Method for multi-element synergistic leaching of phosphogypsum and recovery of by-products

Abstract

This invention discloses a method for the synergistic leaching of multiple elements from phosphogypsum and the recovery of by-products. The method includes pretreatment, leaching, and solid-liquid separation of phosphogypsum to obtain a leachate (mother liquor I) and purified phosphogypsum. In the leaching step, a mixed acid reaction consisting of sulfuric acid and an organic complex acid is used. Mother liquor I is reacted using oxalic acid precipitation, followed by solid-liquid separation to obtain a second precipitate and a filtrate (mother liquor II). The second precipitate is washed, dried, and calcined to obtain a mixed rare earth oxide enrichment, which is further refined to obtain rare earth oxide products. Mother liquor II is adjusted to pH and then subjected to a precipitation reaction, followed by solid-liquid separation to obtain a third precipitate and a filtrate (mother liquor III). The third precipitate is washed and calcined to obtain alumina products. This invention enables the synergistic and efficient leaching of multiple elements from phosphogypsum, and combines gradient precision separation to achieve targeted purification and high-value conversion of each valuable element.

Description

Technical Field

[0001] This invention relates to the field of industrial solid waste resource utilization technology, specifically a method for multi-element synergistic leaching of phosphogypsum and recovery of by-products. Background Technology

[0002] Phosphogypsum is a large industrial solid waste generated during the wet process of phosphoric acid production. Its comprehensive utilization rate is less than 40%. The open-air storage of large amounts of phosphogypsum not only occupies valuable land resources, but the soluble phosphorus, fluorides, trace heavy metals and organic matter it contains will also seep into the ground with leachate, causing soil and groundwater pollution. At the same time, powdered phosphogypsum is prone to generating dust, causing air pollution problems, and has become a key bottleneck restricting the green development of the phosphoric chemical industry.

[0003] Phosphogypsum is not simply a waste product, but a "man-made polymetallic symbiotic mineral" rich in various valuable elements. Its main component is calcium sulfate dihydrate (CaSO4·2H2O, content 85-95%), and it also contains rare earth elements (0.05-0.3% as oxides, mainly light rare earth elements such as lanthanum, cerium, neodymium, and praseodymium), aluminum (0.3-1.5% as Al2O3), and iron (0.2-0.8% as Fe, containing Fe...). 2+ Fe 3+ The phosphogypsum contains various metals, including phosphorus (0.5-2.0% as P2O5, containing soluble phosphorus, eutectic phosphorus, etc.), as well as impurities such as fluorides, silicates, and trace heavy metals. Rare earth elements, due to their ionic radii being similar to calcium, often replace calcium ions in the calcium sulfate lattice in isomorphous form or are adsorbed onto the crystal surface as fine particles, resulting in a complex state of occurrence. Aluminum, iron, and phosphorus are mostly dispersed in the gypsum matrix in adsorbed or combined states. It is evident that multiple metals coexist in phosphogypsum with similar chemical properties. Traditional separation methods (such as selective precipitation and solvent extraction) are ineffective in treating leachates with low concentrations and complex components, often requiring multi-stage processing, resulting in lengthy processes and low recovery rates. Achieving efficient extraction and separation of these valuable elements is the core direction for the high-value resource utilization of phosphogypsum.

[0004] Currently, the utilization technology of phosphogypsum mainly focuses on low-value-added pathways such as the preparation of building materials (e.g., gypsum board, cement retarders) and the production of cement co-produced with sulfuric acid, ignoring or simply discarding high-value elements such as rare earths and aluminum. Existing leaching technologies for valuable elements in phosphogypsum are all focused on the removal of single rare earth elements or impurities such as phosphorus / fluorine, without addressing the synergistic leaching of rare earths, aluminum, iron, and phosphorus. Techniques attempting to recover rare earths mostly employ strong acids (such as concentrated hydrochloric acid and nitric acid) for leaching alone, with the process flow independent of gypsum treatment. This not only results in high energy consumption and severe corrosion but also severely damages the gypsum matrix, making it difficult to achieve the synergistic recovery of gypsum and valuable elements. For example, patent CN105164288A uses a strong acid system of 0.2-8.0M to leach rare earth elements. Although strong acids can destroy the calcium sulfate lattice, they easily lead to the redeposition of metal ions in the solution, and the leaching selectivity for aluminum, iron, and phosphorus is poor, making it difficult to achieve multi-element synergistic leaching. Some technologies use single organic acid leaching, which has good complexation properties, but weak acid hydrolysis ability, and cannot effectively release valuable elements in the lattice, resulting in generally low leaching rates. Moreover, organic acids are expensive, making industrial application difficult. At the same time, existing leaching technologies do not design targeted leaching systems based on the different occurrence states of elements in phosphogypsum, resulting in incomplete leaching of some elements. For example, rare earth elements are isomorphically present in the calcium sulfate lattice, and a single acid system cannot achieve efficient desorption. A few documents, such as CN105543475, mention the use of a "mixed acid" leaching of phosphoric acid and sulfuric acid, but this is only for the enrichment of rare earth elements in phosphate rock and is not applied to phosphogypsum. Furthermore, no organic complexing acids are involved, and calcium, magnesium, aluminum, and phosphorus are only removed as impurities, without addressing the issue of gradual recovery. Summary of the Invention

[0005] The purpose of this invention is to solve the aforementioned technical problems by providing a method for the synergistic leaching of multi-element phosphogypsum and the recovery of by-products. This method can achieve the synergistic and efficient leaching of rare earth elements, aluminum, iron, and phosphorus from phosphogypsum, and, combined with gradient precision separation, complete the targeted purification and high-value conversion of each valuable element, ultimately achieving the full-component, high-value, and green utilization of phosphogypsum, thus overcoming many bottlenecks in existing technologies.

[0006] To achieve the above-mentioned objectives, the technical solution adopted by the present invention is as follows: A method for multi-element synergistic leaching of phosphogypsum and recovery of by-products includes pretreatment, leaching, and solid-liquid separation of phosphogypsum to obtain a leachate as mother liquor I and purified phosphogypsum. In the leaching step, a mixed acid reaction consisting of sulfuric acid and organic complex acid is used. Mother liquor I is reacted using oxalic acid precipitation, and then solid-liquid separation is performed to obtain a second precipitate and a filtrate as mother liquor II. The second precipitate is washed, dried, and calcined to obtain a mixed rare earth oxide enrichment, which is further refined to obtain a rare earth oxide product. Mother liquor II is adjusted to pH 4.5-5.5 for precipitation reaction, and after solid-liquid separation, a third precipitate and a filtrate as mother liquor III are obtained. The third precipitate is washed and calcined to obtain an alumina product.

[0007] In the pretreatment step, the phosphogypsum is purified and dehydrated to obtain pretreated phosphogypsum. Further purification steps may be used as needed. Homogenization includes, but is not limited to, thickening or ball milling; purification includes, but is not limited to, washing or neutralization; purification is not limited to, cyclone or crystallization; and dehydration includes, but is not limited to, pressure filtration or drying. Those skilled in the art can choose the appropriate method based on actual needs.

[0008] In the pretreatment step, when the aluminum content of phosphogypsum exceeds 0.8 wt% (calculated as Al2O3), an alkaline material is added for weak alkaline modification, and the pH of the modified slurry is controlled between 8.0 and 9.5. Examples of alkaline materials include lime milk, calcium hydroxide, carbide slag, ammonia water, fly ash alkaline solution, steel slag or slag suspension, etc.

[0009] In the leaching step, the leaching reaction temperature is 60-95℃ and the time is 2-5 hours.

[0010] In the leaching step, the mixed acid has a sulfuric acid concentration of 15-35 wt% and an organic complex acid selected from at least one of oxalic acid, citric acid, tartaric acid, EDTA, and ascorbic acid, with a total concentration of 0.1-5.0 mol / L; the liquid-to-solid ratio (L:kg) of the mixed acid volume to the dry weight of phosphogypsum is (3-10):1.

[0011] The mixed acid is selected from one of the following combinations: Combination 1: Sulfuric acid concentration 15-30%, organic complex acid 0.25-2.0 mol / L citric acid and / or 0.1-0.8 mol / L EDTA; Combination 2: Sulfuric acid concentration 20-35%, organic complexed acids 0.4-3 mol / L oxalic acid and 0.25-2.0 mol / L ascorbic acid; Combination 3: Sulfuric acid concentration 20-30%, organic complex acids 0.15-1.2 mol / L citric acid and 0.25-1.6 mol / L oxalic acid.

[0012] In the oxalic acid precipitation method, the pH of the mother liquor I is adjusted to 1.5-2.5, an oxalic acid solution with a concentration of 0.5-1.5 mol / L is added, and the mixture is stirred at 40-70℃ for 1-2 hours to precipitate the oxalic acid.

[0013] After obtaining the mixed rare earth oxide concentrate, the rare earth oxide products are obtained by fine separation using continuous ion chromatography or ion exchange methods. The key parameters of the continuous ion chromatography (CIC) device are: resin type: strong acid cation exchange resin, exchange capacity ≥4.5 mmol / g; column length: 10-50 m for a single column, depending on the separation requirements, multiple columns can be connected in series; Eluent: hydrochloric acid solution (0.1-3.0 mol / L) or citric acid / EDTA solution (0.05-0.3 mol / L), using gradient rinsing; flow rate: 1-5 BV / h (bed volume per hour); temperature: 20-60℃.

[0014] The mother liquor II is used to separate the alumina product using any of the following methods: Using Fe 3+ Preferred separation method: Adjust the pH of mother liquor II to 2.5-3.5 to allow Fe... 3+ The aluminum ions are preferentially hydrolyzed to ferric hydroxide or co-precipitated with phosphate to ferric phosphate. After solid-liquid separation, iron precipitate and mother liquor II-A are obtained. Then, the pH of mother liquor II-A is adjusted to 4.5-5.5 to precipitate aluminum ions as aluminum hydroxide or basic aluminum carbonate. After solid-liquid separation, the precipitate is washed and calcined to obtain alumina product with a purity ≥98%. The filtrate after aluminum separation contains Fe. 2+ and P's mother liquor III-A; Alternatively, an aluminum-preferred separation method can be used: reducing organic acids or reducing agents are added in the later stages of leaching or during the rare earth separation step, ultimately ensuring that the iron in mother liquor II remains at the Fe content. 2+ State; Adjust the pH of mother liquor II to 4.5-5.5, so that Al 3+ Hydrolysis precipitates aluminum hydroxide. After solid-liquid separation, the precipitate is washed and calcined to obtain an alumina product with a purity ≥98%. The filtrate after aluminum separation contains Fe. 2+ And the mother liquor III-B of P.

[0015] Fe 3+ In the preferential separation method or aluminum preferential separation method, the pH value is adjusted by adding an alkaline solution or introducing CO2 into mother liquor II or mother liquor II-A, wherein the alkaline solution is sodium carbonate or ammonia water; the reducing organic acid is ascorbic acid; and the reducing agent is sodium sulfite.

[0016] Iron and phosphorus separation is performed on either Mother Liquor III-A or Mother Liquor III-B using any of the following methods: Separation of ferrous oxalate: Add a reducing agent to mother liquor III-A or mother liquor III-B to remove residual Fe in the liquid phase. 3+ Reduced to Fe 2+ Adjust the pH to 2.5-3.5, add oxalic acid solution to precipitate, control the reaction temperature at 25-40℃, stir the reaction for 1-3 hours, then perform solid-liquid separation. After washing and drying, the precipitate yields Fe≥30wt% Fe 3+ <0.5wt% battery-grade ferrous oxalate, the filtrate is a phosphorus-containing mother liquor; Alternatively, ferric phosphate can be separated by adding an oxidizing agent to mother liquor III-A or mother liquor III-B to remove Fe. 2+ Oxidized to Fe 3+ Then, an alkaline precipitant is added to adjust the pH to 1.8-2.5, and the mixture is stirred for 1-2 hours to produce iron phosphate through co-precipitation. After aging and solid-liquid separation, the precipitate is washed and dried and used directly as an iron phosphate precursor, or it can be further calcined to obtain crystalline iron phosphate with Fe≥28wt% and P / Fe molar ratio of 0.95-1.05 for the preparation of lithium iron phosphate cathode materials. The filtrate is a phosphorus-containing mother liquor.

[0017] The mother liquor III-B can also be separated from ferric phosphate using a one-step method: An oxidant and an alkaline precipitant are simultaneously added to mother liquor III-B, and the Fe is precipitated at 50-70℃. 2+ Oxidized to Fe 3+ Simultaneously, it immediately co-precipitates with phosphate to form iron phosphate. After aging and solid-liquid separation, the precipitate is washed and dried and used directly as an iron phosphate precursor, or further calcined to obtain crystalline iron phosphate with Fe≥28wt% and P / Fe molar ratio0.95-1.05 for the preparation of lithium iron phosphate cathode materials. The filtrate is a phosphorus-containing mother liquor.

[0018] In the step of separating ferrous oxalate, the reducing agent is at least one of iron powder, zinc powder, ascorbic acid, and sodium sulfite, and the concentration of the oxalic acid solution is 0.5-1.5 mol / L; in the step of separating ferric phosphate or the one-step separation of ferric phosphate, the oxidizing agent is air, pure oxygen, hydrogen peroxide, or sodium hypochlorite; and the alkaline precipitant is ammonia or sodium carbonate.

[0019] In the step of separating ferrous oxalate, after adding a reducing agent to mother liquor III-A or mother liquor III-B, the reaction is carried out at a controlled temperature of 40-60℃ for 0.5-2 hours to remove residual Fe. 3+ All reduced to Fe 2+ In the step of separating ferric phosphate, an oxidant is added to mother liquor III-A or mother liquor III-B, and the reaction is carried out at a controlled temperature of 50-70℃ for 0.5-1.5 hours to separate the Fe... 2+ All oxidized to Fe 3+ .

[0020] The phosphorus-containing mother liquor is further reacted with ammonia or urea to prepare ammonium phosphate or urea phosphate compound fertilizer; or the phosphorus-containing mother liquor is concentrated, and phosphoric acid or phosphate is crystallized and recovered as a supplementary phosphorus source for wet-process phosphoric acid; or the pH of the phosphorus-containing mother liquor is adjusted to 5.0-6.5 and then directly used as a soluble phosphorus source for subsequent phosphorus chemical product production.

[0021] The washing water generated in each process is filtered. The primary filtrate is dephosphorylated or used to prepare phosphate. The filtered acid is sent to the return water tank. After membrane treatment or neutralization and precipitation, the supernatant is reused for phosphogypsum washing or mixed acid preparation. The precipitated sludge is dewatered and solidified, and then sent to the building materials industry for use or safely landfilled.

[0022] To address the problems existing in the background art, this application makes the following improvements: By pretreating phosphogypsum to remove surface-soluble phosphorus, fluorides, some organic matter, and coarse particulate impurities from the raw material, pretreated phosphogypsum with uniform particle size and improved purity is obtained, laying the foundation for subsequent mixed acid synergistic leaching and avoiding interference from impurities on the leaching system and subsequent elemental separation. For phosphogypsum with high aluminum content, weak alkaline modification can be carried out to promote the transformation of aluminum hydrate oxide form, thereby improving the subsequent leaching efficiency. This application is the first to employ a sulfuric acid-organic complex acid composite leaching system, combining the lattice-destructive ability of sulfuric acid with the metal ion stabilizing ability of organic complex acids. Sulfuric acid provides a strong proton environment, disrupting the calcium sulfate lattice and allowing isomorphically occurring rare earth elements and adsorbed aluminum and iron to enter the liquid phase. The organic complex acid forms stable water-soluble complexes with rare earth, aluminum, and iron ions, inhibiting their redeposition and maintaining a suitable pH environment through buffering, thus regulating the form of phosphorus. The synergistic effect of both systems ensures the leaching of rare earth, aluminum, and iron elements, significantly improving leaching efficiency and reducing co-leaching of impurities. Compared to existing single-acid leaching technologies, this method effectively improves leaching efficiency, significantly shortens leaching time, and allows for reaction at atmospheric pressure, reducing energy consumption and equipment corrosion.

[0023] Furthermore, through in-depth research on the types of mixed acids, this invention has screened out organic acids that are particularly suitable for multi-element leaching. Their main functions and applicable scenarios are as follows: Accordingly, based on different valuable element recovery targets and the characteristics of different types of organic complex acids, the inventors formulated three optimal combinations to achieve synergistic effects and realize targeted adaptation of the leaching system: When adapting to the recovery of rare earth and aluminum elements, the following combination is used: sulfuric acid concentration of 15-30wt%, organic complexing acid of 0.25-2.0mol / L citric acid and / or 0.1-0.8mol / L EDTA (ethylenediaminetetraacetic acid), which utilize the strong complexing ability of citric acid and EDTA on rare earth and aluminum ions to improve the leaching rate of both. When adapting to the recovery of iron and phosphorus, combination two is adopted: sulfuric acid concentration of 20-35wt%, sulfuric acid concentration of 20-35%, organic complexing acid of 0.4-3mol / L oxalic acid and 0.25-2.0mol / L ascorbic acid. Ascorbic acid maintains the reduced state of iron, while oxalic acid has the dual function of complexation and subsequent precipitation, which facilitates the directional separation of iron and phosphorus. When adapting to the synergistic recovery of all elements, combination three is adopted: sulfuric acid concentration of 20-30wt%, and organic complexing acids of 0.15-1.2mol / L citric acid and 0.25-1.6mol / L oxalic acid. Citric acid has both buffering and complexing functions, which can stabilize rare earth and aluminum ions over a wide pH range and prevent their redeposition; oxalic acid has both complexing and subsequent selective precipitation functions, which can assist in the complexation leaching of iron and provide a basis for the subsequent preparation of ferrous oxalate; the synergy of the two can cover the simultaneous leaching requirements of rare earth, aluminum, iron and phosphorus, and by combining the advantages of the two organic complexing agents, the simultaneous and efficient leaching of rare earth, aluminum, iron and phosphorus can be achieved.

[0024] Furthermore, after leaching, solid-liquid separation is performed to obtain leachate as mother liquor I and purified phosphogypsum. After washing, the purified phosphogypsum can be used to prepare high-strength α-hemihydrate gypsum or co-produce sulfuric acid and cementing materials as needed. This is existing technology and will not be described in detail.

[0025] Another important innovation of this application is the design of a gradient separation method for rare earth oxalic acid precipitation → selective aluminum precipitation → directional iron-phosphorus conversion. Specifically, the rare earth separation employs coarse enrichment followed by fine separation using oxalic acid precipitation. The fine separation can utilize continuous ion chromatography (CIC) to replace traditional solvent extraction, solving the problems of low recovery rate and heavy contamination in low-concentration rare earth solutions. The rare earth recovery rate is increased to ≥85%, and the purity to ≥99%. In the aluminum-iron separation process, considering the impact of iron valence state differences on aluminum-iron separation, and based on actual production needs, the following approach is proposed: using "Fe... 3+ Dual path of "preferential separation" or "aluminum preferential separation", Fe 3+ The preferred separation path is more suitable for phosphogypsum leachates with a high Fe3+ content, utilizing Fe 3+ With Al 3+ To address the slight differences in precipitation pH, a segmented operation of "low pH iron removal - medium pH aluminum precipitation" was employed: first, the pH was precisely controlled between 2.5 and 3.5 to allow Fe... 3+Preferred hydrolysis precipitation yields ferric hydroxide, or co-precipitation with phosphate ions in the feed solution yields ferric phosphate (which can be used to prepare low-end iron products such as iron oxide pigments and magnetic materials), achieving Fe... 3+ Complete removal, at this time Al 3+ Because the pH level had not been reached for precipitation, Al remained stable in the liquid phase; the pH was then adjusted to 4.5-5.5 to allow Al to precipitate. 3+ Selective precipitation occurs as aluminum hydroxide / basic aluminum carbonate. This method breaks through the traditional "one-step pH precipitation" approach of existing technologies, precisely separating the previously overlapping aluminum and iron precipitation zones, thus fundamentally avoiding Fe precipitation. 3+ With Al 3+ Co-precipitation; preferential separation of aluminum is more suitable for Fe 2+ High-content phosphogypsum leachate utilizes Fe 2+ The hydrolysis precipitation pH (>7.0) is much higher than that of Al. 3+ The characteristics of this process allow for pre-reduction and valence control during the leaching or rare earth separation stages, followed by a one-step pH precipitation process: ascorbic acid, sodium sulfite, and other reducing organic acids / reducing agents are added in the later stages of leaching or the rare earth separation step to reduce the Fe in the feed solution. 3+ All reduced to Fe 2+ And locked to Fe throughout the process 2+ At this point, directly adjust the pH to 4.5-5.5 to precipitate aluminum and Fe. 2+ Because the precipitation pH was not reached and the iron was completely retained in the liquid phase, a single, interference-free separation of aluminum and iron was achieved, simplifying the process and avoiding reagent waste. In summary, Fe... 3+ The preferred separation path is suitable for oxidative leaching conditions or processes where the reducing environment is not strictly controlled; the "aluminum-preferred separation" path is suitable for reducing leaching conditions and can achieve cleaner aluminum-iron separation. Both paths have their advantages and can be selected according to the characteristics of the raw materials and product requirements.

[0026] Furthermore, regarding the current situation of low-concentration Fe-P feed solutions containing sulfate impurities in phosphogypsum, existing technologies in the iron-phosphorus treatment of phosphogypsum and similar industrial solid waste leachates all aim at impurity removal, without specifically recovering iron and phosphorus as high-value elements after aluminum separation. This application overcomes the technical bias that low-concentration, impurity-containing industrial solid waste leachates cannot be used to prepare high-purity battery-grade new energy materials. It also improves the iron-phosphorus treatment process, coupling it with the aforementioned steps, and flexibly selecting different preparation routes depending on the target product. For Fe 3+ Preferential separation method / Residual Fe after preferential separation of aluminum 3+ + main Fe 2+Based on the mother liquors III-A / III-B, a directional preparation route for separating ferrous oxalate was designed: first, the residual Fe... 3+ Reduced to Fe 2+ , using Fe 2+ With oxalate at pH 2.5-3.5 (this pH range ensures that Fe... 2+ This method utilizes the property of Fe to form sparingly soluble ferrous oxalate (while ensuring sufficient precipitation of oxalate and avoiding co-precipitation of phosphorus), achieving precise precipitation and separation of iron, while phosphorus remains in the liquid phase for subsequent recovery. 2+ The selective precipitation characteristics of Fe with oxalate at pH 2.5-3.5 ensure that Fe is within this pH range. 2+ The process effectively precipitates oxalate while simultaneously inhibiting phosphate co-precipitation, achieving precise separation of iron and phosphorus. It offers advantages such as excellent iron-phosphorus separation, high product purity, and mild operating conditions. The separated ferrous oxalate crystals grow to meet battery-grade requirements (Fe ≥ 30 wt%, Fe...). 3+ <0.5wt%).

[0027] Alternatively, a directional preparation route for separating iron phosphate can be adopted: adding an oxidant to Fe 2+ All oxidized to Fe 3+ Then adjust the pH to 1.8-2.5 (this pH range achieves Fe...) 3+ Calculation of coprecipitation with phosphate ions (while avoiding the formation of Fe(OH)3 impurity phase) yields iron phosphate. This method involves using Fe... 2+ All oxidized to Fe 3+ Subsequently, it is co-precipitated with phosphate in the pH range of 1.8-2.5, and the Fe / P ratio can be precisely controlled within the range of 0.95-1.05. This process offers advantages such as good product crystallinity, precise Fe / P ratio control, and meeting the requirements of a precursor for composite lithium iron phosphate cathode materials. Specifically for the mother liquor III-B after aluminum preferential separation, a one-step oxidation co-precipitation method can be used, where the oxidant and alkaline precipitant are added simultaneously to allow Fe... 2+ Oxidized to Fe 3+ with Fe 3+ The co-precipitation of phosphate ions to form ferric phosphate occurs simultaneously, significantly improving separation efficiency and avoiding the loss of Fe due to separate oxidation. 3+ The problem of localized overconcentration leading to Fe(OH)3 formation has improved the purity and crystallinity of ferric phosphate products.

[0028] Compared with the prior art, the present invention has the following outstanding advantages: (1) One agent with multiple effects and synergistic efficiency: The innovative "sulfuric acid-organic complex acid" mixed system can simultaneously achieve gypsum purification and efficient leaching of multiple elements (rare earth, aluminum, iron, phosphorus) under mild conditions (60-95℃, normal pressure). Compared with traditional single strong acid leaching, the leaching rate of valuable metals is increased by 15-40%, the leaching time is shortened by 30-50%, and the energy consumption is reduced by about 25%.

[0029] (2) Element gradient separation and precise purification: Based on the characteristics of the feed liquid and the differences in the chemical properties of rare earth, aluminum, iron, phosphorus and other elements, a gradient separation and purification strategy of "diversion-stepping" was creatively designed: It solved the problem of low recovery rate and heavy environmental pollution when the traditional solvent extraction method is used to process low concentration complex feed liquid. The total recovery rate of rare earth is ≥85%, and the purity of single rare earth oxide is ≥99.0%; aluminum is selectively precipitated by precise pH control, the aluminum recovery rate is ≥80%, and the alumina purity is ≥98%; iron and phosphorus are directionally converted into ferrous oxalate or ferric phosphate by redox-pH coupling control, and high-value by-products are produced to meet the requirements of new energy materials, realizing the "full component and full value chain" development of phosphogypsum.

[0030] (3) The process is compact and environmentally friendly: the method of the present invention involves less material transfer and low energy consumption. The organic complex acid and extractant used in the production can be recycled and the reagent cost is low. Wastewater is reused after treatment and solid waste can be fully recycled. It does not use highly toxic reagents (such as cyanide) and does not produce persistent organic pollutants, making it safe and environmentally friendly.

[0031] (4) Waste resource utilization, safety and environmental protection: All components of phosphogypsum are fully utilized, and gypsum is collectively transformed into high-value-added building materials such as α-hemihydrate gypsum. Rare earth, aluminum, iron, and phosphorus are transformed into high-value products. Washing water from each process is treated and recycled. Sedimented sludge is dewatered and solidified before being sent to the building materials industry for use or safely landfilled. The entire process does not use highly toxic reagents (such as cyanide) and does not produce persistent organic pollutants, thus realizing the resource utilization of solid waste. (5) Mature technology and easy to industrialize: The product structure can be flexibly adjusted according to the raw material composition and market demand (focusing on rare earth, aluminum or iron and phosphorus), which is highly adaptable; it is suitable for forming industrial coupling with existing phosphorus chemical enterprises and new energy material enterprises, local conversion, and reducing transportation costs. Attached Figure Description

[0032] Figure 1 Pretreatment process for phosphogypsum Figure 2 This is a flowchart of the present invention. Figure 3 A comparative graph showing the effects of different organic acids on the leaching rates of rare earth elements and aluminum.

[0033] As can be seen from the figure, under the same base solution (20% H2SO4, 75℃, 3h), the leaching rate of rare earth elements was only 58% and the leaching rate of aluminum was only 52% when leaching with sulfuric acid alone, which was the lowest among all systems. This proves that although sulfuric acid can provide a strong proton environment to destroy the calcium sulfate lattice, it is not good at complexing and extracting rare earth elements and aluminum that are isomorphically present in the lattice, and there is a significant leaching shortcoming.

[0034] Adding different organic complex acids to sulfuric acid significantly improved the leaching rates of both rare earth elements and aluminum, with noticeable differences in the improvement effects of different organic acids. Overall, the addition of organic complex acids increased the leaching rate of rare earth elements by 22%–38% and the leaching rate of aluminum by 20%–38% compared to the single sulfuric acid system, fully validating the technical superiority of the "sulfuric acid-organic complex acid" composite leaching system in this application for the synergistic and efficient leaching of multiple elements.

[0035] Figure 4 This is a graph showing the effect of pH on the selective precipitation of aluminum and iron.

[0036] The figure contains four curves, representing the aluminum precipitation rate, iron (Fe) precipitation rate, and iron precipitation rate. 3+ Precipitation rate, iron (Fe) 2+ The variation of precipitation rate and phosphorus precipitation rate with pH: (1) Fe 3+ With Al 3+ There is an exploitable pH difference window in the precipitation range (Fe). 3+ Theoretical basis for preferential path separation): Fe 3+ The precipitation curve (solid red line) begins at approximately pH 1.0, and the precipitation rate increases sharply from about 6% to over 50% in the pH 2.0–3.5 range, reaching nearly 100% at pH 4.0. Meanwhile, the aluminum precipitation rate (solid blue line) is only about 20% at pH 2.0, but increases rapidly in the pH 3.5–5.5 range, reaching approximately 99% at pH 4.5. This indicates that Fe precipitation at pH 2.0–3.5... 3+ Priority separation interval (marked by red box in the figure), Fe 3+ It can be preferentially precipitated, while Al 3+ Most of the aluminum remained in the liquid phase, achieving effective stepwise separation; subsequently, the pH was raised to the optimal range for aluminum separation (marked in orange). 3+ The sedimentation rate can reach over 95%.

[0037] (2) Fe 2+ The pH of the precipitate is much higher than that of Al. 3+ This provides theoretical support for the preferential separation pathway of aluminum: iron (Fe) 2+The precipitation curve (orange dashed line) shows a precipitation rate of only about 25% at pH 4.5, approaching 100% at pH 6.0. Its complete precipitation occurs at a pH lower than that of Al. 3+ This is approximately 1.5–2.0 pH units higher. This indicates that when iron in the system is pre-reduced to Fe, 2+ Subsequently, when adjusting the aluminum precipitation within the pH range of 4.5–5.5, Fe... 2+ Because the pH value for precipitation is not reached, the aluminum and iron are completely retained in the liquid phase, enabling one-time, interference-free separation of aluminum and iron, which corresponds to the operational basis of the aluminum-preferred separation path in this application.

[0038] (3) Phosphorus does not precipitate within a wide pH range, which is beneficial for the subsequent stepwise recovery of iron and phosphorus: The phosphorus precipitation rate curve (green solid line) shows an extremely low precipitation rate (close to 0%) when pH ≤ 5.0 and about 35% when pH 6.0. This indicates that throughout the rare earth separation (pH 1.5–2.5) and aluminum-iron separation (pH 2.5–5.5), phosphorus remains in a dissolved state in the liquid mother liquor and will not be lost due to co-precipitation during the iron removal and aluminum removal steps. This provides a sufficient phosphorus concentration basis for the subsequent directional conversion of iron and phosphorus in mother liquor III-A / B (to prepare ferrous oxalate or ferric phosphate).

[0039] Figure 5 This is a schematic diagram of the elution curve for the separation of rare earth elements by continuous ion chromatography (CIC).

[0040] The figure shows the elution separation curves of four rare earth elements, La, Ce, Pr, and Nd, in the CIC system as a function of the eluent concentration gradient (0.15M→0.5M→1.0M→2.0M HCl): (1) Complete baseline separation of the four rare earth elements was achieved, with symmetrical peaks and no overlap: the La (lanthanum) fraction was collected in the range of 150–320 BV, with a peak concentration of 8.2 g / L; the Ce (cerium) fraction was collected in the range of 320–500 BV, with a peak concentration of 7.8 g / L; the Pr (praseodymium) fraction was collected in the range of 500–650 BV, with a peak concentration of 7.5 g / L; and the Nd (neodymium) fraction was collected in the range of 650+ BV, with a peak concentration of 7.9 g / L. The peak and valley concentrations between adjacent fractions were close to 0 g / L, and there was no obvious overlap between adjacent rare earth elements. This indicates that CIC on a 001×7 strongly acidic cation exchange resin, by utilizing the difference in coordination strength between each rare earth ion and the resin (negatively correlated with the ionic radius: La>Ce>Pr>Nd), combined with gradient elution of 0.15M to 2.0M HCl, achieved complete baseline separation of light rare earth elements La, Ce, Pr, and Nd, with good separation selectivity.

[0041] (2) The hydrochloric acid concentration gradient closely matches the rare earth ion radius pattern, proving the rationality of the elution scheme: as the hydrochloric acid concentration gradually increases from 0.15M to 2.0M, rare earth elements are eluted in order of decreasing ion radius (La→Ce→Pr→Nd). The precise gradient switching of the eluent concentration (marked with orange arrows at the beginning of each collection interval) ensures the pure separation of each fraction and avoids cross-contamination of adjacent rare earth components.

[0042] (3) Compared with traditional solvent extraction, CIC achieves efficient and precise separation of low-concentration rare earth feed solution: the total concentration of rare earth in the feed solution is only about 10 g / L (which is a low-concentration leachate). Traditional solvent extraction has a low distribution ratio and unsatisfactory extraction rate at this concentration (as shown in Comparative Example 2). However, CIC can still achieve precise separation of each single rare earth oxide with a purity of ≥99%. Moreover, no organic solvent is used throughout the process, there is no loss of extractant, and the COD of the waste liquid is low, which meets the requirements of green chemical industry. This fully demonstrates the technological innovation value of this application in using CIC to replace the traditional solvent extraction method to treat low-concentration phosphogypsum rare earth leachate.

[0043] Figure 6 The image shows the XRD pattern of ferrous oxalate.

[0044] Figure 7 This is the XRD pattern of iron phosphate. Detailed Implementation

[0045] To make the objectives, technical solutions, and advantages of this invention clearer, the invention will be described in detail below with reference to specific embodiments. These embodiments are for illustrative purposes only and do not constitute a limitation on the scope of protection of this invention.

[0046] The phosphogypsum raw materials used in all embodiments were taken from the same batch, and the main components are as follows: CaSO4·2H2O 90.5wt%, total rare earth oxides (REO) 0.12wt% (of which Nd2O3 accounts for 40wt% of REO, La2O3 accounts for 30wt%, Ce2O3 accounts for 20wt%, and the remaining rare earths account for 10wt%), Al2O3 0.9wt%, Fe 0.65wt% (of which Fe2O3 accounts for 70wt% of the total Fe and FeO accounts for 30wt%), P2O5 1.5wt% (soluble phosphorus 0.6wt%), soluble fluorine 0.5wt%, SiO2 1.2wt%, organic matter 0.3wt%, and the batch feed amount is 500kg (dry basis).

[0047] Example 1: Aluminum-preferred separation path + III-B one-step oxidation co-precipitation to produce ferric phosphate, mixed acid combination two (emphasizing iron + phosphorus) recovery. Step 1: Preprocessing (1-1) Crushing and screening: 500 kg of phosphogypsum (dry basis) is crushed by a jaw crusher and passed through a 2 mm standard sieve. 495 kg of material is undersized and 5 kg of material is oversized and returned to the crusher.

[0048] (1-2) Flotation for impurity removal: Add 1485L of deionized water (liquid-solid ratio 3:1 L / kg) to 495kg of phosphogypsum, stir evenly at room temperature for 15 minutes, add flotation reagent (0.05kg of pine oil), and float for 5 minutes to remove organic impurities and some soluble fluoride. Filter the solution through a plate and frame filter press (30 plates, 800×800mm, pressure 0.5MPa) for 30 minutes. Filtrate analysis: soluble phosphorus removal rate 62wt%, soluble fluoride removal rate 75wt%.

[0049] (1-3) Solid-liquid separation: Pretreated phosphogypsum filter cake with a moisture content of 14wt% and a dry weight of 488kg was obtained and set aside.

[0050] Step 2: Co-leaching with mixed acids (mixed acid combination 2, reducing system) (2-1) Preparation of mixed acid solution: Add 1600L of deionized water to a 3000L acid mixing tank. While stirring (80rpm), slowly add 587kg of concentrated sulfuric acid (98wt%) to prepare a 25wt% sulfuric acid solution. Cool to 40℃. Then, add 315kg (1.0mol / L) of oxalic acid dihydrate (H2C2O4·2H2O) and 352kg (0.8mol / L) of ascorbic acid (C6H8O6) in sequence, stirring for 30 minutes until completely dissolved. Add water to a total volume of 2500L. The final composition of the mixed acid is: 25wt% H2SO4 + 1.0mol / L oxalic acid + 0.8mol / L ascorbic acid, with a liquid-to-solid ratio of 5:1 (L:kg).

[0051] (2-2) Nitrogen protection: Transfer the mixed acid to 3m 3 After leaching the reactor, purge with nitrogen (50L / min) for 30 minutes to remove dissolved oxygen, and maintain a flow rate of 10L / min throughout the process.

[0052] (2-3) Leaching reaction: Add the pretreated phosphogypsum (488 kg) to the mixed acid in 3 batches, stirring evenly with a 5-minute interval between each batch; raise the temperature to 85℃ (heating rate 2℃ / min), and maintain the temperature while stirring (180 rpm) for 3 hours; add 50 L of ascorbic acid solution (0.5 mol / L) in the last 30 minutes of the reaction to ensure that the iron in the system is converted to Fe. 2+ It exists in form.

[0053] (2-4) Leaching effect: Samples were taken for analysis after the reaction was completed. The concentrations of each element in the leachate were: REE 0.48 g / L, Al 1.72 g / L, Fe 1.62 g / L (Fe 2+1.58 g / L, Fe 3+ 0.04 g / L, Fe 2+ (97.5 wt%), P 1.78 g / L; Leaching rate: rare earth 88.2 wt%, aluminum 84.5 wt%, iron 91.6 wt%, phosphorus 86.3 wt%.

[0054] Step 3: Solid-liquid separation and purification of gypsum treatment (3-1) Hot solid-liquid separation: The leaching mixture (75℃) was pumped into a plate and frame filter press (1000×1000mm, 40 plates, pressure 0.8MPa) and filtered for 40 minutes; 2430L of mother liquor I (deep yellow, pH 0.7) was obtained, and 455kg of purified phosphogypsum filter cake (wet basis, moisture content 10wt%) was obtained.

[0055] (3-2) Gypsum washing: The purified phosphogypsum was washed twice with 450L of warm water (50℃), stirring for 15 minutes each time and then filtered to obtain 408kg of purified gypsum (wet basis). The washing water was incorporated into the waste liquid treatment system.

[0056] (3-3) Utilization of gypsum resources: Purified gypsum was converted to crystals by hydrothermal method (120℃, 0.2MPa, 3 hours), and dried at 110℃ for 5 hours to obtain 358kg of α-hemihydrate gypsum (dry basis); Tests: flexural strength 8.5MPa, compressive strength 41MPa, whiteness 83, which meets the GB / T 9776-2008 standard for high-strength gypsum.

[0057] Step 4: Rare Earth Separation (4-1) Oxalic acid precipitation: The pH of the mother liquor I (2430L) was slowly adjusted to 2.0 with 25wt% ammonia water (14kg of ammonia water was consumed, which took 30 minutes), and the temperature was raised to 60℃; oxalic acid solution (1.0mol / L, 189kg of oxalic acid dihydrate dissolved in 150L of water) was slowly added over 1 hour, and the reaction was continued for 1.5 hours; centrifugation (3000rpm, 10 minutes) was performed to obtain 11.5kg of light yellow rare earth oxalate precipitate (wet basis), and the filtrate was mother liquor II 2410L.

[0058] (4-2) Calcination enrichment: The precipitate was washed three times with deionized water (50℃) (300L each time), dried at 80℃ for 8 hours to obtain 3.1kg, and calcined in a muffle furnace at 700℃ for 3 hours (heating rate 5℃ / min) to obtain 0.75kg of mixed rare earth oxide enrichment; the purity of the REO enrichment was tested: the total REO content was 68.5wt%, the main impurities were CaO 12.3wt% and Fe2O 38.2wt%.

[0059] (4-3) CIC fine separation: The enriched product was dissolved in 2 mol / L hydrochloric acid to prepare 48 L of feed solution with a total REE concentration of 10 g / L. This solution was then introduced into a continuous ion chromatography system (001×7 strong acid cation exchange resin, 5 columns in series, column diameter 10 cm × column length 15 m, gradient elution with 0.15–2.0 mol / L hydrochloric acid, flow rate 2.5 BV / h, 45℃). The fractions were collected, precipitated with oxalic acid, and ignited to obtain: Total rare earth recovery rate: 86.7 wt% Step 5: Preferential separation of aluminum (5-1) Confirmation of iron valence state: Detection of mother liquor II: Fe 2+ 1.60 g / L, Fe 3+ 0.04 g / L, Fe 2+ It accounts for 97.6 wt%, with an ORP value of -18 mV (Ag / AgCl), which meets the conditions for preferential aluminum separation.

[0060] (5-2) CO2-assisted alkalization of aluminum precipitation: CO2 (industrial grade, 45L / min) was simultaneously introduced into mother liquor II (2410L) and 18wt% sodium carbonate solution (175kg Na2CO3 dissolved in 820L water) was slowly added. The pH was precisely adjusted to 5.0±0.1 in real time by pH electrode (adjustment time 1.5 hours), the temperature was 55℃, and the stirring was 100rpm. After reaching the pH, the reaction continued for 1 hour, and the mixture was allowed to stand for 1 hour. The mixture was then filtered by plate and frame filter press.

[0061] (5-3) Preparation of alumina: 43 kg of white gel-like aluminum hydroxide precipitate (wet basis) was obtained, washed 4 times (500 L each time) with deionized water (60℃), dried at 110℃ for 12 hours to obtain 6.3 kg, calcined in a rotary kiln (2 rpm) at 650℃ for 3 hours (heating rate 8℃ / min) to obtain 4.0 kg of activated alumina product; Analysis: Al2O3 purity 98.3 wt%, Fe2O3 0.5 wt%, SiO2 0.8 wt%, CaO 0.3 wt%; specific surface area 218 m² 2 / g (BET), pore volume 0.43mL / g, meeting the standards for activated alumina. Aluminum recovery rate: 81.2wt%. (5-4) Mother liquor III-B: 2380L of filtrate after aluminum separation, Fe 2+ 1.58 g / L, Fe 3+ 0.04 g / L, P 1.71 g / L, for later use.

[0062] Step Six: One-step oxidation co-precipitation to prepare ferric phosphate (6-1) One-step co-precipitation: Add hydrogen peroxide (30wt%, 42kg) and 18wt% ammonia solution (added at a controlled rate using a metering pump) simultaneously to mother liquor III-B (2380L), and heat to 60℃; monitor pH in real time, and maintain pH stable at 2.0±0.1 (adjustment time 1 hour), Fe... 2+ Oxidized to Fe 3+ At the same time, immediately with PO4 3- Ferric phosphate was co-precipitated; the reaction was continued at 60°C (120 rpm) for 1.5 hours with stirring, followed by aging for 2 hours (to allow for crystal growth).

[0063] (6-2) Solid-liquid separation: The mixture was filtered using a plate and frame filter press to obtain approximately 28 kg (wet basis) of iron phosphate precipitate (light yellow). This precipitate was washed three times with deionized water (400 L each time) and dried at 80°C for 10 hours to obtain 5.8 kg of iron phosphate precursor. Further calcination was performed in a tube furnace at 600°C for 2 hours (air atmosphere, heating rate 5°C / min) to obtain 5.5 kg of crystalline iron phosphate product. Analysis showed: Fe 29.5 wt% (≥28 wt%), P / Fe molar ratio 0.98 (0.95–1.05), XRD pattern showed FePO4, meeting the requirements for lithium iron phosphate cathode material precursors. Iron recovery rate: 80.3 wt%, Phosphorus recovery rate: 78.5 wt%. (6-3) Phosphorus-containing mother liquor: 2350L of filtrate (containing about 0.36g / L of P) was added to adjust the pH to 6.5, and the mixture was evaporated, concentrated and crystallized to obtain 1.2kg of ammonium phosphate by-product ((NH4)2HPO4) with a purity of 94.2wt%.

[0064] Step 7: Waste liquid recycling The 2600L of washing water from each process is settled, neutralized (with lime milk added to pH 7.0), and filtered. The supernatant is reused for washing phosphogypsum and preparing mixed acids. The settled sludge is dewatered and solidified before being sent to building materials companies for use.

[0065] Example 2: Fe 3+ The preferred separation path + mother liquor III-A reduction precipitation to produce ferrous oxalate, mixed acid combination one (focusing on rare earth + aluminum). Step 1: Preprocessing (1-1) Crushing and screening: 500kg of phosphogypsum is crushed to a particle size of <2mm and passed through a 2mm standard sieve. 496kg of material is undersized and 4kg of material is oversized and returned to the crusher.

[0066] (1-2) Flotation to remove impurities: Add 1488L of deionized water (liquid-solid ratio 3:1 L / kg), stir at room temperature for 15 minutes, remove impurities by flotation, and filter by plate and frame filter press (pressure 0.5MPa) for 30 minutes; test: soluble phosphorus removal rate 64wt%, soluble fluoride removal rate 77wt%, pretreated phosphogypsum filter cake is obtained with a moisture content of 13wt% and a dry basis weight of 490kg, for later use.

[0067] Step 2: Co-leaching with mixed acids (mixed acid combination 1, oxidizing system) (2-1) Preparation of mixed acid solution: Add 1750L of deionized water to a 3000L acid mixing tank, and slowly add 424kg of concentrated sulfuric acid (98wt%) to prepare a 20wt% sulfuric acid solution (the lower sulfuric acid concentration facilitates selective leaching of rare earth elements and aluminum, reducing excessive leaching of iron). Cool to 40℃; add 420kg (0.8mol / L) of citric acid monohydrate (C6H8O7·H2O) and 186kg (0.2mol / L) of disodium EDTA (Na2EDTA·2H2O), and stir for 40 minutes to dissolve; add water to a final volume of 2450L. The final composition of the mixed acid is: 20wt% H2SO4 + 0.8mol / L citric acid + 0.2mol / L EDTA, with a liquid-to-solid ratio of 5:1 (L:kg). This mixture does not contain ascorbic acid, and the system is oxidizing. Fe is leached as Fe2+. 3+ Mainly.

[0068] (2-2) Leaching reaction (atmospheric pressure, without nitrogen protection): Add the pretreated phosphogypsum (490 kg) to the mixed acid in 3 batches, heat to 80℃ (heating rate 2℃ / min), and stir (180 rpm) for 3 hours.

[0069] (2-3) Leaching effect: Sampling analysis: REE 0.52g / L, Al 1.81g / L, Fe 1.55g / L (Fe 2+ 0.25 g / L, Fe 3+ 1.30 g / L, Fe 3+ (83.9 wt%), P 1.68 g / L; Leaching rate: rare earth 91.0 wt%, aluminum 87.5 wt%, iron 87.8 wt%, phosphorus 82.4 wt%.

[0070] Step 3: Solid-liquid separation and purification of gypsum treatment (3-1) Solid-liquid separation: The leaching mixture (80℃) was filtered by plate and frame filter press while hot (pressure 0.8MPa, 40 minutes) to obtain mother liquor I 2420L (pH 0.8) and purified phosphogypsum 452kg (wet basis, water content 10wt%).

[0071] (3-2) Washing and resource utilization of gypsum: Washed twice with 440L of warm water (50℃) to obtain 406kg of purified gypsum (wet basis); after hydrothermal crystallization (120℃, 0.2MPa, 3 hours) and drying, 356kg of α-hemihydrate gypsum (dry basis) was obtained; test results: flexural strength 8.3MPa, compressive strength 40MPa, whiteness 84.

[0072] Step 4: Rare Earth Separation (4-1) Oxalic acid precipitation: Adjust the pH of mother liquor I (2420L) to 2.0 with 25wt% ammonia water (consuming 15kg of ammonia water, 30 minutes), add oxalic acid solution (1.0mol / L, 199kg of oxalic acid dihydrate dissolved in 155L of water, addition time 1 hour) at 60℃, and continue the reaction for 1.5 hours; centrifuge (3000rpm, 10 minutes) to obtain 12.2kg of rare earth oxalate precipitate (wet basis), and the filtrate is mother liquor II 2400L.

[0073] (4-2) Calcination enrichment: The precipitate was washed 3 times (50℃ deionized water, 300L each time), dried at 80℃ for 8 hours to obtain 3.3kg, and calcined at 700℃ for 3 hours to obtain 0.81kg of mixed rare earth oxide enrichment; REO content 69.8wt%.

[0074] (4-3) CIC fine separation: The concentrate was dissolved in 2 mol / L hydrochloric acid to prepare 52 L of feed solution with a total REE concentration of 10 g / L, which was then introduced into the CIC system (with the same parameters as in Example 1); the result was: Total rare earth recovery rate: 89.2 wt% (higher than Example 1, demonstrating the advantages of combining rare earth recovery). Step 5: Fe 3+ Priority separation (5-1) Confirmation of iron valence state: Mother liquor II: Fe 3+ 1.28 g / L, Fe 2+ 0.24 g / L, Fe 3+ The content is 84.2 wt%, and the ORP value is +185 mV (Ag / AgCl), which is suitable for iron. 3+ Prioritize separating paths.

[0075] (5-2) Low pH iron precipitation: Slowly add 18wt% sodium carbonate solution to mother liquor II (2400L), precisely control the pH to 3.0±0.1 (consuming 85L of Na2CO3 solution, adjustment time 45 minutes), temperature 55℃, stir (100rpm) for 1 hour; Fe 3+ Preferentially hydrolyzed to Fe(OH)3 or with PO4 3-The co-precipitate was FePO4; plate and frame filtration yielded 12.8 kg (wet basis) of iron precipitate (brownish-yellow), and the filtrate was 2370 L of mother liquor II-A; the mother liquor II-A was analyzed for Fe. 3+ 0.08 g / L, Fe 2+ 0.22g / L, Al 1.79g / L, P 1.45g / L.

[0076] (5-3) Precipitation of aluminum at pH: Add 18wt% sodium carbonate solution to mother liquor II-A (2370L), precisely controlling the pH to 4.8±0.1 (adjustment time 45 minutes), temperature 55℃, stirring 100rpm, reaction for 1 hour; Al 3+ Selective precipitation yielded aluminum hydroxide; plate and frame filtration was performed to obtain 46 kg of aluminum hydroxide precipitate (wet basis), and the filtrate was 2340 L of mother liquor III-A.

[0077] (5-4) Alumina preparation: The precipitate was washed four times with deionized water (60℃) (500L each time), dried at 110℃ for 12 hours to obtain 6.7kg, and calcined in a rotary kiln at 650℃ for 3 hours (heating rate 8℃ / min) to obtain 4.2kg of alumina product; Analysis: Al2O3 purity 98.5wt%, Fe2O3 0.4wt%, SiO2 0.7wt%, CaO 0.3wt%; specific surface area 221m² 2 / g, pore volume 0.44mL / g. Aluminum recovery rate: 83.1wt% (5-5) Mother liquor III-A: 2340L, Fe 2+ 0.21 g / L, Fe 3+ 0.08 g / L, P 1.43 g / L, for later use.

[0078] Step Six: Reduction and precipitation to prepare ferrous oxalate (6-1) Reduction of iron valence state: Add sodium sulfite (Na2SO3, 15kg, dissolved in 100L water) to mother liquor III-A (2340L), and stir at 50℃ for 1 hour under nitrogen protection (flow rate 10L / min) to remove residual Fe. 3+ (0.08 g / L) was completely reduced to Fe. 2+ Detect ORP to -25mV (Ag / AgCl) to confirm Fe 3+ <0.01g / L; adjust the pH to 3.0 with dilute ammonia (consuming 8kg of ammonia).

[0079] (6-2) Oxalic acid precipitation: Transfer mother liquor III-A to a precipitation vessel and maintain the temperature at 35℃; calculate Fe. 2+Total amount: 0.29 g / L × 2340 L = 679 g (12.1 mol). Weigh 1.83 kg (14.5 mol) of oxalic acid dihydrate (H2C2O4·2H2O) according to 1.2 times the amount used, dissolve it in 80 L of warm water (50 °C) to prepare an oxalic acid solution (approximately 0.17 mol / L, volume 85 L). Add the solution dropwise at a constant rate of 0.7 L / min using a metering pump for 2 hours. After the addition is complete, continue stirring for 1 hour, then reduce the stirring speed to 30 rpm and age for 3 hours.

[0080] (6-3) Solid-liquid separation and drying: Centrifugation (2500 rpm, 10 minutes) was performed. The precipitate was washed four times (300 L each time, under N2 protection) with deoxygenated deionized water (pre-purged with N2 for 30 minutes to remove oxygen, dissolved oxygen <0.5 mg / L), and then vacuum dried (65℃, vacuum degree -0.09 MPa, 6 hours, with continuous trace N2 purging). 2.1 kg of battery-grade ferrous oxalate product (light yellow fine powder) was obtained. Detection: Fe 30.8 wt% (≥30 wt%), Fe 3+ 0.3wt% (<0.5wt%), oxalate content 57.2wt%, XRD shows pure FeC2O4·2H2O crystalline phase, meeting battery-grade requirements. Iron recovery rate: 77.4wt%. (6-4) Phosphorus-containing mother liquor: 2320 L of filtrate (P 1.41 g / L) was reacted with ammonia (pH adjusted to 6.8), evaporated, concentrated, and crystallized to obtain 4.5 kg of ammonium phosphate byproduct ((NH4)2HPO4) with a purity of 95.1 wt%. Phosphorus recovery rate: 76.8 wt%. Step 7: Waste liquid recycling Each process's 2500L of washing water is neutralized (pH 7.0), precipitated, and filtered. The supernatant is reused, and the solidified sludge is sent to building materials companies.

[0081] Example 3: Aluminum-preferred separation path + mother liquor III-B two-step oxidation co-precipitation to produce ferric phosphate, mixed acid combination three (all elements synergistic). The main differences from Example 1 are: a mixed acid combination of three was used, the sulfuric acid concentration was reduced to 22 wt%, and citric acid and oxalic acid were used for synergistic leaching; the leaching temperature was increased to 90°C; the pH of aluminum precipitation was controlled at 4.5; and ferric phosphate was precipitated using a two-step method (oxidation followed by pH adjustment and co-precipitation) instead of a one-step method.

[0082] Step 1: Preprocessing (1-1) Crushing and screening: 500kg of phosphogypsum is crushed to a particle size of <2mm and passed through a 2mm standard sieve. 494kg of material is undersized and 6kg of material is oversized and returned to the crusher.

[0083] (1-2) Weakly alkaline modification (aluminum content 0.9wt% > 0.8wt%): Add lime milk (Ca(OH)2 content 10wt%, the amount is 1.5wt% of the dry basis of phosphogypsum, i.e., add 74.1kg of lime milk, containing 7.4kg of Ca(OH)2) to 494kg of phosphogypsum, stir at room temperature for 30 minutes, the pH of the slurry rises to 8.5, which promotes the conversion of aluminum from aluminum phosphate to Al(OH)3, and improves the subsequent leaching efficiency; filter through a plate and frame filter press (pressure 0.5MPa) to obtain modified phosphogypsum filter cake with a moisture content of 14wt% and a dry basis weight of 486kg, for later use.

[0084] Step 2: Co-leaching with mixed acids (mixed acid combination 3, reducing system) (2-1) Preparation of mixed acid: Add 1700L of deionized water to a 3000L acid mixing tank, add 465kg of concentrated sulfuric acid (98wt%) to prepare a 22wt% sulfuric acid solution, and cool to 40℃; add 158kg (0.3mol / L) of citric acid monohydrate (C6H8O7·H2O) and 252kg (0.8mol / L) of oxalic acid dihydrate (H2C2O4·2H2O), and stir for 30 minutes to dissolve; add water to a final volume of 2430L. The final composition of the mixed acid is: 22wt% H2SO4 + 0.3mol / L citric acid + 0.8mol / L oxalic acid, with a liquid-to-solid ratio of 5:1 (L:kg).

[0085] (2-2) Nitrogen protection: Purge with N2 (50L / min) for 30 minutes, and maintain a flow rate of 10L / min throughout the process.

[0086] (2-3) Leaching reaction: The modified phosphogypsum (486 kg) was added to the mixed acid in three batches, and the temperature was raised to 90℃ (heating rate 2℃ / min to further improve the leaching rate of all elements of combination three). The mixture was stirred (180 rpm) for 2.5 hours. In the last 30 minutes of the reaction, 50 L of ascorbic acid solution (0.4 mol / L) was added to maintain the reducing properties of the system.

[0087] (2-4) Leaching effect: Sampling analysis: REE 0.50g / L, Al 1.88g / L, Fe 1.68g / L (Fe 2+ 1.64 g / L, Fe 3+ 0.04 g / L, Fe 2+ The proportion of rare earth elements was 97.6 wt%, and the phosphorus content was 1.85 g / L. The leaching rates were: rare earth 89.8 wt%, aluminum 90.2 wt% (the leaching rate of aluminum was improved due to the weak alkaline modification), iron 94.8 wt%, and phosphorus 89.1 wt%. The leaching rates of all elements were the highest in the three examples, demonstrating the synergistic advantages of combination three.

[0088] Step 3: Solid-liquid separation and purification of gypsum treatment (3-1) Solid-liquid separation: The leaching mixture (85℃) was filtered by plate and frame filter press while hot (pressure 0.8MPa, 40 minutes) to obtain mother liquor I 2410L (deep yellow, pH 0.7) and purified phosphogypsum filter cake 448kg (wet basis, water content 10wt%).

[0089] (3-2) Washing and resource utilization of gypsum: Washed twice with 440L of warm water (50℃) to obtain 401kg of purified gypsum (wet basis); hydrothermal crystallization (120℃, 0.2MPa, 3 hours) and drying to obtain 352kg of α-hemihydrate gypsum (dry basis); test results: flexural strength 8.4MPa, compressive strength 41MPa, whiteness 82.

[0090] Step 4: Rare Earth Separation (4-1) Oxalic acid precipitation: Adjust the pH of mother liquor I (2410L) to 2.0 with 25wt% ammonia water (consuming 14kg of ammonia water, 30 minutes), add oxalic acid solution (1.2mol / L, 226kg of oxalic acid dihydrate dissolved in 150L of water, added over 1 hour) at 65℃, and continue the reaction for 1.5 hours; centrifuge (3000rpm, 10 minutes) to obtain 12.0kg of rare earth oxalate precipitate (wet basis), and the filtrate is mother liquor II 2390L.

[0091] (4-2) Calcination enrichment: The precipitate was washed 3 times (300L each time at 50℃), dried at 80℃ for 8 hours to obtain 3.2kg, and calcined at 700℃ for 3 hours to obtain 0.79kg of mixed rare earth oxide enrichment; REO content 70.2wt%.

[0092] (4-3) CIC fine separation: The concentrate was dissolved in 2 mol / L hydrochloric acid to prepare 50 L of 10 g / L feed solution, which was then introduced into the CIC system (with the same parameters as in Example 1); the result was: Total rare earth recovery rate: 87.9 wt% Step 5: Aluminum preferential separation (pH 4.5, lower than 5.0 in Example 1) (5-1) Confirmation of iron valence state: Mother liquor II: Fe 2+ 1.62 g / L, Fe 3+ 0.04 g / L, Fe 2+ It accounts for 97.6 wt%, with an ORP of -20 mV, meeting the conditions for preferential aluminum separation.

[0093] (5-2) Sodium carbonate alkalization for aluminum precipitation (pH 4.5): Slowly add 18wt% sodium carbonate solution (dissolving 168kg Na2CO3 in 785L of water) to mother liquor II (2390L), precisely controlling the pH to 4.5±0.1 (lower than 5.0 in Example 1, to reduce Fe).2+ (To mitigate the risk of co-precipitation, while ensuring sufficient aluminum precipitation), temperature 50℃, stirring 100rpm, adjustment time 1.2 hours, reaction time 1 hour after reaching pH, stand for 1 hour, and then filter by plate and frame press.

[0094] (5-3) Preparation of alumina: 41 kg of aluminum hydroxide precipitate (wet basis) was obtained, washed 4 times (500 L each time) with deionized water (60℃), dried at 110℃ for 12 hours to obtain 6.0 kg, and calcined in a rotary kiln at 700℃ for 3 hours (heating rate 8℃ / min, calcination temperature higher than 650℃ in Example 1 to obtain γ-Al2O3), yielding 3.8 kg of alumina product; Analysis: Al2O3 purity 98.6 wt%, Fe2O3 0.3 wt%, SiO2 0.7 wt%, CaO 0.3 wt%; specific surface area 195 m² 2 / g, pore volume 0.38mL / g (phase transformation at 700℃ is γ-Al₂O₃, specific surface area is lower than that of the product at 650℃). Aluminum recovery rate: 79.5wt%. (5-4) Mother liquor III-B: 2360L, Fe 2+ 1.60 g / L, Fe 3+ 0.04 g / L, P 1.78 g / L, for later use.

[0095] Step Six: Two-step oxidation co-precipitation to prepare ferric phosphate (6-1) Oxidation step: Add hydrogen peroxide (30wt%, 42kg) to mother liquor III-B (2360L), stir at 60℃ (120rpm) for 1 hour, and react with Fe. 2+ All oxidized to Fe 3+ The ORP was detected to rise to +350mV (Ag / AgCl), Fe 2+ <0.01g / L, confirming complete oxidation.

[0096] (6-2) Co-precipitation step: Slowly add 18wt% ammonia water (with speed controlled by metering pump), accurately adjust pH to 2.2±0.1 (adjustment time 1 hour, slightly higher than 2.0 in Example 1, to promote complete co-precipitation of ferric phosphate), temperature 60℃, stir (120 rpm) for 1.5 hours, and then age for 2 hours.

[0097] (6-3) Solid-liquid separation and product preparation: Plate and frame filter press yielded 26 kg of iron phosphate precipitate (wet basis), which was washed three times with deionized water (400 L each time), dried at 80℃ for 10 hours to obtain 5.6 kg, and calcined in a tube furnace at 600℃ for 2 hours (air atmosphere, heating rate 5℃ / min) to obtain 5.3 kg of crystalline iron phosphate product; Detection: Fe 29.1 wt% (≥28 wt%), P / Fe molar ratio 1.02 (0.95–1.05), XRD showed pure FePO4 orthorhombic phase, meeting the requirements for lithium iron phosphate cathode material precursor. Iron recovery rate: 79.2 wt%, phosphorus recovery rate: 79.8 wt%.

[0098] (6-4) Phosphorus-containing mother liquor: 2330L of filtrate (P 0.36g / L) was adjusted to pH 6.5 with ammonia gas, concentrated and crystallized to obtain 1.1kg of ammonium phosphate by-product with a purity of 93.8wt%.

[0099] Step 7: Waste liquid recycling The 2550L of washing water from each process is neutralized, settled, and filtered. The supernatant is reused, and the settled sludge is solidified and sent to building materials companies.

[0100] Comparative Example 1: Single sulfuric acid leaching + traditional one-step pH precipitation of aluminum + iron and phosphorus impurity removal This comparative example uses the most common single sulfuric acid leaching process in the existing technology, without using organic complexing acids, without distinguishing the valence state of iron for stepwise separation, and removes iron and phosphorus as impurities.

[0101] Step 1: Preprocessing Following the same procedure as in Example 1, a pretreated phosphogypsum filter cake of 488 kg (dry basis) was obtained for later use.

[0102] Step 2: Single sulfuric acid leaching (2-1) Preparation of sulfuric acid solution: Prepare 2440L of 25wt% sulfuric acid solution in a 3000L acid mixing tank (add 587kg of concentrated sulfuric acid (98wt%) to 1853L of water), with a liquid-to-solid ratio of 5:1 (L:kg), without adding any organic complex acid.

[0103] (2-2) Leaching reaction: The pretreated phosphogypsum (488 kg) was added to the sulfuric acid solution in batches, heated to 85°C, and stirred (180 rpm) for 3 hours (the temperature and time were the same as in Example 1, except that the organic complex acid was removed to exclude other variables).

[0104] (2-3) Leaching effect: Sampling analysis: REE 0.28 g / L, Al 1.05 g / L, Fe 1.18 g / L, P 1.52 g / L; Leaching rate: rare earth 51.4 wt%, aluminum 51.5 wt%, iron 66.8 wt%, phosphorus 74.2 wt%. The leaching rates of rare earth and aluminum decreased by 36.8 and 33.0 percentage points respectively compared with Example 1 (88.2 wt% / 84.5 wt%), and the leaching rate of iron decreased by 24.8 percentage points, proving that organic complex acids have a significant synergistic promoting effect on multi-element leaching.

[0105] Step 3: Solid-liquid separation Plate and frame filtration yielded 2400L of mother liquor I (pH 0.6), and 456kg of purified gypsum filter cake (wet basis); gypsum was subjected to hydrothermal crystallization to obtain 355kg of α-hemihydrate gypsum (dry basis), with a flexural strength of 8.2MPa.

[0106] Step 4: Rare Earth Separation Same as step four in Example 1, oxalic acid precipitation (pH 2.0, 60℃, 1.0 mol / L oxalic acid, 1.5 hours), followed by centrifugation; due to the low rare earth leaching concentration, only 6.8 kg of rare earth oxalate precipitate (wet basis) was obtained, which, after calcination, yielded 0.46 kg of mixed rare earth oxide concentrate with a REO content of 65.2 wt%; after CIC fine separation: Total rare earth recovery: 51.1 wt% (35.6 percentage points lower than in Example 1) Step 5: Traditional pH one-step aluminum precipitation (iron and aluminum mixed precipitation) Lime slurry (Ca(OH)2, 10wt%) was added directly to mother liquor II (2380L) to adjust the pH to 5.5 in one step (Fe2+ is not distinguished). 3+ And Al 3+ Precipitation pH difference (traditional method); Fe 3+ (ORP +180mV, Fe) 3+ (mainly) and Al 3+ Simultaneous precipitation at this pH yielded 48 kg of mixed iron-aluminum hydroxide precipitate (wet basis). After washing, drying at 110℃, and calcination at 650℃, 6.1 kg of mixed iron-aluminum oxide was obtained. Testing revealed that the Al2O3 content was only 61.2 wt% (containing a large amount of Fe2O3 impurities, 32.5 wt%), making it unsuitable for use as an alumina product and only suitable for disposal as low-value waste. Mother liquor III (2340 L): Al 0.42 g / L (aluminum loss rate reached 76.3 wt%, with a large amount of aluminum remaining in the solution), Fe 0.38 g / L, P 1.50 g / L.

[0107] The aluminum products were substandard (Al2O3 purity 61.2wt%, far below the ≥98wt% standard), iron and aluminum were not effectively separated, resulting in a large loss of aluminum resources.

[0108] Step Six: Iron and Phosphorus Impurity Removal Lime slurry was added to Mother Liquor III to adjust the pH to 8.0, causing residual Fe and P to co-precipitate as iron-phosphorus slag (iron-phosphorus mixed precipitate). After filtration, it was disposed of as solid waste, without recovering any iron or phosphorus products. The phosphorus-containing mother liquor (P 0.22 g / L) was directly discharged into the wastewater treatment system. Iron recovery rate: 0 wt%, phosphorus recovery rate: 0 wt% (both were discarded as waste).

[0109] Comparative Example 2: Separation of rare earth elements by citric acid leaching + solvent extraction (single organic acid system) This comparative example uses citric acid as the leaching agent (without sulfuric acid) and employs traditional solvent extraction to separate rare earth elements.

[0110] Step 1: Preprocessing Following the same procedure as in Example 1, a pretreated phosphogypsum filter cake of 488 kg (dry basis) was obtained for later use.

[0111] Step 2: Single citric acid extraction (2-1) Preparation of citric acid solution: Dissolve 525 kg of citric acid monohydrate (C6H8O7·H2O) in 2440 L of deionized water in a 3000 L acid mixing tank to prepare a 1.0 mol / L citric acid solution (without adding sulfuric acid, the concentration of citric acid is taken as the upper limit of the scope of the claims of this invention to maximize the leaching capacity of a single organic acid), with a liquid-to-solid ratio of 5:1 (L:kg).

[0112] (2-2) Leaching reaction: The pretreated phosphogypsum (488 kg) was added to the citric acid solution in batches, heated to 85°C, and stirred (180 rpm) for 3 hours (the parameters are the same as in Example 1).

[0113] (2-3) Leaching effect: Sampling analysis: REE 0.19 g / L, Al 0.82 g / L, Fe 0.76 g / L, P 1.21 g / L; Leaching rate: rare earth 34.8 wt%, aluminum 40.2 wt%, iron 43.0 wt%, phosphorus 59.3 wt%. Although citric acid alone has a strong complexing ability, it lacks the strong proton environment provided by sulfuric acid and cannot effectively destroy the calcium sulfate lattice. The leaching rates of each element are much lower than those of the mixed acid system of this invention (Example 1: rare earth 88.2 wt%, aluminum 84.5 wt%, iron 91.6 wt%). The fundamental reason for the low leaching rate is that the acidity of citric acid (pKa1=3.13) is much weaker than that of sulfuric acid, which is insufficient to fully dissolve the rare earth elements that are isomorphically present in the calcium sulfate lattice.

[0114] Step 3: Solid-liquid separation Plate and frame filtration yielded 2390L of mother liquor I (light yellow, pH 2.8) and 453kg of purified gypsum (wet basis); the gypsum was subjected to hydrothermal crystallization to obtain 354kg of α-hemihydrate gypsum (dry basis), with a flexural strength of 8.1MPa, which meets the standard.

[0115] Step 4: Solvent extraction to separate rare earth elements (4-1) Extraction system: The extraction system of P204 (di(2-ethylhexyl)phosphoric acid, 30wt%) + sulfonated kerosene (70wt%) was adopted, with a saponification degree of 60wt% and a ratio of O / A = 1:3. The extraction was carried out in a mixing and clarifying tank with 5 stages of extraction at 20℃.

[0116] (4-2) Extraction Results: Due to the low REE concentration of only 0.19 g / L in mother liquor I (far lower than 0.48 g / L in Example 1), and the high stability constant of the complex formed by citric acid and rare earth elements, the distribution of rare earth elements to the organic phase was severely inhibited (the distribution ratio D decreased to 0.8, while D should be >10 under normal conditions); after five stages of extraction, the rare earth extraction rate was only 52.3 wt%, with a large amount of rare earth elements unable to enter the organic phase due to citric acid complexation; after back-extraction, a rare earth solution was obtained, which was precipitated with oxalic acid and calcined to obtain 0.31 kg of mixed REO with a REO content of 62.5 wt%; after CIC fine separation, a total of 474 g of single rare earth oxides was obtained. The total rare earth recovery rate was 34.8 wt% × 52.3 wt% = 18.2 wt% (the overall loss was extremely large).

[0117] (4-3) Extractant pollution problem: The residual P2O4 in the raffinate reaches 85mg / L, and the organic solvent loss is about 12kg; the wastewater COD is 1850mg / L, requiring special organic wastewater treatment facilities, generating a large amount of sludge containing organic solvents, and the disposal cost is high.

[0118] Step 5 and beyond: Aluminum, iron, and phosphorus separation Because the concentrations of aluminum (0.82 g / L) and iron (0.76 g / L) in the leachate are low, and the raffinate contains a large amount of citric acid (residual complexing agent), the citric acid has a negative effect on Al precipitation when the pH is adjusted. 3+ The continuous complexation leads to incomplete precipitation: under pH 5.5 conditions, the aluminum precipitation rate is only 55.2wt% (Al in the mother liquor is still 0.37g / L), and the purity of the precipitate Al2O3 is 85.3wt% (caused by co-precipitation of calcium citrate and iron citrate), which cannot meet the purity requirements of alumina products (≥98wt%); the iron and phosphorus concentrations are too low (Fe 0.74g / L, P 1.18g / L), making it difficult to reach the concentration threshold for the economical production of ferrous oxalate or iron phosphate products (iron ion concentration needs to be ≥1.0g / L), and therefore cannot be recovered and is disposed of as waste.

[0119] Summary of key data from Examples 1-3 and Comparative Examples 1 and 2: 。

Claims

1. A method for multi-element synergistic leaching of phosphogypsum and recovery of by-products, comprising pretreatment, leaching, and solid-liquid separation of phosphogypsum to obtain a leachate as mother liquor I and purified phosphogypsum, characterized in that, In the leaching step, a mixed acid reaction consisting of sulfuric acid and organic complex acid is used; the mother liquor I is reacted by oxalic acid precipitation, and then a second precipitate and a filtrate serving as mother liquor II are obtained after solid-liquid separation. The second precipitate is washed, dried, and calcined to obtain a mixed rare earth oxide enrichment, which is then further refined to obtain a rare earth oxide product; the pH of mother liquor II is adjusted to 4.5-5.5 for precipitation reaction, and a third precipitate and a filtrate serving as mother liquor III are obtained after solid-liquid separation. The third precipitate is washed and calcined to obtain an alumina product.

2. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 1, characterized in that, In the pretreatment step, the phosphogypsum is purified and dehydrated to obtain pretreated phosphogypsum.

3. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 1 or 2, characterized in that, In the pretreatment step, when the aluminum content of phosphogypsum exceeds 0.8 wt% (calculated as Al2O3), an alkaline material is added for weak alkaline modification, and the pH of the modified slurry is controlled between 8.0 and 9.

5.

4. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 1, characterized in that, In the leaching step, the leaching reaction temperature is 60-95℃ and the time is 2-5 hours.

5. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 1 or 4, characterized in that, In the leaching step, the sulfuric acid concentration in the mixed acid is 15-35 wt%; the organic complex acid is selected from at least one of oxalic acid, citric acid, tartaric acid, EDTA, and ascorbic acid, with a total concentration of 0.1-5.0 mol / L; the liquid-to-solid ratio (L:kg) of the mixed acid volume to the dry weight of phosphogypsum is (3-10):

1.

6. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 1 or 5, characterized in that, The mixed acid is selected from one of the following combinations. Combination 1: Sulfuric acid concentration 15-30%, organic complex acid 0.25-2.0 mol / L citric acid and / or 0.1-0.8 mol / L EDTA; Combination 2: Sulfuric acid concentration 20-35%, organic complexed acids 0.4-3 mol / L oxalic acid and 0.25-2.0 mol / L ascorbic acid; Combination 3: Sulfuric acid concentration 20-30%, organic complex acids 0.15-1.2 mol / L citric acid and 0.25-1.6 mol / L oxalic acid.

7. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 1, characterized in that, In the oxalic acid precipitation method, the pH of the mother liquor I is adjusted to 1.5-2.5, an oxalic acid solution with a concentration of 0.5-1.5 mol / L is added, and the mixture is stirred at 40-70℃ for 1-2 hours to precipitate the oxalic acid.

8. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 1 or 7, characterized in that, After obtaining the mixed rare earth oxide concentrate, the rare earth oxide products are obtained by fine separation using continuous ion chromatography or ion exchange methods.

9. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 1 or 6, characterized in that, The mother liquor II is used to separate the alumina product using any of the following methods: Fe 3+ Prior separation method: adjust the pH value of mother liquor II to 2.5-3.5, make Fe 3+ Prior hydrolysis and precipitation as iron hydroxide or coprecipitation with phosphate as iron phosphate, after solid-liquid separation, iron precipitate and mother liquor II-A are obtained; then adjust the pH value of mother liquor II-A to 4.5-5.5, make aluminum ions precipitate as aluminum hydroxide or basic aluminum carbonate, after solid-liquid separation, the precipitate is washed and calcined to obtain aluminum oxide product with purity ≥98%; the filtrate after separation of aluminum is mother liquor III-A containing Fe 2+ and P. or using aluminum preferential separation method: adding reducing organic acid or reducing agent in the later stage of leaching or in the rare earth separation step, finally keeping the iron in mother liquor II as Fe 2+ state; adjusting the pH of mother liquor II to 4.5-5.5, making Al 3+ hydrolysis and precipitation as aluminum hydroxide, after solid-liquid separation, the precipitate is washed and calcined to obtain aluminum oxide product with purity ≥98%, and the filtrate after aluminum separation is mother liquor III-B containing Fe 2+ and P.

10. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 9, characterized in that, Fe 3+ In the preferential separation method or the aluminum preferential separation method, the pH value is adjusted by adding a basic solution or passing CO2 into mother liquor II or mother liquor II-A, the basic solution is sodium carbonate or ammonia water; the reducing organic acid is ascorbic acid, and the reducing agent is sodium sulfite.

11. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 9, characterized in that, Iron and phosphorus separation is performed on either Mother Liquor III-A or Mother Liquor III-B using any of the following methods: Separation of ferrous oxalate: Add a reducing agent to mother liquor III-A or mother liquor III-B to remove residual Fe in the liquid phase. 3+ Reduced to Fe 2 + Adjust the pH to 2.5-3.5, add oxalic acid solution to precipitate, control the reaction temperature at 25-40℃, stir the reaction for 1-3 hours, then perform solid-liquid separation. After washing and drying, the precipitate yields Fe≥30wt% Fe 3+ <0.5wt% battery-grade ferrous oxalate, the filtrate is a phosphorus-containing mother liquor; Alternatively, ferric phosphate can be separated by adding an oxidizing agent to mother liquor III-A or mother liquor III-B to remove Fe. 2+ Oxidized to Fe 3+ Then, an alkaline precipitant is added to adjust the pH to 1.8-2.5, and the mixture is stirred for 1-2 hours to produce iron phosphate through co-precipitation. After aging and solid-liquid separation, the precipitate is washed and dried and used directly as an iron phosphate precursor, or it can be further calcined to obtain crystalline iron phosphate with Fe≥28wt% and P / Fe molar ratio0.95-1.05 for the preparation of lithium iron phosphate cathode materials. The filtrate is a phosphorus-containing mother liquor.

12. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 11, characterized in that, The mother liquor III-B can also be separated from ferric phosphate using a one-step method: An oxidant and an alkaline precipitant are simultaneously added to mother liquor III-B, and the Fe is precipitated at 50-70℃. 2+ Oxidized to Fe 3+ Simultaneously, it immediately co-precipitates with phosphate to form iron phosphate. After aging and solid-liquid separation, the precipitate is washed and dried and used directly as an iron phosphate precursor, or further calcined to obtain crystalline iron phosphate with Fe≥28wt% and P / Fe molar ratio0.95-1.05 for the preparation of lithium iron phosphate cathode materials. The filtrate is a phosphorus-containing mother liquor.

13. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 11, characterized in that, In the step of separating ferrous oxalate, the reducing agent is at least one of iron powder, zinc powder, ascorbic acid, and sodium sulfite, and the concentration of the oxalic acid solution is 0.5-1.5 mol / L; in the step of separating ferric phosphate or the one-step separation of ferric phosphate, the oxidizing agent is air, pure oxygen, hydrogen peroxide, or sodium hypochlorite, and the alkaline precipitant is ammonia or sodium carbonate.

14. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in claim 11, characterized in that, In the step of separating ferrous oxalate, after adding a reducing agent to mother liquor III-A or mother liquor III-B, the reaction is carried out at a controlled temperature of 40-60℃ for 0.5-2 hours to remove residual Fe. 3+ All reduced to Fe 2+ In the step of separating ferric phosphate, an oxidant is added to mother liquor III-A or mother liquor III-B, and the reaction is carried out at a controlled temperature of 50-70℃ for 0.5-1.5 hours to separate the Fe... 2+ All oxidized to Fe 3+ .

15. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in any one of claims 11-14, characterized in that, The phosphorus-containing mother liquor is further reacted with ammonia or urea to prepare ammonium phosphate or urea phosphate compound fertilizer; or the phosphorus-containing mother liquor is concentrated, and phosphoric acid or phosphate is crystallized and recovered as a supplementary phosphorus source for wet-process phosphoric acid; or the pH of the phosphorus-containing mother liquor is adjusted to 5.0-6.5 and then used directly as a soluble phosphorus source for subsequent phosphorus chemical product production.

16. The method for multi-element synergistic leaching of phosphogypsum and recovery of by-products as described in any one of claims 11-14, characterized in that, The washing water generated in each process is filtered. The primary filtrate is dephosphorylated or used to prepare phosphate. The filtered acid is sent to the return water tank. After membrane treatment or neutralization and precipitation, the supernatant is reused for phosphogypsum washing or mixed acid preparation. The precipitated sludge is dewatered and solidified, and then sent to the building materials industry for use or safely landfilled.

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