Method for separating and recovering phosphorus and iron in phosphorus-iron slag as a byproduct of battery recycling

Abstract

The present application relates to the technical field of battery recycling, in particular to a method for separating and recovering phosphorus and iron in by-product phosphorus-iron slag in battery recycling, comprising the following steps: step 1, high-temperature reduction: high-temperature reduction of the phosphorus-iron slag; step 2, acid leaching separation: acid leaching of the reduced phosphorus-iron slag with sulfuric acid to obtain graphite through filtration; step 3, crystallization separation: concentration, crystallization, centrifugal separation and drying of the filtrate in step 2 to obtain ferrous sulfate heptahydrate; step 4, extraction: acidification of the centrifugal mother liquor and then extraction, washing of the organic phase after extraction to remove sulfate radicals; step 5, ammonia neutralization: adding pure water and ammonia gas to the washed organic phase for neutralization, and the water phase separated after neutralization is supersaturated after multiple cycles to precipitate crystals, which are centrifuged, dried and then industrial-grade monoammonium phosphate is obtained. The present application solves the technical problem that the product purity and the yield of phosphorus and iron are difficult to be compatible in the existing phosphorus-iron slag recovery technology.

Description

Technical Field

[0001] This invention relates to the field of battery recycling technology, specifically to a method for separating and recovering phosphorus and iron from phosphorus and iron slag, a byproduct of battery recycling. Background Technology

[0002] Currently, to achieve large-scale recycling of retired lithium iron phosphate (LFP) battery cells, the industry typically employs a "full crushing" process for both positive and negative electrodes. Then, copper and aluminum are separated using methods such as air separation, gravity separation, color sorting, and sieving to obtain lithium iron phosphate black powder (graphite content 25%–40%). This black powder undergoes wet selective lithium extraction to obtain lithium carbonate or lithium phosphate recycling materials. The residue after lithium extraction from the black powder is iron phosphate slag, which typically contains 0.15%–0.5% aluminum, 0.2%–0.6% copper, 25%–40% graphite, and 45%–60% iron phosphate, representing a rich mineral resource. Recycling this iron phosphate slag would significantly conserve non-renewable mineral resources.

[0003] To recover ferrophosphate from ferrophosphate slag, most existing technologies directly regenerate it into ferric phosphate for recovery. For example, patent CN114804048A discloses a method for extracting and separating iron and phosphorus from ferric phosphate slag to prepare ferric phosphate. This method uses inorganic acid to directly leach the ferrophosphate slag, then extracts iron and phosphorus separately through an extraction process, and then reacts the purified iron-rich solution and phosphorus-rich solution to prepare battery-grade ferric phosphate. This method recovers graphite with low purity (≤90%), requires a high number of extraction stages (5-20 stages), has a long process, high production costs, and the preparation of the extractant is complex, making recovery difficult and limiting its industrial application prospects. For example, patent CN 119218951A discloses a method for regenerating battery-grade iron phosphate from lithium-extracting iron phosphate slag. The iron phosphate slag is regenerated into iron phosphate after wet reduction leaching and removal of copper and aluminum. The above method uses hydrazine hydrate for reduction, which has a large dosage and high cost (80% content, 16,000 yuan / ton). Fluorides are used for aluminum removal, and fluoride ions will affect subsequent water treatment (corrosion of film materials), so the technical and economic feasibility is not high.

[0004] As can be seen from the two technologies mentioned above, although the existing technology for recycling iron phosphate from iron phosphate slag can recover and utilize the iron and phosphorus elements in the iron phosphate slag, aluminum is difficult to remove during the recycling process because its chemical properties are similar to those of iron. If it is not removed, the aluminum impurity content in the recycled iron phosphate material will be too high, resulting in low quality (the aluminum content of iron phosphate products is generally controlled below 80 ppm), and low economic value. If it is removed, it will affect the yield of iron phosphate. Experimental data shows that the phosphorus yield in the iron phosphate recycling process after removing aluminum from iron phosphate slag raw materials will be below 30% or even lower. This puts the technical solution of directly regenerating iron phosphate from iron phosphate slag into a dilemma. Therefore, some processes use a graded recycling method to recover phosphorus and iron elements from phosphorus iron slag. For example, patent publication number CN 119725837 A discloses a graded recycling method for lithium, iron, phosphorus and graphite in waste lithium iron phosphate batteries. The phosphorus iron slag after lithium extraction is ball-milled, and then acid-leached. The leachate after acid leaching is directly added to sodium hydroxide to prepare iron hydroxide, and the remaining phosphorus-containing solution is further added to sodium hydroxide solution to prepare sodium phosphate.

[0005] Although the above technology can be used to grade and recycle ferric phosphate slag, the above method does not have a copper and aluminum removal process, the purity of the recovered ferric hydroxide is not high, and ferric phosphate will also precipitate at the same time. In addition, ferric hydroxide has colloidal properties, making it difficult to separate, and the purity of the prepared sodium phosphate is also not high. Summary of the Invention

[0006] The purpose of this invention is to provide a method for separating and recovering phosphorus and iron from phosphorus and iron slag, a byproduct of battery recycling, in order to solve the technical problems mentioned above in the prior art of phosphorus and iron slag regeneration iron phosphate technology where it is difficult to achieve both product purity and phosphorus and iron recovery rates, and in the phosphorus and iron slag graded recycling process where the recovered material has low purity and is difficult to separate.

[0007] To solve the above problems, the technical solution adopted by the present invention is as follows: A method for separating and recovering phosphorus and iron from phosphorus and iron slag, a by-product of battery recycling, comprising the following steps:

[0008] Step 1, High-temperature reduction: The by-product phosphorus iron slag from the wet recovery of lithium iron phosphate black powder is calcined at high temperature under nitrogen protection to obtain solid powder.

[0009] Step 2, acid leaching separation: The solid powder from Step 1 is leached with dilute sulfuric acid. After the acid leaching is completed, the acid leaching solution is filtered to obtain graphite solid and filtrate.

[0010] Step 3, Crystallization Separation: The filtrate from Step 2 is concentrated under vacuum until fine crystals begin to appear. Then, it is cooled and crystallized. The cooled crystallized slurry is centrifuged to obtain the mother liquor and solid phase. The solid phase is dried under vacuum to obtain ferrous sulfate heptahydrate.

[0011] Step 4, Secondary Countercurrent Extraction: Add sulfuric acid to the centrifuged mother liquor from Step 3 for acidification. After acidification, perform secondary countercurrent extraction. After extraction, an organic phase and an aqueous phase are obtained. The organic phase is washed with monoammonium phosphate solution to remove sulfate impurities. The aqueous phase is concentrated and dried to obtain cement filler.

[0012] Step 5, ammonia neutralization: Ammonia gas is introduced into the washed organic phase for neutralization. After the reaction reaches the endpoint, the organic phase and aqueous phase are separated. The organic phase is reused as the extractant for secondary countercurrent extraction, and the aqueous phase is dried to obtain industrial-grade monoammonium phosphate.

[0013] Furthermore, in step 3, the filtrate is concentrated by vacuum concentration, with a vacuum degree of -0.07 to -0.1 MPa and a temperature of 65 to 75°C; the cooling crystallization temperature is -15 to 0°C.

[0014] Furthermore, nitrogen gas is introduced for protection during the acid leaching process in step 2.

[0015] Furthermore, in step 3, the vacuum drying temperature of ferrous sulfate heptahydrate is 75–90°C, and the drying time is 1.5–3 hours.

[0016] Furthermore, in step 4, the concentration of monoammonium phosphate used for organic phase washing is 20% to 31%, and the aqueous phase after washing and separation is concentrated by evaporation to obtain ammonium sulfate.

[0017] Furthermore, the extraction in step 4 adopts a two-stage countercurrent extraction, and the extractant used is a composite extractant composed of tributyl phosphate, methyl isobutyl ketone and kerosene in a mass ratio of 5-7:1-4:0-2.

[0018] Furthermore, in step 4, the acidified liquid is added to the secondary extractor, and the extractant is added to the primary extractor.

[0019] Furthermore, the aqueous phase after the secondary countercurrent extraction in step 4 is mainly composed of a mixture of elements such as iron, phosphorus, sulfur, copper, and aluminum, and can be used as a cement retarder after drying.

[0020] Furthermore, in step 5, when the aqueous phase does not reach a supersaturated state, it is returned to the washing process of the secondary extraction organic phase. After multiple cycles, it reaches supersaturation and crystals precipitate before entering the centrifugation process. The centrifuged mother liquor is returned to step 4 to wash the sulfate ions in the fresh organic phase.

[0021] The beneficial effects of this implementation plan are as follows:

[0022] 1. The main components of the by-product iron phosphate slag from the wet recovery of lithium iron phosphate battery black powder are iron phosphate (45%-60%) and graphite (25%-40%), as well as high levels of aluminum (0.15%-0.5%) and copper (0.2%-0.6%) impurities. Other impurities include calcium, magnesium, manganese, titanium, potassium, sodium, etc. (all ≤1000ppm). The most conceivable way to utilize the by-product iron phosphate slag is through iron phosphate regeneration. However, the biggest challenge restricting the direct regeneration of iron phosphate slag is the removal of aluminum impurities. This is because iron and aluminum have similar chemical properties. The conventional method is to acid leach the iron phosphate slag and then add a reducing agent (such as iron powder) to remove the Fe. 3+ Reduced to Fe 2+ Then, by increasing the pH value to precipitate aluminum, this process causes phosphorus loss, especially when the aluminum content is reduced from ≥3000ppm to ≤30ppm in the acid leaching solution (this range ensures that the aluminum content in the prepared ferric phosphate is ≤80ppm), the phosphorus loss rate can reach over 70%. This invention further returns phosphorus and iron to the "raw material end," separating them from the ferric phosphate slag. Iron is recovered using high-purity ferrous sulfate heptahydrate, and phosphorus is recovered using industrial-grade monoammonium phosphate. Both can then be resynthesized into battery-grade ferric phosphate, or applied independently in other fields, solving the problem of large-scale, value-added utilization of carbon, iron, and phosphorus in ferric phosphate slag in the industry.

[0023] 2. Existing separation and recovery technologies involve direct acid leaching of ferric phosphate slag, followed by the preparation of different substances. However, the lack of copper and aluminum removal processes, as well as the absence of sulfate removal, results in low purity of the recovered materials. Furthermore, the generated ferric hydroxide exhibits colloidal properties, making separation difficult. This invention, however, first utilizes the graphite inherent in the ferric phosphate slag to reduce the leaching rate (from 92% to over 98.5%), thereby increasing graphite purity. Then, through cooling and crystallization, iron is preferentially separated as ferrous sulfate to prepare high-purity ferrous sulfate. Finally, the filtrate is extracted to recover phosphorus. The organic phase after extraction still contains a certain amount of SO4. 2- Impurities can affect the final purity of monoammonium phosphate. This invention utilizes the high selectivity of the extractant for phosphoric acid to remove SO4 by adding monoammonium phosphate to the organic phase. 2- The principle is that PO4 in monoammonium phosphate 3- With H + SO4 combines with and enters the organic phase 2- With NH4 + SO4 is introduced into the aqueous phase and realized through the "oil-water" interface. 2- The invention achieves highly efficient removal of pollutants. Compared with existing separation and recovery technologies, the purity and yield of various products obtained by this invention are both at a higher level.

[0024] 3. Due to the diverse types and compositions of retired batteries, the phosphorus-iron slag raw material obtained from battery recycling is unstable. Simple direct acid leaching for separation and recovery results in significant fluctuations in the impurity content of the leachate due to the variability of raw materials. Furthermore, the impurity removal processes used for direct regeneration of ferric phosphate generally have poor applicability. This invention first utilizes the graphite contained in the phosphorus-iron slag to reduce ferric phosphate. Then, through multiple staged extraction processes including cooling crystallization, extraction recovery, and washing purification, the recovered ferrous sulfate and industrial-grade monoammonium phosphate are of high purity. This achieves comprehensive utilization of carbon, iron, and phosphorus in the phosphorus-iron slag byproduct of battery recycling, and the adopted technical method has high applicability to phosphorus-iron slag raw materials.

[0025] 4. This invention reduces the by-product phosphorus-iron slag at high temperature, so that the iron exists in the form of divalent iron. During acid leaching, the leaching rates of phosphorus and iron are both above 98.5%, the leaching conditions are mild, and the purity of graphite after acid leaching is above 95%.

[0026] 5. In addition to cooling and crystallizing to recover iron, the present invention also adds sulfuric acid to the filtrate for acidification, which increases the content of "H3PO4" in the solution and can greatly improve the extraction rate of phosphorus. Attached Figure Description

[0027] Figure 1 This is a process flow diagram of the present invention.

[0028] Figure 2 This is the XRD phase diagram of the phosphorus-iron slag after high-temperature reduction according to the present invention. Detailed Implementation

[0029] The following detailed description illustrates the specific implementation method:

[0030] A method for separating and recovering phosphorus and iron from by-product phosphorus and iron slag in battery recycling, wherein the by-product phosphorus and iron slag has a water content of 45% to 50%, and the composition of the raw materials after crushing and drying is shown in Table 1.

[0031] Table 1. Main chemical composition of phosphorus-iron slag, a byproduct of battery recycling (unit: %)

[0032] content 20.52 12.37 0.1802 0.0316 0.0966 0.1837 0.1579 0.0457 0.2156 name Ti Ca Ni Zn Co Sn As Pb C content 0.0867 0.0872 0.0332 0.017 0.0016 0.0061 0.0102 0.0005 35.45

[0033] Example 1

[0034] 200g of ferrophosphate slag (dry pulverized material) was placed in a high-temperature tube furnace and subjected to high-temperature reduction calcination in a nitrogen atmosphere. The heating rate was 10℃ / min, the final temperature was 600℃, and the holding time was 3 hours. After cooling, 190.6g of solid powder was obtained (mainly ferrous pyrophosphate, with an iron reduction rate of 98.72%).

[0035] The mixture was transferred to a glass-jacketed reactor and 597g of 15% dilute sulfuric acid was added for acid leaching at 65℃. Nitrogen gas was introduced as a protective gas. After reacting for 2 hours, the mixture was filtered. The filter cake was then subjected to a two-stage countercurrent washing process (the washing liquid was used to prepare the dilute sulfuric acid). After drying, the filter cake weighed 69.27g, with a graphite purity of 95.87% and a graphite yield of 93.18%. The main chemical composition of the graphite is shown in Table 2.

[0036] The above filtrate was concentrated under vacuum at a vacuum of -0.095 MPa and a temperature of 65°C until it reached a near-saturated solution state (a crystal film appeared). Then, cooling and crystallization were initiated (using a jacketed glass crystallizer with 45% polyethylene glycol as the refrigerant). The final crystallization temperature was -10°C. After centrifuging the crystal slurry, the ferrous sulfate crystals were transferred to a vacuum oven for drying at 80°C for 2.5 hours to obtain 170.23 g of ferrous sulfate heptahydrate with a purity of 98.23% and an iron yield of 83.55%. The main chemical composition of ferrous sulfate heptahydrate is shown in Table 3.

[0037] The filtrate after centrifugation was acidified with 31.8g of 98% industrial-grade concentrated sulfuric acid. The acidified solution was then subjected to a two-stage countercurrent extraction. The extractant used was a mixture of tributyl phosphate (TBP), methyl isobutyl ketone (MIBK), and kerosene in a mass ratio of 7:2.5:0.5, with a total extractant amount of 2725g. The extraction temperature was 50℃, and the extraction time was 20 minutes. After standing and separation, the liquid was separated to obtain 2780.48g of organic phase and 523.39g of raffinate. The raffinate mainly consisted of a mixture of elements such as iron, phosphorus, sulfur, copper, and aluminum. After evaporation and concentration, it could be added to cement fillers to prepare cement, where it could act as a "retarder."

[0038] 102.5 g of a 30% monoammonium phosphate solution was added to the organic phase to wash the sulfate ions at a washing temperature of 50°C. After separation, the aqueous phase was evaporated, concentrated, and dried to obtain 17.29 g of ammonium sulfate.

[0039] After separation, an appropriate amount of water was added to the organic phase, and ammonia gas was introduced for neutralization. The final pH was 4.85. After standing and separating the layers, the organic phase was returned to the extraction system. The aqueous phase was mixed with the fresh organic phase after secondary extraction, and the above washing, sulfate and ammonia neutralization processes were repeated. After multiple cycles, a steady state was reached, and a fixed amount of monoammonium phosphate crystals of 90±1.0g were precipitated. The main chemical composition of monoammonium phosphate is shown in Table 4.

[0040] Example 2

[0041] 200g of ferrophosphate slag was placed in a high-temperature tube furnace and subjected to high-temperature reduction calcination in a nitrogen atmosphere. The heating rate was 10℃ / min, the final temperature was 850℃, and the holding time was 1 hour. After cooling, 187.2g of solid powder was obtained (iron reduction rate 99.72%).

[0042] The mixture was transferred to a glass-jacketed reactor and 590g of 30% dilute sulfuric acid was added for acid leaching at 45℃. Nitrogen gas was introduced as a protective gas. After reacting for 3 hours, the mixture was filtered. The filter cake was then subjected to a two-stage countercurrent washing process (the washing liquid was used to prepare the dilute sulfuric acid). After drying, the filter cake weighed 68.54g, with a graphite purity of 96.79% and a graphite yield of 91.09%. The main chemical composition of the graphite is shown in Table 2.

[0043] The above filtrate was concentrated under vacuum at a vacuum of -0.085 MPa and a temperature of 70°C until it reached a near-saturated solution. Then, it was cooled and crystallized (using 45% polyethylene glycol as the refrigerant). The final crystallization temperature was -6°C. After centrifuging the crystal slurry, the ferrous sulfate crystals were transferred to a vacuum oven and dried at 75°C for 3 hours to obtain 167.32 g of ferrous sulfate heptahydrate with a purity of 98.67% and an iron yield of 82.13%. The main chemical composition of ferrous sulfate heptahydrate is shown in Table 3.

[0044] 30.1g of 98% industrial grade concentrated sulfuric acid was added to the filtrate after centrifugation for acidification. The acidified solution was then subjected to two-stage countercurrent extraction. The extractant used was a mixture of tributyl phosphate (TBP), methyl isobutyl ketone (MIBK), and kerosene in a mass ratio of 6:3:1. The amount of extractant added was 2700g, the extraction temperature was 50℃, and the extraction time was 20 minutes. After standing and separating the layers, 2754.36g of organic phase and 526.02g of raffinate were obtained.

[0045] 107.8 g of a 28% monoammonium phosphate solution was added to the organic phase to wash the sulfate ions at 50°C. After separation, the aqueous phase was evaporated, concentrated, and dried to obtain 16.78 g of ammonium sulfate. An appropriate amount of water was added to the separated organic phase, and ammonia gas was introduced for neutralization. The final pH was 5.20. After standing and separating the layers, the organic phase was returned to the extraction system. The aqueous phase was mixed with the fresh organic phase from the secondary extraction, and the above process of washing sulfate ions and neutralizing with ammonia was repeated. After multiple cycles, a steady state was reached, and a fixed amount of monoammonium phosphate crystals (90.5 ± 1.0 g) precipitated. The main chemical composition of monoammonium phosphate is shown in Table 4.

[0046] Example 3

[0047] 200g of ferrophosphate slag was placed in a high-temperature tube furnace and subjected to high-temperature reduction calcination in a nitrogen atmosphere. The heating rate was 10℃ / min, the final temperature was 700℃, and the holding time was 2 hours. After cooling, 189.2g of solid powder was obtained (iron reduction rate 99.32%).

[0048] The mixture was transferred to a glass-jacketed reactor and 595g of 20% dilute sulfuric acid was added for acid leaching at 75℃. Nitrogen gas was introduced as a protective gas. After reacting for 1.5 hours, the mixture was filtered. The filter cake was then subjected to a two-stage countercurrent washing process (the washing liquid was used to prepare the dilute sulfuric acid). After drying, the filter cake weighed 69.04g, with a graphite purity of 96.23% and a graphite yield of 92.37%. The main chemical composition of the graphite is shown in Table 2.

[0049] The above filtrate was concentrated under vacuum at a vacuum of -0.090 MPa and a temperature of 75°C until it reached a near-saturated solution. Then, it was cooled and crystallized (using 45% polyethylene glycol as the refrigerant). The final crystallization temperature was -2°C. After centrifuging the crystal slurry, the ferrous sulfate crystals were transferred to a vacuum oven and dried at 85°C for 2 hours to obtain 162.35 g of ferrous sulfate heptahydrate with a purity of 98.96% and an iron yield of 79.69%. The main chemical composition of ferrous sulfate heptahydrate is shown in Table 3.

[0050] 30.5g of 98% industrial grade concentrated sulfuric acid was added to the filtrate after centrifugation for acidification. The acidified solution was then subjected to two-stage countercurrent extraction. The extractant used was a mixture of tributyl phosphate (TBP) and methyl isobutyl ketone (MIBK) in a mass ratio of 6.5:3.5. The amount of extractant added was 2700g, the extraction temperature was 50℃, and the extraction time was 20 minutes. After standing and separating the layers, 2753.31g of organic phase and 522.13g of raffinate were obtained.

[0051] 129.6 g of a 23% monoammonium phosphate solution was added to the organic phase to wash the sulfate ions at 50°C. After separation, the aqueous phase was evaporated, concentrated, and dried to obtain 16.37 g of ammonium sulfate. An appropriate amount of water was added to the separated organic phase, and ammonia gas was introduced for neutralization. The final pH was 4.65. After standing and separating the layers, the organic phase was returned to the extraction system. The aqueous phase was mixed with the fresh organic phase from the secondary extraction, and the above process of washing sulfate ions and neutralizing with ammonia was repeated. After multiple cycles, a steady state was reached, and a fixed amount of monoammonium phosphate crystals (88.5 ± 1.0 g) precipitated. The main chemical composition of monoammonium phosphate is shown in Table 4.

[0052] The specific data tables for Examples 1-3 are as follows:

[0053] Table 2 Main Chemical Composition of Graphite Products in Examples 1-3

[0054] C(％) 95.87 96.79 96.23 Fe (%) 0.67 0.59 0.68 P(％) 0.36 0.31 0.36 Cu (ppm) 235.82 236.71 243.26 Al (ppm) 78.43 72.89 88.25 Ti (ppm) 165.7 143.9 178.4 Li (ppm) 657.8 772.1 598.3 Ni (ppm) 1.97 3.28 5.67

[0055] Table 3. Product indicators of ferrous sulfate heptahydrate in Examples 1-3

[0056] Fe (%) 19.78 19.87 19.93 S(％) 11.51 11.57 11.63 Na (ppm) 285.62 218.43 234.98 Mg (ppm) 3.87 5.98 7.23 K (ppm) 7.88 12.39 16.28 Ca (ppm) 17.73 21.12 ND Zn (ppm) 7.16 11.33 17.92 Cu (ppm) 218.2 264.82 198.79 Al (ppm) 31.36 48.92 41.45 Mn (ppm) 37.85 31.81 24.64 Ni (ppm) 13.89 10.85 21.02 Pb (ppm) 4.48 7.71 3.41 Li (ppm) 38.92 47.23 22.89 Ti (ppm) 22.21 26.21 12.15

[0057] Table 4. Product Indicators of Monoammonium Phosphate in Examples 1-3

[0058] Main content (%) 98.60 98.55 98.43 Total N (%) 11.56 11.58 11.81 <![CDATA[P2O5(%)]]> 60.84 60.81 60.73 <![CDATA[SO4 2- (%)]]> 0.67 0.62 1.02

[0059] The above descriptions are merely embodiments of the present invention, and common knowledge regarding specific structures and characteristics is not elaborated upon here. It should be noted that those skilled in the art can make various modifications and improvements without departing from the structure of the present invention, and these should also be considered within the scope of protection of the present invention. These modifications and improvements will not affect the effectiveness of the present invention or the practicality of the patent. The scope of protection claimed in this application should be determined by the content of its claims, and the specific embodiments described in the specification can be used to interpret the content of the claims.

Claims

1. A method for separating and recovering phosphorus and iron from phosphorus-iron slag, a byproduct of battery recycling, characterized in that, Includes the following steps: Step 1, High-temperature reduction: The by-product phosphorus iron slag from the wet recovery of lithium iron phosphate black powder is calcined at high temperature under nitrogen protection to obtain solid powder. The high-temperature reduction temperature is 600-850℃ and the time is 1-3 hours. Step 2, acid leaching separation: The solid powder from Step 1 is leached with dilute sulfuric acid. After the acid leaching is completed, the acid leaching solution is filtered to obtain graphite solid and filtrate. The concentration of dilute sulfuric acid is 15% to 30%, the acid leaching temperature is 45 to 75°C, and the time is 1.5 to 3 hours. Step 3, Crystallization Separation: The filtrate from Step 2 is concentrated until fine crystals begin to appear. Then, it is cooled and crystallized. The cooled crystallized slurry is centrifuged to obtain the mother liquor and solid phase. The solid phase is dried under vacuum to obtain ferrous sulfate heptahydrate. Step 4, Extraction: Add sulfuric acid to the centrifuged mother liquor from Step 3 for acidification. After acidification, extract the solution. After extraction, obtain an organic phase and a raffinate phase. Wash the organic phase with monoammonium phosphate solution to remove sulfate impurities. Step 5, ammonia neutralization: Add pure water to the washed organic phase and pass ammonia gas to neutralize it. The pH value at the end of the neutralization reaction is 4.5-5.

4. After the reaction reaches the end, separate the liquid to obtain the organic phase and the aqueous phase. The organic phase is reused as the extractant for the secondary countercurrent extraction. The aqueous phase is dried to obtain industrial grade monoammonium phosphate. In step 4, the extraction is carried out using a two-stage countercurrent extraction method. The extractant used is a compound extractant composed of tributyl phosphate, methyl isobutyl ketone, and kerosene in a mass ratio of 5-7:1-4:0-2.

2. The method for separating and recovering phosphorus and iron from by-product phosphorus and iron slag in battery recycling according to claim 1, characterized in that: The acid leaching process in step 2 is protected by nitrogen gas.

3. The method for separating and recovering phosphorus and iron from the by-product phosphorus and iron slag of battery recycling according to claim 1, characterized in that: In step 3, the filtrate is concentrated by vacuum concentration, with a vacuum degree of -0.07 to -0.1 MPa and a temperature of 65 to 75°C; the cooling crystallization temperature is -15 to 0°C.

4. The method for separating and recovering phosphorus and iron from the by-product phosphorus and iron slag of battery recycling according to claim 3, characterized in that: In step 3, the vacuum drying temperature of ferrous sulfate heptahydrate is 75–90°C, and the drying time is 1.5–3 hours.

5. The method for separating and recovering phosphorus and iron from by-product phosphorus and iron slag in battery recycling according to claim 1, characterized in that: In step 4, the concentration of monoammonium phosphate used for organic phase washing is 20% to 31%, and the aqueous phase after washing and separation is concentrated by evaporation to obtain ammonium sulfate.

6. The method for separating and recovering phosphorus and iron from by-product phosphorus and iron slag in battery recycling according to claim 1, characterized in that: In step 4, the acidified solution is added to the secondary extractor, and the extractant is added to the primary extractor.

7. The method for separating and recovering phosphorus and iron from by-product phosphorus and iron slag in battery recycling according to claim 1, characterized in that: The raffinate phase after the secondary countercurrent extraction in step 4 is mainly composed of a mixture of iron, phosphorus, sulfur, copper and aluminum elements, and is used as a cement retarder after drying.

8. The method for separating and recovering phosphorus and iron from by-product phosphorus and iron slag in battery recycling according to claim 1, characterized in that: In step 5, when the aqueous phase does not reach a supersaturated state, it is returned to the washing process of the secondary extraction organic phase. After multiple cycles, it reaches supersaturation and crystals precipitate before entering the centrifugation process. The centrifuged mother liquor is returned to step 4 to wash sulfate ions from the fresh organic phase.

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