Method for recovering rubidium and cesium from acid-process lithium extraction slag

Abstract

The invention provides a method for recovering rubidium and cesium from acid-process lithium extraction slag, which comprises the following steps: S1, drying and crushing the acid-process lithium extraction slag to obtain crushed slag, and removing impurities from the crushed slag to obtain activated roasting clinker; s2, the activated roasting clinker is subjected to water leaching treatment, and leaching liquid and leaching residues are obtained; s3, cesium extraction and rubidium extraction are sequentially conducted on the leachate, and a cesium extraction product and a rubidium extraction product are obtained; and S4, the cesium extraction product and the rubidium extraction product are concentrated, crystallized, filtered and dried to obtain a cesium salt product and a rubidium salt product, and the method has the effects of being high in extraction recovery rate and low in cost.

Description

Technical Field

[0001] This invention belongs to the technical field of hydrometallurgy and comprehensive utilization of secondary resources, specifically relating to a method for recovering rubidium and cesium from lithium extraction slag by acid extraction. Background Technology

[0002] Lithium extraction from ores such as lepidolite using the sulfuric acid process is a crucial technology for lithium salt production. This process generates a large amount of acidic lithium extraction slag, which, in addition to containing incompletely leached lithium, is typically enriched with rare metals such as rubidium and cesium that were present in the original ore. Due to their unique physicochemical properties, rubidium and cesium have irreplaceable and vital applications in high-tech fields such as aerospace, energy, medicine, and chemical engineering.

[0003] Currently, lithium extraction slag is often stockpiled as industrial waste, occupying land and posing environmental risks due to its heavy metals and acidic substances, resulting in a serious waste of valuable rubidium and cesium resources. Existing technologies for recovering rubidium and cesium from such waste slag or low-grade ores mainly include: 1. Calcination-Leaching Method: The residue is mixed with additives such as calcium and sodium salts and calcined at high temperature to destroy its crystal structure, followed by leaching with water or dilute acid. This method is energy-intensive, and the additives introduce a large number of impurities, increasing the difficulty of subsequent separation.

[0004] 2. Direct acid leaching method: This method uses high-concentration strong acid for leaching under high temperature and pressure, which causes severe corrosion to equipment, consumes a large amount of reagents, is costly, has poor selectivity, and results in the co-leaching of a large number of impurity ions.

[0005] 3. Ion exchange / solvent extraction method: Although it can effectively separate and enrich rubidium and cesium, it usually requires complex pretreatment to obtain a pure solution, and the process is long.

[0006] Therefore, developing a new method for recovering rubidium and cesium from lithium extraction slag that is simple in process, has a high recovery rate, low cost, and is environmentally friendly has extremely high economic value and environmental significance. Summary of the Invention

[0007] To address the aforementioned technical problems, this invention provides a method for recovering rubidium and cesium from acid-process lithium extraction slag, thereby resolving the technical issues in the prior art.

[0008] In a first aspect, the present invention provides the following technical solution: a method for recovering rubidium and cesium from acid-process lithium extraction slag, comprising: S1. The acid process lithium extraction slag is dried and crushed to obtain crushed slag. The crushed slag is then subjected to impurity removal to obtain activated roasted clinker. S2. The activated roasted clinker is subjected to water leaching treatment to obtain leachate and leachate residue; S3. The leachate is subjected to cesium extraction and rubidium extraction in sequence to obtain cesium extraction product and rubidium extraction product. S4. The cesium extract and the rubidium extract are concentrated, crystallized, filtered, and dried respectively to obtain cesium salt products and rubidium salt products.

[0009] Compared with existing technologies, the beneficial effects of this invention are as follows: 1. High recovery rate: Through low-temperature activation with concentrated sulfuric acid, rubidium and cesium, which are located in sparingly soluble structures, are efficiently converted into soluble salts, with a total recovery rate of over 85% for both rubidium and cesium; 2. Good selectivity: The water leaching process effectively separates rubidium and cesium from major impurities (silicon, aluminum, and iron), greatly reducing the burden on subsequent extraction processes; 3. Efficient process: Utilizing the difference in t-BAMBP extraction performance for cesium and rubidium (at different pH levels), efficient separation and purification of cesium and cesium are achieved, resulting in a simple process flow and high product purity; 4. Low cost and environmentally friendly: The activation temperature is much lower than that of the traditional calcination method, resulting in low energy consumption; water leaching replaces strong acid leaching, reducing reagent consumption and corrosiveness; the final waste residue is inert silicon slag, which can be disposed of harmlessly or utilized as a resource. The entire method achieves resource utilization and high-value utilization of the waste residue, resulting in significant economic and environmental benefits.

[0010] Preferably, step S1 specifically includes: The lithium extraction residue from the acid process is air-dried or dried at 60-110℃, and then crushed into a particle size of less than 150μm. The crushed residue is mixed with water or dilute sulfuric acid for slurry washing to remove some soluble impurities. The washed residue is then filtered to obtain the washed residue. The washed residue is uniformly mixed with concentrated sulfuric acid and activated and roasted at a low temperature of 150-300℃ for 0.5-3 hours to convert rubidium and cesium into soluble sulfates, while solidifying some impurities to obtain activated roasted clinker.

[0011] Preferably, the liquid-to-solid ratio of the crushed residue to water or dilute sulfuric acid is 1 to 4:1, and the mass ratio of the washed residue to concentrated sulfuric acid is 1:0.5 to 1.5.

[0012] Preferably, step S2 specifically includes: The activated roasted clinker is mixed with deionized water and leached under normal pressure and at 60-90°C for 0.5-2 hours to fully dissolve the rubidium and cesium sulfates. The mixture is then filtered to obtain a leachate and a leachate residue.

[0013] Preferably, the liquid-to-solid ratio of the activated roasted clinker to deionized water is 2-6:1.

[0014] Preferably, step S3 specifically includes: The pH of the leachate is adjusted to neutral or weakly alkaline using sodium hydroxide or potassium hydroxide. An organic phase with a concentration of 0.5-1.5 mol / L is prepared using an extractant and a diluent. Multi-stage countercurrent extraction is performed according to a first preset O / A ratio. Cesium ions are selectively extracted into the organic phase, while rubidium, potassium, and sodium remain in the raffinate. The cesium-loaded organic phase is back-extracted using dilute hydrochloric acid or dilute nitric acid to obtain the cesium salt product. The pH of the raffinate was adjusted to strongly acidic using hydrochloric acid. The extractant and diluent were used as the organic phase, and multi-stage countercurrent extraction was performed at a second preset O / A ratio. The rubidium-loaded organic phase was back-extracted using dilute hydrochloric acid or dilute nitric acid to obtain the rubidium salt product.

[0015] Preferably, the first preset O / A ratio is 1:3-1:1, and the second preset O / A ratio is 1:2-2:1.

[0016] Preferably, the extractant is t-BAMBP and the diluent is sulfonated kerosene.

[0017] Preferably, aluminum salts or antimony salts are added as co-extractants during the multi-stage countercurrent extraction process to improve the extraction rate and separation coefficient of rubidium and cesium.

[0018] Preferably, the concentration of dilute hydrochloric acid or dilute nitric acid is 0.1-1.0 mol / L. Attached Figure Description

[0019] To more clearly illustrate the technical solutions in the embodiments of the present invention, the drawings used in the description of the embodiments or the prior art will be briefly introduced below. Obviously, the drawings described below are only some embodiments of the present invention. For those skilled in the art, other drawings can be obtained based on these drawings without creative effort.

[0020] Figure 1 This is a flowchart of a method for recovering rubidium and cesium from acid-process lithium extraction slag, provided in Embodiment 1 of the present invention.

[0021] The embodiments of the present invention will be further described below with reference to the accompanying drawings. Detailed Implementation

[0022] Embodiments of the present invention are described in detail below, examples of which are illustrated in the accompanying drawings, wherein the same or similar reference numerals denote the same or similar elements or elements having the same or similar functions throughout. The embodiments described below with reference to the accompanying drawings are exemplary and intended to explain embodiments of the present invention, and should not be construed as limiting the present invention.

[0023] Example 1 In Embodiment 1 of the present invention, as Figure 1As shown, a method for recovering rubidium and cesium from lithium residue obtained through acid extraction includes: S1. The acid process lithium extraction slag is dried and crushed to obtain crushed slag. The crushed slag is then subjected to impurity removal to obtain activated roasted clinker. Specifically, step S1 includes: The lithium extraction residue from the acid process is air-dried or dried at 60-110℃, and then crushed into a particle size of less than 150μm. The crushed residue is mixed with water or dilute sulfuric acid for slurry washing to remove some soluble impurities. The washed residue is then filtered to obtain the washed residue. The washed residue is uniformly mixed with concentrated sulfuric acid and activated and roasted at a low temperature of 150-300℃ for 0.5-3 hours to convert rubidium and cesium into soluble sulfates, while solidifying some impurities to obtain activated roasted clinker.

[0024] The liquid-to-solid ratio of the crushed residue to water or dilute sulfuric acid is 1 to 4:1, and the mass ratio of the washed residue to concentrated sulfuric acid is 1:0.5 to 1.5.

[0025] S2. The activated roasted clinker is subjected to water leaching treatment to obtain leachate and leachate residue; Specifically, step S2 includes: The activated roasted clinker is mixed with deionized water and leached under normal pressure and 60-90°C for 0.5-2 hours to fully dissolve rubidium and cesium sulfates. The mixture is then filtered to obtain leachate and leach residue. Specifically, the activated roasted clinker obtained in step S1 is added to deionized water at a liquid-to-solid ratio of (2~6):1 (mL / g), and leached under normal pressure and at 60-90℃ with stirring for 0.5-2 hours to fully dissolve rubidium and cesium sulfates into the solution, while most impurities such as silicon, aluminum, and iron remain in the slag. The solution is then filtered to obtain a rubidium- and cesium-rich leachate and leaching residue (mainly composed of silicon slag, which can be used as a building material raw material).

[0026] S3. The leachate is subjected to cesium extraction and rubidium extraction in sequence to obtain cesium extraction product and rubidium extraction product. Specifically, step S3 includes: The pH of the leachate is adjusted to neutral or weakly alkaline using sodium hydroxide or potassium hydroxide. An organic phase with a concentration of 0.5-1.5 mol / L is prepared using an extractant and a diluent. Multi-stage countercurrent extraction is performed according to a first preset O / A ratio. Cesium ions are selectively extracted into the organic phase, while rubidium, potassium, and sodium remain in the raffinate. The cesium-loaded organic phase is back-extracted using dilute hydrochloric acid or dilute nitric acid to obtain the cesium salt product. The pH of the raffinate was adjusted to strongly acidic using hydrochloric acid. The extractant and diluent were used as the organic phase, and multi-stage countercurrent extraction was performed at a second preset O / A ratio. The rubidium-loaded organic phase was back-extracted using dilute hydrochloric acid or dilute nitric acid to obtain the rubidium salt product.

[0027] Specifically, the pH of the leachate is adjusted to neutral or weakly alkaline (pH 6-9) using sodium hydroxide or potassium hydroxide. 4-tert-butyl-2-(α-methylbenzyl)phenol (t-BAMBP) is used as the extractant, and sulfonated kerosene as the diluent. An organic phase with a concentration of 0.5-1.5 mol / L is prepared. Multi-stage countercurrent extraction is performed at an O / A (organic phase / aqueous phase) ratio of (1:3)-(1:1), meaning the organic and aqueous phases flow in opposite directions at a (1:3)-(1:1) ratio in multiple intermediate mixing and clarifying tanks. Cesium ions are selectively extracted into the organic phase, while rubidium, potassium, sodium, etc., mostly remain in the aqueous phase (raffinate). The cesium-loaded organic phase is then back-extracted with dilute hydrochloric acid or dilute nitric acid to obtain a pure cesium chloride or cesium nitrate enriched solution. The raffinate (rich in rubidium) was adjusted to a strongly acidic pH (pH < 2) with hydrochloric acid. A sulfonated kerosene solution containing t-BAMBP (concentration 0.2-0.8 mol / L) was used as the organic phase, and multi-stage countercurrent extraction was performed at an O / A ratio of (1:2)-(2:1) to separate rubidium from residual potassium, sodium, and other impurities. The rubidium-loaded organic phase was then back-extracted with dilute hydrochloric acid or dilute nitric acid to obtain a pure rubidium chloride or rubidium nitrate enriched solution.

[0028] In the multi-stage countercurrent extraction process, aluminum salts or antimony salts are added as co-extractants to improve the extraction rate and separation coefficient of rubidium and cesium. The concentration of dilute hydrochloric acid or dilute nitric acid is 0.1-1.0 mol / L.

[0029] S4. The cesium extract and the rubidium extract are concentrated, crystallized, filtered, and dried respectively to obtain cesium salt products and rubidium salt products.

[0030] To further verify the beneficial effects of this invention, several experimental groups and control groups are provided below: Experimental Group 1 Experimental Group 1 used the method provided in Implementation 1 above to extract rubidium and cesium. 10 kg of lithium extraction residue (dry basis, composition: Rb₂O 0.85%, Cs₂O 0.42%, SiO₂ 65%, Al₂O₃ 18%) from a company after sulfuric acid treatment of lepidolite was taken. The extraction process included the following steps: 1. Mix the slag with 98% concentrated sulfuric acid at a mass ratio of 1:1, place it in a muffle furnace and activate and roast at 220°C for 1.5 hours.

[0031] 2. The roasted clinker was leached with deionized water at a liquid-to-solid ratio of 4:1 mL / g at 80°C with stirring for 1 hour. After filtration, approximately 3.8 L of leachate and leaching residue were obtained. Analysis showed that the leaching rates of rubidium and cesium were 91.5% and 93.2%, respectively.

[0032] 3. Adjust the pH of the leachate to 8 with NaOH. Perform a three-stage countercurrent extraction of cesium using a 1.0 mol / L t-BAMBP / sulfonated kerosene solution at an O / A ratio of 1:2. Back-extract with 0.5 mol / L HCl to obtain a cesium-rich solution. Adjust the pH of the remaining cesium extract to 1 with HCl. Perform a four-stage countercurrent extraction of rubidium using a 0.5 mol / L t-BAMBP / sulfonated kerosene solution at an O / A ratio of 1:1. Back-extract with 0.5 mol / L hydrochloric acid to obtain a rubidium-rich solution. S4. The cesium and rubidium enriched solutions were evaporated and concentrated, cooled and crystallized, filtered, and dried at 150°C to obtain 36.9g of cesium chloride product (purity >99.0%) and 77.1g of rubidium chloride product (purity >98.5%).

[0033] The total recovery rate for cesium was calculated to be 87.1%, and the total recovery rate for rubidium was 89.3%.

[0034] Control group 1: Calcination and leaching method Take 1 kg of the same acid-process lithium extraction slag as in Example 1 of this invention.

[0035] 1. Batching and calcination: The lithium extraction slag and sodium carbonate (NaCO3) solid were mixed uniformly at a mass ratio of 1:0.8. The mixture was placed in a muffle furnace and calcined at 850°C for 2 hours.

[0036] 2. Water leaching: The calcined clinker was cooled to room temperature, and then deionized water was added at a liquid-to-solid ratio of 4:1 (mL / g). The mixture was stirred and leached at 80°C for 1.5 hours. After filtration, approximately 3.8 L of leachate and leaching residue were obtained. The leaching rate of rubidium was 78.5%, and the leaching rate of cesium was 81.3%.

[0037] 3. Extraction and back-extraction: Same as in Example 1, 27.5 g of cesium chloride product (purity >99.0%) and 66.1 g of rubidium chloride product (purity >98.5%) were obtained. The total recovery rate of cesium was calculated to be 76.0%, and the total recovery rate of rubidium was 76.6%.

[0038] Control group 2: Direct acid leaching method Take 1 kg of the same acid-process lithium extraction slag as in the embodiment of the present invention.

[0039] 1. Direct acid leaching: Without any pretreatment, the lithium extraction residue is directly mixed with a 4 mol / L sulfuric acid solution at a liquid-to-solid ratio of 4:1 (mL / g). Leaching is carried out with vigorous stirring at 95℃ for 4 hours. The leaching rate of rubidium is 72.3%, and the leaching rate of cesium is 75.5%.

[0040] 2. Extraction and Back-Extraction: Without extraction, aluminum ions (Al³⁺) are directly extracted from the leachate using acid leaching. + The concentration was as high as 15 g / L, with iron ions (Fe³⁺) +The concentration is about 2 g / L. Such a high impurity solution must undergo a complicated impurity removal pretreatment before t-BAMBP extraction. The process will consume a lot of alkali and generate a lot of impurities. Therefore, extraction is not performed. The leaching rate of rubidium and cesium is used as its recovery rate. The cesium recovery rate is 75.5% and the rubidium recovery rate is 72.3%.

[0041] Based on the above experimental and control groups, it can be seen that the total recovery rate of cesium in experimental group 1 was 87.1%, and the total recovery rate of rubidium was 89.3%. Therefore, it had the highest extraction recovery rate and also had the effect of low cost.

[0042] The method for recovering rubidium and cesium from acid-process lithium extraction residue provided in Embodiment 1 of the present invention has the following advantages: 1. High recovery rate: Through low-temperature activation with concentrated sulfuric acid, rubidium and cesium, which are located in sparingly soluble structures, are efficiently converted into soluble salts, and the total recovery rate of rubidium and cesium can reach more than 85%; 2. Good selectivity: The water immersion process effectively separates rubidium and cesium from the main impurities (silicon, aluminum, iron), greatly reducing the burden on subsequent extraction processes; 3. Highly efficient process: By utilizing the difference in t-BAMBP extraction performance for cesium and rubidium (at different pH levels), highly efficient separation and purification of cesium and rubidium are achieved. The process is simple and the product purity is high. 4. Low cost and environmentally friendly: The activation temperature is much lower than that of the traditional calcination method, resulting in low energy consumption; water leaching replaces strong acid leaching, reducing reagent consumption and corrosiveness; the final waste residue is inert silica slag, which can be disposed of harmlessly or utilized as a resource. The entire method achieves resource utilization and high-value utilization of the waste residue, resulting in significant economic and environmental benefits.

[0043] The technical features of the above embodiments can be combined in any way. For the sake of brevity, not all possible combinations of the technical features in the above embodiments are described. However, as long as there is no contradiction in the combination of these technical features, they should be considered to be within the scope of this specification.

[0044] The embodiments described above are merely illustrative of several implementations of the present invention, and while the descriptions are relatively specific and detailed, they should not be construed as limiting the scope of the invention patent. It should be noted that those skilled in the art can make various modifications and improvements without departing from the concept of the present invention, and these all fall within the protection scope of the present invention. Therefore, the protection scope of this invention patent should be determined by the appended claims.

Claims

1. A method for recovering rubidium and cesium from lithium residue obtained by acid extraction, characterized in that, include: S1. The acid process lithium extraction slag is dried and crushed to obtain crushed slag. The crushed slag is then subjected to impurity removal to obtain activated roasted clinker. S2. The activated roasted clinker is subjected to water leaching treatment to obtain leachate and leachate residue; S3. The leachate is subjected to cesium extraction and rubidium extraction in sequence to obtain cesium extraction product and rubidium extraction product. S4. The cesium extract and the rubidium extract are concentrated, crystallized, filtered, and dried respectively to obtain cesium salt products and rubidium salt products.

2. The method for recovering rubidium and cesium from lithium residue extracted by acid process according to claim 1, characterized in that, Step S1 specifically includes: The lithium extraction residue from the acid process is air-dried or dried at 60-110℃, and then crushed into a particle size of less than 150μm. The crushed residue is mixed with water or dilute sulfuric acid for slurry washing to remove some soluble impurities. The washed residue is then filtered to obtain the washed residue. The washed residue is uniformly mixed with concentrated sulfuric acid and activated and roasted at a low temperature of 150-300℃ for 0.5-3 hours to convert rubidium and cesium into soluble sulfates, while solidifying some impurities to obtain activated roasted clinker.

3. The method for recovering rubidium and cesium from lithium residue extracted by acid process according to claim 2, characterized in that, The liquid-to-solid ratio of the crushed residue to water or dilute sulfuric acid is 1 to 4:1, and the mass ratio of the washed residue to concentrated sulfuric acid is 1:0.5 to 1.

5.

4. The method for recovering rubidium and cesium from lithium residue extracted by acid process according to claim 1, characterized in that, Step S2 specifically includes: The activated roasted clinker is mixed with deionized water and leached under normal pressure and at 60-90°C for 0.5-2 hours to fully dissolve the rubidium and cesium sulfates. The mixture is then filtered to obtain a leachate and a leachate residue.

5. The method for recovering rubidium and cesium from lithium residue extracted by acid process according to claim 4, characterized in that, The liquid-to-solid ratio of the activated roasted clinker to deionized water is 2-6:

1.

6. The method for recovering rubidium and cesium from lithium residue extracted by acid process according to claim 1, characterized in that, Step S3 specifically includes: The pH of the leachate is adjusted to neutral or weakly alkaline using sodium hydroxide or potassium hydroxide. An organic phase with a concentration of 0.5-1.5 mol / L is prepared using an extractant and a diluent. Multi-stage countercurrent extraction is performed according to a first preset O / A ratio. Cesium ions are selectively extracted into the organic phase, while rubidium, potassium, and sodium remain in the raffinate. The cesium-loaded organic phase is back-extracted using dilute hydrochloric acid or dilute nitric acid to obtain the cesium salt product. The pH of the raffinate was adjusted to strongly acidic using hydrochloric acid. The extractant and diluent were used as the organic phase, and multi-stage countercurrent extraction was performed at a second preset O / A ratio. The rubidium-loaded organic phase was back-extracted using dilute hydrochloric acid or dilute nitric acid to obtain the rubidium salt product.

7. The method for recovering rubidium and cesium from lithium residue extracted by acid process according to claim 6, characterized in that, The first preset O / A ratio is 1:3-1:1, and the second preset O / A ratio is 1:2-2:

1.

8. The method for recovering rubidium and cesium from lithium residue extracted by acid process according to claim 6, characterized in that, The extractant is t-BAMBP, and the diluent is sulfonated kerosene.

9. The method for recovering rubidium and cesium from lithium residue extracted by acid process according to claim 6, characterized in that, In the multi-stage countercurrent extraction process, aluminum salts or antimony salts are added as co-extractants to improve the extraction rate and separation coefficient of rubidium and cesium.

10. The method for recovering rubidium and cesium from lithium residue extracted by acid process according to claim 6, characterized in that, The concentration of dilute hydrochloric acid or dilute nitric acid is 0.1-1.0 mol / L.

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