68.23%.

[0048] 9% mass concentration of sodium carbonate was added as a precipitant to the cobalt-rich solution, and the pH value of the solution was adjusted to 8. After stirring at room temperature for 4 hours, solid-liquid separation was performed to obtain cobalt precipitate slag, and the cobalt recovery rate was 59.25%.

[0049] The gold-containing leaching residue was mixed with 0.3 mol / L sodium thiosulfate solution at a solid-liquid mass ratio of 1:4, a system pH of 10.5, a stirring time of 6 hours, and a stirring temperature of 40°C. The mixture was then separated into solid and liquid to obtain a gold leaching solution, which was then subjected to activated carbon adsorption at an activated carbon dosage of 10 g / L and a stirring time of 24 hours. The mixed solution was further separated into solid and liquid to obtain a gold-loaded carbon solid phase, which was then desorbed using a mixed desorption solution of 5 wt% NaCl and 2 wt% NaOH at a desorption dosage of 0.05 L / g for 24 hours to obtain a gold-rich solution. The gold-rich solution was electrolytically treated to obtain gold mud, wherein the gold recovery rate was 62.25%.

[0050] Example 2

[0051] This embodiment still uses the above-mentioned sulfuric acid slag, with an iron grade of 60.87%. The other main components are shown in Table 1. This is the raw material for the phosphoric acid reduction leaching test. The molar ratio of Fe2O3, phosphoric acid and reducing agent in the sulfuric acid slag is controlled to be 1:10:1.5, the reducing agent is reduced iron powder, the phosphoric acid concentration is 4 mol / L, the liquid-solid ratio is 15 ml / g, and the leaching temperature is controlled at 90°C. After 5 hours of leaching, the iron components in the sulfuric acid slag have basically been converted into soluble ferrous dihydrogen phosphate, and the main components of the leached slag are siliceous gangue and insoluble metals such as gold, copper and cobalt, which can be used for subsequent extraction of valuable metals. After solid-liquid separation, a slightly excess of 25% mass concentration hydrogen peroxide is added to the ferrous dihydrogen phosphate solution, and the molar ratio of Fe2O3 to hydrogen peroxide is 1:1.7. The reaction temperature of the hydrogen peroxide oxidation step is controlled to 60°C and stirred for 4 hours until the iron ion content of the supernatant is less than 0.4g / L. After solid-liquid separation, the obtained dihydrated iron phosphate product (( Figure 4 The ferric phosphate product was finally obtained by calcining at 600°C for 2 hours. The chemical composition of the ferric phosphate product is shown in Table 4.

[0052] Table 4 Main chemical components of iron phosphate / %

[0053]

[0054] The leaching residue was mixed with 1 mol / L sulfuric acid at a solid-liquid mass ratio of 1:4 and stirred for 4 hours to obtain a copper-cobalt leachate and a gold-containing leaching residue. Sodium thiosulfate with a mass concentration of 10% was added to the leachate, and the amount was 2.5 times the copper ion concentration. Sodium carbonate was used as a pH adjuster to adjust the pH value of the mixed solution to 5. The stirring temperature was 90°C and the stirring time was 3 hours. Finally, copper precipitate residue and cobalt-rich solution were obtained, and the copper recovery rate was 70.68%.

[0055] 9% mass concentration of sodium carbonate was added to the cobalt-rich solution as a precipitant, and the pH value of the solution was adjusted to 9. After stirring at room temperature for 4 hours, solid-liquid separation was performed to obtain cobalt precipitate slag, and the cobalt recovery rate was 61.23%.

[0056] The gold-containing leaching residue was mixed with 0.4 mol / L sodium thiosulfate solution at a solid-liquid mass ratio of 1:4, a system pH of 10.5, a stirring time of 6 hours, and a stirring temperature of 50°C. The mixture was then separated into solid and liquid to obtain a gold leaching solution, which was then subjected to activated carbon adsorption at an activated carbon dosage of 10 g / L and a stirring time of 12 hours. The mixed solution was further separated into solid and liquid to obtain a gold-loaded carbon solid phase, which was then desorbed using a mixed desorption solution of 5 wt% NaCl and 2 wt% NaOH at a desorption dosage of 0.05 L / g for 12 hours to obtain a gold-rich solution. The gold-rich solution was electrolytically treated to obtain gold mud, wherein the gold recovery rate was 60.17%.

[0057] Example 3

[0058] This example uses sulfuric acid slag from a certain factory with an iron grade of 63.45%. The other main components are shown in Table 5.

[0059] Table 5 Main chemical components of sulfate slag / %

[0060]

[0061] Phosphoric acid reduction leaching experiments were conducted using this raw material. The molar ratio of Fe₂O₃, phosphoric acid, and reducing agent in the iron concentrate was controlled at 1:9:1.5. Reduced iron powder was used as the reducing agent, the phosphoric acid concentration was 6 mol / L, the liquid-to-solid ratio was 6 ml / g, and the leaching temperature was controlled at 90°C. After 5 hours of leaching, the iron components in the sulfate residue were essentially converted into soluble ferrous dihydrogen phosphate. The leached residue was primarily composed of siliceous gangue and insoluble metals such as gold, copper, and cobalt, suitable for subsequent extraction of valuable metals. After solid-liquid separation, air was introduced into the ferrous dihydrogen phosphate solution. The oxidation reaction temperature was controlled at 70°C and stirred until the supernatant contained less than 0.4 g / L iron ions. The resulting ferric phosphate dihydrate was calcined at 600°C for 2 hours to yield the final ferric phosphate product. The chemical composition of the ferric phosphate product is shown in Table 6.

[0062] Table 6 Main chemical components of iron phosphate / %

[0063]

[0064] The leaching residue was mixed with 1 mol / L sulfuric acid in a solid-liquid mass ratio of 1:2 and stirred for 3 hours to obtain a copper-cobalt leachate and a gold-containing leaching residue. Sodium thiosulfate with a mass concentration of 8% was added to the leachate, and the amount was twice the copper ion concentration. Sodium carbonate was used as a pH adjuster to adjust the pH value of the mixed solution to 5. The stirring temperature was 80°C and the stirring time was 3 hours. Finally, copper precipitate residue and cobalt-rich solution were obtained, and the copper recovery rate was 65.52%.

[0065] 8% mass concentration of sodium carbonate was added as a precipitant to the cobalt-rich solution, and the pH value of the solution was adjusted to 8. After stirring at room temperature for 3 hours, solid-liquid separation was performed to obtain cobalt precipitate slag, and the cobalt recovery rate was 56.53%.

[0066] The gold-containing leaching residue was mixed with 25 kg / t thiourea solution at a solid-liquid mass ratio of 1:4, a system pH of 1.5, a stirring time of 5 hours, and a stirring temperature of 30°C. The mixture was then separated into solid and liquid to obtain a gold leaching solution, which was then adsorbed on activated carbon at an amount of 15 g / L and a stirring time of 24 hours. The mixed solution was further separated into solid and liquid to obtain a gold-loaded carbon solid phase, which was then desorbed using a mixed desorption solution of 5% NaCl and 2% NaOH at a desorption amount of 0.1 L / g for 24 hours to obtain a gold-rich solution. The gold-rich solution was then electrolytically treated to obtain gold mud, wherein the gold recovery rate was 69.42%.

[0067] Comparative Example 1

[0068] This comparative example uses the sulfuric acid slag in Example 1 as raw material to carry out phosphoric acid reduction leaching test. The molar ratio of Fe2O3, phosphoric acid and reducing agent in the sulfuric acid slag is controlled to be 1:8:1, the reducing agent is elemental iron powder, the phosphoric acid concentration is 5 mol / L, the liquid-solid ratio is 8 ml / g, the leaching temperature is controlled at 70°C, and the leaching time is 4 hours. In this comparative example, the molar ratio of Fe2O3 to elemental iron powder in the sulfuric acid slag is 1:1, which is the theoretical ratio. In the actual leaching process, since the elemental iron powder itself will react with phosphoric acid to generate ferrous dihydrogen phosphate, there is not enough elemental iron to reduce the trivalent iron ions generated by the reaction of Fe2O3 and phosphoric acid, resulting in insufficient reaction of the sulfuric acid slag and some iron remaining in the leaching slag ( Figure 5 The chemical composition of the leached residue is shown in Table 7.

[0069] Table 7 Chemical composition analysis of reduction leaching residue / %

[0070]

[0071] After solid-liquid separation, a slight excess of 25% mass concentration of hydrogen peroxide was added to the ferrous dihydrogen phosphate solution, with a molar ratio of Fe₂O₃ to H₂O₂ of 1:1.7. The hydrogen peroxide oxidation step was carried out at a temperature of 60°C and stirred for 3 hours until the supernatant's iron ion content was less than 0.4 g / L. The resulting ferric phosphate dihydrate was calcined at 550°C for 2 hours to yield the final ferric phosphate product, with an iron recovery of 73.68%. The chemical composition of the ferric phosphate product is shown in Table 8.

[0072] Table 8 All element analysis of iron phosphate / %

[0073]

[0074] The leaching residue obtained during the phosphoric acid leaching process is used to recover gold, copper and cobalt:

[0075] The leaching residue was mixed with 1 mol / L sulfuric acid at a solid-liquid mass ratio of 1:3 and stirred for 3 hours to obtain a copper-cobalt leachate and a gold-containing leaching residue. Sodium thiosulfate with a mass concentration of 10% was added to the copper-cobalt leachate, and the amount was twice the copper ion concentration. Sodium carbonate was used as a pH adjuster to adjust the pH value of the mixed solution to 5. The stirring temperature was 80°C and the stirring time was 3 hours. Finally, copper precipitate residue and cobalt-rich solution were obtained, and the copper recovery rate was 29.36%.

[0076] 9% mass concentration of sodium carbonate was added to the cobalt-rich solution as a precipitant, and the pH value of the solution was adjusted to 8. After stirring at room temperature for 4 hours, solid-liquid separation was performed to obtain cobalt precipitate slag, and the cobalt recovery rate was 19.63%.

[0077] The gold-containing leaching residue was mixed with 0.3 mol / L sodium thiosulfate solution at a solid-liquid mass ratio of 1:4, a system pH of 10.5, a stirring time of 6 hours, and a stirring temperature of 40°C. The mixture was then separated into solid and liquid to obtain a gold leaching solution, which was then subjected to activated carbon adsorption at an activated carbon dosage of 10 g / L and a stirring time of 24 hours. The mixed solution was further separated into solid and liquid to obtain a gold-loaded carbon solid phase, which was then desorbed using a mixed desorption solution of 5 wt% NaCl and 2 wt% NaOH at a desorption dosage of 0.05 L / g for 24 hours to obtain a gold-rich solution. The gold-rich solution was electrolytically treated to obtain gold mud, wherein the gold recovery rate was 24.65%.

[0078] Various modifications to these embodiments will be readily apparent to those skilled in the art, and the general principles defined herein may be implemented in other embodiments without departing from the spirit or scope of the invention. Therefore, the present invention is not limited to the embodiments shown herein, but is to be construed in the widest manner consistent with the principles and novel features disclosed herein.

Claims

1. A method for preparing battery-grade iron phosphate using sulfuric acid slag and recovering gold, copper and cobalt, characterized in that: The sulfuric acid residue is mixed with a phosphoric acid solution and elemental iron for leaching to obtain a ferrous dihydrogen phosphate leaching solution and a leaching residue; an oxidant is added to the ferrous dihydrogen phosphate leaching solution for precipitation reaction to obtain a dihydrate ferric phosphate precipitate and a phosphoric acid solution; the dihydrate ferric phosphate precipitate is calcined to obtain battery-grade ferric phosphate; the leaching residue is used to recover gold, copper and cobalt; The molar ratio of Fe2O3 to phosphoric acid and elemental iron in the sulfuric acid slag is 1:9-10:1.5-3; The process of recovering gold, copper and cobalt comprises: leaching the leaching residue with sulfuric acid and then performing solid-liquid separation to obtain a copper-cobalt leaching solution and a gold-containing leaching residue; mixing the copper-cobalt leaching solution with a sodium thiosulfate solution and adjusting the pH to perform a copper precipitation reaction, and then performing solid-liquid separation to obtain a copper slag and a cobalt-containing solution; and leaching the gold-containing leaching residue with a gold leaching agent to obtain a gold-containing leaching solution.

2. The method for preparing battery-grade iron phosphate and recovering gold, copper and cobalt by using sulfuric acid slag according to claim 1, characterized in that: During the leaching process, the concentration of the phosphoric acid solution is 3-8 mol / L, the liquid-solid ratio is 6-15 ml / g, the temperature is 30-100° C., and the time is 2-6 h.

3. The method for preparing battery-grade iron phosphate and recovering gold, copper and cobalt by using sulfuric acid slag according to claim 1, characterized in that: The oxidant includes at least one of hydrogen peroxide, air and oxygen.

4. A method for preparing battery-grade iron phosphate and recovering gold, copper and cobalt using sulfuric acid slag according to claim 1 or 3, characterized in that: The precipitation reaction process is: stirring the reaction at a temperature of 60-80° C. for 1-3 hours.

5. The method for preparing battery-grade iron phosphate and recovering gold, copper and cobalt by using sulfuric acid slag according to claim 1, characterized in that: The calcination conditions are: temperature of 500-700° C. and time of 60-180 min.

6. The method for preparing battery-grade iron phosphate and recovering gold, copper and cobalt by using sulfuric acid slag according to claim 1, characterized in that: The gold leaching agent includes at least one of cyanide, thiosulfate and thiourea.

7. A method for preparing battery-grade iron phosphate and recovering gold, copper and cobalt using sulfuric acid slag according to claim 1 or 6, characterized in that: The gold-containing leaching solution is adsorbed by activated carbon to obtain gold-loaded carbon, the gold-loaded carbon is desorbed by a desorption solution to obtain a gold-rich solution, and the gold-rich solution is electrolytically treated to obtain gold mud.

8. The method for preparing battery-grade iron phosphate and recovering gold, copper and cobalt by using sulfuric acid slag according to claim 7, characterized in that: The desorption liquid is a mixed solution of sodium chloride and sodium hydroxide.

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