Method for recovering valuable metals in sodium electric material

Abstract

The invention provides a method for recovering valuable metals in a sodium electric material, which comprises the following steps: firstly, carrying out primary slurrying treatment on the sodium electric material, and carrying out solid-liquid separation to obtain callback filtrate and callback filter residue; secondly, carrying out oxidation iron removal treatment on the callback filtrate, and carrying out solid-liquid separation to obtain iron-removed slag and a nickel original material; then, the iron-removed residues and the callback filter residues are mixed and then subjected to secondary pulpifying treatment, and pulpified slurry is obtained; then, mixing the pulpified slurry with sulfide, and performing stirring treatment to obtain mixed slurry; and finally, carrying out pressurized acid leaching treatment on the mixed slurry to obtain pressurized slag and pressurized liquid, carrying out secondary countercurrent washing on the pressurized slag, and packaging and warehousing. According to the method, firstly, the callback filtrate is subjected to oxidation iron removal treatment by adding the oxidizing agent, and then the pulpified slurry is subjected to reduction treatment through the sulfide, so that the using amount of the oxidizing agent can be reduced by 90%-95% under the condition that other reducing agents are not used, and the recovery cost of the valuable metal in the sodium electric material is remarkably reduced.

Description

Technical Field

[0001] The invention relates to the technical field of waste recycling, and in particular to a method for recycling valuable metals in sodium electric materials. Background Art

[0002] At present, sodium electric materials have attracted much attention due to their outstanding advantages such as abundant sodium ion resources in the earth's crust, low price, and similar working principles to mainstream lithium-ion batteries (LIBs), making them the most likely substitute for lithium-ion batteries in grid-scale energy storage. Since iron is the main component of sodium electric materials, when recycling the production tailings or waste of sodium electric materials, they are generally first reduced and leached with a reducing agent, and then oxidized and removed with an oxidant. In this process of reduction and oxidation, a large amount of auxiliary materials (acids, alkalis, reducing agents, oxidants) will be consumed, thereby increasing the recycling cost. Summary of the invention

[0003] The purpose of the present invention is to provide a method for recovering valuable metals in sodium electric materials, so as to improve the technical problem of high recovery cost when the prior art adopts a reduction-then-oxidation process to recover sodium electric materials.

[0004] In order to solve the above technical problems, the present invention provides a method for recovering valuable metals in sodium electric materials, comprising: S10, performing a slurry treatment on the sodium electrode material, and obtaining a filtrate and a residue after solid-liquid separation; S20, performing an oxidation deironing treatment on the callback filtrate, and obtaining deironing slag and nickel raw material after solid-liquid separation; S30, mixing the iron removal slag and the adjusted filter residue and performing secondary slurry treatment to obtain a slurry; S40, mixing the slurry and the sulfide and stirring them to obtain a mixed slurry; S50, the mixed slurry is subjected to pressurized acid leaching treatment to obtain pressurized slag and pressurized liquid, and the pressurized slag is packaged and stored after secondary countercurrent washing.

[0005] Preferably, in step S10: the sodium electrode material is subjected to a slurry treatment using a pressurized liquid.

[0006] Preferably, in the primary slurry treatment: the leaching time is 1 to 2 hours, the leaching temperature is 70 to 80°C, the solid-liquid ratio is 1:(4 to 10), and the pH value at the leaching endpoint is 3.5 to 6.5.

[0007] Preferably, in step S20: the pH value of the callback filtrate after the oxidative iron removal treatment is 3.5~6.0; when the oxidant selected for the oxidative iron removal treatment is a solid, the amount of the oxidant is 0.33~2 times the mass content of iron in the callback filtrate; when the oxidant is a gas, the gas flow rate of the oxidant is 5~10L / h; when the oxidant is a liquid, the amount of the oxidant is 10~20 times the mass content of iron in the callback filtrate.

[0008] Preferably, in step S30: the mixture of the iron removal slag and the callback filter residue is subjected to secondary slurry treatment by slag washing water, the slag washing water is obtained by secondary countercurrent washing of the pressurized slag; the solid-liquid ratio of the secondary slurry treatment is 1: (4~10).

[0009] Preferably, in step S40: the sulfide is at least one of nickel sulfide, copper sulfide tailings, nickel-sulfur tailings and high nickel matte; Among them, copper sulfide tailings are waste residues precipitated from manganese chloride solution containing copper, cobalt and nickel with sodium sulfide. The copper sulfide tailings include the following components by mass percentage: copper 8%~12%, cobalt 0.5%~2%, nickel 0.1%~1%, manganese 3%~7%, zinc 1%~4%, sulfur 15%~20%, chlorine 1%~1.5%, and water 40%~55%; nickel-sulfur tailings are waste residues produced by sodium sulfide precipitation of heavy metals in wastewater. The nickel-sulfur tailings include the following components by mass percentage: nickel 7%~10%, cobalt 0.5%~2%, manganese 0.5%~1%, sulfur 5%~8%, chlorine 1%~1.5%, and water 40%~60%.

[0010] Preferably, before performing step S40, the method further comprises: pre-treating the sulfide; Among them, when the sulfide is copper sulfide tailings or nickel-sulfur tailings, the pretreatment includes: storing the sulfide for 10 to 15 days to reduce the moisture in the sulfide to 20% to 30% after natural evaporation, and then pickling the sulfide once with a weak acid with a pH of 4.0 to 6.5, the washing time is 2 to 4 hours, the washing temperature is 50-70°C, and the washing water cycle times are 10 to 15 times; when the sulfide is nickel sulfide or high nickel matte, the pretreatment includes: obtaining a powder with a 300 mesh screening rate greater than 95% through a ball milling process.

[0011] Preferably, in step S40: the mass ratio of the slurry to the sulfide is 1:(1-5).

[0012] Preferably, step S50 specifically includes: S501, adjusting the pH value of the mixed slurry to 1.0-1.5 with concentrated sulfuric acid, the reaction temperature is 70-85° C., and the reaction is continued for 4-6 hours after the pH value of the mixed slurry is stabilized to obtain a prepreg slurry; S502, the prepreg slurry is subjected to a pressurized leaching treatment, the pressurized temperature is 140-180°C, the leaching time is 4-10 hours, and after the reaction is completed, pressurized slag and pressurized liquid are obtained by solid-liquid separation.

[0013] Preferably, the pressurized slag comprises the following components by mass percentage: cobalt < 0.1%, nickel < 0.2%, manganese < 0.1%, and copper less than 0.5%.

[0014] The beneficial effects of the present invention are as follows: Different from the prior art, the present invention provides a method for recovering valuable metals in sodium electric materials, comprising: firstly, subjecting the sodium electric material to a pulping treatment, and obtaining a callback filtrate and a callback filter residue after solid-liquid separation; secondly, subjecting the callback filtrate to an oxidative iron removal treatment, and obtaining an iron removal slag and a nickel raw material after solid-liquid separation; thereafter, mixing the iron removal slag with the callback filter residue and subjecting the mixture to a secondary pulping treatment to obtain a pulped slurry; thereafter, mixing the pulped slurry with a sulfide and subjecting the mixture to a stirring treatment to obtain a mixed slurry; finally, subjecting the mixed slurry to a pressurized acid leaching treatment to obtain a pressurized slag and a pressurized liquid, and packaging the pressurized slag for storage after a secondary countercurrent washing; the present invention firstly subjecting the callback filtrate after the sodium electric material is subjected to a pulping treatment to an oxidative iron removal treatment by adding an oxidant, and then subjecting the pulped slurry formed by mixing the iron removal slag and the callback filter residue to a reduction treatment by a sulfide, thereby being able to reduce the amount of the oxidant by 90% to 95% without using other reducing agents, and significantly reducing the cost of recovering valuable metals in the sodium electric material. BRIEF DESCRIPTION OF THE DRAWINGS

[0015] Figure 1 A flow chart of a method for recovering valuable metals in sodium electric materials provided by an embodiment of the present invention; Figure 2 This is a process flow chart of the method for recovering valuable metals in sodium electrical materials provided in Example 1 of the present invention. DETAILED DESCRIPTION

[0016] The following will be combined with the embodiments of the present invention to clearly and completely describe the technical solutions in the embodiments of the present invention. Obviously, the described embodiments 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 creative work are within the scope of protection of the present invention.

[0017] The purpose of the present invention is to overcome the above technical deficiencies, provide a method for preparing battery-grade manganese sulfate solution from cobalt-nickel intermediates, and solve the technical problem that in the prior art, when conventional acid leaching processes are used to recover cobalt and nickel, the leaching residue contains a large amount of high-valent manganese impurities, resulting in excessively high extraction costs in subsequent extraction processes.

[0018] See also Figure 1 , Figure 1A flow chart of a method for recovering valuable metals in sodium electric materials provided by an embodiment of the present invention; wherein the method for preparing a battery-grade manganese sulfate solution from the above-mentioned cobalt-nickel intermediate specifically comprises: S10, performing a slurry treatment on the sodium electrode material, and obtaining a filtrate and a residue after solid-liquid separation.

[0019] Specifically, step S10 also includes: First, a sodium electric material is provided, wherein the sodium electric material is a mixture of one or more of a sodium ion battery precursor, a sodium ion battery positive electrode material, and a sodium ion battery crushed black powder; wherein the sodium ion battery precursor generally refers to an intermediate product in the process of preparing a positive and negative electrode material of a sodium ion battery; the sodium ion battery positive electrode material is of various types, and the common ones are layered oxides, polyanion compounds, etc., which play a key role in storing and releasing sodium ions during the charging and discharging process; the sodium ion battery crushed black powder may be a powder substance containing sodium ion battery-related components generated during the battery production, recycling or treatment process; After that, the sodium electrode material is slurried by adding a slurrying agent, and the solid-liquid separation is performed by filter pressing to obtain the adjusted filtrate and adjusted filter residue; filter pressing solid-liquid separation is an efficient separation method that can quickly separate the solid and liquid in the slurry. The filter pressing process can control the separation effect and the quality of the filtrate by adjusting parameters such as pressure and the pore size of the filter medium.

[0020] Specifically, the slurrying agent is sulfuric acid or a pressurized liquid produced in a subsequent step. The pressurized liquid produced in a subsequent step can be used as a slurrying agent to realize the recycling of resources and reduce costs.

[0021] Specifically, in the primary slurry treatment: the leaching time is 1 to 2 hours, the leaching temperature is 70 to 80°C, the solid-liquid ratio is 1: (4 to 10), and the pH value at the leaching end point is 3.5 to 6.5; Among them, if the leaching time is less than 1 hour, the components in the material may not be completely leached, affecting the effect of subsequent treatment and product quality; if the leaching time is greater than 2 hours, it may increase unnecessary costs and may cause some undesirable side reactions, affecting the purity of the product.

[0022] Among them, when the leaching temperature is less than 70°C, the reaction speed will slow down, and a longer leaching time may be required to achieve the ideal effect; when the leaching temperature is greater than 80°C, the excessively high temperature will damage the material structure and lead to excessive consumption of energy.

[0023] Among them, when the solid-liquid ratio is greater than 1:4, that is, there are too many solid materials, which may cause uneven slurrying and some materials cannot fully react with the slurrying agent, affecting the leaching effect; when the solid-liquid ratio is less than 1:10, the amount of slurrying agent will increase, increasing the cost and may make subsequent separation treatment more difficult.

[0024] Among them, when the pH value of the leaching end point is 3.5-6.5, it can ensure that the useful components in the sodium electrode material can be stably leached, and avoid the damage to the material structure caused by an overly acidic or alkaline environment.

[0025] Furthermore, when the sodium electrolyte material does not contain aluminum or chromium, the pH value of the leaching endpoint is 3.5~6.0; within this range, the slurrying agent can effectively react with the components in the material to promote the leaching of useful substances.

[0026] Furthermore, when the sodium-based material contains aluminum, the pH value of the leaching endpoint is 4.0-6.0; slightly increasing the pH value of the leaching endpoint can reduce the excessive dissolution of aluminum or adverse reactions with other components. This can not only ensure the smooth leaching of the main components in the sodium-based material, but also control the behavior of aluminum to a certain extent, avoiding its negative impact on the leaching process and product quality.

[0027] Furthermore, when the sodium electrolyte contains aluminum and chromium, the pH value at the leaching endpoint is 4.5~6.0; this is because the presence of chromium will also be affected by the pH value at the leaching endpoint, and complex interactions may occur between aluminum and chromium. By increasing the pH value at the leaching endpoint, the stability of aluminum and chromium, as well as the reaction with other components, can be better balanced. Within this range, the dissolution and precipitation of aluminum and chromium can be effectively controlled to ensure the smooth progress of the leaching process, while improving the purity and quality of the product.

[0028] Specifically, the callback filtrate contains sodium electrolyte material components dissolved in the slurry and some impurities. The callback filter residue is mainly undissolved or undispersed sodium electrolyte material and some insoluble impurities.

[0029] S20, performing oxidation deironing treatment on the callback filtrate, and obtaining deironing slag and nickel raw material after solid-liquid separation.

[0030] Specifically, step S20 also includes: Firstly, an oxidant is added to the callback filtrate to oxidize the divalent iron in the callback filtrate, and the pH value of the callback filtrate after the oxidation and iron removal treatment is 3.5-6.0. After the reaction is completed, the reaction is continued for 20-40 minutes; then, iron removal slag and nickel raw material are obtained after solid-liquid separation, and the nickel raw material is collected and transferred to the downstream workshop; wherein the oxidant is any one of persulfate, chlorate, sodium hypochlorite solution, calcium hypochlorite, air, chlorine, oxygen, manganese dioxide and hydrogen peroxide.

[0031] Specifically, since divalent iron is relatively difficult to remove in the callback filtrate, trivalent iron can be separated from the nickel raw material by forming a precipitate through methods such as adjusting the pH value.

[0032] Specifically, the iron removal slag is mainly composed of ferric iron precipitation, which can be further processed or recycled. The nickel raw material is the product after the iron impurities are removed, and its purity is improved, which can be used in the subsequent sodium ion battery material preparation process.

[0033] Specifically, when the oxidant used in the oxidation iron removal treatment is solid, the amount of the oxidant used is 0.33 to 2 times the mass content of iron in the callback filtrate; when the oxidant is gas, the gas flow rate of the oxidant is 5 to 10 L / h; when the oxidant is liquid, the amount of the oxidant used is 10 to 20 times the mass content of iron in the callback filtrate.

[0034] S30, mixing the iron removal slag and the adjusted filter residue and performing secondary slurry treatment to obtain slurry.

[0035] Specifically, step S30 also includes: First, pure water or slag washing water produced in subsequent steps is used to slurry the filtered slag and iron removal slag to obtain slurry; the slag washing water is obtained by secondary countercurrent washing of pressurized slag, and the solid-liquid ratio of the secondary slurry treatment is 1: (4~10); wherein, the slag washing water is obtained by secondary countercurrent washing of pressurized slag. The utilization of this slag washing water realizes the recycling of resources, reduces the dependence on pure water, and reduces production costs. At the same time, secondary countercurrent washing can improve the quality of slag washing water, remove impurities therein, and make it more suitable for slurrying process.

[0036] S40, mixing the slurry and the sulfide and stirring them to obtain a mixed slurry.

[0037] Specifically, step S40 also includes: First, a sulfide is provided, wherein the sulfide is at least one of nickel sulfide, copper sulfide tailings, nickel-sulfur tailings and high nickel matte; wherein the copper sulfide tailings are waste residues precipitated from a manganese chloride solution containing copper, cobalt and nickel using sodium sulfide, and the copper sulfide tailings include the following components by mass percentage: copper 8%~12%, cobalt 0.5%~2%, nickel 0.1%~1%, manganese 3%~7%, zinc 1%~4%, sulfur 15%~20%, chlorine 1%~1.5%, and water 40%~55%; the nickel-sulfur tailings are waste residues produced by sodium sulfide precipitation of heavy metals in wastewater, and the nickel-sulfur tailings include the following components by mass percentage: nickel 7%~10%, cobalt 0.5%~2%, manganese 0.5%~1%, sulfur 5%~8%, chlorine 1%~1.5%, and water 40%~60%.

[0038] Afterwards, the sulfide is pretreated; wherein, when the sulfide is copper sulfide tailings or nickel sulfide tailings, the pretreatment includes: storing the sulfide for 10 to 15 days to reduce the moisture in the sulfide to 20% to 30% after natural evaporation, and then pickling the sulfide once with a weak acid with a pH of 4.0 to 6.5, the washing time is 2 to 4 hours, the washing temperature is 50 to 70°C, and the washing water is cycled 10 to 15 times; wherein, the weak acid with a pH of 4.0 to 6.5 can remove impurities and oxides on the surface of the sulfide without causing serious damage to the main structure of the sulfide.

[0039] Specifically, when the sulfide is nickel sulfide or nickel matte, the pretreatment includes: obtaining a powder with a 300-mesh screening rate greater than 95% through a ball milling process; wherein, the ball milling can break the nickel sulfide or nickel matte into fine particles, increase its surface area, and facilitate subsequent chemical reactions and treatment processes. The requirement of a 300-mesh screening rate greater than 95% ensures the fineness of the powder, allowing the powder to participate in the reaction more evenly in subsequent treatments, thereby improving the reaction rate and product quality.

[0040] Finally, the pulped slurry and the sulfide are mixed and stirred to obtain a mixed slurry; wherein the mass ratio of the pulped slurry to the sulfide is 1: (1-5); wherein some metal ions in the pulped slurry, such as copper ions and lead ions, can form insoluble metal sulfide precipitates with sulfur ions in the sulfide, and through this reaction, these metal ions can be effectively removed to improve the purity of the pulped slurry; at the same time, the sulfur element in the sulfide has a certain reducing property, and can react with the oxidizing substances in the pulped slurry to change the redox state of the system and create favorable conditions for subsequent treatment processes.

[0041] Specifically, when the mass ratio of slurry to sulfide is greater than 1:1, too little sulfide will make it difficult to remove some impurity metal ions in the slurry, such as copper ions and lead ions; when the mass ratio of slurry to sulfide is less than 1:5, too much sulfide may require a larger capacity reaction vessel and more powerful stirring equipment to ensure that the materials can be fully mixed.

[0042] S50, the mixed slurry is subjected to pressurized acid leaching treatment to obtain pressurized slag and pressurized liquid, and the pressurized slag is packaged and stored after secondary countercurrent washing.

[0043] Specifically, step S50 also includes: S501, adjusting the pH value of the mixed slurry to 1.0-1.5 with concentrated sulfuric acid, the reaction temperature is 70-85° C., and the reaction is continued for 4-6 hours after the pH value of the mixed slurry is stabilized to obtain a prepreg slurry; S502, the prepreg slurry is moved to the autoclave for pressure leaching treatment, the pressure temperature is 140~180℃, the leaching time is 4~10h, after the reaction is completed, the pressure slag and the pressure liquid are obtained by solid-liquid separation, the pressure slag is washed twice in countercurrent and then packed and stored, the washing water after the pressure slag is washed twice in countercurrent is reused in step S30 for secondary slurry treatment; the pressure liquid is reused in step S10 as a slurrying agent for primary slurry treatment of the sodium electrode material.

[0044] Specifically, the secondary countercurrent washing can effectively remove impurities in the slag and improve the purity of the slag. The slag washing water after washing is recycled to step S30 for secondary slurry treatment, which realizes the recycling of water resources and reduces production costs; the pressurized liquid is recycled to step S10 as a slurrying agent for primary slurrying of sodium electric materials, which also realizes the recycling of resources, reduces the demand for fresh slurrying agents, reduces production costs, and also reduces waste emissions.

[0045] Specifically, the pressurized slag includes the following components by mass percentage: cobalt <0.1%, nickel <0.2%, manganese <0.1%, and copper less than 0.5%.

[0046] The technical solution of the present application is now further described in detail in conjunction with specific embodiments.

[0047] Embodiment 1: See also Figure 2 , Figure 2 The process flow chart of the method for recovering valuable metals in sodium electric materials provided in Example 1 of the present invention is as follows: Step 1, taking an appropriate amount of copper sulfide tailings, and placing the material in a cool and ventilated place for 10 days to obtain the copper sulfide tailings after standing; washing the material once with a weak acid, adjusting the pH of the solution to be stable between 4.5 and 6.5 with sulfuric acid during the washing process, washing for 2 hours, the washing temperature is 65°C, and solid-liquid separation is performed to obtain weak acid washing water and pretreated copper sulfide tailings; Step 2: Take an appropriate amount of sodium electric material so that the mass ratio of the sodium electric material to the copper sulfide tailings after standing is 2.5:1, take the pressurized liquid slurry sodium electric material so that the solid-liquid ratio is 1:6, control the leaching end point pH to be between 5.0 and 6.0, the leaching time is 2 hours, the leaching temperature is 75°C, and the solid-liquid separation is performed to obtain the callback residue and the callback filtrate; Step 3, titrate the divalent iron in the callback filtrate, add 1.5 times the mass content of manganese dioxide according to the total iron content in the callback filtrate, react for 30 minutes, control the pH of the solution between 3.5 and 6.0 with liquid alkali, separate the solid and liquid to obtain iron-removing slag and nickel raw material, and transfer the nickel raw material to the downstream workshop; Step 4: Mix the iron removal slag and the adjusted filter residue and slurry them with slag washing water (if insufficient, pure water is used to supplement), with a solid-liquid ratio of 1:6 to obtain slurry; Step 5, adding the pretreated copper sulfide tailings into the slurry to obtain a mixed slurry; Step 6: Use acid to adjust the pH of the mixed slurry to 1.0-1.5, the reaction temperature is 70-85°C, and the reaction is continued for 4 hours after the pH is stable to obtain a prepreg slurry; Step 7: Pump (move) the prepreg slurry to the autoclave, the pressurization temperature is 170°C, and the leaching time is 6 hours. After the reaction is completed, the solid-liquid separation is performed to obtain pressurized slag and pressurized liquid, and the pressurized liquid is reused; Step eight, the pressurized slag is packaged and stored after secondary countercurrent washing, and the slag washing water is returned to the slurrying process as bottom water; the filter residue after washing contains 0.1834% nickel, 0.1471% cobalt, 0.3478% copper and 0.1674% manganese; the recovery rate of nickel, cobalt, manganese and copper reaches 99.5%.

[0048] The cobalt, manganese and copper recovery rate of the method for recovering valuable metals in sodium electric materials provided in Example 1 of the present invention is 99%, which reduces the amount of alkali used by 95% and the amount of acid used by 10%~15%; other reducing agents are no longer used in the above-mentioned recovery process, and the amount of oxidant used is reduced by 90%~95%, which significantly reduces the recovery cost.

[0049] In summary, the present invention provides a method for recovering valuable metals in sodium electric materials, comprising: first, performing a slurry treatment on the sodium electric material, and obtaining a callback filtrate and a callback filter residue after solid-liquid separation; secondly, performing an oxidative iron removal treatment on the callback filtrate, and obtaining iron removal slag and nickel original material after solid-liquid separation; thereafter, mixing the iron removal slag with the callback filter residue and performing a secondary slurry treatment to obtain a slurry; thereafter, mixing the slurry with a sulfide and performing a stirring treatment to obtain a mixed slurry; finally, performing a pressurized acid leaching treatment on the mixed slurry to obtain a pressurized slag and a pressurized liquid, and the pressurized slag is packaged and stored after secondary countercurrent washing; the present invention first performs an oxidative iron removal treatment on the callback filtrate after the sodium electric material is subjected to a primary slurry treatment by adding an oxidant, and then performs a reduction treatment on the slurry formed by mixing the iron removal slag and the callback filter residue by using a sulfide, thereby reducing the amount of the oxidant by 90% to 95% without using other reducing agents, and significantly reducing the recovery cost of the valuable metals in the sodium electric material.

[0050] It should be noted that the above embodiments all belong to the same inventive concept, and the description of each embodiment has its own focus. For matters that are not described in detail in some embodiments, reference may be made to the description in other embodiments.

[0051] The above embodiments only express the implementation methods of the present invention, and the descriptions thereof are relatively specific and detailed, but they cannot be understood as limiting the scope of the invention patent. It should be pointed out that, for those of ordinary skill in the art, several variations and improvements can be made without departing from the concept of the present invention, and these all belong to the protection scope of the present invention. Therefore, the protection scope of the patent of the present invention shall be subject to the attached claims.

Claims

1. A method for recovering valuable metals in sodium electric materials, characterized in that: include: S10, performing a slurry treatment on the sodium electrode material, and obtaining a filtrate and a residue after solid-liquid separation; S20, performing an oxidative iron removal treatment on the callback filtrate, and obtaining iron removal slag and nickel raw material after solid-liquid separation; S30, mixing the iron removal slag and the callback filter residue and performing secondary slurry treatment to obtain slurry; S40, mixing the slurry and the sulfide and stirring to obtain a mixed slurry; S50, the mixed slurry is subjected to pressurized acid leaching treatment to obtain pressurized slag and pressurized liquid, and the pressurized slag is packaged and stored after secondary countercurrent washing.

2. The method for recovering valuable metals in sodium electric materials according to claim 1, characterized in that: In the step S10: the sodium electrode material is subjected to the primary slurrying treatment using the pressurized liquid.

3. The method for recovering valuable metals in sodium electric materials according to claim 2, characterized in that: In the primary slurry treatment, the leaching time is 1 to 2 hours, the leaching temperature is 70 to 80° C., the solid-liquid ratio is 1:(4 to 10), and the pH value at the leaching endpoint is 3.5 to 6.

5.

4. The method for recovering valuable metals in sodium electric materials according to claim 1, characterized in that: In the step S20: the pH value of the callback filtrate after the oxidation and iron removal treatment is 3.5-6.0; when the oxidant selected for the oxidation and iron removal treatment is a solid, the amount of the oxidant is 0.33-2 times the mass content of iron in the callback filtrate; when the oxidant is a gas, the gas flow rate of the oxidant is 5-10 L / h; when the oxidant is a liquid, the amount of the oxidant is 10-20 times the mass content of iron in the callback filtrate.

5. The method for recovering valuable metals in sodium electric materials according to claim 1, characterized in that: In the step S30: the mixture of the iron removal slag and the callback filter residue is subjected to the secondary slurry treatment by slag washing water, the slag washing water is obtained by the pressurized slag after the secondary countercurrent washing; the solid-liquid ratio of the secondary slurry treatment is 1: (4-10).

6. The method for recovering valuable metals in sodium electric materials according to claim 1, characterized in that: In the step S40: the sulfide is at least one of nickel sulfide, copper sulfide tailings, nickel sulfur tailings and high nickel matte; The copper sulfide tailings are waste residues precipitated from a manganese chloride solution containing copper, cobalt and nickel using sodium sulfide, and the copper sulfide tailings include the following components by mass percentage: 8% to 12% copper, 0.5% to 2% cobalt, 0.1% to 1% nickel, 3% to 7% manganese, 1% to 4% zinc, 15% to 20% sulfur, 1% to 1.5% chlorine, and 40% to 55% water; the nickel-sulfur tailings are waste residues produced by sodium sulfide precipitation of heavy metals in wastewater, and the nickel-sulfur tailings include the following components by mass percentage: 7% to 10% nickel, 0.5% to 2% cobalt, 0.5% to 1% manganese, 5% to 8% sulfur, 1% to 1.5% chlorine, and 40% to 60% water.

7. The method for recovering valuable metals in sodium electric materials according to claim 6, characterized in that: Before performing the step S40, the method further includes: pre-treating the sulfide; Among them, when the sulfide is the copper sulfide tailings or the nickel-sulfur tailings, the pretreatment includes: storing the sulfide for 10 to 15 days to reduce the moisture in the sulfide to 20% to 30% after natural evaporation, and then pickling the sulfide once with a weak acid with a pH of 4.0 to 6.5, the washing time is 2 to 4 hours, the washing temperature is 50-70°C, and the number of washing water cycles is 10 to 15 times; when the sulfide is nickel sulfide or high nickel matte, the pretreatment includes: obtaining a powder with a 300 mesh screening rate greater than 95% through a ball milling process.

8. The method for recovering valuable metals in sodium electric materials according to claim 1, characterized in that: In the step S40: the mass ratio of the slurry to the sulfide is 1:(1-5).

9. The method for recovering valuable metals in sodium electric materials according to claim 1, characterized in that: The step S50 specifically includes: S501, adjusting the pH value of the mixed slurry to 1.0-1.5 with concentrated sulfuric acid, the reaction temperature is 70-85° C., and after the pH value of the mixed slurry is stabilized, the reaction is continued for 4-6 hours to obtain a prepreg slurry; S502, performing a pressurized leaching treatment on the prepreg slurry, the pressurized temperature is 140-180°C, the leaching time is 4-10 hours, and after the reaction is completed, the pressurized slag and the pressurized liquid are obtained by solid-liquid separation.

10. The method for recovering valuable metals in sodium electric materials according to claim 9, characterized in that: The pressurized slag comprises the following components by mass percentage: cobalt <0.1%, nickel <0.2%, manganese <0.1%, and copper less than 0.5%.

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