A method for recovering rare earth elements from yttrium lutetium silicate crystal waste and its application

Abstract

This invention provides a method for recovering rare earth elements from lutetium yttrium silicate crystal waste and its application. The method includes the following steps: mixing lutetium yttrium silicate crystal waste with an alkaline substance for alkali fusion, followed by water leaching and acid leaching to obtain an acid leaching solution; mixing the acid leaching solution with an inorganic polymer flocculant and an alkaline solution for preliminary desiliconization to obtain a first desiliconized solution; mixing the first desiliconized solution with a phosphorus-based extractant for extraction to remove impurities, obtaining an extract and a raffinate; performing deep desiliconization on the raffinate using resin adsorption or heteropolyacid extraction to obtain a second desiliconized solution; mixing the second desiliconized solution with a precipitant solution for precipitation to obtain a precipitate product, and then calcining the precipitate product to obtain rare earth oxides. This method can reduce the silicon content to below 1 mg / L, comprehensively recover rare earth elements such as Y and Lu, effectively reduce acid and alkali consumption, improve the comprehensive utilization rate of waste materials, and increase economic benefits.

Description

Technical Field

[0001] This invention belongs to the field of hydrometallurgy and solid waste resource utilization technology, specifically relating to a method for recovering rare earth elements from yttrium silicate lutetium crystal waste and its application. Background Technology

[0002] Rare-earth silicate scintillation crystals can convert high-energy rays / particles into ultraviolet or visible light, possessing advantages such as high density, high light output, and short decay time, and are widely used in nuclear medicine, safety detection, and high-energy physics. Lutetium yttrium silicate is the inorganic scintillation crystal material with the best overall performance, making it an ideal replacement for NaI(Tl) and BGO for SPECT and PET applications. During the production and processing of lutetium yttrium silicate, the crystal material often requires cutting, grinding, and polishing, inevitably producing defective products and scraps. These materials, like crucible bottom materials, are contaminated with organic glue, cutting oil, dust, and other metallic impurities, and cannot be directly reused for crystal preparation. If lutetium yttrium and other rare earth elements in the production waste can be recycled, production costs can be reduced, the ecological environment can be protected, and the sustainable development of the rare earth industry can be promoted.

[0003] For example, patent CN103436719 first adds sodium hydroxide or potassium hydroxide to the waste for roasting, then leaches the roasted material with water, and obtains filter residue after solid-liquid separation. Then, it is dissolved and leached with nitric acid and oxidant. Cerium is then extracted from the leachate with an extractant, while lutetium remains in the raffinate. Oxalic acid is then added to the lutetium-containing solution to obtain lutetium oxalate precipitate. After filtration and calcination, lutetium oxide is obtained. Patent CN10691478 proposes a method for recovering lutetium. This method first decomposes the waste with acid, then leaches it with water, filters it, precipitates and removes impurities, and then extracts and separates yttrium and lutetium. Yttrium chloride-rich solution is precipitated with carbonate to obtain yttrium carbonate-rich solution, while high-purity lutetium chloride solution is precipitated with oxalic acid and calcined to obtain 4N high-purity lutetium oxide. Patent CN114318018 discloses a method for removing silicon from lutetium yttrium silicate leachate. This method involves first alkaline roasting, adding hot water and stirring to react, filtering to obtain rare earth hydroxide precipitate, then washing the rare earth hydroxide 2-3 times with liquid alkali, adding pure water and stirring, heating in a water bath at 70-90℃, slowly adding hydrochloric acid to dissolve the rare earth, filtering to remove silica gel, and then extracting the leachate with an extractant prepared from N235, isooctanol, and kerosene to remove iron. After heating and concentration, sodium chloride crystallizes out, and filtration yields a pure lutetium yttrium chloride solution. Patent CN110306059 discloses a method for recovering rare earth from cerium-doped lutetium yttrium silicate waste. This method involves first alkaline fusion to obtain enriched material, then acid dissolution, followed by extraction of the acid leachate using an extractant comprising a mixture of etheramide functional ionic liquid, additives, and diluents. The organic phase is then back-extracted with water and ammonium oxalate to obtain... The rare earth enrichment solution is then roasted to obtain rare earth oxides. Patent CN110042245 discloses a method for purifying lutetium from scintillation crystals. The technical route involves first performing alkaline fusion, with the alkaline fusion agent being a mixture of any two of NaOH, NaNO3, Na2O2, and Na2CO3, followed by hydrochloric acid leaching. The leaching solution is then controlled to be a supersaturated sodium chloride solution to remove a large amount of sodium chloride. Next, the pH value is adjusted to 2.5-3.5 with liquid alkali to remove a large amount of silicon. Then, lutetium-yttrium solution is extracted and separated using P507 and C272. The purified lutetium chloride solution is then subjected to oxalic acid precipitation, calcination, and other steps to synthesize lutetium oxide. Patent CN116926352 provides a method for recycling rare earth silicate scintillation crystal waste that can reduce the roasting temperature. The method involves ball milling the waste and alkaline substances together, then roasting the ball-milled product. The roasted product is then subjected to water leaching and acid leaching treatments to obtain a rare earth-containing liquid.

[0004] In summary, most current methods for recovering rare earth elements from scintillation crystal waste involve preliminary silicon removal followed by extraction to enrich the rare earth elements. These methods do not involve direct, deep silicon removal from the rare earth leachate, and some even require complex separation processes for the rare earth elements. These recovery processes all suffer from drawbacks such as high acid and alkali consumption and high costs.

[0005] Therefore, there is an urgent need to develop a new method for recycling yttrium lutetium silicate crystal waste, in order to help with resource recycling and reduce the manufacturing cost of scintillation crystals. Summary of the Invention

[0006] To address the shortcomings of existing technologies, the present invention aims to provide a method for recovering rare earth elements from lutetium yttrium silicate crystal waste and its application. This invention utilizes a synergistic approach of preliminary silicon removal, extraction for impurity removal, and deep silicon removal to achieve a silicon content below 1 mg / L, thereby producing rare earth oxides with a purity exceeding 99.9%. This process can comprehensively recover rare earth elements such as Y and Lu, enabling the comprehensive utilization of rare earth elements in lutetium yttrium silicate crystal waste. Furthermore, this method can effectively reduce the overall acid and alkali consumption, lower the manufacturing cost of scintillation crystals, and improve the comprehensive utilization rate of waste, exhibiting significant economic cost advantages.

[0007] To achieve this objective, the present invention adopts the following technical solution:

[0008] In a first aspect, the present invention provides a method for recovering rare earth elements from yttrium lutetium silicate crystal waste, the method comprising the following steps:

[0009] (1) Mix yttrium silicate lutetium crystal waste with alkaline substances, perform alkaline melting, and then perform water leaching and acid leaching to obtain acid leaching solution;

[0010] (2) The acid leaching solution, inorganic polymer flocculant and alkaline solution are mixed to perform preliminary desiliconization to obtain the first desiliconization solution;

[0011] (3) Mix the first desiliconizing solution and phosphorus extractant, and extract and remove impurities to obtain extract and raffinate;

[0012] (4) The raffinate is subjected to deep desiliconization by resin adsorption or heteropolyacid extraction to obtain a second desiliconized solution.

[0013] (5) The second desiliconizing liquid and the precipitant solution are mixed to carry out a precipitation reaction to obtain a precipitate product. Then the precipitate product is calcined to obtain rare earth oxides.

[0014] This invention utilizes a synergistic approach of preliminary silicon removal, extraction for impurity removal, and deep silicon removal to achieve a silicon content below 1 mg / L, thereby producing rare earth oxides with a purity exceeding 99.9%. This process comprehensively recovers rare earth elements such as Y and Lu, enabling the full utilization of rare earth elements in yttrium lutetium silicate crystal waste. Furthermore, this method effectively reduces the overall acid and alkali consumption, lowers the manufacturing cost of scintillation crystals, and improves the comprehensive utilization rate of waste materials, demonstrating significant economic cost advantages.

[0015] It should be noted that the present invention does not limit the alkaline substances; for example, they may be sodium hydroxide or potassium hydroxide.

[0016] As a preferred technical solution of the present invention, the mixing process in step (1) is accompanied by ball milling.

[0017] In this invention, ball milling helps to mix the yttrium silicate lutetium crystal waste and alkaline substances more thoroughly, which is beneficial for subsequent alkaline melting.

[0018] Preferably, the rotational speed of the ball mill is 150-850 rpm, for example, it can be 150 rpm, 250 rpm, 350 rpm, 450 rpm, 550 rpm, 650 rpm, 750 rpm or 850 rpm.

[0019] Preferably, the mixing time in step (1) is 0.5-4h, for example, it can be 0.5h, 1h, 1.5h, 2h, 2.5h, 3h, 3.5h or 4h.

[0020] Preferably, the molar ratio of rare earth elements and hydroxide ions in alkaline substances in the yttrium silicate crystal waste in step (1) is 1:(3.5-6), for example, it can be 1:3.5, 1:4, 1:4.5, 1:5, 1:5.5 or 1:6, etc.

[0021] Preferably, the alkali fusion temperature in step (1) is 300-400℃, for example, it can be 300℃, 320℃, 340℃, 360℃, 380℃ or 400℃, and the alkali fusion time is 0.5-4h, for example, it can be 0.5h, 1h, 1.5h, 2h, 2.5h, 3h, 3.5h or 4h.

[0022] In this invention, alkaline melting at 300-400℃ enables yttrium silicate crystal waste to undergo a chemical reaction with alkaline substances, generating rare earth hydroxides and silicates.

[0023] Preferably, the liquid-to-solid ratio of the water immersion in step (1) is (5-20) mL:1g, for example, it can be 5 mL:1g, 10 mL:1g, 15 mL:1g or 20 mL:1g, etc.

[0024] Preferably, the water immersion temperature in step (1) is 25-60℃, for example, it can be 25℃, 30℃, 35℃, 40℃, 45℃, 50℃, 55℃ or 60℃, and the water immersion time is 10-60min, for example, it can be 10min, 20min, 30min, 40min, 50min or 60min.

[0025] In this invention, soaking in water at 25-60°C for 10-60 minutes helps to form a white precipitate of rare earth hydroxides, while most silicates and excess alkali dissolve in the aqueous solution.

[0026] Preferably, in the acid leaching process described in step (1), the acid used is an inorganic acid, which includes any one or a combination of at least two of hydrochloric acid, sulfuric acid, or nitric acid.

[0027] Preferably, the liquid-to-solid ratio of the acid leaching in step (1) is (0.5-10) mL:1g, for example, it can be 0.5mL:1g, 1mL:1g, 3mL:1g, 5mL:1g, 7mL:1g, 9mL:1g or 10mL:1g, etc.

[0028] Preferably, the acid leaching temperature in step (1) is 25-45℃, for example, it can be 25℃, 30℃, 35℃, 40℃ or 45℃, and the acid leaching time is 1-30min, for example, it can be 1min, 5min, 10min, 15min, 20min, 25min or 30min.

[0029] In this invention, acid leaching is performed at 25-45℃ for 1-30 minutes, where rare earth hydroxides react with acid to generate a rare earth acid solution, and a small amount of precious metals enter the filter residue.

[0030] As a preferred technical solution of the present invention, the inorganic polymer flocculant in step (2) includes aluminum salt-based inorganic polymer flocculant and / or iron salt-based inorganic polymer flocculant.

[0031] In this invention, aluminum salt-based inorganic polymer flocculants and / or iron salt-based inorganic polymer flocculants are used. These flocculants can directly form aluminum silicate or iron silicate precipitates with silicate ions, or they can hydrolyze to form positively charged polymers, which then react with negatively charged silica colloids to generate precipitates, thus helping to improve the desiliconization effect.

[0032] Preferably, the aluminum salt-based inorganic polymeric flocculant includes polyaluminum sulfate and / or polyaluminum chloride.

[0033] Preferably, the iron salt-based inorganic polymeric flocculant includes polyferric sulfate and / or polyferric chloride.

[0034] Preferably, the alkaline solution in step (2) includes liquid alkali.

[0035] Preferably, the volume ratio of the acid leaching solution, inorganic polymer flocculant and alkaline solution in step (2) is 1000:(3-8):(0.5-1.5), wherein the inorganic polymer flocculant can be selected from the range of "3-8", for example, 3, 4, 5, 6, 7 or 8, and the alkaline solution can be selected from the range of "0.5-1.5", for example, 0.5, 0.75, 1, 1.25 or 1.5.

[0036] In this invention, if the volume ratio of the acid leaching solution to the inorganic polymer flocculant is too small, that is, if the amount of inorganic polymer flocculant used is too large, the solution pH will decrease, the desiliconization effect will decrease, the required alkali consumption will increase, and the loss of rare earth elements entrained in the flocculant will also increase. If the volume ratio of the acid leaching solution to the inorganic polymer flocculant is too large, that is, if the amount of inorganic polymer flocculant used is too small, the desiliconization effect will decrease, increasing the difficulty of subsequent deep desiliconization.

[0037] Preferably, the reaction temperature for the initial silicon removal in step (2) is 50-80℃, for example, 50℃, 60℃, 70℃ or 80℃, and the reaction time is 0.5-2h, for example, 0.5h, 1h, 1.5h or 2h.

[0038] In this invention, if the reaction temperature for initial silicon removal is too low, the degree of polymerization of silicic acid will be low, which is not conducive to flocculation and silicon removal; if the reaction temperature for initial silicon removal is too high, the degree of polymerization of the flocculant itself will decrease, resulting in increased entrainment loss of rare earth elements.

[0039] Preferably, the pH value of the preliminary desiliconization reaction in step (2) is 3.5-5, for example, it can be 3.5, 4, 4.5 or 5, etc.

[0040] In this invention, the pH value of the reaction system is 3.5-5, which helps the flocculant and silicon to react fully and form precipitates.

[0041] As a preferred technical solution of the present invention, the phosphorus extractant in step (3) includes P204 and / or P507.

[0042] It should be noted that P204 refers to di(2-ethylhexyl) phosphate, and P507 refers to 2-ethylhexyl phosphate.

[0043] It should be noted that the solution after preliminary silicon removal still contains impurities such as Fe, Al, Zn, and Ca, so phosphorus-based extractants are needed to extract these impurities.

[0044] Preferably, a diluent is added during the mixing process in step (3), and the specific mixing method includes:

[0045] The phosphorus extractant and diluent are mixed, and the resulting mixture is then added to the first desiliconizing solution.

[0046] In this invention, the phosphorus extractant needs to be diluted with a diluent before being mixed with the first desiliconizing solution for extraction; otherwise, the flowability and phase separation clarification performance of the organic phase will deteriorate, which is not conducive to operation.

[0047] Preferably, the diluent includes any one or a combination of at least two of sulfonated kerosene, 260# solvent oil, or Escaid 110.

[0048] Preferably, the volume percentage of phosphorus extractant in the mixture is 5-25%, for example, it can be 5%, 10%, 15%, 20% or 25%.

[0049] In this invention, if the volume ratio of phosphorus extractant in the mixture is too small, the extraction effect will be unsatisfactory and the purity of rare earth will decrease; if the volume ratio of phosphorus extractant in the mixture is too large, the loss rate of rare earth during the extraction process will be large.

[0050] It should be noted that phosphorus extractants do not require saponification with alkali; they can directly extract impurities.

[0051] It should be noted that the extract is subsequently mixed with a back-extraction agent to regenerate the loaded organic phase. This back-extraction agent can be 15-30% hydrochloric acid or sulfuric acid. If the resulting back-extraction solution contains H... + Concentrations higher than 4.5 mol / L can be recycled as a back-extraction agent. If H... + If the concentration is below 4.5 mol / L, the back-extraction solution will be sent to the downstream for wastewater treatment.

[0052] As a preferred technical solution of the present invention, the specific steps of the resin adsorption method in step (4) include:

[0053] The raffinate was desiliconized by adsorption resin, resulting in a desiliconized solution and a silicon-loaded adsorption resin.

[0054] In this invention, the resin adsorption method can achieve deep silicon removal, and has the advantages of not introducing impurity ions and low energy consumption.

[0055] It should be noted that the silica-loaded adsorption resin is desorbed using an alkaline solution of 150-200 g / L (such as sodium hydroxide solution) to obtain regenerated adsorption resin and metasilicate solution. The regenerated resin can be reused, while the metasilicate solution is sent to the downstream wastewater treatment process.

[0056] Preferably, the adsorbent resin includes a zirconium-containing resin.

[0057] Preferably, the zirconium-containing resin includes HP4500 resin and / or LX-860 resin.

[0058] Preferably, the adsorption temperature for silicon removal is 25-50℃, for example, it can be 25℃, 30℃, 35℃, 40℃, 45℃ or 50℃.

[0059] In this invention, if the adsorption temperature for silicon removal is too low, the adsorption reaction rate will be reduced, resulting in an unsatisfactory silicon removal effect; if the adsorption temperature for silicon removal is too high, the resin will expand, soften, and its structure will be damaged, reducing the service life of the resin and making it difficult to stably control the adsorption effect.

[0060] Preferably, the flow rate of the adsorption resin is 0.5-2 BV / h, for example, it can be 0.5 BV / h, 1 BV / h, 1.5 BV / h or 2 BV / h.

[0061] In this invention, if the flow rate of the adsorption resin is too small, the silicon removal time will be too long and the efficiency per unit time will be low; if the flow rate of the adsorption resin is too large, the time required for the adsorption reaction will be insufficient, resulting in a decrease in its silicon removal rate.

[0062] It should be noted that BV refers to the volume of a resin column filled per unit volume.

[0063] As a preferred technical solution of the present invention, the specific steps of the heteropolyacid extraction method in step (4) include:

[0064] The raffinate and complexing agent are mixed to carry out a complexation reaction to form a heteropolyacid. The heteropolyacid is then extracted with an amine extractant to obtain a rare earth solution with deep desiliconization and a silicon-loaded organic phase.

[0065] In this invention, the heteropolyacid extraction method can achieve deep desiliconization, and has the advantages of high production efficiency and low acid and alkali consumption.

[0066] Preferably, the organic phase supporting silicon is post-processed using the following steps:

[0067] First, wash with an acidic solution with a pH of 1-4 (e.g., 1, 2, 3, or 4, etc.), then back-extract and regenerate with an alkaline solution of 50-200 g / L (e.g., 50 g / L, 100 g / L, 150 g / L, or 200 g / L, etc.). After the organic phase is regenerated, it is acidified again (e.g., with sulfuric acid or hydrochloric acid, etc.) for transformation, and then reused. The back-extraction liquid is sent to the downstream for wastewater treatment.

[0068] Preferably, the complexing agent comprises a metal element, which includes any one or a combination of at least two of the elements Mo, W, or V.

[0069] It should be noted that the complexing agent can be a sodium salt or an ammonium salt, such as sodium molybdate, ammonium molybdate, sodium tungstate, ammonium tungstate, sodium vanadate, or ammonium vanadate.

[0070] In this invention, the above-mentioned elements are used to form heteropoly acids by complexing with the raffinate, which is beneficial to significantly improve the depth of silicon removal during extraction.

[0071] Preferably, the ratio of the number of moles of the metal element to the number of moles of Si element in the raffinate is (1.3-1.5):1, for example, it can be 1.3:1, 1.35:1, 1.4:1, 1.45:1 or 1.5:1, etc.

[0072] In this invention, the ratio of the number of moles of metal elements to the number of moles of Si elements in the raffinate is (1.3-1.5):1, which can completely convert Si in the solution into heteropolysilicate.

[0073] Preferably, the temperature of the complexation reaction is 20-45℃, for example, 20℃, 25℃, 30℃, 35℃, 40℃ or 45℃, and the time is 1-2h, for example, 1h, 1.2h, 1.4h, 1.6h, 1.8h or 2h.

[0074] In this invention, if the temperature of the complexation reaction is too low, the rate of the complexation reaction will slow down and the time to transform into heteropolyacid salts will be too long; if the temperature of the complexation reaction is too high, the loss of rare earth elements during extraction will increase and the required energy consumption will increase.

[0075] Preferably, the pH value of the complexation reaction is 2-5, for example, it can be 2, 2.5, 3, 3.5, 4, 4.5 or 5.

[0076] Preferably, the amine extractant comprises any one or a combination of at least two of the following: acidified 10% N1923-20% isooctanol-sulfonated kerosene, 10% N235-20% isooctanol-sulfonated kerosene, 10% N263-20% isooctanol-sulfonated kerosene, 10% N1923-20% TBP-sulfonated kerosene, 10% N235-20% TBP-sulfonated kerosene, or 10% N263-20% TBP-sulfonated kerosene.

[0077] Preferably, the volume ratio of the amine extractant to the aqueous phase is (3-1):(1-3), wherein the amine extractant is selected in the range of "3-1", for example, 3, 2.5, 2, 1.5 or 1, and the aqueous phase is selected in the range of "1-3", for example, 1, 1.5, 2, 2.5 or 3.

[0078] It should be noted that the aqueous phase refers to the solution in which the raffinate and the complexing agent are mixed and complexed to form a heteropoly acid.

[0079] Preferably, the extraction time for extracting heteropoly acids is 5-15 min, for example, 5 min, 10 min, or 15 min.

[0080] In this invention, the extraction time for extracting heteropoly acids is 5-15 minutes, which can achieve deep removal of heteropoly acids by amine extractants.

[0081] As a preferred technical solution of the present invention, before mixing the second desiliconizing liquid and the precipitant solution in step (5), the pH value of the second desiliconizing liquid is adjusted so that the pH value of the second desiliconizing liquid is 1-2, for example, it can be 1, 1.5 or 2.

[0082] In this invention, the purpose of adjusting the pH of the second desiliconizing solution is to enable the rare earth elements in the solution to form stable precipitates with oxalic acid within the specified pH range, which is beneficial for subsequent processes to ensure the purity and yield of rare earth oxides.

[0083] Preferably, the precipitant solution in step (5) includes an oxalic acid solution.

[0084] Preferably, the concentration of the precipitant solution in step (5) is 1-1.5 mol / L, for example, it can be 1 mol / L, 1.1 mol / L, 1.2 mol / L, 1.3 mol / L, 1.4 mol / L or 1.5 mol / L, etc.

[0085] In this invention, the concentration of the precipitant solution is 1-1.5 mol / L, which can reduce the amount of precipitant consumed.

[0086] Preferably, the amount of precipitant solution used in step (5) is 1.15-1.5 times the theoretical amount, for example, it can be 1.15 times, 1.2 times, 1.35 times, 1.4 times, 1.45 times or 1.5 times, etc.

[0087] In this invention, the amount of precipitant solution used is 1.15-1.5 times the theoretical amount, which can achieve a higher rare earth precipitation rate. The theoretical amount refers to the amount of precipitant solution used to completely precipitate the second desiliconizing solution.

[0088] Preferably, the temperature of the precipitation reaction in step (5) is 60-90℃, for example, 60℃, 70℃, 80℃ or 90℃, and the precipitation reaction time is 1-6h, for example, 1h, 2h, 3h, 4h, 5h or 6h.

[0089] In this invention, if the temperature of the precipitation reaction is too low, the particle size of the precipitate will be fine and the filtration will be slow; if the temperature of the precipitation reaction is too high, the particle size of the precipitate will be too coarse.

[0090] As a preferred technical solution of the present invention, before calcination in step (5), the precipitated product is dried.

[0091] In this invention, the purpose of drying the precipitate is to remove the water physically entrained in the precipitate.

[0092] Preferably, the drying temperature is 80-120℃, for example, 80℃, 90℃, 100℃, 110℃ or 120℃, and the drying time is 4-8h, for example, 4h, 5h, 6h, 7h or 8h.

[0093] Preferably, the calcination temperature in step (5) is 850-950℃, for example, 850℃, 875℃, 900℃, 925℃ or 950℃, and the calcination time is 2-4h, for example, 2h, 2.5h, 3h, 3.5h or 4h.

[0094] In this invention, if the calcination temperature is too low, the precipitated product will not decompose completely, reducing the yield of rare earth oxides; if the calcination temperature is too high, the energy consumption will increase, and the requirements for the calcination equipment will increase.

[0095] As a preferred technical solution of the present invention, the method includes the following steps:

[0096] (1) Crush the waste yttrium lutetium silicate crystals to 30-50 mesh, then add alkaline substances and solvents and perform wet ball milling for 0.5-4 hours at a speed of 150-850 rpm. After ball milling to 200-350 mesh, perform alkaline fusion at a temperature of 300-400℃ for 0.5-4 hours to obtain the alkaline fusion product.

[0097] (2) The alkaline fusion product is subjected to water immersion at 25-60℃ for 10-60 min, with a liquid-to-solid ratio of (5-20) mL:1 g, to obtain the water-immersed product;

[0098] (3) Dry the water-soaked product, and then add inorganic acid for acid leaching. The liquid-to-solid ratio of acid leaching is (0.5-10) mL:1g, the acid leaching temperature is 25-45℃, and the acid leaching time is 1-30min to obtain acid leaching solution.

[0099] (4) The acid leaching solution, inorganic polymer flocculant and liquid alkali are mixed at a volume ratio of 1000:(3-8):(0.5-1.5) and subjected to preliminary desiliconization at 50-80℃ for 0.5-2h. The pH value of the preliminary desiliconization reaction is 3.5-5, and the first desiliconization solution is obtained.

[0100] Among them, inorganic polymeric flocculants include aluminum salt-based inorganic polymeric flocculants and / or iron salt-based inorganic polymeric flocculants;

[0101] (5) Mix the phosphorus extractant and the diluent, and then add the resulting mixture to the first desiliconizing solution for extraction and impurity removal to obtain the extract and the raffinate.

[0102] The phosphorus extractant includes P204 and / or P507, and the diluent includes any one or a combination of at least two of sulfonated kerosene, 260# solvent oil or Escaid 110. The volume percentage of the phosphorus extractant in the mixture is 5-25%.

[0103] (6) The raffinate is subjected to deep desiliconization by resin adsorption or heteropolyacid extraction to obtain a second desiliconized solution.

[0104] The specific steps of the resin adsorption method include: using zirconium-containing resin as the adsorption resin, connecting 2-4 resin columns in series, and adsorbing and removing silicon from the raffinate at 25-50℃ and a flow rate of 0.5-2 BV / h to obtain the desiliconized liquid after adsorption and the adsorption resin loaded with silicon.

[0105] The specific steps of the heteropolyacid extraction method include:

[0106] The raffinate is mixed with a complexing agent containing a metal element, wherein the metal element includes any one or a combination of at least two of the elements Mo, W or V. The complexing reaction is carried out for 1-2 hours under the conditions of pH 2-5 and temperature 20-45℃ to form a heteropoly acid. Then, an amine extractant is added at room temperature to extract the heteropoly acid for 5-15 minutes. After that, the mixture is allowed to stand and separate into layers to obtain a rare earth solution with deep desiliconization and a silicon-loaded organic phase.

[0107] The ratio of the number of moles of the metal element to the number of moles of Si element in the raffinate is (1.3-1.5):1, and the volume ratio of the amine extractant to the aqueous phase is (3-1):(1-3).

[0108] (7) Adjust the pH value of the second desiliconizing solution to 1-2, then add oxalic acid solution with a concentration of 1-1.5 mol / L and mix to carry out precipitation reaction. After filtration, wash to obtain precipitate product. Then dry the precipitate product and calcine to obtain rare earth oxides.

[0109] The amount of oxalic acid solution used is 1.15-1.5 times the theoretical amount. The precipitation reaction temperature is 60-90℃, the precipitation reaction time is 1-6h, the drying temperature is 80-120℃, the drying time is 4-8h, the calcination temperature is 850-950℃, and the calcination time is 2-4h.

[0110] Secondly, the present invention provides an application of the method described in the first aspect in the field of rare earth recycling from solid waste.

[0111] The numerical range described in this invention includes not only the point values ​​listed above, but also any point values ​​within the numerical ranges not listed above. Due to space limitations and for the sake of brevity, this invention will not exhaustively list all the specific point values ​​included in the range.

[0112] Compared with the prior art, the present invention has the following beneficial effects:

[0113] This invention utilizes a synergistic approach of preliminary silicon removal, extraction for impurity removal, and deep silicon removal to achieve a silicon content below 1 mg / L, thereby producing rare earth oxides with a purity exceeding 99.9%. This process comprehensively recovers rare earth elements such as Y and Lu, enabling the full utilization of rare earth elements in yttrium lutetium silicate crystal waste. Furthermore, this method effectively reduces the overall acid and alkali consumption, lowers the manufacturing cost of scintillation crystals, and improves the comprehensive utilization rate of waste materials, demonstrating significant economic cost advantages. Attached Figure Description

[0114] Figure 1 This is a process flow diagram of rare earth elements in yttrium silicate crystal waste provided in Example 1 of the present invention. Detailed Implementation

[0115] The technical solution of the present invention will be further illustrated below through specific embodiments. Those skilled in the art should understand that the embodiments described are merely illustrative of the present invention and should not be construed as limiting the invention in any way.

[0116] Example 1

[0117] This embodiment provides a method for recovering rare earth elements from lutetium yttrium silicate crystal waste, and its process flow diagram is shown below. Figure 1 As shown, the method includes the following steps:

[0118] (1) 10g of yttrium silicate lutetium crystal waste was crushed to 40 mesh, and then sodium hydroxide and 0.5mL of anhydrous ethanol were added for 1h of planetary mechanical wet ball milling at 500rpm. After ball milling to 300 mesh, the waste was poured into a crucible and the crucible was placed in a muffle furnace for alkali fusion at 400℃ for 2h to obtain the alkali fusion product.

[0119] Among them, the molar ratio of rare earth elements in yttrium silicate crystal waste to hydroxide ions in sodium hydroxide is 1:5;

[0120] (2) The alkaline fusion product was ground and then soaked in pure water at 50°C for 30 minutes. The liquid-to-solid ratio of the water was 20 mL: 1 g. The product was filtered and washed 4 times to obtain the water-soaked product.

[0121] (3) The water-leached product is dried at 60°C for 1 hour, and then hydrochloric acid with a concentration of 1 mol / L is added for acid leaching. The liquid-solid ratio of acid leaching is 10 mL: 1 g, the acid leaching temperature is 45°C, the acid leaching time is 10 min, and the stirring rate is 300 rpm to obtain acid leaching solution.

[0122] (4) The acid leaching solution, polyaluminum sulfate and liquid alkali are mixed in a volume ratio of 1000:5:1 and subjected to preliminary desiliconization at 60°C for 2 hours. The pH value of the preliminary desiliconization reaction is 4, and the first desiliconization solution is obtained.

[0123] (5) Mix P204 and sulfonated kerosene, then add the resulting mixture to the first desiliconizing solution for extraction and impurity removal to obtain extract and raffinate. Then, use 15% hydrochloric acid as a back-extraction agent to back-extract and regenerate the extract.

[0124] The volume percentage of P2O4 in the mixture is 15%.

[0125] (6) The raffinate is subjected to deep desiliconization by resin adsorption method. The specific steps include: using HP4500 resin as adsorption resin, three resin columns are connected in series, and the raffinate is adsorbed and desiliconized at 40°C and a flow rate of 1 BV / h to obtain a second desiliconized liquid with a Si content of less than 1 mg / L and a silicon-loaded adsorption resin. The silicon-loaded adsorption resin is desorbed by 175 g / L sodium hydroxide solution. The solid-liquid ratio of the loaded adsorption resin and the sodium hydroxide solution is 1 kg: 3 L to obtain the regenerated adsorption resin and sodium metasilicate solution.

[0126] (7) Adjust the pH value of the second desiliconizing solution to 1.5, then add oxalic acid solution with a concentration of 1.2 mol / L and mix to carry out precipitation reaction. After filtration, wash with deionized water 10 times to obtain precipitated product. Then dry the precipitated product and calcine it to obtain rare earth oxides.

[0127] The amount of oxalic acid solution used was 1.5 times the theoretical amount. The precipitation reaction temperature was 80℃ and the precipitation reaction time was 6h. The drying temperature was 100℃ and the drying time was 8h. The calcination temperature was 950℃ and the calcination time was 4h.

[0128] Example 2

[0129] This embodiment provides a method for recovering rare earth elements from yttrium lutetium silicate crystal waste, the method comprising the following steps:

[0130] (1) 10g of yttrium lutetium silicate crystal waste was crushed to 30 mesh, and then potassium hydroxide and 0.5mL of anhydrous ethanol were added for 1h of planetary mechanical wet ball milling at 500rpm. After ball milling to 300 mesh, the waste was poured into a crucible and the crucible was placed in a muffle furnace for alkali fusion at 400℃ for 2h to obtain the alkali fusion product.

[0131] Among them, the molar ratio of rare earth elements in yttrium silicate lutetium crystal waste to hydroxide ions in potassium hydroxide is 1:4.5;

[0132] (2) The alkaline fusion product was ground and then soaked in pure water at 50°C for 30 minutes. The liquid-to-solid ratio of the water was 20 mL: 1 g. The product was filtered and washed 4 times to obtain the water-soaked product.

[0133] (3) The water-leached product is dried at 60°C for 1 hour, and then hydrochloric acid with a concentration of 1 mol / L is added for acid leaching. The liquid-solid ratio of acid leaching is 10 mL: 1 g, the acid leaching temperature is 45°C, the acid leaching time is 10 min, and the stirring rate is 300 rpm to obtain acid leaching solution.

[0134] (4) The acid leaching solution, polyferric sulfate and liquid alkali are mixed in a volume ratio of 1000:5:1 and subjected to preliminary desiliconization at 60°C for 2 hours. At the same time, sodium hydroxide solution is added to make the pH value of the preliminary desiliconization reaction 5, and the first desiliconization solution is obtained.

[0135] (5) Mix P204 and 260# solvent oil, then add the resulting mixture to the first desiliconizing solution for extraction and impurity removal to obtain extract and raffinate. Then, use 15% hydrochloric acid as a back-extraction agent to back-extract and regenerate the extract.

[0136] The volume percentage of P2O4 in the mixture is 25%.

[0137] (6) The raffinate is subjected to deep desiliconization by resin adsorption method. The specific steps include: using Lanxiao resin LX-860 as adsorption resin, two resin columns are connected in series, and the raffinate is adsorbed and desiliconized at 25°C and a flow rate of 2 BV / h to obtain a second desiliconized liquid with a Si content of less than 1 mg / L and a silicon-loaded adsorption resin. The silicon-loaded adsorption resin is desorbed by 150 g / L sodium hydroxide solution. The solid-liquid ratio of the loaded adsorption resin and the sodium hydroxide solution is 1 kg: 3 L to obtain the regenerated adsorption resin and sodium metasilicate solution.

[0138] (7) Adjust the pH value of the second desiliconizing solution to 1, then add oxalic acid solution with a concentration of 1 mol / L and mix to carry out precipitation reaction. After filtration, wash with deionized water 10 times to obtain precipitated product. Then dry the precipitated product and calcine it to obtain rare earth oxides.

[0139] The amount of oxalic acid solution used was 1.5 times the theoretical amount. The precipitation reaction temperature was 80℃ and the precipitation reaction time was 6h. The drying temperature was 100℃ and the drying time was 8h. The calcination temperature was 950℃ and the calcination time was 4h.

[0140] Example 3

[0141] This embodiment provides a method for recovering rare earth elements from yttrium lutetium silicate crystal waste, the method comprising the following steps:

[0142] (1) 10g of yttrium silicate lutetium crystal waste was crushed to 50 mesh, then potassium hydroxide and 0.5mL of anhydrous ethanol were added and subjected to planetary mechanical wet ball milling for 1h at a speed of 500rpm. After ball milling to 300 mesh, the waste was poured into a crucible and placed in a muffle furnace for alkali fusion at a temperature of 400℃ for 2h to obtain the alkali fusion product.

[0143] Among them, the molar ratio of rare earth elements in yttrium silicate lutetium crystal waste to hydroxide ions in potassium hydroxide is 1:5;

[0144] (2) The alkaline fusion product was ground and then soaked in pure water at 50°C for 30 minutes. The liquid-to-solid ratio of the water was 20 mL: 1 g. The product was filtered and washed 4 times to obtain the water-soaked product.

[0145] (3) The water-leached product is dried at 60°C for 1 hour, and then hydrochloric acid with a concentration of 1 mol / L is added for acid leaching. The liquid-solid ratio of acid leaching is 10 mL: 1 g, the acid leaching temperature is 45°C, the acid leaching time is 10 min, and the stirring rate is 300 rpm to obtain acid leaching solution.

[0146] (4) The acid leaching solution, polyaluminum chloride and liquid alkali are mixed in a volume ratio of 1000:5:1 and subjected to preliminary desiliconization at 60°C for 2 hours. The pH value of the preliminary desiliconization reaction is 4, and the first desiliconization solution is obtained.

[0147] (5) Mix P204 and Escaid 110, then add the resulting mixture to the first desiliconizing solution for extraction and impurity removal to obtain extract and raffinate. Then, use 15% hydrochloric acid as a back-extraction agent to back-extract and regenerate the extract.

[0148] The volume percentage of P2O4 in the mixture is 5%.

[0149] (6) The raffinate is subjected to deep desiliconization by heteropolyacid extraction. The specific steps include: mixing the raffinate with sodium molybdate and performing a complexation reaction for 1.5 h at pH 3 and temperature 30°C to form heteropolyacid. Then, amine extractant is added at room temperature (25°C) to extract the heteropolyacid for 10 min. After standing and separating, a second desiliconization liquid with Si content less than 1 mg / L and a silicon-loaded organic phase are obtained. The silicon-loaded organic phase is first washed with dilute hydrochloric acid solution with pH 3, and then back-extracted and regenerated with 100 g / L sodium hydroxide solution. After regeneration, the organic phase is acidified and transformed again with sulfuric acid and then reused. The back-extracted liquid is sent to the downstream for wastewater treatment.

[0150] The ratio of the number of moles of Mo to the number of moles of Si in the raffinate is 1.4:1, the volume ratio of the amine extractant to the heteropoly acid is 2:2, and the amine extractant is 10% N1923-20% isooctanol-sulfonated kerosene that has undergone acidification transformation.

[0151] (7) Adjust the pH value of the second desiliconizing solution to 1.5, then add oxalic acid solution with a concentration of 1.5 mol / L and mix to carry out precipitation reaction. After filtration, wash with deionized water 10 times to obtain precipitated product. Then dry the precipitated product and calcine it to obtain rare earth oxides.

[0152] The amount of oxalic acid solution used was 1.5 times the theoretical amount. The precipitation reaction temperature was 80℃ and the precipitation reaction time was 6h. The drying temperature was 100℃ and the drying time was 8h. The calcination temperature was 950℃ and the calcination time was 4h.

[0153] Example 4

[0154] This embodiment provides a method for recovering rare earth elements from yttrium lutetium silicate crystal waste, the method comprising the following steps:

[0155] (1) 10g of yttrium lutetium silicate crystal waste was crushed to 30 mesh, and then sodium hydroxide and 0.5mL of anhydrous ethanol were added for 1h of planetary mechanical wet ball milling at a speed of 500rpm. After ball milling to 300 mesh, the waste was poured into a crucible and the crucible was placed in a muffle furnace for alkali fusion at a temperature of 400℃ for 2h to obtain the alkali fusion product.

[0156] Among them, the molar ratio of rare earth elements in yttrium silicate crystal waste to hydroxide ions in sodium hydroxide is 1:5;

[0157] (2) The alkaline fusion product was ground and then soaked in pure water at 50°C for 30 minutes. The liquid-to-solid ratio of the water was 20 mL: 1 g. The product was filtered and washed 4 times to obtain the water-soaked product.

[0158] (3) The water-leached product is dried at 60°C for 1 hour, and then hydrochloric acid with a concentration of 1 mol / L is added for acid leaching. The liquid-solid ratio of acid leaching is 10 mL: 1 g, the acid leaching temperature is 45°C, the acid leaching time is 10 min, and the stirring rate is 300 rpm to obtain acid leaching solution.

[0159] (4) The acid leaching solution, polyferric sulfate and liquid alkali are mixed in a volume ratio of 1000:5:1 and subjected to preliminary desiliconization at 60°C for 2 hours. At the same time, sodium hydroxide solution is added to make the pH value of the preliminary desiliconization reaction 5, and the first desiliconization solution is obtained.

[0160] (5) Mix P204 and sulfonated kerosene, then add the resulting mixture to the first desiliconizing solution for extraction and impurity removal to obtain extract and raffinate. Then, use 15% hydrochloric acid as a back-extraction agent to back-extract and regenerate the extract.

[0161] The volume percentage of P2O4 in the mixture is 20%.

[0162] (6) The raffinate is subjected to deep desiliconization by heteropolyacid extraction. The specific steps include: mixing the raffinate with sodium tungstate and performing a complexation reaction for 2 hours at pH 5 and temperature 45°C to form heteropolyacid. Then, amine extractant is added at room temperature (25°C) to extract the heteropolyacid for 15 minutes. After standing and separating, a second desiliconization liquid with Si content less than 1 mg / L and a silicon-loaded organic phase are obtained. The silicon-loaded organic phase is first washed with dilute hydrochloric acid solution with pH 4, and then back-extracted and regenerated with 200 g / L sodium hydroxide solution. After regeneration, the organic phase is acidified and transformed again with hydrochloric acid and then reused. The back-extracted liquid is sent to the downstream for wastewater treatment.

[0163] The ratio of the number of moles of Mo to the number of moles of Si in the raffinate is 1.4:1, the volume ratio of the amine extractant to the heteropoly acid is 1:3, and the amine extractant is 10% N1923-20% isooctanol-sulfonated kerosene that has undergone acidification transformation.

[0164] (7) Adjust the pH value of the second desiliconizing solution to 2, then add oxalic acid solution with a concentration of 1 mol / L and mix to carry out precipitation reaction. After filtration, wash with deionized water 10 times to obtain precipitated product. Then dry the precipitated product and calcine it to obtain rare earth oxides.

[0165] The amount of oxalic acid solution used was 1.5 times the theoretical amount. The precipitation reaction temperature was 80℃ and the precipitation reaction time was 6h. The drying temperature was 100℃ and the drying time was 8h. The calcination temperature was 950℃ and the calcination time was 4h.

[0166] Example 5

[0167] The difference between this embodiment and embodiment 1 is that the volume ratio of acid leaching solution, polyaluminum sulfate and liquid alkali in step (4) is 1000:10:1.

[0168] The remaining preparation methods and parameters are consistent with those in Example 1.

[0169] Example 6

[0170] The difference between this embodiment and embodiment 1 is that the volume ratio of acid leaching solution, polyaluminum sulfate and liquid alkali in step (4) is 1000:1:1.

[0171] The remaining preparation methods and parameters are consistent with those in Example 1.

[0172] Example 7

[0173] The difference between this embodiment and embodiment 1 is that the initial silicon removal temperature in step (4) is 30°C.

[0174] The remaining preparation methods and parameters are consistent with those in Example 1.

[0175] Example 8

[0176] The difference between this embodiment and embodiment 1 is that the temperature for the initial silicon removal in step (4) is 100°C.

[0177] The remaining preparation methods and parameters are consistent with those in Example 1.

[0178] Example 9

[0179] The difference between this embodiment and embodiment 1 is that the volume percentage of P2O4 in the mixture in step (5) is 1%.

[0180] The remaining preparation methods and parameters are consistent with those in Example 1.

[0181] Example 10

[0182] The difference between this embodiment and embodiment 1 is that the volume percentage of P2O4 in the mixture in step (5) is 30%.

[0183] The remaining preparation methods and parameters are consistent with those in Example 1.

[0184] Example 11

[0185] The difference between this embodiment and embodiment 1 is that the adsorption and desiliconization temperature in step (6) is 80°C.

[0186] The remaining preparation methods and parameters are consistent with those in Example 1.

[0187] Example 12

[0188] The difference between this embodiment and embodiment 1 is that the flow rate of adsorption and desilicon removal in step (6) is 5 BV / h.

[0189] The remaining preparation methods and parameters are consistent with those in Example 1.

[0190] Example 13

[0191] The difference between this embodiment and embodiment 3 is that the ratio of the number of moles of Mo to the number of moles of Si in the raffinate in step (6) is 1:1.

[0192] The remaining preparation methods and parameters are consistent with those in Example 3.

[0193] Example 14

[0194] The difference between this embodiment and embodiment 3 is that the ratio of the number of moles of Mo to the number of moles of Si in the raffinate in step (6) is 2:1.

[0195] The remaining preparation methods and parameters are consistent with those in Example 3.

[0196] Example 15

[0197] The difference between this embodiment and embodiment 3 is that the temperature of the complexation reaction in step (6) is 50°C.

[0198] The remaining preparation methods and parameters are consistent with those in Example 3.

[0199] Comparative Example 1

[0200] The difference between this comparative example and Example 1 is that step (4) is omitted, and the acid leaching solution is directly extracted to remove impurities.

[0201] The remaining preparation methods and parameters are consistent with those in Example 1.

[0202] Comparative Example 2

[0203] The difference between this comparative example and Example 1 is that step (6) is omitted.

[0204] The remaining preparation methods and parameters are consistent with those in Example 1.

[0205] Performance testing

[0206] The Si content of the products obtained by the preparation methods provided in the above embodiments and comparative examples was determined, and their purity and yield were also determined.

[0207] The test method refers to the method in GB / T 12690.7-2021 "Chemical Analysis Methods for Non-Rare Earth Impurities in Rare Earth Metals and Their Oxides - Part 7: Determination of Silicon Content", namely the molybdenum blue spectrophotometric method.

[0208] The test results are shown in Table 1.

[0209] Table 1

[0210]

[0211]

[0212] analyze:

[0213] As shown in the table above, this invention, through the synergistic combination of preliminary silicon removal, extraction for impurity removal, and deep silicon removal, achieves a silicon content below 1 mg / L, thereby producing rare earth oxides with a purity exceeding 99.9%. This process can comprehensively recover rare earths such as Y and Lu with high yields, effectively reducing the overall acid and alkali consumption, lowering the manufacturing cost of scintillation crystals, and improving the comprehensive utilization rate of waste materials, thus exhibiting significant economic cost advantages.

[0214] As can be seen from Examples 1 and 5-6, if the volume ratio of acid leaching solution to polyaluminum sulfate is too small, it will lead to a decrease in solution pH, a decrease in desiliconization effect, an increase in the loss of rare earth elements entrained in the flocculant, and a decrease in rare earth yield. If the volume ratio of acid leaching solution to polyaluminum sulfate is too large, it will lead to a decrease in desiliconization effect, an increase in the difficulty of subsequent deep desiliconization, and a decrease in the purity of rare earth oxides.

[0215] As can be seen from Examples 1 and 7-8, if the initial silicon removal temperature is too low, the degree of polymerization of silicic acid will be low, which is not conducive to flocculation and silicon removal, and the purity of rare earth oxides will decrease; if the initial silicon removal temperature is too high, the degree of polymerization of the flocculant itself will decrease, resulting in increased entrainment loss of rare earth and decreased rare earth yield.

[0216] As can be seen from Examples 1 and 9-10, if the volume ratio of P2O4 in the mixture is too small, the extraction effect will be unsatisfactory and the purity of rare earth oxides will decrease; if the volume ratio of P2O4 in the mixture is too large, the loss rate of rare earth during the extraction process will be large and the rare earth yield will decrease.

[0217] As can be seen from Examples 1 and 11-12, if the temperature for adsorption and desilication is too high, it will reduce the adsorption reaction rate, resulting in an unsatisfactory desilication effect and a decrease in the purity of rare earth oxides. If the flow rate for adsorption and desilication is too high, it will cause the resin to expand, soften, and be structurally damaged, reducing the service life of the resin, making it difficult to stably control the adsorption effect, and causing a decrease in the purity of rare earth oxides.

[0218] As can be seen from Examples 3 and 13-14, if the ratio of the number of moles of Mo to the number of moles of Si in the raffinate is too small, the purity of rare earth oxides will decrease; if the ratio of the number of moles of Mo to the number of moles of Si in the raffinate is too large, the rare earth yield will decrease.

[0219] As can be seen from Examples 1 and 15, if the temperature of the complexation reaction is too high, it will lead to increased rare earth loss during extraction, decreased rare earth yield, and increased energy consumption.

[0220] As can be seen from Example 1 and Comparative Example 1, if preliminary silicon removal is not performed, the purity of rare earth oxides will decrease significantly.

[0221] As can be seen from Example 1 and Comparative Example 2, if deep desiliconization is not performed, the purity of rare earth oxides will decrease significantly.

[0222] The applicant declares that the present invention is illustrated by the above embodiments, but the present invention is not limited to the above process steps, that is, it does not mean that the present invention must rely on the above process steps to be implemented. Those skilled in the art should understand that any improvements to the present invention, equivalent substitutions of the raw materials used in the present invention, addition of auxiliary components, selection of specific methods, etc., all fall within the protection scope and disclosure scope of the present invention.

Claims

1. A method for recovering rare earth elements from yttrium lutetium silicate crystal waste, characterized in that, The method includes the following steps: (1) Mix yttrium lutetium silicate crystal waste with alkaline substances, perform alkali fusion, and then perform water leaching and acid leaching to obtain an acid leaching solution; the temperature of alkali fusion in step (1) is 300-400℃; (2) The acid leaching solution, inorganic polymer flocculant and alkaline solution are mixed to perform preliminary desiliconization to obtain the first desiliconization solution; the volume ratio of the acid leaching solution, inorganic polymer flocculant and alkaline solution in step (2) is 1000:(3-8):(0.5-1.5); the inorganic polymer flocculant in step (2) includes aluminum salt inorganic polymer flocculant and / or iron salt inorganic polymer flocculant; the pH value of the preliminary desiliconization reaction in step (2) is 3.5-5; (3) The first desiliconizing solution and phosphorus extractant are mixed and extracted to remove impurities, resulting in extract and raffinate; the phosphorus extractant includes P204 and / or P507. (4) The raffinate is subjected to deep desiliconization by resin adsorption or heteropolyacid extraction to obtain a second desiliconized solution; the specific steps of the resin adsorption method in step (4) include: using an adsorption resin to adsorb and desiliconize the raffinate to obtain a desiliconized solution after adsorption and an adsorption resin loaded with silicon; the adsorption resin is a zirconium-containing resin; the specific steps of the heteropolyacid extraction method in step (4) include: mixing the raffinate and a complexing agent to carry out a complexation reaction to form a heteropolyacid, and then using an amine extractant to extract the heteropolyacid to obtain a rare earth solution after deep desiliconization and an organic phase loaded with silicon; the complexing agent includes a metal element, and the metal element includes any one or at least two combinations of Mo, W or V elements. (5) The second desiliconizing liquid and the precipitant solution are mixed to carry out a precipitation reaction to obtain a precipitated product. Then the precipitated product is calcined to obtain rare earth oxides.

2. The method according to claim 1, characterized in that, The mixing process described in step (1) involves ball milling.

3. The method according to claim 1, characterized in that, The alkali fusion time in step (1) is 0.5-4 hours.

4. The method according to claim 1, characterized in that, The water immersion temperature in step (1) is 25-60℃, and the immersion time is 10-60min.

5. The method according to claim 1, characterized in that, The acid immersion temperature in step (1) is 25-45℃, and the acid immersion time is 1-30min.

6. The method according to claim 1, characterized in that, The reaction temperature for the initial silicon removal in step (2) is 50-80℃, and the reaction time is 0.5-2h.

7. The method according to claim 1, characterized in that, In step (3), a diluent is also added during the mixing process. The specific mixing methods include: The phosphorus extractant and diluent are mixed, and the resulting mixture is then added to the first desiliconizing solution.

8. The method according to claim 7, characterized in that, The volume percentage of phosphorus extractant in the mixture is 5-25%.

9. The method according to claim 1, characterized in that, The adsorption temperature for silicon removal is 25-50℃.

10. The method according to claim 1, characterized in that, The flow rate of the adsorption resin is 0.5-2 BV / h.

11. The method according to claim 1, characterized in that, The ratio of the number of moles of the metal element to the number of moles of Si element in the raffinate is (1.3-1.5):

1.

12. The method according to claim 1, characterized in that, The complexation reaction is carried out at a temperature of 20-45℃ for 1-2 hours.

13. The method according to claim 1, characterized in that, The extraction time for the heteropolyacids is 5-15 min.

14. The method according to claim 1, characterized in that, Step (5) Before mixing the second desiliconizing solution and the precipitant solution, the pH value of the second desiliconizing solution is adjusted so that the pH value of the second desiliconizing solution is 1-2.

15. The method according to claim 1, characterized in that, The concentration of the precipitant solution in step (5) is 1-1.5 mol / L.

16. The method according to claim 1, characterized in that, The amount of precipitant solution used in step (5) is 1.15-1.5 times the theoretical amount.

17. The method according to claim 1, characterized in that, The precipitation reaction in step (5) is carried out at a temperature of 60-90℃ and for a time of 1-6 hours.

18. The method according to claim 1, characterized in that, Before calcination in step (5), the precipitated product is dried.

19. The method according to claim 1, characterized in that, The calcination temperature in step (5) is 850-950℃, and the calcination time is 2-4h.

20. The method according to claim 1, characterized in that, The method includes the following steps: (1) Crush the yttrium lutetium silicate crystal waste to 30-50 mesh, then add alkaline substances and solvents and perform wet ball milling for 0.5-4 hours at a speed of 150-850 rpm. After ball milling to 200-350 mesh, perform alkaline fusion at a temperature of 300-400℃ for 0.5-4 hours to obtain the alkaline fusion product. (2) The alkaline fusion product is subjected to water immersion at 25-60℃ for 10-60 min, with a liquid-to-solid ratio of (5-20) mL:1 g, to obtain the water-immersed product; (3) Dry the water-soaked product, and then add inorganic acid for acid leaching. The liquid-solid ratio of acid leaching is (0.5-10) mL:1g, the acid leaching temperature is 25-45℃, and the acid leaching time is 1-30min to obtain acid leaching solution. (4) The acid leaching solution, inorganic polymer flocculant and alkaline solution are mixed at a volume ratio of 1000:(3-8):(0.5-1.5) and subjected to preliminary desiliconization at 50-80℃ for 0.5-2h. The pH value of the preliminary desiliconization reaction is 3.5-5, and the first desiliconization solution is obtained. Among them, inorganic polymeric flocculants include aluminum salt-based inorganic polymeric flocculants and / or iron salt-based inorganic polymeric flocculants; (5) Mix the phosphorus extractant and diluent, and then add the resulting mixture to the first desiliconizing solution for extraction and impurity removal to obtain the extract and raffinate. The phosphorus extractant includes P204 and / or P507, and the diluent includes any one or a combination of at least two of sulfonated kerosene, 260# solvent oil, or Escaid110. The volume percentage of the phosphorus extractant in the mixture is 5-25%. (6) The raffinate is subjected to deep desiliconization by resin adsorption or heteropolyacid extraction to obtain a second desiliconized solution; The specific steps of the resin adsorption method include: using zirconium-containing resin as the adsorption resin, connecting 2-4 resin columns in series, and adsorbing and removing silicon from the raffinate at 25-50℃ and a flow rate of 0.5-2 BV / h to obtain the desiliconized liquid after adsorption and the adsorption resin loaded with silicon. The specific steps of the heteropolyacid extraction method include: The raffinate is mixed with a complexing agent containing a metal element, wherein the metal element includes any one or a combination of at least two of the elements Mo, W or V. The complexing reaction is carried out for 1-2 hours under the conditions of pH 2-5 and temperature 20-45℃ to form a heteropoly acid. Then, an amine extractant is added at room temperature to extract the heteropoly acid for 5-15 minutes. After that, the mixture is allowed to stand and separate into layers to obtain a rare earth solution with deep desiliconization and a silicon-loaded organic phase. The ratio of the number of moles of the metal element to the number of moles of Si element in the raffinate is (1.3-1.5):1, and the volume ratio of the amine extractant to the aqueous phase is (3-1):(1-3). (7) Adjust the pH of the second desiliconizing solution to 1-2, then add oxalic acid solution with a concentration of 1-1.5 mol / L and mix to carry out precipitation reaction. After filtration, wash to obtain precipitate product. Then dry the precipitate product and calcine to obtain rare earth oxides. The amount of oxalic acid solution used is 1.15-1.5 times the theoretical amount. The precipitation reaction temperature is 60-90℃, the precipitation reaction time is 1-6h, the drying temperature is 80-120℃, the drying time is 4-8h, the calcination temperature is 850-950℃, and the calcination time is 2-4h.

21. The application of the method as described in any one of claims 1-20 in the field of rare earth recycling from solid waste.

Images