Method for recovering lithium, rubidium, cesium and potassium from low-grade iron lepidolite

Abstract

The invention discloses a method for recovering lithium, rubidium, cesium and potassium from low-grade iron lepidolite. Comprising the following steps: preparing materials, drying, dispensing and grinding, granulating, roasting, leaching, filtering and washing, recycling leached residues, purifying and removing impurities from leachate, concentrating and evaporating liquid after impurity removal, separating out sylvite at low temperature, filtering and washing sylvite cores, filtering and washing leached residues, precipitating lithium carbonate from liquid after salt separation, refining lithium carbonate, extracting cesium from liquid after lithium precipitation, and extracting rubidium from liquid after cesium extraction. And removing carbonate radicals from the rubidium-extracted solution, and returning to the steps of evaporation and concentration and the like. The overall operation of the process of the method has the advantages of low-cost roasting, high comprehensive recovery rate of lithium, rubidium, cesium and potassium and the like, the whole-section process does not need to adjust the acid-base of the feed liquid, the chemical cost is saved, the economic benefit is increased, and the process with high efficiency and high comprehensive recovery rate is formed.

Description

Technical Field

[0001] The invention relates to the technical field of metal smelting, and in particular to a method for recovering lithium, rubidium and cesium potassium from low-grade lithium iron mica. Background Art

[0002] Lepidolite is an important and scarce strategic resource. Some lepidolite deposits have low grade, and most of them are strongly hematized, and a few are corroded into chlorite, which makes it impossible to enrich the weakly magnetic target minerals in the magnetic separation process, and the chlorite part is easy to mud in the flotation process, resulting in the loss of target minerals in the tailings in the beneficiation process. Therefore, after magnetic separation or flotation, the Li2O content is only 1.5%-3.0%, and it is accompanied by high content of iron and fluorine, including Fe2O3 content of 13%-15%, F content of 5%-7%, which further increases the difficulty of extracting lithium, rubidium and cesium, and makes it more difficult to replace them. In addition, fluorine is released into the gas phase or water phase during roasting or leaching, which will cause corrosion to the equipment, increase the economic investment of the equipment, and easily pollute the environment and affect the product quality. Therefore, the research of economical, efficient and environmentally friendly processing technology is an important direction for the development of such resources.

[0003] The metallurgical extraction of lithium mica concentrate mainly includes sulfuric acid method, fluorine chemical method, alkali method, pressure cooking method and sulfate method in the laboratory stage. These methods all have some problems in economy, environmental protection, and equipment operation, and have not been industrialized. Chinese patent application CN107475537A discloses a method for extracting lithium, rubidium and cesium salts from lithium mica raw materials, using lithium mica as raw material, including raw material pretreatment, two-stage roasting, mechanical activation treatment, pressure cooking leaching, and separation and extraction processes for lithium mica ore; including the following method steps: raw material pretreatment is to crush the lithium mica ore and mix it with a roasting additive, wherein the roasting additive includes a mixture of sodium humate, sodium hydroxide, sodium sulfate, and calcium compounds, and by controlling low-temperature roasting and plasma high-temperature roasting combined with defluorination and mechanical activation treatment of roasted sand, the metal elements in the lithium mica ore can be separated and extracted to the greatest extent, which can improve the utilization rate and economic benefits of lithium, rubidium and cesium in the lithium mica ore. However, the scheme is difficult to operate (sodium hydroxide will deliquesce, it is difficult to mix the ore and the medicine uniformly, and F entering the gas phase and liquid phase has a great corrosion effect on the equipment). For enterprises that actually produce, the process flow is complicated and not adaptable. Chinese patent application CN107937733A discloses a process for extracting lithium potassium, rubidium and cesium from iron lithium mica, wherein the iron lithium mica ore is ground into iron lithium mica powder, and -100 mesh accounts for more than 80wt%; after the ore and the medicine are mixed, they are roasted at 700℃~1000℃, and the water extract obtained by water immersion is added with alkali to adjust the pH value to 8.5~9.5, and solid-liquid separation is performed; after the filtrate is concentrated, carbonate is added, and solid-liquid separation is performed to remove calcium, and after calcium removal, the liquid is evaporated and concentrated, and carbonate is added to precipitate lithium, and after lithium precipitation, hydroxide ions are added to the liquid to extract cesium, and hydroxide is added to the residual liquid of cesium extraction, and rubidium is extracted, and the residual liquid of rubidium extraction is evaporated and concentrated to prepare potassium salt. This solution can extract lithium, rubidium, cesium and potassium from lithium mica ore, but the system needs to continuously add sodium hydroxide, which has high reagent costs. The solution entering the rubidium and cesium extraction has high potassium and sodium content, which causes serious competitive adsorption on the extraction and affects the product quality of rubidium and cesium salts. The obtained potassium salt will carry a large amount of unprecipitated lithium, which greatly reduces the comprehensive recovery rate of lithium and causes economic waste.Chinese patent application CN104649302A discloses a method for obtaining lithium carbonate from iron lithium mica, wherein the material and the reagent are mixed at a ratio of 1:0.2-1.5, roasted at 800-1100°C for 15min-120min, subjected to primary leaching, and subjected to four-stage countercurrent washing, wherein the obtained leachate is impurity-removed using a 5% NaOH solution, potassium-sodium complex salt is precipitated from the impurity-removed liquid, the precipitated liquid is concentrated, lithium carbonate is precipitated, and rubidium-cesium potassium is extracted from the precipitated liquid, wherein the Li2O content of the leachate is 5-10g / L, the liquid-to-solid ratio of the water leaching liquid is 0.5-3, and its process defects are as follows: ① The Li2O content of the leaching liquid is 5-10g / L. According to its liquid-to-solid ratio, the washing efficiency is -91%, and -9% will be carried away by the leaching residue, causing entrainment loss; ② The system needs to add 5% NaOH, the reagent cost is high, and the economic benefit is poor; ③ The leaching liquid directly precipitates potassium and sodium salts, the salt concentration is low, and the potassium and sodium salt removal rate is low. It needs to be concentrated first and then potassium and sodium are precipitated. ④ The separation of potassium and sodium salts will entrain a large amount of lithium, and the process fails to achieve the effective separation of lithium, rubidium and cesium. Summary of the invention

[0004] In view of the deficiencies in the prior art, the present invention aims to provide a method for recovering lithium, rubidium and cesium potassium from low-grade lithium iron mica.

[0005] In order to achieve the above object, the present invention adopts the following technical solution:

[0006] A method for recovering lithium, rubidium and cesium potassium from low-grade lithium iron mica comprises the following steps:

[0007] S1. Grinding low-grade lithium iron mica ore to obtain lithium mica concentrate by magnetic separation or flotation;

[0008] S2, drying the lepidolite concentrate obtained in step S1;

[0009] S3, blending, mixing and dry-grinding the lepidolite concentrate, CaSO4·2H2O, Ca(OH)2 and CaCO3 to obtain a mixed material;

[0010] S4, granulating the mixed material obtained in step S3 to obtain ore pellets;

[0011] S5, drying the ore mass obtained in step S4 and then roasting it;

[0012] S6, taking the roasted sand obtained after roasting in step S5 out of the furnace and cooling it, then grinding and slurrying it, and pumping the obtained slurry to a leaching tank for leaching;

[0013] S7, the leached pulp obtained after leaching in step S6 is subjected to solid-liquid separation, the precious leached liquid obtained by solid-liquid separation is transferred to step S9 for purification and impurity removal, the leached residue is washed, the washing liquid is returned to step S6 to participate in leaching, and the leached washing residue is transferred to step S8;

[0014] S8, beneficiating the leached washing residue obtained in step S7, flotating gypsum, and drying the gypsum and returning it to step S3 for use;

[0015] S9, sending the leached precious liquid obtained in step S7 to a primary purification tank, adding lime slurry to preliminarily remove F and Mg, and then further sending it to a secondary purification tank, adding carbonate-containing materials to remove calcium, to obtain a purified slurry; after the purified slurry is filtered, the purified liquid obtained by filtration is sent to step S10, the filter residue obtained by filtration is washed and dried, and then returned to step S5 for roasting, and the washed liquid obtained by washing the filter residue is returned to the primary purification tank;

[0016] S10, evaporating and concentrating the purified liquid obtained in step S9, and when evaporation causes crystals to precipitate, refrigerating to precipitate potassium salt, and then filtering, and the obtained filtrate enters the treatment of step S11, the precipitated potassium salt is centrifuged, and the liquid obtained by centrifugation is returned for evaporation and concentration, and the separated solid is washed with a saturated potassium sulfate solution, and the washing liquid is returned for evaporation and concentration;

[0017] S11, the filtrate obtained in step S10 is heated and potassium carbonate is added, and after the reaction is completed, the crude lithium carbonate is obtained by centrifugal filtration, and the crude lithium carbonate is refined, and the lithium precipitation mother liquor enters step S12;

[0018] S12, the lithium precipitation mother liquor is mixed with the extracted organic phase, and then extracted, washed and stripped in sequence to obtain a cesium-rich solution, and the cesium-rich solution is evaporated and concentrated to obtain a cesium sulfate product;

[0019] S13, the cesium-extracted liquid obtained after the extraction in step S12 is mixed with the extracted organic phase and then extracted, washed and stripped in sequence, and the obtained rubidium-rich liquid is evaporated and concentrated to obtain a rubidium sulfate product;

[0020] S14: After the extraction in step S13, a part of the rubidium extract residue obtained is returned to step S9 as a carbonate-containing material for decalcification after oil removal, and the other part is added with sulfuric acid after oil removal and then returned to step S10 for evaporation and concentration.

[0021] Furthermore, in step S1, the mass content of Li2O in the lithium mica concentrate is 1-3%.

[0022] Furthermore, in step S2, the lepidolite concentrate is dried to obtain a moisture content of less than 5%.

[0023] Furthermore, in step S3, the mass ratio of lithium mica concentrate: CaSO4·2H2O: Ca(OH)2: CaCO3 is 1: 0.3-0.6: 0-0.3: 0-0.5; the fineness of the mixture is -15--25um, accounting for 85%.

[0024] Furthermore, in step S3, the particle size of the ore mass is 1-20 mm.

[0025] Furthermore, in step S5, the ore mass is roasted in a rotary kiln at 850-1050° C. for 0.5-2 h.

[0026] Furthermore, in step S6, the roasted sand is cooled to 100-200°C after being taken out of the furnace; and leached for 0.5-2h at a leaching temperature of 50-90°C.

[0027] Further, in step S11, the filtrate obtained in step S10 is heated and potassium carbonate in an amount of 1-1.2 times the theoretical molar amount is added, and the reaction is carried out for 1 hour; the crude lithium carbonate is sequentially hydrogenated, ion exchanged and pyrolyzed to obtain battery-grade lithium carbonate.

[0028] Furthermore, the extracted organic phases described in step S12 and step S13 both contain an extractant t-BAMBP and a diluent kerosene, and the volume concentration of t-BAMBP is 10-30%; the washing water used for washing in step S12 is a 5-20 g / L dilute sulfuric acid solution, and the washing water used for washing in step S13 is a 20-30 g / L dilute sulfuric acid solution; the stripping solution used for back-extraction in step S12 is a 10-30 g / L dilute sulfuric acid solution, and the stripping solution used for back-extraction in step S13 is a 20-40 g / L sulfuric acid solution.

[0029] The beneficial effects of the present invention are:

[0030] (1) By using the method of the present invention, lithium, rubidium, cesium and potassium in lepidolite can be efficiently leached, which has the advantages of efficient recovery and can reduce the secondary treatment process.

[0031] (2) In the method of the present invention, F in the ore is fixed in the form of CaF2, thereby avoiding corrosion of equipment and loss of lithium due to roasting.

[0032] (3) Compared with conventional leaching, the method of the present invention realizes the simultaneous recovery of lithium, rubidium, cesium and potassium, and through centrifugal filtration and saturated potassium sulfate washing, it ensures that the entrainment loss of lithium, rubidium and cesium in potassium salt is less than 0.05%.

[0033] (4) In the method of the present invention, calcium sulfate from the leached slag is returned to the ore blending through flotation, which can save reagents and reduce the amount of tailings, making the tailings mainly composed of silicon and iron, which is easier to use in the construction industry, such as brick making and cement making.

[0034] (5) The calcium sulfate slag and calcium carbonate slag obtained in the impurity removal process of the method of the present invention participate in ore blending, which reduces the loss caused by the entrainment of lithium, rubidium and cesium in the impurity removal slag, and can improve the comprehensive recovery rate of lithium, rubidium and cesium.

[0035] (6) In the method of the present invention, part of the liquid after rubidium extraction is returned for calcium removal, which utilizes the remaining carbonate ions from lithium precipitation with sodium carbonate and reduces the amount of reagents used.

[0036] (7) In the method of the present invention, part of the rubidium extraction liquid is added with dilute sulfuric acid and then returned to the evaporation and concentration section, and the excess sulfate and residual carbonate in the system are removed by acid-base neutralization, thereby simplifying the process flow and improving the comprehensive recovery rate of lithium and the comprehensive utilization rate of potassium.

[0037] (8) The overall operation of the process of the present invention has the advantages of low-cost roasting and high comprehensive recovery rate of lithium, rubidium and cesium potassium. The whole process does not require acid-base adjustment of the feed liquid, which reduces the need to increase sodium hydroxide in the potassium sulfate method to increase the pH of the leaching solution to meet the reagent consumption of subsequent extraction stages. In a true sense, process innovation and improvement are used to save reagent costs while increasing economic benefits, forming a process with high efficiency and comprehensive recovery rate. BRIEF DESCRIPTION OF THE DRAWINGS

[0038] Figure 1 This is a flow chart of the method of Embodiment 1-2 of the present invention. DETAILED DESCRIPTION

[0039] The present invention will be further described below in conjunction with the accompanying drawings. It should be noted that this embodiment is based on the technical solution and provides a detailed implementation method and specific operation process, but the protection scope of the present invention is not limited to this embodiment.

[0040] Example 1

[0041] This embodiment provides a method for recovering lithium, rubidium, and cesium potassium from low-grade iron lithium mica, wherein the low-grade iron lithium mica is from a lithium mica mine in Hunan. Figure 1 As shown, the following steps are included:

[0042] S1. Low-grade lepidolite is subjected to magnetic separation or flotation to obtain lepidolite concentrate. The main components of the lepidolite concentrate are shown in Table 1. The contents of Li2O, Rb2O, Cs2O and K2O are 1.55%, 0.72%, 0.048% and 8.37% respectively. The contents of iron and fluorine are relatively high, which are 9.77% and 5.52% respectively. The main gangue is SiO2 and Al2O3, with contents of 50.58% and 19.28% respectively.

[0043] Table 1

[0044]

[0045] S2, sending the lithium mica concentrate obtained in step S1 into a drying kiln for drying until the moisture content is less than 5%.

[0046] S3. The lepidolite concentrate, CaSO4·2H2O, Ca(OH)2 and CaCO3 are blended, mixed and dry-ground according to the mass ratio of lepidolite concentrate: CaSO4·2H2O: Ca(OH)2: CaCO3 = 1:0.6:0.25:0.4. The dry-ground fineness is -15 um, accounting for 85%, to obtain a mixed material.

[0047] It should be noted that CaSO4·2H2O, Ca(OH)2, and CaCO3 are all added agents at the start-up. After the process is running, CaSO4·2H2O can come from the flotation product of the subsequent step S8, or it can be supplemented with an external agent, and CaCO3 can come from the precipitate obtained by decalcification in step S9, or it can be supplemented with an external agent.

[0048] S4, granulating the mixed material obtained in step S3 to obtain ore pellets with a particle size of 1 mm.

[0049] S5. After drying the ore mass obtained in step S4, it is roasted in a rotary kiln at 1050° C. for 2 hours. During the roasting process, lithium, rubidium and cesium potassium are replaced, and fluorine is fixed in the roasted sand in the form of CaF2.

[0050] S6, the roasted sand obtained after roasting in step S5 is cooled to 100 ° C, and then ground and slurried, the liquid-solid mass ratio is 2: 1, the obtained slurry is pumped to the leaching tank, without heating, and leached at 50 ° C for 1 hour, and lithium rubidium cesium potassium is leached into the solution. It should be noted that in the startup phase, clean water is used for slurry mixing; after the process is running, the CCD1 overflow of the subsequent process is used for slurry mixing.

[0051] S7, pump the leached slurry obtained after leaching in step S6 to the thickener for thickening, and the thickened overflow (i.e., leached precious liquid) enters the solution purification tank and enters step S9 for purification and impurity removal. The thickened overflow contains 3g / L Li, 1.6g / L Rb, 100mg / L Cs, 15g / L K, and 15g / L Na. The thickened underflow undergoes three-stage thickening countercurrent washing and two-stage belt filtration washing. The thickening underflow enters the first-stage washing and thickening machine (CCD1), is mixed with the overflow of the second-stage washing and thickening machine (CCD2) and then settled. The overflow of CCD1 is returned for pulping in step S6. The underflow of CCD1 enters CCD2, is mixed with the overflow of the third-stage washing and thickening machine (CCD3) and then settled. The underflow of CCD2 goes to CCD3, is mixed with the filtrate after filtering and washing by the first-stage belt filter and then settled. The underflow of CCD3 goes to the first-stage belt filter for filtering and washing. The filter cake of the first-stage belt filter goes to the second-stage belt filter for filtering and washing. The second-stage belt filter uses fresh industrial water for filtering and washing. The filtrate of the second-stage belt filter goes to the first-stage belt filter for belt filtering and washing. The moisture content of the filter cake is about 40%.

[0052] S8, beneficiating the leached washing residue obtained in step S7, flotating the gypsum, drying the gypsum until the moisture content is less than 5%, and then returning to step S3 for use.

[0053] S9, the leached precious liquid obtained in step S7 is sent to a primary purification tank, and lime slurry is added to preliminarily remove F and Mg, the amount of calcium oxide used is 1 / 8 of the theoretical amount of all sulfates in the precipitated leached precious liquid, and the reaction is stirred at 90°C for 1h, and then further sent to a secondary purification tank, and carbonate-containing materials are added, and the mixture is stirred and purified at 90°C for 0.5h to remove calcium, to obtain a purified slurry, and the molar ratio of carbonate to calcium is 1; the purified slurry is filtered through a plate and frame filter press, and the obtained purified liquid is sent to step S10, the obtained filter residue is washed and dried, and then returned to step S5 for roasting, and the washed liquid is returned to the primary purification tank.

[0054] S10, evaporating and concentrating the purified liquid obtained in step S9, and when evaporation causes crystals to precipitate, refrigerating at 5° C. for 12 hours to precipitate potassium salt, and then filtering, and the obtained filtrate enters the treatment of step S11, the precipitated potassium salt is centrifuged, and the separated liquid is returned for evaporation and concentration, and the separated solid is washed with a saturated potassium sulfate solution, and the washing liquid is returned for evaporation and concentration;

[0055] S11. The filtrate obtained in step S10 is heated to 70° C. and then 1 times the theoretical amount of potassium carbonate is added. After reacting for 1 hour, the crude lithium carbonate is obtained by centrifugal filtration. The crude lithium carbonate is further refined by hydrogenation decomposition, ion exchange, and pyrolysis processes to obtain battery-grade lithium carbonate. The purity of the battery-grade lithium carbonate is >99.9%. The lithium precipitation mother liquor enters step S12.

[0056] S12, the lithium precipitation mother liquor is mixed with the extracted organic phase, and then three-stage extraction, five-stage washing and three-stage stripping are carried out in sequence. The obtained cesium-rich solution is evaporated and concentrated to obtain the cesium sulfate product; the extracted organic phase contains the extractant t-BAMBP and the diluent kerosene, and the volume concentration of t-BAMBP is 18%; the washing water used for washing is 13g / L dilute sulfuric acid solution, and the stripping solution is 16g / L dilute sulfuric acid solution. The cesium empty organic phase obtained after stripping is returned to continue to be used for cesium extraction. The cesium extraction rate is greater than 95%, the lithium extraction rate is less than 1%, the potassium and sodium extraction rates are less than 0.5%, and the cesium sulfate product quality is greater than 99.9%.

[0057] S13, the cesium extraction liquid obtained after step S12 is mixed with the extracted organic phase and then subjected to three-stage extraction, three-stage washing and three-stage stripping in sequence, and the obtained rubidium-rich liquid is evaporated and concentrated to obtain a rubidium sulfate product; the extracted organic phase contains an extractant t-BAMBP and a diluent kerosene, and the volume concentration of t-BAMBP is 18%; the washing water is a 30g / L dilute sulfuric acid solution, and the stripping liquid is a 20g / L sulfuric acid solution. Among them, the extraction rate of rubidium is greater than 95%, the extraction rate of lithium is less than 1%, the extraction rate of potassium and sodium is less than 0.5%, and the quality of the rubidium sulfate product is greater than 99.9%. The rubidium-free organic phase obtained after stripping is returned to continue to be used for rubidium extraction.

[0058] S14, the rubidium extraction residual solution obtained after step S13, a part of which is returned to step S9 as a carbonate-containing material for decalcification after degreasing, and the other part is added with dilute sulfuric acid solution after degreasing, and the excess sulfate and residual carbonate are removed by acid-base neutralization, and then returned to step S10 for evaporation and concentration. The dilute sulfuric acid solution comes from the washing liquid of cesium extraction and rubidium extraction.

[0059] The method of this embodiment can obtain the process indicators of lithium, rubidium, and cesium leaching rates> about 90%, lithium comprehensive recovery rate of 85%, and rubidium and cesium comprehensive recovery rate of 82%. The water leaching slag Li2O and Rb2O content is less than 0.1%, Cs2O content is less than 0.005%, and K2O content is less than 0.2%.

[0060] Example 2

[0061] This embodiment provides a method for recovering lithium, rubidium, and cesium potassium from low-grade iron lithium mica, wherein the low-grade iron lithium mica is from a certain lithium mica mine in Hunan Province, which is the same as that in Example 1. Figure 1 As shown, the following steps are included:

[0062] S1. Low-grade iron lithium mica is subjected to magnetic separation or flotation to obtain lithium mica concentrate, and the main components of the lithium mica concentrate are the same as those in Example 1.

[0063] S2, sending the lithium mica concentrate obtained in step S1 into a drying kiln for drying until the moisture content is less than 5%.

[0064] S3, the lepidolite concentrate, CaSO4·2H2O, Ca(OH)2 and CaCO3 are blended, mixed and dry-ground according to the mass ratio of lepidolite concentrate: CaSO4·2H2O:Ca(OH)2:CaCO3=1:0.6:0.2:0.5, and the dry-ground fineness is -25um accounting for 85% to obtain a mixed material. It should be noted that CaSO4·2H2O, Ca(OH)2 and CaCO3 are all added agents at the start-up. After the process is running, CaSO4·2H2O can come from the flotation product of the subsequent step S8, or it can be supplemented with an added agent, and CaCO3 can come from the precipitate obtained by decalcification in step S9, or it can be supplemented with an added agent.

[0065] S4, granulating the mixed material obtained in step S3 to obtain ore pellets with a particle size of 20 mm.

[0066] S5. After drying the ore mass obtained in step S4, it is roasted in a rotary kiln at 1050° C. for 2 hours. During the roasting process, lithium, rubidium and cesium potassium are replaced, and fluorine is fixed in the roasted sand in the form of CaF2.

[0067] S6, the roasted sand obtained after roasting in step S5 is cooled to 200 ° C, and then ground and slurried, the liquid-solid mass ratio is 2: 1, the obtained slurry is pumped to the leaching tank, without heating, and leached at 90 ° C for 1 hour, and lithium rubidium cesium potassium is leached into the solution. It should be noted that in the startup phase, clean water is used for slurry mixing; after the process is running, the CCD1 overflow of the subsequent process is used for slurry mixing.

[0068] S7, pump the leached slurry obtained after leaching in step S6 to the thickener for thickening, and the thickened overflow (i.e., leached precious liquid) enters the solution purification tank and enters step S9 for purification and impurity removal. The thickened overflow contains 2.5g / L Li, 1.4g / L Rb, 100mg / L Cs, 10g / L K, and 200mg / L Na. The thickened underflow undergoes three-stage thickening countercurrent washing and two-stage belt filtration washing. The thickening underflow enters the first-stage washing and thickening machine (CCD1), is mixed with the overflow of the second-stage washing and thickening machine (CCD2) and then settled. The overflow of CCD1 is returned for pulping in step S6. The underflow of CCD1 enters CCD2, is mixed with the overflow of the third-stage washing and thickening machine (CCD3) and then settled. The underflow of CCD2 goes to CCD3, is mixed with the filtrate after filtering and washing by the first-stage belt filter and then settled. The underflow of CCD3 goes to the first-stage belt filter for filtering and washing. The filter cake of the first-stage belt filter goes to the second-stage belt filter for filtering and washing. The second-stage belt filter uses fresh industrial water for filtering and washing. The filtrate of the second-stage belt filter goes to the first-stage belt filter for belt filtering and washing. The moisture content of the filter cake is about 40%.

[0069] S8, beneficiating the leached washing residue obtained in step S7, flotating the gypsum, drying the gypsum until the moisture content is less than 5%, and then returning to step S3 for use.

[0070] S9, the leached precious liquid obtained in step S7 is sent to a primary purification tank, and lime slurry is added to preliminarily remove F and Mg, the amount of calcium oxide used is 1 / 8 of the theoretical amount of all sulfates in the precipitated leached precious liquid, and the reaction is stirred at 90°C for 1h, and then further sent to a secondary purification tank, and carbonate-containing materials are added, and the mixture is stirred and purified at 90°C for 0.5h to remove calcium, to obtain a purified slurry, and the molar ratio of carbonate to calcium is 1; the purified slurry is filtered through a plate and frame filter press, and the obtained purified liquid is sent to step S10, the obtained filter residue is washed and dried, and then returned to step S5 for roasting, and the washed liquid is returned to the primary purification tank.

[0071] S10, evaporating and concentrating the purified liquid obtained in step S9, and when evaporation causes crystals to precipitate, refrigerating at 5° C. for 12 hours to precipitate potassium salt, and then filtering, and the obtained filtrate enters the treatment of step S11, the precipitated potassium salt is centrifuged, and the separated liquid is returned for evaporation and concentration, and the separated solid is washed with a saturated potassium sulfate solution, and the washing liquid is returned for evaporation and concentration;

[0072] S11. The filtrate obtained in step S10 is heated to 95° C. and 1.2 times the theoretical amount of potassium carbonate is added. After reacting for 1 hour, the crude lithium carbonate is obtained by centrifugal filtration. The crude lithium carbonate is further refined by hydrogenation decomposition, ion exchange, and pyrolysis processes to obtain battery-grade lithium carbonate. The purity of the battery-grade lithium carbonate is >99.9%. The lithium precipitation mother liquor enters step S12.

[0073] S12, the lithium precipitation mother liquor is mixed with the extracted organic phase, and then three-stage extraction, five-stage washing and three-stage stripping are carried out in sequence. The cesium-rich solution finally obtained is evaporated and concentrated to obtain the cesium sulfate product; the extracted organic phase contains the extractant t-BAMBP and the diluent kerosene, and the volume concentration of t-BAMBP is 10%; the washing water used for washing is 5g / L dilute sulfuric acid solution, and the stripping solution is 10g / L dilute sulfuric acid solution; the cesium empty organic phase obtained after stripping is returned to continue to be used for cesium extraction. The cesium extraction rate is greater than 95%, the lithium extraction rate is less than 1%, the potassium and sodium extraction rates are less than 0.5%, and the cesium sulfate product quality is greater than 99.9%.

[0074] S13, the cesium extraction liquid obtained after step S12 is mixed with the extracted organic phase and then subjected to three-stage extraction, three-stage washing and three-stage stripping in sequence, and the obtained rubidium-rich liquid is evaporated and concentrated to obtain a rubidium sulfate product; the extracted organic phase contains an extractant t-BAMBP and a diluent kerosene, and the volume concentration of t-BAMBP is 30%; the washing water is 30g / L dilute sulfuric acid solution, and the stripping liquid is 40g / L sulfuric acid solution. Among them, the extraction rate of rubidium is greater than 95%, the extraction rate of lithium is less than 1%, the extraction rate of potassium and sodium is less than 0.5%, and the quality of the rubidium sulfate product is greater than 99.9%. The rubidium-free organic phase obtained after stripping is returned to continue to be used for rubidium extraction.

[0075] S14, the rubidium extraction residual solution obtained after step S13, a part of which is returned to step S9 as a carbonate-containing material for decalcification after degreasing, and the other part is added with dilute sulfuric acid solution after degreasing, and the excess sulfate and residual carbonate are removed by acid-base neutralization, and then returned to step S10 for evaporation and concentration. The dilute sulfuric acid solution comes from the washing liquid of cesium extraction and rubidium extraction.

[0076] The method of this embodiment can obtain the process indicators of lithium, rubidium and cesium leaching rate of about 92%, lithium comprehensive recovery rate of 88%, rubidium comprehensive recovery rate of 85%, and cesium comprehensive recovery rate of 83%. The content of Li2O and Rb2O in the water leaching slag is less than 0.1%, the content of Cs2O is less than 0.005%, and the content of K2O is less than 0.2%.

[0077] For those skilled in the art, various corresponding changes and modifications can be made according to the above technical solutions and concepts, and all of these changes and modifications should be included in the protection scope of the claims of the present invention.

Claims

1. A method for recovering lithium, rubidium and cesium potassium from low-grade iron lithium mica, characterized in that: The steps include: S1. Grinding low-grade lithium iron mica ore to obtain lithium mica concentrate by magnetic separation or flotation; S2, drying the lepidolite concentrate obtained in step S1; S3, blending, mixing and dry-grinding the lepidolite concentrate, CaSO4·2H2O, Ca(OH)2 and CaCO3 to obtain a mixed material; S4, granulating the mixed material obtained in step S3 to obtain ore pellets; S5, drying the ore mass obtained in step S4 and then roasting it; S6, taking the roasted sand obtained after roasting in step S5 out of the furnace and cooling it, then grinding and slurrying it, and pumping the obtained slurry to a leaching tank for leaching; S7, the leached pulp obtained after leaching in step S6 is subjected to solid-liquid separation, the precious leached liquid obtained by solid-liquid separation is transferred to step S9 for purification and impurity removal, the leached residue is washed, the washing liquid is returned to step S6 to participate in leaching, and the leached washing residue is transferred to step S8; S8, beneficiating the leached washing residue obtained in step S7, flotating gypsum, and drying the gypsum and returning it to step S3 for use; S9, sending the leached precious liquid obtained in step S7 to a primary purification tank, adding lime slurry to preliminarily remove F and Mg, and then further sending it to a secondary purification tank, adding carbonate-containing materials to remove calcium, to obtain a purified slurry; after the purified slurry is filtered, the purified liquid obtained by filtration is sent to step S10, the filter residue obtained by filtration is washed and dried, and then returned to step S5 for roasting, and the washed liquid obtained by washing the filter residue is returned to the primary purification tank; S10, evaporating and concentrating the purified liquid obtained in step S9, and when evaporation causes crystals to precipitate, refrigerating to precipitate potassium salt, and then filtering, and the obtained filtrate enters the treatment of step S11, the precipitated potassium salt is centrifuged, and the liquid obtained by centrifugation is returned for evaporation and concentration, and the separated solid is washed with a saturated potassium sulfate solution, and the washing liquid is returned for evaporation and concentration; S11, the filtrate obtained in step S10 is heated and potassium carbonate is added, and after the reaction is completed, the crude lithium carbonate is obtained by centrifugal filtration, and the crude lithium carbonate is refined, and the lithium precipitation mother liquor enters step S12; S12, the lithium precipitation mother liquor is mixed with the extracted organic phase, and then extracted, washed and stripped in sequence to obtain a cesium-rich solution, and the cesium-rich solution is evaporated and concentrated to obtain a cesium sulfate product; S13, the cesium-extracted liquid obtained after the extraction in step S12 is mixed with the extracted organic phase and then extracted, washed and stripped in sequence, and the obtained rubidium-rich liquid is evaporated and concentrated to obtain a rubidium sulfate product; S14: After the extraction in step S13, a part of the rubidium extract residue obtained is returned to step S9 as a carbonate-containing material for decalcification after oil removal, and the other part is added with sulfuric acid after oil removal and then returned to step S10 for evaporation and concentration.

2. The method according to claim 1, characterized in that In step S1, the mass content of Li2O in the lepidolite concentrate is 1-3%.

3. The method according to claim 1, characterized in that In step S2, the lepidolite concentrate is dried to obtain a moisture content of less than 5%.

4. The method according to claim 1, characterized in that In step S3, the mass ratio of lithium mica concentrate: CaSO4·2H2O: Ca(OH)2: CaCO3 is 1: 0.3-0.6: 0-0.3: 0-0.5; the fineness of the mixed material is -15--25um, accounting for 85%.

5. The method according to claim 1, characterized in that In step S3, the particle size of the ore mass is 1-20 mm.

6. The method according to claim 1, characterized in that In step S5, the ore mass is roasted in a rotary kiln at 850-1050° C. for 0.5-2 h.

7. The method according to claim 1, characterized in that In step S6, the roasted sand is cooled to 100-200°C after being taken out of the furnace; and leached for 0.5-2h at a leaching temperature of 50-90°C.

8. The method according to claim 1, characterized in that In step S11, the filtrate obtained in step S10 is heated and potassium carbonate in an amount of 1-1.2 times the theoretical molar amount is added, and the reaction is carried out for 1 hour; the crude lithium carbonate is sequentially hydrogenated, ion exchanged and pyrolyzed to obtain battery-grade lithium carbonate.

9. The method according to claim 1, characterized in that: The extracted organic phases described in step S12 and step S13 both contain an extractant t-BAMBP and a diluent kerosene, and the volume concentration of t-BAMBP is 10-30%; the washing water used for washing in step S12 is a 5-20 g / L dilute sulfuric acid solution, and the washing water used for washing in step S13 is a 20-30 g / L dilute sulfuric acid solution; the stripping solution used for back extraction in step S12 is a 10-30 g / L dilute sulfuric acid solution, and the stripping solution used for back extraction in step S13 is a 20-40 g / L sulfuric acid solution.

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