Method for recovering rare earth and other valuable elements from solid waste

Abstract

The invention discloses a method for recovering rare earth and other valuable elements from solid waste and a complete system thereof. The method sequentially comprises the six steps of pretreatment-phase change enrichment, selective leaching, multi-stage solvent extraction, fractional precipitation and sintering, product refining and residue harmlessness, and according to the separation sequence of precious metal, rare earth, high-melting-point metal and basic metal, Nd, Pr, Dy, Ti, Zr, Sc, Fe, Au, Pd and other elements can be efficiently recycled in the same process. The treatment cost is obviously reduced; and secondary wastes are reduced. Through performance verification, the overall recovery rate of rare earth is not less than 95%, the recovery rate of precious metals is not less than 98%, and the recovery rate of matrix metals such as iron, titanium, zirconium and scandium is not less than 90%. The method disclosed by the invention is wide in applicable objects including waste magnets, electronic wastes, metallurgical slag, tailings and the like, realizes high-valued and clean production of resources, and provides a new way for green and sustainable utilization of rare earth resources.

Description

Technical Field

[0001] The present invention relates to the technical field of resource utilization of solid waste and metallurgical separation, and specifically to a multi-element collaborative recovery method and its complete process system for metallurgical slag and tailings containing rare earth and other valuable elements (such as Au, Pd, Ti, Zr, Sc, etc.). Background Art

[0002] Due to their unique 4f electron configuration, rare earth elements have excellent optical, electrical, and magnetic properties and are widely used in high-tech fields such as permanent magnet materials, luminescent materials, catalysts, and hydrogen storage alloys, and are known as the "industrial vitamins". However, the global high-grade rare earth ore resources are gradually exhausted, and a large amount of radiation and acidic wastewater are easily generated during the mining and separation processes, resulting in a significant increase in environmental pressure. At the same time, a large amount of rare earth and precious / harmful elements such as palladium, platinum, gold, titanium, zirconium, and scandium are enriched in secondary resources such as metallurgical slag and tailings, and their contents are often higher than those of the original ore, making them "urban mines" with great economic value and huge environmental protection potential.

[0003] Traditional rare earth recycling mostly adopts a single-element recovery route, where solid substances are first roasted at high temperature or acid-dissolved, and then rare earth products are obtained through solvent extraction-precipitation. This route has two major pain points: 1) Low element utilization rate: Fe, Co, Ni, Cu, Ti, Zr, precious metals, etc. in the waste materials other than rare earth are discharged as impurities or new secondary waste liquids are generated; 2) High economic and environmental protection pressure: Multiple acid-base washings result in high-salt wastewater, and the calcination method / sulfuric acid roasting method discharges fluorine-containing or S-SO x gases, and the investment and energy consumption of multi-stage separation equipment are extremely high.

[0004] In recent years, the concept of "urban mine" emphasizes the full-component recovery of multi-element-containing waste from the perspective of comprehensive utilization, turning "end treatment" into "whole-process value addition". However, due to the significant differences in the chemical properties of different elements, if they are synchronously separated in the same system, problems such as mutual interference, low separation coefficient, long process, large reagent consumption, and difficulty in large-scale production are likely to occur.

[0005] Therefore, there is an urgent need to develop a complete system with a "noble metal → rare earth → high-melting-point metal → base metal" cascade separation idea: that is, it can first extract noble metals with high priority value and high toxicity, then finely fractionate and enrich rare earth, and can also adjust the remaining liquid to suitable conditions to recover iron and base metals and make building fillers, achieving true zero waste discharge and providing a new path for the green and sustainable utilization of key metal resources such as rare earth. Summary of the Invention

[0006] The present invention aims to provide a multi-element collaborative recovery method for solid waste and its supporting system that is simple in process, highly selective, low in energy consumption, and environmentally friendly, realizing one-stop efficient separation and recovery of noble metals, rare earth, high-melting-point metals, and base metals.

[0007] The object of the present invention is achieved by the following technical solutions: "A method for recovering rare earths and other valuable elements from solid waste" includes the following steps.

[0008] S1. Pretreatment - Phase change enrichment: The rare earth-containing solid waste is crushed, pulverized, and subjected to magnetic separation / sieving to obtain a uniform material with a particle size ≤ 2 mm. Sodium sulfate and calcium fluoride fluxes are added, and it is melted at 900 - 1100 °C for 1 - 2 h, and then water quenched / gas quenched to obtain an enriched phase;

[0009] S2. Selective leaching: The enriched phase is pulverized to about 200 mesh, and a composite acid system containing chloride - sulfuric acid is used for selective leaching at 80 - 90 °C and pH 0 - 1 for 2 h, so that precious metals and matrix metals enter the leaching solution, and rare earths remain in the solid phase;

[0010] S3. Stepwise extraction:

[0011] The first extraction stage: Use an organic phase containing N - N - dioctylthiourea (DOS) to extract and recover precious metals such as gold (Au), palladium (Pd), and platinum (Pt) from the leaching solution;

[0012] The second extraction stage: Adjust the pH of the leaching solution to 1.5 - 2.0, and extract and recover rare earth elements with bis(2 - ethylhexyl) phosphate (P204);

[0013] The third extraction stage: Adjust the pH of the mother liquor to 2.5 - 3.0, and extract and recover iron, titanium, zirconium, scandium, etc. with β - diketone ligand;

[0014] The residual liquid is used to prepare Cu / Ni / Co metal powder by electrowinning / chemical replacement.

[0015] S4. Fractional precipitation and sintering: The precious metal solution, rare earth solution, and matrix metal solution obtained after back extraction in each extraction stage are respectively passed through selected precipitants to form corresponding precipitates, and the products are prepared through washing, drying, and roasting.

[0016] Add sodium sulfite / sodium bisulfite to the back extraction solution of the first extraction stage for reduction, and adjust the pH to 4.5 to precipitate Au, Pd, and Pt;

[0017] Dropwise add oxalic acid / ammonium oxalate to the rare earth back extraction solution of the second extraction stage, keep the pH = 1.8, and keep warm at 60 °C for 1 h to obtain rare earth oxalate;

[0018] Dropwise add NH3·H2O to the filtrate of the third extraction stage until the pH = 7.5 to precipitate Fe(OH)3, Ti(OH)4, Zr(OH)4, Sc(OH)4;

[0019] S5. Product Refining: The precious metal precipitate is purified by co-dissolution with HNO3-HCl. Rare earth oxalate is calcined to rare earth oxide. The precipitates of Fe, Ti, Zr, and Sc are obtained as high-purity FeCl3, TiCl4, ZrCl4, and ScCl4 through chlorination and distillation.

[0020] S6. Residue Innocuity: The residue after solid-liquid separation is ball-milled with the iron precipitation residue in a ratio of 3:7, and 10% curing agent (active C-S-H) is added to produce building materials such as subgrade fillers, realizing the full resource utilization of solid waste.

[0021] The beneficial effects of the present invention are as follows:

[0022] (1) High resource efficiency: The rare earth recovery rate is ≥95%, the precious metal recovery rate is ≥98%, and the base metal recovery rate is ≥90%.

[0023] (2) High reagent circulation rate: The recycling rates of the chlorinating agent, extractant, and cleaning solution all exceed 90%, significantly reducing the operating costs.

[0024] (3) High process integration degree: Through microchannel-modular design, the traditional 4-6h extraction is shortened to <5min, and the total pipeline length is shortened by 30%.

[0025] (4) Environmentally friendly: The waste liquid circulation rate in the whole process is >90%. After the solid waste is detoxified, it can be completely resourcefully utilized, and the SO x / NO x emission is reduced by 80%.

[0026] (5) Strong versatility: The leaching-separation conditions can be adjusted by regulating the extractant ratio and pH gradient, and it is applicable to various materials such as magnets, smelting slag, circuit boards, and waste catalysts.

[0027] Explanation of attached tables

[0028] Table 1 Comparison of Key Parameters between Examples and Comparative Examples Specific Embodiments

[0029] The technical solutions of the present invention will be described in detail below. However, it is necessary to point out here that the following embodiments are only for further illustration of the present invention and should not be construed as limiting the protection scope of the present invention. Some non-essential improvements and adjustments made by those skilled in the art based on the content of the present invention still fall within the protection scope of the present invention.

[0030] Example 1 (Oxidizing Blast Furnace Slag / Rare Earth Evaporation Residue Dust)

[0031] S1 Pretreatment - Phase Change Enrichment

[0032] 1 kg of the mixture (Fe 35%, REO 3.1%, Au 9 ppm, Pd 6 ppm, Ti 4.8%, Zr 0.6%, Sc 240 ppm) is successively subjected to mechanical crushing and vibration milling to make the particle size all less than 2 mm. Subsequently, 300 g of Na2SO4 (28 wt%) and 50 g of CaF2 (6 wt%) are successively dry-mixed into the powder. After dry-mixing for 5 min, it is put into a ceramic crucible and sent into a box-type electric furnace, and kept at 1000 °C in an air atmosphere for 1.5 h to complete melting, and then quickly poured into 20 L of cooling water for water quenching to obtain 1.22 kg of glassy enriched phase.

[0033] S2 Selective Leaching

[0034] The enriched phase is ground to about 200 mesh (about 75 μm), and a composite acid composed of "6 mol·L -1 HCl + 1 mol·L -1 H2SO4" is used as the leaching solution, and the solid-liquid ratio is 10 mL·g -1 . The leaching temperature is 85 °C, and the pH of the system is adjusted to 0.5 with a small amount of concentrated HCl, and mechanical stirring (400 rpm) is used for continuous leaching for 2 h. After filtration, about 5.8 L of leaching solution is obtained (>99% of Au / Pd / Pt and matrix Fe, etc. enter the leaching solution), and the rare earths (Nd, Pr, Dy) in the filter residue remain in the solid phase.

[0035] S3 Stepwise Extraction

[0036] 1) The first stage (noble metals)

[0037] A sulfonated kerosene organic phase with 30% volume fraction of DOS (N-N-dioctylthiourea) is used, and it is mixed with the aqueous phase at a ratio of organic phase:aqueous phase = 2:1, and countercurrent extraction is carried out at room temperature for two stages to make the extraction rates of Au, Pd, and Pt reach >99%, >99%, and 97% respectively. Then, 30 wt% HNO3 + 0.1 mol·L -1 thiourea is used for stripping at 60 °C for 10 min to obtain the noble metal stripping solution.

[0038] 2) The second stage (rare earths)

[0039] The pH of the raffinate is adjusted to 1.8, and then an organic phase of 25% volume fraction of P204 (bis(2-ethylhexyl) phosphate) and kerosene is used. The organic phase and the aqueous phase are in a ratio of 2:1, and rare earths are extracted by four-stage countercurrent extraction at 30 °C to make the total extraction rate of Nd, Pr, and Dy ≥98%. Stripping is carried out with 1 mol·L -1 HCl + 0.1 mol·L -1 (NH4)2C2O4.

[0040] 3) The third stage (Fe / Ti / Zr / Sc)

[0041] Then adjust the pH of the mother liquor to 2.8 and add 0.7 mol·L -1 The organic phase of β-diketone ligand (HDEHP) was subjected to three-stage extraction at 40°C with the ratio of organic to aqueous phase = 1.5:1, extracting 94% of Fe, 91% of Ti, 92% of Zr, and 88% of Sc. 1 mol·L was used for back extraction. - 1 H2SO4.

[0042] 4) The remaining mother liquor was continuously electrolyzed for 4 h (cathode current density 200 A·m -2 ) to obtain 18g of Cu-Ni powder.

[0043] S4 graded precipitation and sintering

[0044] For precious metal stripping solution, first add 5g·L -1 Na2SO3 and 2g·L -1 NaHSO3, then add NaOH dropwise to make the pH reach 4.5, and let it stand at 60℃ for 1h to precipitate Au-Pd-Pt co-precipitate powder;

[0045] To the rare earth stripping solution, ammonium oxalate was added dropwise to pH 1.8 and kept at 60°C for 1 hour to generate rare earth oxalates;

[0046] To the matrix metal raffinate, NH3·H2O was slowly added until the pH reached 7.5, and stirred at room temperature for 1 h to obtain a mixed precipitate of Fe(OH)3, Ti(OH)4, Zr(OH)4, and Sc(OH)4.

[0047] After filtering and washing, the three types of precipitates are calcined at 900°C, 800°C and 600°C for 2-3 hours respectively to obtain corresponding oxidation (or alloy) products.

[0048] S5 product refinement

[0049] The precious metal precipitate is reduced by HNO3-HCl co-dissolution, and then purified stepwise by electrochemical or chemical reduction to obtain 99.95% Au, 99.9% Pd, and 99.9% Pt; rare earth oxalate is calcined at 900°C for 3 hours to generate rare earth oxides with TREO≥99%; the mixed hydroxide is chlorinated and distilled, and after rectification, FeCl3, TiCl4, ZrCl4, and ScCl4 with a purity of ≥99.5% are obtained respectively.

[0050] S6 residue detoxification

[0051] 200g of the solid-liquid separation residue and 470g of the third stage iron precipitate were ball-milled at a mass ratio of 3:7 for 3 hours. 10wt% of active CSH curing agent was then added to form a 20mm diameter green body. After pressing, the green body was sintered at 950℃ for 3 hours. After 28 days of curing, the finished product had a compressive strength of 25.4MPa and a total α activity of 0.06Bq·g -1, can be used as the subgrade filler for secondary highways.

[0052] Example 2 (Baotou rare earth tailings)

[0053] S1: Crush 2 kg of tailings (REO 2.8%, Fe 26%, Ti 3.2%, Zr 0.9%, Sc 110 ppm) to ≤2 mm, mix it with Na2SO4 (30 wt%) and CaF2 (5 wt%), load it into a graphite crucible, keep it at 1050 °C for 2 h, and then let the melt flow into a 300 °C nitrogen gas cooling tower for gas quenching to obtain 1.08 kg of vitreous enriched phase.

[0054] S2: Grind the enriched phase to 200 mesh, use the composite acid composed of "5 mol·L -1 HCl + 0.8 mol·L -1 H2SO4" as the leaching solution, with a solid-liquid ratio of 8 mL·g -1 , leaching temperature of 90 °C, adjust the pH of the system to 0.7 with a small amount of concentrated HCl, and continuously leach for 2 h with mechanical stirring (400 rpm). There are no precious metals.

[0055] S3: Omit the precious metal extraction section; for the rare earth section, use 20% P204, O / A 2:1, and extract rare earth in 4 stages; for the Fe / Ti / Zr / Sc section, reduce the concentration of β-diketone ligand to 0.6 mol·L -1 , extract in 3 stages, and keep the remaining parameters unchanged.

[0056] S4 - S5: The same as Example 1 (the process without Au / Pd / Pt)

[0057] S6: Vitrify and water quench the enriched slag at 1250 °C, then sinter at 980 °C, and the compressive strength of the finished product is 23.8 MPa. The rare earth recovery rate is 95.2%.

[0058] Example 3 (Ni-Cu smelting electric furnace slag)

[0059] S1: Add 1.5 kg of Ni-Cu smelting electric furnace slag (Fe 38%, Ni 1.2%, Cu 0.9%, Au 11 ppm, PGMs 17 ppm, Sc 80 ppm), 27 wt% of Na2SO4 + 7 wt% of CaF2, load it into the furnace, melt it at 990 °C for 2 h and then water quench.

[0060] S2: After grinding the enriched phase, use "7 mol·L -1 HCl + 1.2 mol·L -1 H2SO4" as the leaching solution, leach at 80 °C, with a liquid-solid ratio of 12, pH 0.6, and leach for 2 h. The subsequent solid-liquid separation operation is the same as in Example 1.

[0061] S3: 35% DOS for 2-stage extraction of Au / PGMs; 30% P204 for 4-stage extraction of rare earths; 0.8 mol·L-1 β-diketone for extraction of Fe / Ti / Zr / Sc; electrowinning of Ni / Cu / Co from the raffinate.

[0062] S4 - S5: Same as Example 1. Recycling process of Ni / Cu anode slime for Co extraction.

[0063] S6: Sintering the residue at 930 °C, with a compressive strength of 23.1 MPa. Recovery rate of precious metals is 98.8%, and that of rare earths is 95.8%.

[0064] Example 4 (Red mud - high titanium slag mixed tailings)

[0065] S1: 1 kg of mixed tailings (Fe 31%, Ti 8.1%, REO 1.4%, Sc 260 ppm, Zr 0.5%): Add 32% Na2SO4 + 4% CaF2, melt at 1080 °C for 1.2 h and then water quench to obtain the enriched phase.

[0066] S2: 4 mol·L -1 HCl + 1.0 mol·L -1 H2SO4, at 88 °C, liquid - solid ratio of 10, pH 0.9, for 2 h.

[0067] S3: Without the precious metal section; 22% P204, O / A 2:1, 3 - stage extraction of rare earths; 0.7 mol·L -1 , 3 - stage extraction of Fe / Ti / Zr / Sc.

[0068] S4 - S5: Same as Example 1.

[0069] S6: Vitrification and water quenching of the residue at 1300 °C followed by sintering at 1000 °C, with a compressive strength of 24.7 MPa; recovery rate of rare earths is 96.9%.

[0070] Comparative Example 1 (Traditional direct acid leaching method)

[0071] The raw materials are the same as those in Example 1, crushed to ≤2 mm and ground to 200 mesh.

[0072] S1: Skip the phase transformation enrichment.

[0073] S2: Single 6 mol·L -1 HCl, at 85 °C, liquid - solid ratio of 10, pH ≈ 1.0, leaching for 2 h.

[0074] S3: Only use 25 wt% P204, O / A 2:1 for one - step extraction of rare earths, without adjusting pH in sections, nor extracting Au / Pd / Pt, Fe / Ti, etc.; Neutralize the raffinate with lime and discharge.

[0075] S4 - S5: Roast rare earth oxalate to obtain REO, and noble metals and high - melting - point metals are lost with the slag.

[0076] S6: Roast the residue in air at 850 °C without vitrification - solidification treatment. The compressive strength of the finished product is only 12.3 MPa, and the α activity is 0.11 Bq·g -1 。

[0077] Appendix Table:

[0078] Table 1 Comparison of key parameters between examples and comparative examples

[0079]

[0080]

[0081] Performance test:

[0082] Content / recovery rate of rare earth elements and TREO: ICP - OES (GB / T 18115 - 2020);

[0083] Noble metals: Fire assay - AAS (GB / T 17418.6 - 2010);

[0084] Radioactivity of the residue: Gamma spectrometry (GB / T 34500.3 - 2017);

[0085] 28 - day compressive strength of the solidified material: GB / T 50081 - 2019;

[0086] TCLP leaching toxicity: HJ / T 299 - 2007;

[0087] COD emission: Fast digestion spectrophotometry (HJ / T 399 - 2007);

[0088] The purity of rare earth and noble metal products is re - confirmed according to the corresponding chemical analysis standards of the GB / T 15000 series.

[0089] Table 2 Performance of examples and comparative examples

[0090] Components Example 1 Example 2 Example 3 Example 4 Comparative Example 1 Total rare earth recovery rate / % 96.5 95.2 95.8 96.9 72.1 Total precious metal recovery rate / % 99.1 -- 98.8 -- 65.0 Total recovery rate of Fe / Ti / Zr / Sc / % 99.95 99.93 99.9 99.94 56.5 Recycling rate of leaching solution 95-96％ 95-96％ 95-96％ 95-96％ All excluded <![CDATA[Residual radioactivity (Bq·g -1 )]]> 0.06 0.05 0.06 0.05 0.11 Compressive strength of solidified residue / MPa (28d) 25.4 23.8 23.1 24.7 12.3 <![CDATA[Residue TCLP-Pb / mg·L -1 > 0.09 0.12 0.11 0.10 1.4 <![CDATA[Residue TCLP-F / mg·L -1 > 1.1 1.3 1.2 1.2 16 Reduction of COD emission ≥40％ ≥40％ ≥40％ ≥40％ Uncontrolled

[0091] According to the test results of Comparative Example 1 and Examples 1-4 in Table 2, the Examples showed excellent performance in metal recovery, achieving high recovery rates of rare earth ≥95%, precious metals ≥98%, and Fe / Ti / Zr / Sc ≥90%, which were significantly better than those of the Comparative Example (rare earth 72%, precious metals 65%, matrix metals 56%). In terms of resource and environmental indicators, the recycling rate of the leachate in the Examples reached 95-96%, the loss of extractant was ≤1.8%, and the COD emission reduction was ≥40%, while the leachate of the Comparative Example was all discharged externally. In addition, the residue utilization effect of the Examples was good, the compressive strength of the sintered body after 28 days was 23-25 MPa, and the radioactivity was ≤0.07 Bq·g -1 , meeting the subgrade material standard, while the Comparative Example did not meet the standards in terms of strength and radioactivity. This Comparative Example verified the key role of the fluxing agents Na2SO4 / CaF2 and the melting-quenching process in forming a fragile enrichment phase and improving the selectivity of multiple elements.

[0092] The above Examples 1-4 comprehensively demonstrated the replicability and high efficiency of the method of "a method for recovering rare earth and other valuable elements from solid waste" of the present invention in different solid waste scenarios; the Comparative Example provided counter-evidence data for the key process parameters of the claims. All operations can be scaled up to pilot or industrial scale according to the original ratio.

[0093] The above description is only a preferred embodiment of the present invention. It should be understood that the present invention is not limited to the form disclosed herein, should not be regarded as excluding other embodiments, but can be used in various other combinations, modifications, and environments, and can be changed within the scope of the concept described herein through the above teachings or the techniques or knowledge in related fields. Any changes and modifications made by those skilled in the art without departing from the spirit and scope of the present invention shall fall within the protection scope of the appended claims of the present invention.

Claims

1. A method for recovering rare earths and other valuable elements from solid waste, characterized in that, It includes the following steps: S1. Pretreatment - Phase change enrichment; S2. Selective leaching is carried out at a pH of 0 - 1 in a composite acid system containing chloride - sulfuric acid; S3. The first extraction stage is carried out using DOS to recover precious metals such as gold, palladium, and platinum; S4. Adjust the pH of the leaching solution to 1.5 - 2.0, and the second extraction stage is carried out using P204 to recover rare earth elements; S5. Adjust the pH of the mother liquor to 2.5 - 3.0, and the third extraction stage is carried out to recover iron, titanium, zirconium, and scandium; S6. Precipitate and calcine each stripping solution respectively to obtain the corresponding oxide products; S7. Carry out alkali fusion - vitrification treatment on the residue to realize the resource utilization of solid waste.

2. According to the process described in claim 1, its pretreatment step is: crushing the solid waste to a particle size ≤ 2 mm and removing impurities by magnetic separation, or roasting at 350 - 450 °C in an atmosphere containing 5 - 15 vol% oxygen for 0.5 - 2 h, and adding fluxes of Na2SO4 and CaF2 with a mass ratio of (25 - 35):(3 - 7) during roasting to generate a magnetic γ - Fe2O3 phase and remove the organic coating.

3. The selective leaching conditions of the process according to claim 1 are as follows: the concentration of HCl is 4 - 10 mol·L -1 , the concentration of H2SO4 is 0.5 - 2 mol·L -1 , the liquid-solid ratio is 5 - 15 mL·L -1 , the leaching temperature is 20 - 90 °C, the time is 0.5 - 3 h, and the stirring rate is 200 - 500 rpm.

4. According to the process described in claim 1, the DOS concentration in the first extraction stage is 20 - 50%, the organic phase / water phase volume ratio is (1 - 3):1, and the number of extraction stages is 1 - 3, which is used to recover Au, Pd, and Pt.

5. According to the process described in claim 1, the P204 concentration in the second extraction stage is 10 - 40%, the organic phase / water phase volume ratio is (1 - 3):1, the operating temperature is 15 - 40 °C, and the number of extraction stages is 3 - 5, which is used to enrich rare earth elements such as La - Lu, Y, and Sc.

6. According to the process described in claim 1, the third extraction section uses a ferric sulfate-kerosene extractant system to recover Fe, Ti, Zr, and Sc, and performs back-extraction with 1 mol·L -1 H2SO4 to obtain corresponding metal ion solutions.

7. According to the process described in claim 1, its stripping - precipitation - calcination process is: a) After stripping the precious metals with concentrated nitric acid - thiourea, precipitate with Na2SO3 / NaHSO3 at pH 4 - 5, and then reduce and prepare metal powder under a hydrogen atmosphere at 1200 °C; b) The rare earths are back-extracted with 1 mol·L -1 HCl - 0.1 mol·L -1 (NH4)2C2O4, then precipitated at 50 - 70 °C with an oxalic acid / rare earth molar ratio of (2.5 - 3):1, and calcined in an oxidizing atmosphere at 800 - 1000 °C for 2 - 4 h to obtain rare earth oxides; c) Adjust the pH of Fe, Ti, Zr, and Sc to 7 - 8 with NH3·H2O to precipitate as hydroxides, and then calcine into oxides.

8. According to the process described in claim 1, the leaching residue is first subjected to alkali fusion at 450 - 550 °C and then washed with water, followed by vitrification at 1200 - 1300 °C and water quenching. Subsequently, a C-S-H activator is incorporated, and after high-speed ball milling, it is pressed and sintered at 900 - 1000 °C. The resulting solidified body has a 28-day compressive strength ≥ 20 MPa and a total alpha activity of radioactive substances ≤ 0.1 Bq·g -1 .

9. According to the process described in any one of claims 1 - 8, its comprehensive indicators are: the overall rare earth recovery rate ≥ 95%, the precious metal recovery rate ≥ 98%, the Fe / Ti / Zr / Sc recovery rate ≥ 90%; the recycling rate of the leaching solution ≥ 95%, the loss of the extractant after washing and regeneration ≤ 2%, the COD emission reduction compared with the traditional process ≥ 40%, and no fluorine - containing calcium slag is generated in the whole process.

10. A multi - element closed - loop collaborative recovery system applying the process described in any one of claims 1 - 9, which includes a pretreatment unit, a phase change furnace, a leaching kettle, a multi - stage extraction tower, a sedimentation tank, a refining furnace, and a residue solidification device connected in sequence, and each unit is connected in a loop through pipelines to enable the multi - stage reuse of the leaching solution, fluxes, and energy.

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