Method for separating and recovering valuable elements from rare earth and titanium associated resources

Abstract

The present application belongs to the field of metal resource recovery, and particularly relates to a method for separating and recovering valuable elements of rare earth and titanium associated resources; wherein the separation method is as follows: after the associated raw material containing rare earth and titanium is mixed with sulfuric acid and an additive and briquetted, roasting is performed, and then water immersion treatment is performed, to obtain rare earth-rich water immersion liquid and titanium-rich water immersion residue; the additive is organic sulfate or a water-soluble salt thereof; the weight ratio of the associated raw material, sulfuric acid and the additive is 1:0.2-0.8:0.03-0.2; the roasting temperature is 150-350 DEG C. The present application innovatively uses the additive to cooperate with sulfuric acid to perform roasting treatment on the rare earth and titanium associated resources, and the composition ratio of the additive and the roasting temperature are synergistically controlled, so that the reaction mechanism and the physicochemical structure can be optimized, and the high-selectivity water immersion separation of rare earth and titanium can be unexpectedly realized.

Description

Technical Field

[0001] This invention relates to the field of metal resource recycling technology, specifically to methods for separating and recycling valuable elements in rare earth and titanium-associated resources. Background Technology

[0002] Rare earth elements possess a unique 4f electron shell structure and excellent optical, electrical, magnetic, and catalytic properties, making them indispensable key metallic resources supporting strategic emerging industries such as modern industry, new energy, new materials, aerospace, and electronic information. However, long-term mining and smelting of rare earth ores have generated a large amount of associated and secondary rare earth and titanium resources. These materials generally exhibit characteristics of multi-element symbiosis, complex composition, diverse mineral phases, fine particle size, and dispersed occurrence. In addition to rare earth elements, they typically contain multiple valuable impurities such as titanium and silicon, offering high comprehensive utilization value but also posing significant challenges to extraction and separation.

[0003] Existing rare earth secondary resource recovery technologies still suffer from problems such as high acid consumption and high temperatures, and typically focus on constructing process routes around a single target element. For example, patent application CN1246540A discloses a new acid decomposition process for Baotou rare earth ore, but its sulfuric acid consumption is 1.1 to 1.7 times the concentrate mass, and it requires roasting at 500 to 600°C. Another example is patent application CN109112293A, which discloses a method for selectively recovering scandium from red mud, specifically reporting the recovery of Sc from red mud.

[0004] In summary, most existing technologies are designed for the extraction of a single rare earth element or the recovery of a single associated element, lacking an integrated process capable of simultaneously achieving selective separation and synergistic recovery of rare earth elements from associated elements such as titanium and silicon. Therefore, for complex secondary resource materials where rare earth and titanium coexist, there is an urgent need to establish a process method that can achieve selective separation and efficient recovery of rare earth and titanium. Summary of the Invention

[0005] In view of the problems existing in the prior art, the first objective of the present invention is to provide a method for separating valuable elements of rare earth and titanium associated resources, aiming to achieve selective separation of rare earth and titanium in the resources.

[0006] The second objective of this invention is to provide a method for recovering valuable elements from rare earth and titanium associated resources, aiming to achieve selective separation and recovery of rare earth and titanium.

[0007] This invention aims to process complex rare earth resources containing multiple coexisting components such as rare earth elements, titanium, silicon, phosphorus, and calcium. These resources have diverse components, more complex mineral phases, and closer coexistence, unlike the single rare earth concentrates or single titanium-containing materials commonly found in existing technologies. During sulfuric acid roasting / leaching, they are more prone to problems such as silica gelation, high acid consumption, difficulty in selectively separating rare earth and titanium during leaching, and mutual interference among multiple elements. This invention provides the following solution for such complex rare earth resources:

[0008] A method for separating valuable elements from rare earth and titanium associated resources involves mixing associated raw materials containing rare earth and titanium with sulfuric acid and additives, pressing them into pellets, roasting them, and then treating them with water to obtain a rare earth-rich water extract and a titanium-rich water extract residue.

[0009] The additive is an organic sulfate ester or its water-soluble salt;

[0010] The weight ratio of by-products, sulfuric acid, and additives is 1:0.2~0.8:0.03~0.2;

[0011] The roasting temperature is 150~350℃.

[0012] This invention innovatively uses the additives in combination with sulfuric acid to perform pelleting and roasting treatment on rare earth and titanium associated resources. By coordinating the proportion of additive components and the roasting temperature, the reaction mechanism and physicochemical structure can be optimized, and a highly selective water leaching separation of rare earth and titanium can be unexpectedly achieved.

[0013] The process of this invention can achieve excellent results based on multiple combined principles: First, the combination of additives and sulfuric acid can transform and improve the reaction environment of rare earth components during acidic roasting, promoting the efficient conversion of rare earth into soluble sulfates under mild conditions; Second, the additives can form an isolation film to prevent the phase transformation caused by the reaction of titanium components with sulfuric acid, thereby helping to reduce the dissolution of titanium components in the subsequent water immersion process; Third, the surface activity of the additives can reduce the liquid-solid interfacial tension, improve the wetting and dispersion uniformity of sulfuric acid in complex mineral powders, reduce the formation of unfavorable phases caused by local over-acidity and overheating, and thus establish a more uniform local reaction environment.

[0014] In this invention, the associated raw material containing rare earth and titanium is at least one of solid waste and minerals containing rare earth and titanium.

[0015] The rare earth elements include at least one element selected from Ce, La, Pr, Nd, Sm, Gd, Y, Dy, Eu, and Er.

[0016] In this invention, the associated raw materials containing rare earth and titanium also contain silicon, and the silicon, through the roasting and water leaching treatment, is enriched together with titanium in the water leaching residue.

[0017] Furthermore, the associated raw materials are high-silicon and high-phosphorus systems that are difficult to process in the industry. For example, by mass content, REO: 25~50%, TiO2: 5~12%, SiO2: 10~18%, P2O5: 28~38%, CaO: 1~5%, TFe: 2~6%; preferably REO: 35~40%, TiO2: 7~9%, SiO2: 12~15%, P2O5: 32~35%, CaO: 2~3%, TFe: 3~4%.

[0018] In this invention, the sulfuric acid is concentrated sulfuric acid;

[0019] The additive is at least one of diol organic sulfate esters and polyol organic sulfate esters.

[0020] The diol organic sulfate ester refers to a sulfate ester formed from a diol and sulfuric acid. The polyol organic sulfate ester refers to a sulfate ester formed from a polyol and sulfuric acid. The diol refers to an alcohol with 2 to 8 carbon atoms and 2 hydroxyl groups; further, it can be propylene glycol. The polyol refers to a polyol with 2 to 8 carbon atoms (preferably 3 to 4) and 3 to 6 hydroxyl groups; further, it can be glycerol.

[0021] Preferably, the use of polyol sulfates can help to further improve the thermal modification effect of associated raw materials and further improve the separation selectivity of rare earth and titanium.

[0022] Preferably, the organic sulfate esters include monosulfates and / or disulfates of polyols. Further, the organic sulfate esters include glycerol monosulfates and / or glycerol disulfates.

[0023] More preferably, the organic sulfate esters include monosulfates and disulfates of polyols in a weight ratio of 55-95:5-45, or more preferably 75-85:15-25. Preferred combinations of organic sulfate esters can achieve better synergistic modification effects, contributing to further enhancing the separation selectivity of rare earth elements and titanium.

[0024] In this invention, the weight ratio of the associated raw material to sulfuric acid and additives is 1:0.3~0.6:0.05~0.15; more preferably, it is 1:0.45~0.55:0.08~0.12. Within this preferred range, a better synergistic modification effect can be obtained, which helps to further enhance the separation selectivity of rare earth elements and titanium.

[0025] The thickness of the briquettes is controlled at 5~20 mm, the particle size is controlled at 5~15 mm, the briquetting pressure is 100~500 MPa, and the holding time is 30~60 s.

[0026] In this invention, the calcination temperature is 180~220℃; in the preferred range, good selectivity for the separation of rare earth elements and titanium can be obtained at a low temperature.

[0027] And / or, the roasting time is 0.5 to 3 hours; further, it can be 1 to 1.5 hours.

[0028] In this invention, a pre-calcination process is included before calcination, wherein the pre-calcination temperature is 80~120℃. The preferred two-stage low-temperature sintering process, combined with the additives described in this invention, helps to further synergistically optimize the modification effect and further enhance the separation selectivity of rare earth elements and titanium.

[0029] In this invention, the pre-calcination time can be 0.2~2h; more specifically, it can be 0.5~1h.

[0030] In this invention, with the aid of the additives, the combination of pre-calcination and calcination can further improve the thermal modification effect of the additives, which helps to further improve the separation selectivity of rare earth and titanium.

[0031] In this invention, the liquid-to-solid ratio during the water immersion process is 5~20mL / g, and can be further 8~12mL / g;

[0032] The water immersion temperature is 10~95℃; further, it can be room temperature.

[0033] The water immersion is carried out under mechanical stirring, wherein the mechanical stirring speed is 400~600 r / min.

[0034] This invention also provides a method for recovering valuable elements from rare earth and titanium associated resources. The method described in this invention is used to roast and leach associated raw materials of rare earth and titanium to obtain a water leaching solution and a water leaching residue.

[0035] The aqueous extract is subjected to precipitation treatment, and the solid-liquid separation is performed to obtain rare earth products;

[0036] The water-leached residue was subjected to alkaline leaching treatment, and the solid-liquid separation was performed to obtain sodium titanate product.

[0037] In this invention, the pH during the precipitation process is 2-3.5;

[0038] The precipitant used in the precipitation process is at least one of oxalic acid, water-soluble oxalate, water-soluble carbonate, and water-soluble bicarbonate.

[0039] The precipitant used in the precipitation process is 1 to 2 times the theoretical molar amount of rare earth precipitation; further, it can be 1.4 to 1.6 times.

[0040] In this invention, the alkaline solution for alkaline leaching is an aqueous solution of alkali metal hydroxide; the solute concentration of the alkaline solution can be reasonably adjusted as needed, for example, it can be 1~10M.

[0041] The alkaline immersion temperature is 25~120℃, and the immersion time is 0.5~4h;

[0042] And / or, the alkaline leaching process is carried out with microwave assistance; the microwave power can be 500~1000W.

[0043] And / or, the water-leached residue also contains silicon; after alkaline leaching treatment, the silicon is enriched in the alkaline leaching solution;

[0044] And / or, acid treatment of the alkaline leaching solution yields silica.

[0045] Beneficial effects

[0046] For difficult-to-process rare earth and titanium associated resources, an innovative method is used to combine additives and sulfuric acid for roasting, which helps to improve the uniformity of the roasting and activation process, reduce acid consumption, and achieve effective selective activation of rare earth components in complex rare earth and titanium associated resources at a lower temperature.

[0047] Based on the separation of valuable elements, this invention enables the value-added utilization of elements such as titanium, silicon, and rare earths, avoiding the loss of other valuable elements caused by the current process of only recovering one element. At the same time, it reduces the amount of final solid waste and improves the comprehensive utilization level of rare earth and titanium associated resources. Attached Figure Description

[0048] Figure 1 This is a process flow diagram of the present invention.

[0049] Figure 2 The images shown are SEM-EDS diagrams of the raw materials in Example 1; where (a) is an SEM image; (b) is an EDS image of Ce; (c) is an EDS image of La; (d) is an EDS image of Pr; (e) is an EDS image of Sm; (f) is an EDS image of Gd; (g) is an EDS image of Nd; (h) is an EDS image of Na; (i) is an EDS image of Ti; (j) is an EDS image of P; (k) is an EDS image of O; and (l) is an EDS image of Si.

[0050] Figure 3 The images shown are SEM-EDS images of the titanium-rich slag after water immersion in Example 1; where (a) is an SEM image; (b) is an EDS image of Si; (c) is an EDS image of O; (d) is an EDS image of Ti; (e) is an EDS image of P; and (f) is an EDS image of Na. Detailed Implementation

[0051] The present invention will be further described below with reference to embodiments.

[0052] This invention provides an optional method for the synergistic separation and recovery of valuable elements from complex rare earth and titanium associated resources (see flowchart). Figure 1 The method includes the following steps:

[0053] Step (1): The raw material is mixed with concentrated sulfuric acid and additives, pressed into a pellet, and then roasted to obtain the roasted product;

[0054] Step (2) involves leaching the roasted product in water and separating the solid and liquid components to obtain a rare earth-rich leachate (water leaching solution) and a titanium-rich leachate residue (water leaching residue).

[0055] Step (3): After adjusting the pH of the rare earth-rich leachate obtained in step (2), a precipitant is added to obtain rare earth products.

[0056] Step (4) involves alkaline leaching and desilication of the titanium-rich leaching residue, as well as directional transformation of the titanium phase, to obtain a silicon-containing alkaline leaching solution and sodium titanate, thereby achieving titanium-silicon separation.

[0057] Step (5): Adjust the pH of the silicon-containing alkaline leaching solution obtained in step (4) to obtain the silica product.

[0058] The complex rare earth and titanium associated resources contain one or more of the following: rare earth phosphates, fluorocarbon cerium ore, sodium titanium phosphate, titanium hydrogen phosphate, silicon dioxide, and silica gel.

[0059] The briquetting conditions described in step (1) are: pressure 100~500 MPa, briquet diameter 10 mm, briquet height 5~20 mm, and holding time 30~60 s. The additive is polyol organic sulfate; the amount of polyol organic sulfate is 0.03~0.2% of the raw material mass; the acid-to-ore ratio is 0.2~0.8:1; and the calcination temperature is 150~350℃.

[0060] In this invention, the water immersion temperature is 10~95℃, the liquid-to-solid ratio is 5~20:1, the pH of the resulting leachate is 0.5~1.0, and titanium and silicon are mainly retained in the water immersion residue.

[0061] In this invention, the rare earth-rich leachate obtained in step (2) is first adjusted to pH 2-3.5, and then a precipitant is added to prepare a mixed rare earth precipitate. The precipitant is oxalic acid, sodium bicarbonate, ammonium bicarbonate, sodium carbonate, magnesium bicarbonate, etc., and the precipitant is 1-2 times the theoretical molar amount of rare earth precipitation.

[0062] In this invention, the alkaline treatment in step (4) uses an alkaline solution, namely NaOH, with a mass fraction of 10-30%, a liquid-to-solid ratio of 5-20:1, a leaching temperature of 25-120℃, and a leaching time of 0.5-4h. Preferably, it is carried out under a microwave field with a microwave output power of 500-1000 W, using continuous microwave or pulsed microwave with a pulse duty cycle of 10%-90%. The titanium phase is directionally converted into sodium titanate and then dried, allowing silicon to enter the solution, thus achieving silicon-titanium separation.

[0063] In this invention, the alkaline leaching solution obtained in step (4) is used to prepare silica by adjusting the pH. The pH is adjusted to 6.5-7.5 by introducing CO2, adding hydrochloric acid, sulfuric acid, etc., and then dried to obtain silica.

[0064] The acid-to-ore ratio refers to the weight ratio of concentrated sulfuric acid to the raw material to be treated, which contains rare earth elements and titanium.

[0065] In this invention, the element content in the raw materials refers to the weight percentage.

[0066] Unless otherwise stated, the starting processing temperature in steps (2) to (5) in the following operations refers to room temperature.

[0067] Example 1

[0068] The main components of the raw material are: REO 37.90%, TiO2 7.98%, SiO2 13.19%, P2O5 33.55%, CaO 2.33%, and TFe 3.45%. SEM-EDS data are shown below. Figure 2 .

[0069] (1) Calcination: The raw materials, concentrated sulfuric acid, and additives are mixed and stirred evenly at a mass ratio of 1:0.5:0.08. The additives include glyceryl monosulfate and glyceryl disulfate at a mass ratio of 8:2. After mixing and stirring evenly, the mixture is pressed into a lump with a diameter of 10 mm and a height of 15 mm under a pressure of 300 MPa and a holding time of 30 s. The lump is then acid-calcined in a muffle furnace. First, the temperature is raised to 100℃ (marked as T1) at 5℃ / min and held for 30 min. Then, the temperature is raised to 200℃ (marked as T2) at 10℃ / min and held for 60 min to obtain the calcined material.

[0070] (2) Water leaching separation: The calcined material was added to ultrapure water at a liquid-to-solid ratio of 10:1 (mL / g), and stirred and leached for 0.5 h at 25℃ and a stirring speed of 500 r / min. Solid-liquid separation was then carried out to obtain rare earth-rich leachate and titanium-rich leachate residue (SEM-EDS figure shown). Figure 3 ).

[0071] (3) Rare earth recovery: Add magnesium oxide to the rare earth-rich leachate to adjust the pH to 2, then add sodium bicarbonate solution as a precipitant, control the molar ratio of sodium bicarbonate to rare earth ions to be 1.5:1, and carry out precipitation, filtration, washing and drying to obtain mixed rare earth carbonate products.

[0072] (4) Separation of silicon and titanium: The obtained titanium-rich leaching residue was added to a 20% NaOH solution for alkaline leaching treatment with a liquid-to-solid ratio of 10:1 (mL / g). The treatment was carried out for 1 hour under microwave power of 800W. After alkaline leaching, silicon-containing alkaline leaching solution and sodium titanate were obtained and dried.

[0073] (5) Silicon recovery: The silica-containing alkaline leaching solution is purged with CO2 to adjust the pH to 7 (±0.5), and after aging, filtration and drying, the silica product is obtained.

[0074] According to the test results, the rare earth leaching rate was 96.68%, the titanium leaching rate was 0.83%, and the silicon leaching rate was 1.67% during the water leaching process in this embodiment; the total rare earth recovery rate in rare earth carbonate products was 92.77%, the total titanium recovery rate in sodium titanate was 93.63%, and the total silicon recovery rate in silica was 92.25%.

[0075] Example 2

[0076] Compared with Example 1, the only difference is that the type and amount of additive in step (1) are changed. All other operations and parameters are the same as in Example 1. The experimental groups and results are as follows:

[0077] Group A: The additive is glyceryl monosulfate;

[0078] Group B: The additive is glyceryl disulfate;

[0079] Group C: The additive is 1,2-propanediol monosulfate;

[0080] Group D: The additive is 1,2-propanediol disulfide;

[0081] The test results are as follows:

[0082] Group A: During the water leaching process, the rare earth leaching rate was 93.76%, the titanium leaching rate was 0.95%, and the silicon leaching rate was 1.49%; the total rare earth recovery rate in rare earth carbonate products was 89.34%, the total titanium recovery rate in sodium titanate was 91.08%, and the total silicon recovery rate in silica was 91.47%.

[0083] Group B: During the water leaching process, the rare earth leaching rate was 96.55%, the titanium leaching rate was 4.31%, and the silicon leaching rate was 5.15%; the total rare earth recovery rate in rare earth carbonate products was 91.21%, the total titanium recovery rate in sodium titanate was 89.42%, and the total silicon recovery rate in silica was 86.29%.

[0084] Group C: During the water leaching process, the rare earth leaching rate was 79.74%, the titanium leaching rate was 3.28%, and the silicon leaching rate was 6.03%; the total rare earth recovery rate in rare earth carbonate products was 78.68%, the total titanium recovery rate in sodium titanate was 89.36%, and the total silicon recovery rate in silica was 85.73%.

[0085] Group D: During the water leaching process, the rare earth leaching rate was 84.86%, the titanium leaching rate was 5.71%, and the silicon leaching rate was 6.54%; the total rare earth recovery rate in rare earth carbonate products was 79.35%, the total titanium recovery rate in sodium titanate was 88.74%, and the total silicon recovery rate in silica was 85.16%.

[0086] As can be seen from Examples 1 and 2, the structure of the additives has certain differences in the modification results of calcination. Using polyol sulfates (such as glyceryl sulfates) can achieve better thermal modification effects. In addition, the combination of glyceryl monosulfate and glyceryl disulfate can further enhance the modification effect and improve the separation selectivity of rare earth, titanium, and silicon.

[0087] Example 3

[0088] Compared with Example 1, the only difference is that the proportion of additives in step (1) is changed, the total amount of additives added, and other operations and parameters are the same as in Example 1; the experimental group is:

[0089] Group A: The mass ratio of glyceryl monosulfate to glyceryl disulfate is 9:1;

[0090] Group B: The mass ratio of glyceryl monosulfate to glyceryl disulfate is 55:45;

[0091] The results for each group are as follows:

[0092] Group A: During the water leaching process, the rare earth leaching rate was 94.41%, the titanium leaching rate was 0.72%, and the silicon leaching rate was 1.75%; the total rare earth recovery rate in rare earth carbonate products was 90.23%, the total titanium recovery rate in sodium titanate was 91.59%, and the total silicon recovery rate in silica was 91.38%.

[0093] Group B: During the water leaching process, the rare earth leaching rate was 97.12%, the titanium leaching rate was 5.58%, and the silicon leaching rate was 6.41%; the total rare earth recovery rate in rare earth carbonate products was 91.03%, the total titanium recovery rate in sodium titanate was 87.62%, and the total silicon recovery rate in silica was 86.81%.

[0094] As can be seen from Examples 1 and 3, using glycerol monosulfate and glycerol disulfate in a weight ratio of 8:2 as additives can significantly improve the rare earth leaching rate and total recovery rate while maintaining low Ti and Si leaching, resulting in the best overall effect.

[0095] Example 4

[0096] Compared with Example 1, the only difference is that the amounts of additives and concentrated sulfuric acid relative to the raw materials in step (1) are changed. The experimental groups are as follows:

[0097] Group A: The raw materials, concentrated sulfuric acid, and additives are mixed in a mass ratio of 1:0.8:0.08;

[0098] Group B: Raw materials, concentrated sulfuric acid, and additives are mixed in a mass ratio of 1:0.5:0.15;

[0099] All other conditions in each group are the same as in Example 1;

[0100] The results of the tests for each group are as follows:

[0101] Group A: During the water leaching process, the rare earth leaching rate was 97.79%, the titanium leaching rate was 6.29%, and the silicon leaching rate was 8.08%; the total rare earth recovery rate in rare earth carbonate products was 92.66%, the total titanium recovery rate in sodium titanate was 85.35%, and the total silicon recovery rate in silica was 84.62%.

[0102] Group B: During the water leaching process, the rare earth leaching rate was 83.31%, the titanium leaching rate was 0.36%, and the silicon leaching rate was 0.82%; the total rare earth recovery rate in rare earth carbonate products was 79.98%, the total titanium recovery rate in sodium titanate was 91.24%, and the total silicon recovery rate in silica was 90.07%.

[0103] Example 5

[0104] Compared with Example 1, the only difference is that the roasting temperature and time in step (1) are changed. The distinguishing features are: the roasting temperature T1 is 110℃; the roasting temperature T2 is 300℃, and the holding time is 30min; all other conditions of each group are the same as those of Example 1.

[0105] The tests showed that the leaching rate of rare earth elements was 97.52%, the leaching rate of titanium was 5.34%, and the leaching rate of silicon was 8.82% during the water leaching process. The total recovery rate of rare earth elements in rare earth carbonate products was 92.41%, the total recovery rate of titanium in sodium titanate was 84.86%, and the total recovery rate of silicon in silica was 81.35%.

[0106] Example 6

[0107] Compared with Example 1, the only difference is that the water immersion liquid-solid ratio in step (2) is 15 ml / g and the time is 0.6 h;

[0108] Step (3) Rare earth recovery: Add magnesium oxide to the rare earth-rich leachate to adjust the pH to 2.5, then add sodium bicarbonate solution as a precipitant, control the molar ratio of sodium bicarbonate to rare earth ions to be 1.15:1, and carry out precipitation, filtration, washing and drying to obtain mixed rare earth carbonate products.

[0109] Step (4) Silicon-titanium separation: The obtained titanium-rich leaching residue is added to a 15% NaOH solution for alkaline leaching treatment with a liquid-to-solid ratio of 15:1 (mL / g). The treatment is carried out for 2 hours under microwave power of 600W. After alkaline leaching, a silicon-containing alkaline leaching solution and sodium titanate are obtained and dried.

[0110] Step (5) Silicon recovery: The silica-containing alkaline leaching solution is purged with CO2 to adjust the pH to 6.5~7.0, and after aging, filtration and drying, the silica product is obtained.

[0111] According to the test results, the rare earth leaching rate was 96.72%, the titanium leaching rate was 0.74%, and the silicon leaching rate was 1.32% during the water leaching process in this embodiment. The total rare earth recovery rate in rare earth carbonate products was 93.10%, the total titanium recovery rate in sodium titanate was 91.05%, and the total silicon recovery rate in silica was 90.27%.

[0112] Example 7

[0113] Compared with Example 1, the only difference is that step (4) is not microwave-assisted, while the other operations and parameters are the same as in Example 1;

[0114] The tests showed that the total recovery rate of titanium in sodium titanate was 86.59%, and the total recovery rate of silicon in silica was 85.42%.

[0115] Example 8

[0116] Compared to Example 1, the only difference is that the T1 heat preservation process was not performed. All other operations and parameters are the same as in Example 1.

[0117] The tests showed that the leaching rate of rare earth elements was 88.74%, the leaching rate of titanium was 0.42%, and the leaching rate of silicon was 1.36% during the water leaching process. The total recovery rate of rare earth elements in rare earth carbonate products was 84.91%, the total recovery rate of titanium in sodium titanate was 91.68%, and the total recovery rate of silicon in silica was 90.91%.

[0118] Comparative Example 1

[0119] Compared with Example 1, the only difference is that the sulfuric acid roasting step (1) is not performed, but the raw material is directly leached with 1M sulfuric acid solution (the liquid-to-solid ratio of the leaching is 10 mL / g), the temperature is 25℃, and the leaching time is 0.5h, while the other operating conditions remain the same.

[0120] According to the test results, the rare earth leaching rate in Comparative Example 1 was 26.68%, the titanium leaching rate was 1.32%, and the silicon leaching rate was 1.43% during the sulfuric acid leaching process.

[0121] Comparative Example 2

[0122] Compared with Example 1, the only difference is that no additives were added in step (1), and the acid-to-ore ratio was changed. The experimental groups were as follows:

[0123] Group A: No additives added, acid-to-mineral ratio is 0.5:1;

[0124] Group B: No additives added, acid-to-mineral ratio is 1.5:1;

[0125] All other conditions for each group were the same as in Example 1.

[0126] Test results:

[0127] In Group A, the leaching rates of rare earth elements were 64.86%, titanium was 1.39%, and silicon was 3.68% during the water leaching process. The total recovery rate of rare earth elements in rare earth carbonate products was 60.32%, the total recovery rate of titanium in sodium titanate was 87.51%, and the total recovery rate of silicon in silica was 85.28%.

[0128] In Group B, the leaching rates of rare earth elements (REEs) were 95.25%, titanium 9.60%, and silicon 15.66% during water leaching. The total rare earth recovery rate in rare earth carbonate products was 94.32%, in sodium titanate 80.87%, and in silica 75.64%. These results indicate that without additives, rare earth activation is insufficient at low acid ratios. While increasing the acid-to-ore ratio improves REE leaching rates, it also increases titanium and silicon leaching rates, suggesting that additives are beneficial for achieving both efficient REE leaching and selectivity at lower acid consumption.

[0129] Comparative Example 3

[0130] Compared with Example 1, the only difference is that concentrated sulfuric acid is not added during roasting in step (1), and all other operations and parameters are the same as in Example 1;

[0131] The tests showed that the leaching rate of rare earth elements was 30.94%, the leaching rate of titanium was 0.54%, and the leaching rate of silicon was 1.83% during the water leaching process. The total recovery rate of rare earth elements in rare earth carbonate products was 27.64%, the total recovery rate of titanium in sodium titanate was 80.62%, and the total recovery rate of silicon in silica was 82.19%.

[0132] Comparative Example 4

[0133] Compared with Example 1, the only difference is that in step (1), ammonium sulfate is used as an additive, and all other operations and parameters are the same as in Example 1;

[0134] The tests showed that the leaching rate of rare earth elements was 59.32%, the leaching rate of titanium was 3.56%, and the leaching rate of silicon was 4.87% during the water leaching process. The total recovery rate of rare earth elements in rare earth carbonate products was 56.97%, the total recovery rate of titanium in sodium titanate was 78.73%, and the total recovery rate of silicon in silica was 77.58%.

Claims

1. A method for separating valuable elements from rare earth and titanium associated resources, characterized in that, By mixing associated raw materials containing rare earth and titanium with sulfuric acid and additives, pressing them into pellets, and then roasting them, followed by water leaching treatment, a rare earth-rich water leaching solution and a titanium-rich water leaching residue are obtained. The additive is an organic sulfate ester or its water-soluble salt; wherein the organic sulfate ester is at least one of a diol organic sulfate ester and a polyol organic sulfate ester; The weight ratio of by-products, sulfuric acid, and additives is 1:0.2~0.8:0.03~0.2; The roasting temperature is 150~350℃.

2. The method for separating valuable elements from rare earth and titanium associated resources as described in claim 1, characterized in that, The associated raw materials containing rare earth and titanium are at least one of solid waste and minerals containing rare earth and titanium. The rare earth elements include at least one element selected from Ce, La, Pr, Nd, Sm, Gd, Y, Dy, Eu, and Er. And / or, by mass content, the associated raw materials include REO: 25~50%, TiO2: 5~12%, SiO2: 10~18%, P2O5: 28~38%, CaO: 1~5%, and TFe: 2~6%.

3. The method for separating valuable elements from rare earth and titanium associated resources as described in claim 1, characterized in that, The sulfuric acid mentioned is concentrated sulfuric acid.

4. The method for separating valuable elements from rare earth and titanium associated resources as described in claim 3, characterized in that, Organic sulfates include monosulfates and / or disulfates of polyols; wherein, polyols refer to polyols with 2 to 8 carbon atoms and 3 to 6 hydroxyl groups.

5. The method for separating valuable elements from rare earth and titanium associated resources as described in claim 4, characterized in that, Organic sulfates include monosulfates and disulfates of polyols in a weight ratio of 55-95:5-45.

6. The method for separating valuable elements from rare earth and titanium associated resources as described in claim 1, characterized in that, The weight ratio of by-products, sulfuric acid, and additives is 1:0.3~0.6:0.05~0.15; The roasting temperature is 180~220℃; And / or, the roasting time is 0.5~3 h; And / or, the roasting process includes a pre-roasting process, wherein the pre-roasting temperature is 80~120℃.

7. The method for separating valuable elements from rare earth and titanium associated resources as described in claim 1, characterized in that, The liquid-to-solid ratio during the water immersion process is 5~20mL / g; The water immersion temperature is 10~95℃; The water immersion is carried out under mechanical stirring, wherein the mechanical stirring speed is 400~600 r / min.

8. A method for recovering valuable elements from rare earth and titanium associated resources, characterized in that, The method described in any one of claims 1 to 7 is used to roast and leach associated raw materials of rare earth and titanium to obtain water leaching solution and water leaching residue; The aqueous extract is subjected to precipitation treatment, and the solid-liquid separation is performed to obtain rare earth products; The water-leached residue was subjected to alkaline leaching treatment, and the solid-liquid separation was performed to obtain sodium titanate product.

9. The method for recovering valuable elements from rare earth and titanium associated resources as described in claim 8, characterized in that, The pH during the precipitation process is 2-3.5; The precipitant used in the precipitation process is at least one of oxalic acid, water-soluble oxalate, water-soluble carbonate, and water-soluble bicarbonate. The precipitant used in the precipitation process is 1 to 2 times the theoretical molar amount of rare earth precipitation.

10. The method for recovering valuable elements from rare earth and titanium associated resources as described in claim 8, characterized in that, The alkaline solution for alkaline leaching is an aqueous solution of alkali metal hydroxide; The alkaline immersion temperature is 25~120℃, and the immersion time is 0.5~4h; And / or, the alkaline leaching process is carried out under microwave assistance; And / or, the water-leached residue also contains silicon; after alkaline leaching treatment, the silicon is enriched in the alkaline leaching solution; And / or, acid treatment of the alkaline leaching solution yields silica.

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