A method for recovering lithium, rubidium and cesium from lepidolite ore

Abstract

The present invention belongs to the technical field of metal smelting, and specifically relates to a method for recovering lithium, rubidium, and cesium from lepidolite ore. The method comprises the following steps: mixing lepidolite concentrate with a reaction agent and roasting the mixture; leaching the obtained roasted sand and an immersion liquid; purifying the obtained leached precious liquid; subjecting the obtained purified slurry to solid-liquid separation and evaporation and concentration to obtain a potassium-sodium mixed salt and a solution after potassium and sodium are removed; using an adsorption resin to remove calcium and magnesium from the solution after potassium and sodium are removed to precipitate lithium, thereby obtaining a crude lithium carbonate product and a lithium precipitation mother liquor; refining the crude lithium carbonate product to obtain battery-grade lithium carbonate; subjecting the lithium precipitation mother liquor to countercurrent continuous extraction and stripping to obtain a cesium-rich strip liquor; subjecting the obtained cesium-extracted solution to countercurrent continuous extraction and stripping to obtain a rubidium-rich strip liquor; and evaporating and crystallizing the cesium-rich strip liquor and the rubidium-rich strip liquor to obtain cesium sulfate and rubidium sulfate products. The invention realizes high leaching of lithium, rubidium and cesium, with a lithium recovery rate of 85% to 90% and a rubidium and cesium recovery rate both greater than 70%. The purity of the prepared lithium carbonate, rubidium sulfate and cesium sulfate products is greater than 99.9%.

Description

Technical Field

[0001] The invention belongs to the technical field of metal smelting, and in particular relates to a method for recovering lithium, rubidium and cesium from lepidolite ore. Background Art

[0002] Lithium, rubidium, and cesium are essential mineral raw materials for the development of industries such as aerospace, defense, and new energy. Lithium is a core raw material for lithium batteries. Rubidium and cesium, with their excellent photoelectric properties, are irreplaceable building blocks for the rubidium (cesium) atomic clocks required by satellite navigation and positioning systems like GPS and Beidou.

[0003] Spodumene and lepidolite are the most widely distributed lithium-containing minerals. Spodumene is the main source of lithium extraction minerals. Its ore composition is simple, the lithium grade is high, and the process is mature.

[0004] Lepidolite is a unique mineral that contains three key minerals: lithium, rubidium, and cesium. It also boasts the highest known rubidium content. Extracting lithium from lepidolite is more challenging than extracting lithium from spodumene. This is due to its low grade, high slag volume, and the presence of a certain amount of fluorine, which easily forms insoluble fluorides during smelting, resulting in low yields and environmental pollution. Furthermore, the rare and valuable potassium, rubidium, and cesium contained in the lepidolite ore are difficult to fully utilize.

[0005] Some lepidolite deposits are low-grade, and most are strongly hematized, with a small amount corroding to chlorite. This prevents the enrichment of weakly magnetic target minerals during magnetic separation, and the chlorite easily becomes mud during flotation, resulting in the loss of target minerals in the tailings during the beneficiation process. As a result, after magnetic separation or flotation, the Li2O content in the concentrate is only 1.8% to 2.0%, accompanied by high levels of iron and fluorine, including Fe2O3 at 13% to 15% and F at 5% to 7%. Therefore, developing cost-effective and environmentally friendly treatment processes is a key direction for the development of lepidolite resources.

[0006] The main extraction processes for lepidolite concentrate include sulfuric acid roasting, fluorine chemical method, alkali dissolution method, pressure cooking method, limestone roasting method, and sulfate roasting method. Currently, the main industrial process for processing lepidolite concentrate is potassium sulfate roasting. While the sulfuric acid roasting method can achieve a lithium, rubidium, and cesium leaching rate of over 90%, this process has a long reaction time, produces acid mist, and creates a poor operating environment. Impurities such as aluminum and magnesium are leached from the ore in large quantities, and the aluminum in the leachate forms Al(OH)3, which absorbs lithium, rubidium, and cesium, making separation difficult. This results in a low overall recovery rate for lithium, rubidium, and cesium. Furthermore, the high magnesium content in the leachate also affects product quality.

[0007] The advantages of the fluorine chemical method are low leaching temperature, short reaction time, low energy consumption and high lithium extraction efficiency, but it has disadvantages such as large acid consumption and the generation of HF gas during the reaction, which pollutes the environment and harms human health.

[0008] The advantages of lithium extraction from lepidolite alkali dissolution are as follows: 1) Li in lepidolite can be completely dissolved; 2) the valuable metal conversion process can be completed in one go; 3) HF corrosion equipment will not be generated; 4) costs can be reduced by selling aluminum silicate sol byproducts. However, its shortcomings are also very obvious. The concentrated alkali waste liquid remaining after the reaction is difficult to recycle, the reaction requires pressure leaching, which has high requirements for equipment, and the impurity cations in the solution (mainly Mg 2+ ) will co-precipitate with the crude lithium extraction product Li2CO3, reducing the purity of lithium carbonate.

[0009] The pressure cooking method for lithium extraction involves first defluorinating and roasting lepidolite concentrate with water vapor at high temperature, which changes the mineral structure of the defluorinated lepidolite. The concentrate is then mixed with a mineral reconstructor (such as alkali, chloride, sulfate, and carbonate) in a certain proportion and subjected to high-pressure leaching in a pressure cooker reactor, resulting in a mother liquor containing lithium and other valuable metal compounds. This mother liquor is then purified, decontaminated, and lithium precipitated to yield the desired lithium salt product. The pressure cooking method offers the advantages of a simple process, high Li2O leaching rate, short pressure cooking time, low material throughput, minimal corrosion to reaction equipment, and good overall utilization. However, the pressure cooking method requires defluorinating and roasting, which poses a pressure on the environment. The reaction must be carried out at high temperature and pressure, which imposes stringent reaction conditions, poses safety risks, and places high demands on equipment and operating procedures.

[0010] The limestone roasting method for lithium extraction generally involves mixing lepidolite concentrate with limestone containing greater than 54% CaO in a mass ratio of 1:3. The mixture is then finely ground and roasted at high temperature. The roasted material is then quenched with water, finely ground, and leached to obtain a lithium-containing solution. Subsequently, Li2CO3 is obtained through purification, impurity removal, and lithium precipitation. The advantages of the limestone method lie in its simple process flow, readily available raw materials, and low cost. However, it also has the disadvantages of large material throughput, low lithium recovery rate, high energy consumption, and large slag production. Since the 1990s, with the continuous application and improvement of other lithium extraction processes, the limestone method has been gradually replaced and eventually eliminated.

[0011] The advantages of the sulfate process are its high applicability and its ability to process lepidolite ores of different grades. Compared with the sulfuric acid process, the probability of sulfate reacting with aluminum in lepidolite to form soluble aluminum salts is low, resulting in less lithium loss during the subsequent chemical precipitation process due to aluminum removal. The process also has short roasting times and a high lithium precipitation rate. However, its disadvantages are also quite obvious, mainly including the following: strict roasting temperature requirements, which can easily cause ring formation in the furnace; high energy consumption; the generation of low-solubility LiKSO4 complex salts, which affects the lithium concentration and precipitation process; some rubidium and cesium remain in the residue and are difficult to utilize; waste gases such as HF and SO2 / SO3 generated by roasting need to be treated to reduce environmental pollution; and the cost of roasting with K2SO4 as sulfate is high. Using Na2SO4 instead of K2SO4 reduces costs, but when a certain amount is reached, a glass phase will be generated, affecting the normal operation of the process.

[0012] Chinese patent CN107937733A discloses a process for extracting lithium, potassium, rubidium, and cesium from lepidolite. The process comprises grinding lepidolite ore into lepidolite powder, wherein -100 mesh accounts for more than 80 wt %; mixing the ore and the powder, calcining the mixture at 700-1000°C, adding alkali to the resulting water extract to adjust the pH to 8.5-9.5, and performing solid-liquid separation; concentrating the filtrate, adding carbonate, performing solid-liquid separation to remove calcium, evaporating and concentrating the solution after decalcification, adding carbonate to precipitate lithium, adding hydroxide ions to the solution after lithium precipitation, adding hydroxide ions to the residual solution after cesium extraction, extracting rubidium, and evaporating and concentrating the residual solution after rubidium extraction to prepare potassium salt. Although CN107937733A can extract lithium, rubidium, cesium, and potassium from lepidolite ore, it uses chloride as an auxiliary agent, which is highly corrosive to equipment. Summary of the Invention

[0013] The present invention aims to provide a method for recovering lithium, rubidium and cesium from lepidolite concentrate. The method provided by the present invention can simultaneously improve the recovery rates of lithium, rubidium and cesium valuable metals, and obtain high-purity lithium salt, rubidium salt and cesium salt products. Therefore, the method provided by the present invention is an efficient method for recovering lithium, rubidium and cesium from lepidolite ore with a high comprehensive recovery rate.

[0014] In order to achieve the above object, the present invention provides the following technical solutions:

[0015] The present invention provides a method for recovering lithium, rubidium and cesium from lepidolite concentrate, comprising the following steps:

[0016] Mixing lepidolite concentrate and a reaction agent and dry grinding them to obtain a mixed material; the reaction agent includes gypsum and CaCO3;

[0017] calcining the mixed material to obtain calcine;

[0018] The calcined sand and the impregnation liquid are mixed and wet-ground to obtain a slurry; the impregnation liquid is water or an aqueous solution containing lithium ions;

[0019] leaching the ore pulp to obtain leached ore pulp;

[0020] performing a first solid-liquid separation on the leached slurry to obtain a leached precious liquid;

[0021] Mixing the leached noble solution and lime slurry, performing pH adjustment and a first purification, to obtain a first purified slurry;

[0022] mixing the first purified slurry and the carbonate-containing solution for a second purification to obtain a second purified slurry;

[0023] separating the second purified slurry into a second solid-liquid phase to obtain a purified liquid and impurity-removed residue;

[0024] evaporating and concentrating the purified liquid to obtain a concentrated liquid;

[0025] Cooling the concentrated liquid and performing a third solid-liquid separation to obtain a mixed salt containing impurities potassium and sodium and a concentrated solution containing lithium;

[0026] The mixed salt containing impurities of potassium and sodium is first washed with tap water or a potassium and sodium solution to obtain a purified potassium and sodium mixed salt and a washed solution; the lithium-containing concentrated solution is subjected to an adsorption resin to remove calcium and magnesium to obtain a calcium and magnesium-removed solution;

[0027] The calcium and magnesium-removed solution is mixed with a sodium carbonate solution to precipitate lithium, thereby obtaining a crude lithium carbonate product and a lithium precipitation mother liquor, respectively;

[0028] The crude lithium carbonate is subjected to slurrying, hydrogenation, filter pressing, pyrolysis and drying to obtain battery-grade lithium carbonate;

[0029] The lithium-precipitated mother liquor is mixed with the cesium-unloaded organic phase and subjected to a first countercurrent continuous extraction to obtain a cesium-loaded organic phase and a cesium-extracted liquid, respectively; the cesium-loaded organic phase is subjected to a first continuous washing to obtain a purified cesium-loaded organic phase; the purified cesium-loaded organic phase is subjected to a first continuous stripping to obtain a cesium-rich stripping solution and a cesium-unloaded organic phase; the cesium-rich stripping solution is evaporated and crystallized to obtain cesium sulfate;

[0030] The cesium-extracted liquid is mixed with the rubidium-free organic phase and subjected to a second countercurrent continuous extraction to obtain a rubidium-loaded organic phase and a rubidium-extracted liquid; the rubidium-loaded organic phase is subjected to a second continuous washing to obtain a purified rubidium-loaded organic phase; the purified rubidium-loaded organic phase is subjected to a second continuous back extraction to obtain a rubidium-rich back extract and a rubidium-free organic phase; the rubidium-rich back extract is evaporated and crystallized to obtain rubidium sulfate.

[0031] Preferably, the reaction reagent further comprises any one or more of lime, Na2SO4 and K2SO4;

[0032] The mass ratio of the lepidolite concentrate to the reaction reagent is 1:(0.2-2.2);

[0033] In the lepidolite concentrate, the mass percentage of lepidolite concentrate with a fineness of less than 0.153 mm is ≥50%.

[0034] Preferably, the reaction reagents include gypsum, CaCO3, lime, Na2SO4 and K2SO4;

[0035] The mass ratio of the gypsum measured as CaSO4·2H2O to the lepidolite concentrate is (0.1-0.6):1;

[0036] The mass ratio of the CaCO3 measured as pure CaCO3 to the lepidolite concentrate is (0.1-0.7):1;

[0037] The mass ratio of the lime, measured as quicklime CaO, to the lepidolite concentrate is (0-0.3):1, and is not 0;

[0038] The mass ratio of the Na2SO4 to the lepidolite concentrate is (0-0.3):1, and is not 0;

[0039] The mass ratio of the K2SO4 and the lepidolite concentrate is (0-0.3):1, and is not 0.

[0040] Preferably, the calcination temperature is 850-1150° C., and the calcination holding time is 0.5-5 h.

[0041] Preferably, the leaching temperature is 5 to 95° C.; the leaching time is 0.1 to 5 hours; and the leaching is performed by pipeline leaching or single-stage or multi-stage stirred tank leaching.

[0042] Preferably, the first solid-liquid separation obtains a filter residue phase; after the first solid-liquid separation, the further step includes: performing a second washing on the filter residue phase to obtain a washed filter cake and a washing liquid; the washed filter cake is used to make a raw material for a gelling material or a raw material for cement;

[0043] The second washing includes one or more combinations of countercurrent thickening washing, belt filtration washing and filter press washing, and the number of stages of the countercurrent thickening washing is 2 to 6; the washing water used in the second washing is water and / or condensed water obtained by the evaporation and concentration; the washing liquid obtained by the washing is reused as the impregnation liquid.

[0044] Preferably, the first washing includes one or more stages of plate and frame filtration washing, belt filtration washing or centrifugal filtration washing; the purified potassium and sodium mixed salt obtained by the first washing is reused as part of the reaction reagent, and the washed liquid obtained by the first washing is returned to the first purification.

[0045] Preferably, when the leached noble solution and lime slurry are mixed for pH adjustment and first purification, the mass concentration of the lime slurry is 5% to 30%;

[0046] The impurity-removing residue is recycled as part of the reaction reagent.

[0047] Preferably, during the first countercurrent continuous extraction and the second countercurrent continuous extraction, the extractant is t-BAMBP and the diluent is kerosene; in the mixed system consisting of the extractant and the diluent, the volume percentage of the extractant is 10% to 30%;

[0048] During the first continuous washing and the second continuous washing, the washing agent is a first sulfuric acid solution, and the concentration of the first sulfuric acid solution is independently 2 to 20 g / L; the first continuous washing and the second continuous washing also respectively obtain a potassium-sodium washing solution with a low cesium concentration and a potassium-sodium washing solution with a low rubidium concentration, and the potassium-sodium washing solution with a low cesium concentration and the potassium-sodium washing solution with a low rubidium concentration are reused as the impregnation solution.

[0049] During the first continuous stripping and the second continuous stripping, the stripping liquid is a second sulfuric acid solution, and the mass concentration of the second sulfuric acid solution is independently 5 to 30 g / L;

[0050] The mass concentration of the second sulfuric acid solution is greater than the mass concentration of the first sulfuric acid solution.

[0051] Preferably, after the second countercurrent continuous extraction, the method further comprises: recycling the rubidium extraction solution as an impregnation solution and / or recycling the rubidium extraction solution as a carbonate-containing solution for the second purification.

[0052] The present invention first mixes lepidolite concentrate and a reaction agent and performs dry grinding. By regulating the type of the reaction agent, the reaction agent can fully react with the lepidolite concentrate during roasting, thereby replacing lithium, rubidium, cesium and potassium in the lepidolite concentrate. Then, the obtained roasted sand is mixed with an immersion liquid and leached, so that lithium, rubidium, cesium and potassium in the lepidolite concentrate can be efficiently leached. The results of the embodiment show that the leaching rates of lithium, rubidium and cesium in the present invention are all 90% to 95%, and the leaching rate of potassium is 70% to 90%. At the same time, it ensures that aluminum, iron and silicon in the lepidolite concentrate are not leached, and F in the lepidolite concentrate is mainly fixed in the form of CaF2, which not only avoids corrosion of equipment but also avoids roasting loss of lithium. The results of the embodiment show that during leaching, only trace amounts of F (20 to 30 mg / L) and Mg (10 to 20 mg / L) are lost. L) is leached into the precious leaching solution, reducing the pressure of subsequent impurity removal; the precious leaching solution is subjected to a first purification by lime slurry to remove F and Mg in the precious leaching solution, and is subjected to a second purification by a carbonate-containing solution to remove Ca in the precious leaching solution, followed by evaporation and concentration to obtain a mixed salt containing impurities of potassium and sodium and a solution after potassium and sodium are removed, and the mixed salt containing impurities of potassium and sodium is washed to obtain a potassium and sodium mixed salt, which can be used as a reaction reagent; the concentrated solution after potassium and sodium are further subjected to an adsorption resin to remove calcium and magnesium, and the adsorbed solution is mixed with a sodium carbonate solution to precipitate lithium to obtain a crude lithium carbonate product; the lithium precipitated mother liquor is subjected to a first countercurrent continuous extraction-washing-strip extraction to obtain a cesium-rich strip solution; the cesium extraction residual liquor is subjected to a second countercurrent continuous extraction-washing-strip extraction to obtain a rubidium-rich strip solution; the results of the embodiment show that the comprehensive lithium recovery rate of the present invention is 85% to 90%, and the rubidium and cesium recovery rates are both greater than 80%. In summary, the method provided by the present invention is a process route of "reaction reagent-roasting-leaching-leachate purification and impurity removal-impurity removal liquid concentration and evaporation-mixed salt precipitation-salt precipitation and lithium carbonate precipitation-lithium precipitation and liquid extraction of cesium-cesium extraction and liquid extraction of rubidium". The leaching rates of lithium, rubidium and cesium are 90% to 95%, the leaching rate of potassium is 70% to 90%, the comprehensive recovery rate of lithium is 85% to 90%, and the comprehensive recovery rate of rubidium and cesium is greater than 80%.

[0053] Furthermore, in the present invention, the impurity-removed slag and potassium-sodium mixed salt are recycled as reaction reagents, thereby reducing the amount of added reagents and saving reagent costs.

[0054] Furthermore, in the present invention, after the second countercurrent continuous extraction, the method further includes: recycling the rubidium extraction solution as an impregnation solution and / or recycling the rubidium extraction solution as a carbonate-containing solution for the second purification. Recycling the rubidium extraction solution reduces the use of external impregnation solution and carbonate reagents, saving reagent costs. The method provided by the present invention achieves acid-base self-balancing without the need for additional pH adjustment reagents.

[0055] Furthermore, in the present invention, the washing process includes one or more combinations of countercurrent thickening washing, belt filtration washing, and filter press washing, with the countercurrent thickening washing having two to six stages. The washing water used in the washing process is water and / or condensed water obtained by the evaporation and concentration process, and the washing liquid obtained by the washing process is reused as the impregnation liquid. By reusing the washing liquid, the present invention reduces the use of external impregnation liquid and recovers the target metal in the washing liquid, saving reagent costs while increasing the target metal recovery rate. BRIEF DESCRIPTION OF THE DRAWINGS

[0056] Figure 1 Flowchart of a method for recovering lithium, rubidium and cesium from lepidolite concentrate provided in an embodiment of the present invention. DETAILED DESCRIPTION

[0057] The present invention provides a method for recovering lithium, rubidium and cesium from lepidolite concentrate, comprising the following steps:

[0058] Mixing lepidolite concentrate and a reaction agent and dry grinding them to obtain a mixed material; the reaction agent includes gypsum and CaCO3;

[0059] calcining the mixed material to obtain calcine;

[0060] The calcined sand and the impregnation liquid are mixed and wet-ground to obtain a slurry; the impregnation liquid is water or an aqueous solution containing lithium ions;

[0061] leaching the ore pulp to obtain leached ore pulp;

[0062] performing a first solid-liquid separation on the leached slurry to obtain a leached precious liquid;

[0063] Mixing the leached noble solution and lime slurry, performing pH adjustment and a first purification, to obtain a first purified slurry;

[0064] mixing the first purified slurry and the carbonate-containing solution for a second purification to obtain a second purified slurry;

[0065] separating the second purified slurry into a second solid-liquid phase to obtain a purified liquid and impurity-removed residue;

[0066] evaporating and concentrating the purified liquid to obtain a concentrated liquid;

[0067] Cooling the concentrated liquid and performing a third solid-liquid separation to obtain a mixed salt containing impurities potassium and sodium and a concentrated solution containing lithium;

[0068] The mixed salt containing impurities of potassium and sodium is first washed with tap water or a solution of mixed salt containing potassium and sodium to obtain a purified mixed salt of potassium and sodium and a washed solution;

[0069] removing calcium and magnesium from the lithium-containing concentrated solution by using an adsorption resin to obtain a calcium- and magnesium-removed solution;

[0070] The calcium and magnesium-removed solution is mixed with a sodium carbonate solution to precipitate lithium, thereby obtaining a crude lithium carbonate product and a lithium precipitation mother liquor, respectively;

[0071] The crude lithium carbonate is subjected to slurrying, hydrogenation, filter pressing, pyrolysis and drying to obtain battery-grade lithium carbonate;

[0072] The lithium-precipitated mother liquor is mixed with the cesium-unloaded organic phase and subjected to a first countercurrent continuous extraction to obtain a cesium-loaded organic phase and a cesium-extracted liquid, respectively; the cesium-loaded organic phase is subjected to a first continuous washing to obtain a purified cesium-loaded organic phase; the purified cesium-loaded organic phase is subjected to a first continuous stripping to obtain a cesium-rich stripping solution and a cesium-unloaded organic phase; the cesium-rich stripping solution is evaporated and crystallized to obtain cesium sulfate;

[0073] The cesium-extracted liquid is mixed with the rubidium-free organic phase and subjected to a second countercurrent continuous extraction to obtain a rubidium-loaded organic phase and a rubidium-extracted liquid; the rubidium-loaded organic phase is subjected to a second continuous washing to obtain a purified rubidium-loaded organic phase; the purified rubidium-loaded organic phase is subjected to a second continuous back extraction to obtain a rubidium-rich back extract and a rubidium-free organic phase; the rubidium-rich back extract is evaporated and crystallized to obtain rubidium sulfate.

[0074] In the present invention, unless otherwise specified, all preparation raw materials / components are commercially available products well known to those skilled in the art.

[0075] The present invention mixes lepidolite concentrate and reaction reagents and performs dry grinding to obtain a mixed material; the reaction reagents include gypsum and CaCO3, and preferably also include any one or more of lime, Na2SO4 and K2SO4; the gypsum can be CaSO4·2H2O, CaSO4 (anhydrous gypsum), or industrial by-product gypsum; the lime can be quicklime (CaO), slaked lime (Ca(OH)2) or industrial stone; the CaCO3 can be analytically pure or industrial limestone; the Na2SO4 and K2SO4 can be analytically pure, or industrial by-products such as sodium sulfate and potassium sulfate, or mixed salt produced in a later stage (mixed salt of sodium sulfate and potassium sulfate).

[0076] In the present invention, the preparation method of the lepidolite concentrate preferably comprises the following steps:

[0077] Crushing the lepidolite ore to obtain lepidolite ore;

[0078] The lepidolite ore is subjected to magnetic separation or flotation, and then dried to obtain the lepidolite concentrate.

[0079] The present invention has no special requirements on the specific implementation process of the crushing, magnetic separation or flotation.

[0080] In the present invention, the mass percentage of Li2O in the lepidolite concentrate is preferably 1 to 4.5%.

[0081] In one or more specific embodiments of the present invention, the mass percentages of Li2O, Rb2O, Cs2O and K2O in the lepidolite concentrate are 1.55%, 0.72%, 0.048% and 8.37% respectively, the mass percentages of iron and fluorine are 9.77% and 5.52% respectively, and the mass percentages of the main gangues are SiO2 and Al2O3, which are 50.58% and 19.28% respectively.

[0082] In one or more specific embodiments of the present invention, the mass percentages of Li2O, Rb2O, Cs2O and K2O in the lepidolite concentrate are 1.75%, 0.72%, 0.048% and 8.37% respectively, the mass percentages of iron and fluorine are 9.77% and 5.52% respectively, and the main gangues are SiO2 and Al2O3, with mass percentages of 50.58% and 19.28% respectively.

[0083] In one or more specific embodiments of the present invention, the mass percentages of Li2O, Rb2O, Cs2O and K2O in the lepidolite concentrate are 1.89%, 1.02%, 0.06% and 3.22% respectively, and the main gangues are SiO2 and Al2O3, with mass percentages of 56.8% and 19.28% respectively.

[0084] In the present invention, the mass percentage of the lepidolite concentrate with a fineness of less than 0.153 mm is ≥ 50%.

[0085] In the present invention, the mass ratio of the lepidolite concentrate to the reaction reagent is preferably 1:(0.2-2.2), more preferably 1:(0.35-1.64).

[0086] In the present invention, the reaction reagent is preferably gypsum, CaCO3, lime, Na2SO4 and K2SO4; the gypsum is calculated by the mass of CaSO4·2H2O,

[0087] The mass ratio of the gypsum to the lepidolite concentrate is preferably (0.1-0.6):1; more preferably (0.35-0.55):1.

[0088] The mass ratio of the CaCO3 measured as pure CaCO3 to the lepidolite concentrate is preferably (0.1-0.7):1; more preferably (0.3-0.5):1.

[0089] The mass ratio of the lime, calculated as quicklime CaO, to the lepidolite concentrate is preferably (0-0.3):1, and is not 0; more preferably (0.05-0.19):1.

[0090] The mass ratio of the Na2SO4 and the lepidolite concentrate is preferably (0-0.3):1, and is not 0; more preferably (0.05-0.2):1.

[0091] The mass ratio of the K2SO4 and the lepidolite concentrate is preferably (0-0.3):1, and is not 0; more preferably (0.05-0.2):1.

[0092] In the present invention, the K2SO4 and Na2SO4 are preferably replaced by a mixture of potassium and sodium obtained by evaporation and concentration.

[0093] In the present invention, part of the gypsum and CaCO3 comes from the impurity removal slag, and part is a new added agent.

[0094] The present invention has no special requirements on the specific implementation of the dry grinding.

[0095] The present invention roasts the mixed material to obtain roasted sand.

[0096] In the present invention, the calcination temperature is preferably 850-1150° C., more preferably 950-1050° C.; the calcination holding time is preferably 0.5-5 h, more preferably 1-2 h.

[0097] In the present invention, the roasting is preferably carried out in a rotary kiln or a tunnel kiln.

[0098] After obtaining the calcined sand, the present invention mixes the calcined sand with an impregnation liquid and performs wet grinding to obtain a slurry; the impregnation liquid includes water or a lithium ion aqueous solution;

[0099] In the present invention, the calcined sand is preferably cooled to 100-200° C. and then mixed with the impregnation liquid.

[0100] In the present invention, the impregnation solution is preferably the rubidium extraction residue and the washing solution obtained by washing. In a specific embodiment of the present invention, the washing solution is preferably the first-stage concentrated overflow solution of the countercurrent concentrated washing.

[0101] In the present invention, the wet grinding is preferably performed in a wet grinder.

[0102] In the present invention, the liquid-to-solid ratio of the slurry is preferably 2:1.

[0103] After obtaining the ore pulp, the present invention leaches the ore pulp to obtain leached ore pulp.

[0104] In the present invention, the leaching temperature is preferably 5 to 95° C., more preferably 35 to 80° C.; the leaching time is preferably 0.5 to 5 h, more preferably 1 to 1.5 h; and the leaching is preferably two-stage stirring leaching.

[0105] After obtaining the leached ore pulp, the present invention performs a first solid-liquid separation on the leached ore pulp to obtain leached precious liquid.

[0106] In the present invention, the specific implementation of the first solid-liquid separation is preferably dense filtration, plate and frame filter press or belt filtration.

[0107] In the present invention, a specific embodiment of the first solid-liquid separation is preferably that during dense filtration, the leached precious liquid is dense overflow.

[0108] In the present invention, the first solid-liquid separation also obtains a filter residue phase; the present invention preferably further comprises: performing a second washing on the filter residue phase to obtain a washed filter cake; the washed filter cake can be used as a raw material and directly applied to the preparation of cementitious materials or cement.

[0109] In the present invention, the second washing preferably includes one or more combinations of countercurrent dense washing (CCD), belt filtration washing and filter press washing. The number of stages of the countercurrent dense washing is preferably 2 to 6, more preferably 2 to 4.

[0110] In a specific embodiment of the present invention, the second washing comprises sequentially performing three-stage countercurrent thickening washing and one-stage belt filtration washing. In the present invention, the three-stage countercurrent thickening washing and one-stage belt filtration washing comprise sequentially performing a first-stage countercurrent thickening washing (CCD1), a second-stage countercurrent thickening washing (CCD2), a third-stage countercurrent thickening washing (CCD3), and a first-stage belt filtration washing. In the present invention, the washing water of the belt filtration washing is preferably water or condensed water obtained by evaporation and concentration; the washing water of the third-stage countercurrent thickening washing is preferably the filtrate of the belt filtration washing; the washing water of the second-stage countercurrent thickening washing is preferably the third-stage thickening overflow of the third-stage countercurrent thickening washing; the washing water of the first-stage countercurrent thickening washing is preferably the second-stage thickening overflow of the second-stage thickening countercurrent washing; the filter residue phase is mixed with the second-stage thickening overflow to perform the first-stage countercurrent thickening washing to obtain the first-stage thickening overflow and the first-stage thickening underflow; the first-stage thickening underflow is mixed with the third-stage thickening overflow to perform the second-stage countercurrent thickening washing to obtain the second-stage thickening overflow and the second-stage thickening underflow; the second-stage thickening underflow is mixed with the filtrate of the first-stage belt filtration to perform the third-stage countercurrent thickening washing to obtain the third-stage thickening overflow and the third-stage thickening underflow; the third-stage thickening underflow is mixed with clean water or condensed water obtained by evaporation and concentration to perform belt filtration washing; and the belt filtration filtrate and belt filtration filter residue are obtained. filter cake.

[0111] In the present invention, the concentration of lithium ions in the first-stage dense overflow is preferably 0.6 to 3 g / L, more preferably 1.5 to 3 g / L.

[0112] In the present invention, the first-stage concentrated overflow is preferably reused as part of the impregnation liquid.

[0113] In the present invention, the concentration of lithium ions in the leaching precious solution is preferably 2.5 to 6 g / L, more preferably 3 to 6 g / L, the concentration of rubidium ions is preferably 1.4 to 3 g / L, more preferably 1.6 to 3 g / L, the concentration of cesium ions is preferably 100 to 200 mg / L, the concentration of potassium ions is preferably 10 to 15 g / L, and the concentration of sodium ions is preferably 200 mg / L to 15 g / L.

[0114] After the precious solution is leached, the present invention mixes the leached precious solution with lime slurry for a first purification to obtain a first purified slurry.

[0115] In the present invention, the mass percentage of the lime slurry is 5-50%, and the preferred concentration of the lime slurry is 20%.

[0116] In the present invention, the lime slurry is calculated based on calcium oxide, and the mass of calcium oxide in the lime slurry is preferably 1 / 8 of the theoretical amount of sulfate in the leaching precious solution.

[0117] In the present invention, the fluoride ions and magnesium ions are preferably removed by reacting the lime slurry with the negative ions and magnesium ions in the leaching precious solution.

[0118] In the present invention, the temperature of the first purification is preferably 50-90° C., and the holding time of the first purification is preferably 1 hour; the first purification is preferably performed under stirring.

[0119] After obtaining the first purified slurry, the present invention mixes the first purified slurry with a carbonate-containing solution to perform a second purification to obtain a second purified slurry and impurity-removed slag.

[0120] In the present invention, the carbonate-containing solution is preferably a portion of the rubidium extraction residue.

[0121] In the present invention, calcium ions in the first purified slurry are preferably removed by the carbonate-containing solution.

[0122] In the present invention, the molar amount of carbonate in the carbonate-containing solution is preferably equal to the molar amount of calcium ions in the first purified slurry.

[0123] In the present invention, the temperature of the second purification is preferably 50-90° C., and the holding time of the second purification is preferably 0.5-1 h; the second purification is preferably carried out under stirring.

[0124] After obtaining the second purified slurry, the present invention performs a second solid-liquid separation on the second purified slurry to obtain a purified liquid.

[0125] In the present invention, the second solid-liquid separation is preferably plate and frame filter pressing.

[0126] In the present invention, the second solid-liquid separation further produces a solid phase residue, and the present invention preferably further comprises: drying the solid phase residue and using it as a reaction reagent.

[0127] After obtaining the purified liquid, the present invention evaporates and concentrates the purified liquid to obtain a concentrated liquid; and after cooling the concentrated liquid, performs a third solid-liquid separation to obtain a mixed salt containing impurities potassium and sodium and a concentrated liquid containing lithium.

[0128] In the present invention, the endpoint of the evaporation concentration is preferably: crystals are precipitated in the purified liquid.

[0129] The present invention preferably cools the purified liquid containing precipitated crystals and then performs a third solid-liquid separation. The cooling is preferably refrigerated crystallization, and the third solid-liquid separation is preferably centrifugal separation. The present invention preferably uses the potassium-sodium mixed salt as a reaction reagent. In the present invention, the potassium-sodium mixed salt preferably replaces K2SO4 and Na2SO4 in the reaction reagent.

[0130] After obtaining the impurity potassium and sodium mixed salt, the present invention performs a first washing on the impurity potassium and sodium mixed salt with tap water or a potassium and sodium mixed salt solution to obtain a purified potassium and sodium mixed salt and a washed liquid.

[0131] In the present invention, the first washing preferably includes one or more stages of plate and frame filtration washing, belt filtration washing or centrifugal filtration washing; the purified potassium and sodium mixed salt obtained by the first washing is preferably reused as part of the reaction reagent, and the washed liquid obtained by the first washing is preferably returned to the first purification.

[0132] After obtaining the lithium-containing concentrated solution, the present invention uses an adsorption resin to remove calcium and magnesium from the lithium-containing concentrated solution to obtain a calcium- and magnesium-removed solution.

[0133] In the present invention, the adsorption resin is preferably LSC-500.

[0134] In the present invention, after the adsorption resin removes calcium and magnesium, an adsorbed resin is obtained. In the present invention, the adsorbed resin is preferably subjected to post-treatment. In the present invention, the post-treatment preferably includes the following steps: washing the adsorbed resin with water, and the resulting washing liquid preferably replaces part of the carbonate-containing solution during the second purification; the washed resin obtained by the washing is subjected to hydrochloric acid desorption, and the desorption liquid obtained by the desorption is evaporated and crystallized to obtain calcium chloride, which is stored for unified treatment.

[0135] After obtaining the calcium and magnesium-removed liquid, the present invention mixes the calcium and magnesium-removed liquid with a sodium carbonate solution to precipitate lithium, thereby obtaining a crude lithium carbonate product and a lithium precipitation mother liquor.

[0136] In the present invention, the mass of the sodium carbonate in the sodium carbonate solution is preferably 1 to 1.2 times the theoretical mass of lithium ions in the solution after precipitation of calcium and magnesium removal.

[0137] In the present invention, the temperature of the lithium deposition is preferably 70-95° C., and the holding time of the lithium deposition is preferably 1 hour.

[0138] In the present invention, after the lithium precipitation, a lithium precipitation solution is obtained, and the present invention preferably performs solid-liquid separation on the lithium precipitation solution to obtain crude lithium carbonate and lithium precipitation mother liquor. In the present invention, the solid-liquid separation is preferably centrifugal separation.

[0139] The invention obtains battery-grade lithium carbonate by subjecting the crude lithium carbonate to slurrying, hydrogenation, filter pressing, pyrolysis and drying.

[0140] In the present invention, the purity of the battery-grade lithium carbonate product is preferably ≥99.9%.

[0141] After obtaining the lithium precipitation mother liquor, the present invention mixes the lithium precipitation mother liquor with a cesium empty organic phase and performs a first countercurrent continuous extraction to respectively obtain a cesium-loaded organic phase and a cesium-extracted liquid; the cesium-loaded organic phase is subjected to a first continuous washing to obtain a pure cesium-loaded organic phase; the pure cesium-loaded organic phase is subjected to a first continuous stripping to obtain a cesium-rich stripping solution and a cesium empty organic phase; the cesium-rich stripping solution is evaporated and crystallized to obtain cesium sulfate.

[0142] In the present invention, during the first countercurrent continuous extraction, the extractant is preferably t-BAMBP, and the diluent is preferably kerosene; in the mixed system consisting of the extractant and the diluent, the volume percentage of the extractant is preferably 10-30%, more preferably 15-25%.

[0143] In the present invention, during the first continuous washing, the detergent is preferably a sulfuric acid solution, and the mass concentration of the sulfuric acid solution is preferably 5 to 20 g / L, more preferably 10 to 15 g / L.

[0144] In the present invention, during the first continuous stripping, the stripping liquid is preferably a sulfuric acid solution, and the mass concentration of the sulfuric acid solution is preferably 10 to 30 g / L, more preferably 15 to 25 g / L.

[0145] In the present invention, the mass concentration of the sulfuric acid solution used in the first continuous stripping is greater than the mass concentration of the sulfuric acid solution used in the first continuous washing.

[0146] In the present invention, the first countercurrent continuous extraction is preferably one to five-stage extraction; the first continuous washing is preferably one to five-stage washing; and the first continuous stripping is preferably one to five-stage stripping.

[0147] After obtaining the cesium-extracted liquid, the present invention mixes the cesium-extracted liquid with a rubidium-free organic phase and performs a second countercurrent continuous extraction to obtain a rubidium-loaded organic phase and a rubidium-extracted liquid. The rubidium-loaded organic phase is subjected to a second continuous washing to obtain a pure rubidium-loaded organic phase. The pure rubidium-loaded organic phase is subjected to a second continuous stripping to obtain a rubidium-rich stripping solution and a rubidium-free organic phase. The rubidium-rich stripping solution is evaporated and crystallized to obtain rubidium sulfate.

[0148] In the present invention, during the second countercurrent continuous extraction, the extractant is preferably t-BAMBP, and the diluent is preferably kerosene; in the mixed system consisting of the extractant and the diluent, the volume percentage of the extractant is preferably 10-30%, more preferably 15-25%.

[0149] In the present invention, during the second continuous washing, the detergent is preferably a sulfuric acid solution, and the mass concentration of the sulfuric acid solution is preferably 5 to 20 g / L, more preferably 10 to 15 g / L.

[0150] In the present invention, during the second continuous stripping, the stripping liquid is preferably a sulfuric acid solution, and the mass concentration of the sulfuric acid solution is preferably 10 to 30 g / L, more preferably 15 to 25 g / L.

[0151] In the present invention, the mass concentration of the sulfuric acid solution used in the second continuous stripping is greater than the mass concentration of the sulfuric acid solution used in the second continuous washing.

[0152] In the present invention, the second countercurrent continuous extraction is preferably one to five-stage extraction; the second continuous washing is preferably one to five-stage washing; and the second continuous stripping is preferably one to five-stage stripping.

[0153] In the present invention, after the second countercurrent continuous extraction, the method further comprises: recycling the rubidium extraction solution as an impregnation solution and / or recycling the rubidium extraction solution as a carbonate-containing solution for the second purification.

[0154] In the present invention, the mass content of lithium ions in the rubidium extraction residue is preferably 2.5 g / L.

[0155] In the present invention, when the impregnation solution is preferably a mixed solution of the first-stage concentrated overflow and the rubidium extraction raffinate, the volume ratio of the first-stage concentrated overflow and the rubidium extraction raffinate is (4-6):1, preferably 6:1.

[0156] In the present invention, before the rubidium extraction residual solution is used, the present invention preferably performs deoiling treatment on the rubidium extraction residual solution. In the present invention, the deoiling treatment is preferably performed by deoiling with an oil separator or deoiling adsorption material.

[0157] In order to further illustrate the present invention, the technical solution provided by the present invention is described in detail below with reference to the accompanying drawings and embodiments, but they should not be construed as limiting the scope of protection of the present invention.

[0158] Example 1

[0159] This embodiment uses the lepidolite ore produced in a lepidolite mine in Hunan as raw material;

[0160] The lepidolite ore is crushed and then floated to obtain lepidolite concentrate. The main components of the lepidolite concentrate are shown in Table 1. It can be concluded from Table 1 that the mass percentages of Li2O, Rb2O, Cs2O and K2O in the lepidolite concentrate are 1.55%, 0.72%, 0.048% and 8.37% respectively. The iron and fluorine contents are relatively high, at 9.77% and 5.52% respectively. The main gangues are SiO2 and Al2O3, at contents of 50.58% and 19.28% respectively.

[0161] Table 1 Main components and contents of lepidolite concentrate in Example 1

[0162] element <![CDATA[Li2O]]> <![CDATA[Rb2O]]> <![CDATA[Cs2O]]> <![CDATA[WO3]]> Sn <![CDATA[SiO2]]> TFe CaO content 1.55 0.72 0.048 <0.05 0.068 50.58 9.77 0.12 element MgO <![CDATA[TiO2]]> <![CDATA[Na2O]]> <![CDATA[K2O]]> <![CDATA[Al2O3]]> F TS Cl content 0.073 0.16 0.15 8.37 19.28 5.52 <0.03 0.029

[0163] 1) After the lepidolite concentrate is dried, a reaction reagent is added according to a mass ratio of lepidolite concentrate: CaSO4·2H2O: CaCO3: CaO: Na2SO4: K2SO4: Na2SO4 of 1:0.5:0.4:0.19:0.2:0.1, and each reagent is an external agent at the time of startup; after the process of Example 1 is operated, K2SO4 and Na2SO4 are mixed salts produced by the process, CaSO4·2H2O and CaCO3 are partly derived from the subsequent section precipitation return material, and partly are new reagents added, and CaO is a new reagent added; after the ore is mixed, the ore is mixed and dry-ground together to obtain a mixed material with a grinding fineness of -45 μm accounting for 85%;

[0164] 2) The mixed material obtained in step 1) is calcined in a rotary kiln at 1050° C. for 2 hours; the calcined sand obtained after being discharged from the kiln is cooled to 100-200° C. and transferred to a wet mill together with an impregnation solution to disperse the calcined sand to obtain a slurry with a liquid-to-solid ratio of 2:1; during the startup phase, the impregnation solution is water; after startup, the impregnation solution is the overflow of CCD1 in step 3) and the rubidium extraction raffinate in step 15), wherein the lithium content of the CCD1 overflow is 1.5 g / L, the lithium content of the rubidium extraction raffinate is 2.5 g / L, and the volume of the rubidium extraction raffinate is 1 / 6 of the volume of the CCD1 overflow;

[0165] 3) The slurry obtained in step 2) is pumped to a leaching tank and subjected to two-stage stirring leaching. During leaching, the lithium, rubidium, and cesium potassium displaced during the roasting process are dissolved in the leaching solution, and fluorine is fixed in the calcined sand in the form of CaF2. The total residence time is 1 hour, the leaching system is not heated, and the leaching temperature is 30°C.

[0166] 4) The leached slurry obtained in step 3) is pumped to a thickener for thickening to obtain a precious leached solution (thickening overflow) and a thickening underflow; the thickening underflow is pumped to a three-stage CCD+belt filter washing process, the thickener 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 settles. The overflow of CCD1 of the first-stage washing and thickening machine (CCD1) returns to step 2) and is mixed with the roasted sand to enter step 3) of the leaching process. The settled underflow of CCD1 enters CCD2, is mixed with the overflow of the third-stage washing and thickening machine (CCD3), and then settles. The underflow of CCD2 goes to CCD3, is mixed with the filtrate of the belt filter, and then settles. The underflow of CCD3 goes to the belt filter for filtration and washing to wash the filter cake (with a moisture content of approximately 40%).

[0167] 5) The washed filter cake obtained in step 4) can be directly used as a raw material for preparing gelling materials or cement.

[0168] 6) The dense overflow (i.e., the leached precious solution) obtained in step 4) enters a solution purification tank for further purification and impurity removal; wherein the leached precious solution contains 3 g / L Li, 1.6 g / L Rb, 100 mg / L Cs, 15 g / L K, and 15 g / L Na; specifically, the solution enters a two-stage purification tank in sequence, wherein a lime slurry with a slurry concentration of 5% is added to the first purification tank to preliminarily remove F and Mg, and the mass of calcium oxide in the lime slurry added to the first purification tank is 1 / 8 of the theoretical amount of sulfate in the dense overflow, and the solution is stirred and reacted at 90°C for 1 hour. The obtained first purified slurry is then pumped into a secondary purification tank, the secondary purification tank is added with the rubidium extraction residual solution of step 14), and the carbonate in the rubidium extraction residual solution is used to remove the calcium ions in the first purified slurry. The amount of rubidium extraction residual solution is based on the removal of calcium ions in the first purified slurry entering the secondary purification tank. The solution is stirred and purified at 90° C. for 0.5 h. After purification, a second purified slurry is obtained. The second purified slurry enters a plate and frame filter press for filtration. After the obtained filter residue is dried, the solution is returned to the roasting process in step 2), and the purified liquid obtained enters the evaporation and concentration process, and the wash water is returned to the first purification process;

[0169] 7) The purified liquid obtained in step 6) is subjected to evaporation and concentration. When the purified liquid is evaporated to the point where crystals precipitate, it is refrigerated at 5° C. for 12 hours to precipitate potassium and sodium salts. The purified liquid is then centrifuged to obtain a filter residue, i.e., a potassium and sodium mixed salt. The potassium and sodium mixed salt is washed and returned to step 1) as a partial reaction reagent. The washing liquid is returned to the first-stage purification tank.

[0170] 8) The potassium and sodium-free solution obtained by centrifugation and filtration in step 7) enters a lithium carbonate precipitation step; wherein the resin is LSC-500;

[0171] 9) washing the adsorbed resin obtained in step 8) with clean water, and returning the washed solution to the first purification step; then desorbing the washed resin with dilute hydrochloric acid, and evaporating and crystallizing the desorbed solution to obtain calcium chloride, which is then stored for unified treatment;

[0172] 10) heating the decalcified and magnesium-removed solution obtained in step 8) to 70° C. and then adding a sodium carbonate solution, wherein the mass of the sodium carbonate in the sodium carbonate solution is 1.2 times the theoretical amount of lithium precipitation required for the decalcified and magnesium-removed solution, reacting for 1 hour, and centrifuging to obtain crude lithium carbonate. The crude lithium carbonate enters the lithium carbonate refining process, and the lithium precipitation mother liquor enters the rubidium and cesium extraction process;

[0173] 11) slurrying, hydrogenating, filtering, pyrolyzing, and drying the crude lithium carbonate obtained in step 10) to obtain battery-grade lithium carbonate, wherein the purity of the battery-grade lithium carbonate is greater than 99.9%;

[0174] 12) The lithium-precipitated mother liquor obtained in step 10) is mixed with the cesium-free organic phase after stripping, and then subjected to three-stage extraction, five-stage washing, and three-stage stripping steps. The resulting cesium-rich stripping solution is evaporated and concentrated to prepare a cesium sulfate product, wherein the extractant is t-BAMBP, the diluent is kerosene, and the t-BAMBP volume concentration is 18%; the wash water is 13 g / L dilute sulfuric acid solution, and the stripping solution is 16 g / L dilute sulfuric acid solution; wherein 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%;

[0175] 13) The cesium-extracted liquid obtained in step 12) is mixed with the rubidium-free organic phase after stripping, and then subjected to three-stage extraction, three-stage washing, and three-stage stripping steps, and the rubidium-rich stripping liquid is evaporated and concentrated to prepare a rubidium sulfate product; wherein the extractant is t-BAMBP, the diluent is kerosene, and the t-BAMBP volume concentration is 18%; the wash water is 30 g / L dilute sulfuric acid solution, and the stripping liquid is 20 g / L sulfuric acid solution; wherein the rubidium extraction rate is greater than 95%, the lithium extraction rate is less than 1%, and the potassium and sodium extraction rates are less than 0.5%, and the rubidium sulfate product quality is greater than 99.9%;

[0176] 14) The rubidium extraction residue obtained in step 13) is deoiled and then returned to the secondary purification tank of step 6) to remove calcium and utilize the excess carbonate in the rubidium extraction residue to reduce the amount of reagents used;

[0177] 15) The rubidium extraction raffinate obtained in step 13) is deoiled and returned to step 2) and, together with the CCD1 overflow in step 3), the CCD1 overflow and rubidium extraction raffinate volume ratio is 6:1; the calcined sand is wet-milled and dispersed at a liquid-to-solid ratio of 2:1, and then leached.

[0178] This embodiment provides a method for recovering lithium, rubidium, and cesium from lepidolite ore, achieving a lithium, rubidium, and cesium leaching rate of approximately 90%, an overall lithium recovery rate of 85%, and an overall rubidium and cesium recovery rate of 82%. During the roasting process, fluorine is fixed in the ore in the form of CaF2, while iron is not leached. The leached residue can be used as a raw material for cement production, further recycling it.

[0179] Example 2

[0180] This embodiment uses the lepidolite ore produced in a lepidolite mine in Jiangxi as raw material;

[0181] The lepidolite ore is crushed and then magnetically separated to obtain lepidolite concentrate; the main components of the lepidolite concentrate are shown in Table 2: It can be concluded from Table 2 that the contents of Li2O, Rb2O, Cs2O, and K2O in the lepidolite concentrate are 1.75%, 0.72%, 0.048%, and 8.37%, respectively. The contents of iron and fluorine are relatively high, at 9.77% and 5.52%, respectively. The main gangues are SiO2 and Al2O3, with contents of 50.58% and 19.28%, respectively.

[0182] Table 2 Main components and contents of lepidolite concentrate in Example 2

[0183]

[0184]

[0185] 1) After drying the lepidolite concentrate, the corresponding reagents are added according to the ratio of lepidolite concentrate: CaSO4·2H2O:CaCO3=1:0.6:0.4, and each reagent is an external reagent at the start-up; after mixing, the ore is mixed and dry-ground together to obtain a mixed material with a grinding fineness of -75μm accounting for 85%;

[0186] 2) The mixed material obtained in step 1) is calcined in a rotary kiln at 1000° C. for 1 hour; the calcined sand obtained after being discharged from the kiln is cooled to 100-200° C. and transferred to a wet mill together with an impregnation solution to disperse the calcined sand to obtain a slurry with a liquid-to-solid ratio of 2:1; during the startup phase, the impregnation solution is water; after startup, the impregnation solution is the overflow of CCD1 in step 3) and the rubidium extraction raffinate in step 15), wherein the lithium content of the CCD1 overflow is 0.6 g / L, the lithium content of the rubidium extraction raffinate is 1.5 g / L, and the volume of the rubidium extraction raffinate is 1 / 6 of the volume of the CCD1 overflow;

[0187] 3) The slurry obtained in step 2) is pumped to a leaching tank and subjected to two-stage stirring leaching. During leaching, the lithium, rubidium, and cesium potassium displaced during the roasting process are dissolved in the leaching solution, and fluorine is fixed in the calcined sand in the form of CaF2. The total residence time is 1 hour, the leaching system is not heated, and the leaching temperature is 90°C.

[0188] 4) The leached slurry obtained in step 3) is pumped to a thickener for thickening to obtain a precious leaching solution (thickening overflow) and a thickening underflow; the thickening underflow is pumped to a three-stage CCD+belt filter washing process, the thickener 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 settles. The overflow of CCD1 of the first-stage washing and thickening machine (CCD1) returns to step 2) and is mixed with the roasted sand to enter step 3) of the leaching process. The settled underflow of CCD1 enters CCD2, is mixed with the overflow of the third-stage washing and thickening machine (CCD3), and then settles. The underflow of CCD2 goes to CCD3, is mixed with the filtrate of the belt filter, and then settles. The underflow of CCD3 goes to the belt filter for filtration and washing to wash the filter cake (with a moisture content of approximately 40%).

[0189] 5) The filter cake obtained in step 4) is used as a raw material for preparing a cementitious material or cement.

[0190] 6) The leached precious solution (i.e., dense overflow) obtained in step 4) enters a solution purification tank for further purification and impurity removal; wherein the leached precious solution contains 2.5 g / L of Li, 1.4 g / L of Rb, 100 mg / L of Cs, 10 g / L of K, and 200 mg / L of Na; specifically, the solution enters a two-stage purification tank in sequence, wherein a lime slurry with a pulp concentration of 20% is added to the first purification tank to preliminarily remove F and Mg, and the mass of the calcium oxide in the lime slurry added to the first purification tank is 1 / 8 of the theoretical amount of sulfate in the dense overflow, and the solution is stirred and reacted at 50°C. The first purified slurry is pumped into a secondary purification tank after 0.5h, and the rubidium extraction residual solution of step 14) is added to the secondary purification tank, and the carbonate in the rubidium extraction residual solution is used to remove the calcium ions in the first purified slurry. The amount of rubidium extraction residual solution is based on the removal of calcium ions in the first purified slurry entering the secondary purification tank. The slurry is stirred and purified at 50°C for 0.5h to obtain a second purified slurry after purification. The second purified slurry enters a plate and frame filter press for filtration. After the obtained filter residue is dried, the calcination process is returned to step 2), and the purified liquid obtained enters an evaporation and concentration process, and the wash water returns to the purification process;

[0191] 7) The purified liquid obtained in step 6) is subjected to evaporation and concentration. When the purified liquid is evaporated to the point where crystals precipitate, it is refrigerated at 5° C. for 18 hours to precipitate potassium and sodium salts. The purified liquid is then centrifuged to obtain a filter residue, i.e., a potassium and sodium mixed salt. The potassium and sodium mixed salt is washed and returned to step 1) as a partial reaction reagent. The washing liquid is returned to the first-stage purification tank.

[0192] 8) The potassium and sodium-removed solution obtained by centrifugation and filtration in step 7) is evaporated and concentrated again, and the concentrated solution is subjected to deep calcium and magnesium removal using a resin, and the resulting calcium and magnesium-removed solution enters a lithium carbonate precipitation step; wherein the resin is LSC-500;

[0193] 9) The adsorbed resin obtained in step 8) is first washed with clean water, and the washed liquid is returned to the purification process; then the washed resin is desorbed with dilute hydrochloric acid, and the desorption liquid is evaporated and crystallized to obtain calcium chloride, which is stored and uniformly treated;

[0194] 10) heating the decalcified and magnesium-removed solution obtained in step 8) to 95° C. and then adding a sodium carbonate solution, wherein the mass of the sodium carbonate in the sodium carbonate solution is 1.2 times the theoretical amount of lithium precipitation required for the decalcified and magnesium-removed solution, reacting for 1 hour, and centrifuging to obtain crude lithium carbonate. The crude lithium carbonate enters the lithium carbonate refining process, and the lithium precipitation mother liquor enters the rubidium and cesium extraction process;

[0195] 11) slurrying, hydrogenating, filtering, pyrolyzing, and drying the crude lithium carbonate obtained in step 10) to obtain battery-grade lithium carbonate, wherein the purity of the battery-grade lithium carbonate is greater than 99.9%;

[0196] 12) The lithium-precipitated mother liquor obtained in step 10) is mixed with the cesium-free organic phase after stripping, and then subjected to five-stage extraction, five-stage washing, and five-stage stripping steps. The resulting cesium-rich stripping solution is evaporated and concentrated to prepare a cesium sulfate product, wherein the extractant is t-BAMBP, the diluent is kerosene, and the t-BAMBP volume concentration is 10%; the wash water is 5 g / L dilute sulfuric acid solution, and the stripping solution is 10 g / L dilute sulfuric acid solution; wherein 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%;

[0197] 13) The cesium-extracted liquid obtained in step 12) is mixed with the rubidium-free organic phase after stripping, and then subjected to a five-stage extraction, five-stage washing, and five-stage stripping process. The rubidium-rich stripping liquid is evaporated and concentrated to prepare a rubidium sulfate product; wherein the extractant is t-BAMBP, the diluent is kerosene, and the t-BAMBP volume concentration is 30%; the wash water is 30 g / L dilute sulfuric acid solution, and the stripping liquid is 40 g / L sulfuric acid solution; wherein the rubidium 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 rubidium sulfate product quality is greater than 99.9%;

[0198] 14) The rubidium extraction residue obtained in step 13) is deoiled and then returned to the secondary purification tank of step 6) to remove calcium and utilize the excess carbonate in the rubidium extraction residue to reduce the amount of reagents used;

[0199] 15) The rubidium extraction raffinate obtained in step 13) is deoiled and returned to step 2) and, together with the CCD1 overflow in step 3), the CCD1 overflow and rubidium extraction raffinate volume ratio is 6:1; the calcined sand is wet-milled and dispersed at a liquid-to-solid ratio of 2:1, and then leached.

[0200] This embodiment provides a method for recovering lithium, rubidium, and cesium from lepidolite ore, achieving a lithium, rubidium, and cesium leaching rate of approximately 92%, with a comprehensive lithium recovery rate of 88%, a comprehensive rubidium recovery rate of 85%, and a comprehensive cesium recovery rate of 83%. During the roasting process, fluorine is fixed in the ore in the form of CaF2, while iron is not leached. The leached residue can be used as a raw material for cementitious materials or cement production.

[0201] Example 3

[0202] This embodiment uses the lepidolite ore produced in a lepidolite mine in Sichuan as raw material;

[0203] The lepidolite ore is crushed and then floated to obtain lepidolite concentrate; the main components of the lepidolite concentrate are shown in Table 3. It can be concluded from Table 3 that the contents of Li2O, Rb2O, Cs2O, and K2O in the lepidolite concentrate are 1.89%, 1.02%, 0.06%, and 3.22%, respectively, and the main gangue is SiO2 and Al2O3, with contents of 56.8% and 19.28%, respectively.

[0204] Table 3 Main components and contents of lepidolite concentrate in Example 3

[0205] element <![CDATA[Li2O]]> <![CDATA[Rb2O]]> <![CDATA[Cs2O]]> <![CDATA[WO3]]> Sn <![CDATA[SiO2]]> TFe CaO content 1.89 1.02 0.06 <0.05 0.032 58.8 1.77 0.22 element MgO <![CDATA[TiO2]]> <![CDATA[Na2O]]> <![CDATA[K2O]]> <![CDATA[Al2O3]]> F TS Cl content 0.089 0.06 0.23 3.22 16.30 3.21 <0.05 0.034

[0206] 1) After the lepidolite concentrate is dried, a reaction reagent is added according to a mass ratio of lepidolite concentrate: CaSO4·2H2O:CaCO3:CaO:K2SO4:Na2SO4 of 1:0.6:0.3:0.1:0.05:0.1, and each reagent is an external agent at the time of startup; after the process of Example 1 is operated, K2SO4 and Na2SO4 are mixed salts produced by the process, CaSO4·2H2O and CaCO3 are partly derived from the subsequent section precipitation return material, and partly are new reagents added, and CaO is a new reagent added; after the ore is mixed, the ore is mixed and dry-ground together to obtain a mixed material with a grinding fineness of -45 μm accounting for 85%;

[0207] 2) The mixed material obtained in step 1) is calcined in a rotary kiln at 950° C. for 1.5 hours; the calcined sand obtained after being discharged from the kiln is cooled to 100-200° C. and transferred to a wet mill together with an impregnation solution to disperse the calcined sand to obtain a slurry with a liquid-to-solid ratio of 2:1; during the startup phase, the impregnation solution is water; after startup, the impregnation solution is the overflow of CCD1 in step 3) and the rubidium extraction residual solution in step 15), wherein the lithium content of the CCD1 overflow is 3 g / L, the lithium content of the rubidium extraction residual solution is 2.5 g / L, and the volume of the rubidium extraction residual solution is 1 / 8 of the volume of the CCD1 overflow;

[0208] 3) The slurry obtained in step 2) is pumped to a leaching tank and subjected to two-stage stirring leaching. During leaching, the lithium, rubidium, and cesium potassium displaced during the roasting process are dissolved in the leaching solution, and fluorine is fixed in the calcined sand in the form of CaF2. The total residence time is 1 hour, the leaching system is not heated, and the leaching temperature is 50°C.

[0209] 4) The leached slurry obtained in step 3) is pumped to a thickener for thickening to obtain a precious leached solution (thickening overflow) and a thickening underflow; the thickening underflow is pumped to a three-stage CCD+belt filter washing process, the thickener 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 settles. The overflow of CCD1 of the first-stage washing and thickening machine (CCD1) returns to step 2) and is mixed with the roasted sand to enter step 3) of the leaching process. The settled underflow of CCD1 enters CCD2, is mixed with the overflow of the third-stage washing and thickening machine (CCD3), and then settles. The underflow of CCD2 goes to CCD3, is mixed with the filtrate of the belt filter, and then settles. The underflow of CCD3 goes to the belt filter for filtration and washing to wash the filter cake (with a moisture content of approximately 40%).

[0210] 5) The washed filter cake obtained in step 4) is used as a raw material for preparing a gelling material or cement.

[0211] 6) The dense overflow (i.e., the leached precious solution) obtained in step 4) enters a solution purification tank for further purification and impurity removal; wherein the leached precious solution contains 6 g / L of Li, 3 g / L of Rb, 200 mg / L of Cs, 10 g / L of K, and 15 g / L of Na; specifically, the solution enters a two-stage purification tank in sequence, wherein a lime slurry with a slurry concentration of 25% is added to the first purification tank to preliminarily remove F and Mg, and the mass of the calcium oxide in the lime slurry added to the first purification tank is 1 / 8 of the theoretical amount of sulfate in the dense overflow, and the reaction is stirred at 90°C for 1 minute. h, the first purified slurry obtained is pumped into a secondary purification tank, the secondary purification tank is added with the rubidium extraction residual solution of step 14), and the calcium ions in the first purified slurry are removed by carbonic acid in the rubidium extraction residual solution. The amount of rubidium extraction residual solution is based on the removal of calcium ions in the first purified slurry entering the secondary purification tank. The slurry is stirred and purified at 90° C. for 0.5 h. After purification, a second purified slurry is obtained, and the second purified slurry is filtered in a plate and frame filter press. After the obtained filter residue is dried, the calcination process is returned to step 2), and the purified liquid obtained enters the evaporation concentration process, and the wash water is returned to the purification process;

[0212] 7) The purified liquid obtained in step 6) is subjected to evaporation and concentration. When the purified liquid is evaporated to the point where crystals precipitate, it is refrigerated at 5° C. for 12 hours to precipitate potassium and sodium salts. The purified liquid is then centrifuged to obtain a filter residue, i.e., a potassium and sodium mixed salt. The potassium and sodium mixed salt is washed and returned to step 1) as a partial reaction reagent. The washing liquid is returned to the first-stage purification tank.

[0213] 8) The potassium and sodium-removed solution obtained by centrifugation and filtration in step 7) is evaporated and concentrated again, and the concentrated solution is subjected to deep calcium and magnesium removal using a resin, and the resulting calcium and magnesium-removed solution enters a lithium carbonate precipitation step; wherein the resin is LSC-500;

[0214] 9) The adsorbed resin obtained in step 8) is first washed with clean water, and the washed liquid is returned to the purification process; then the washed resin is desorbed with dilute hydrochloric acid, and the desorption liquid is evaporated and crystallized to obtain calcium chloride, which is stored and uniformly treated;

[0215] 10) heating the decalcified and magnesium-removed solution obtained in step 8) to 95° C. and then adding a sodium carbonate solution, wherein the mass of the sodium carbonate in the sodium carbonate solution is 1.2 times the theoretical amount of lithium precipitation required for the decalcified and magnesium-removed solution, reacting for 1 hour, and centrifuging to obtain crude lithium carbonate. The crude lithium carbonate enters the lithium carbonate refining process, and the lithium precipitation mother liquor enters the rubidium and cesium extraction process;

[0216] 11) slurrying, hydrogenating, filtering, pyrolyzing, and drying the crude lithium carbonate obtained in step 10) to obtain battery-grade lithium carbonate, wherein the purity of the battery-grade lithium carbonate is greater than 99.9%;

[0217] 12) The lithium-precipitated mother liquor obtained in step 10) is mixed with the cesium-free organic phase after stripping, and then subjected to a primary extraction, a primary washing, and a primary stripping process. The resulting cesium-rich stripping solution is evaporated and concentrated to prepare a cesium sulfate product, wherein the extractant is t-BAMBP, the diluent is kerosene, and the t-BAMBP volume concentration is 28%; the wash water is a 13 g / L dilute sulfuric acid solution, and the stripping solution is a 16 g / L dilute sulfuric acid solution; wherein the cesium extraction rate is >95%, the lithium extraction rate is <1%, the potassium and sodium extraction rates are <0.5%, and the cesium sulfate product quality is >99.9%;

[0218] 13) The cesium-extracted liquid obtained in step 12) is mixed with the rubidium-free organic phase after stripping, and then subjected to a primary extraction, a primary washing, and a primary stripping process. The rubidium-rich stripping liquid is evaporated and concentrated to prepare a rubidium sulfate product; wherein the extractant is t-BAMBP, the diluent is kerosene, and the t-BAMBP volume concentration is 28%; the wash water is 30 g / L dilute sulfuric acid solution, and the stripping liquid is 20 g / L sulfuric acid solution; wherein the rubidium 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 rubidium sulfate product quality is greater than 99.9%;

[0219] 14) The rubidium extraction residue obtained in step 13) is deoiled and then returned to the secondary purification tank of step 6) to remove calcium and utilize the excess carbonate in the rubidium extraction residue to reduce the amount of reagents used;

[0220] 15) The rubidium extraction raffinate obtained in step 13) is deoiled and returned to step 2) and, together with the CCD1 overflow in step 3), the CCD1 overflow and rubidium extraction raffinate volume ratio is 6:1; the calcined sand is wet-milled and dispersed at a liquid-to-solid ratio of 2:1, and then leached.

[0221] This embodiment provides a method for recovering lithium, rubidium, and cesium from lepidolite ore, achieving process indicators of approximately 93% lithium, rubidium, and cesium leaching rates, 91% lithium recovery rates, 87% rubidium recovery rates, and 83% cesium recovery rates. During the roasting process, fluorine is fixed in the ore in the form of CaF2, while iron is not leached. The leached residue can be used as a raw material for cementitious materials or cement production.

[0222] Beneficial effects of the present invention:

[0223] The present invention adopts for the first time the process of "material preparation - roasting - leaching - dense washing - leach residue resource utilization - leachate purification and impurity removal - leachate concentration and evaporation after impurity removal - low-temperature precipitation of mixed salt - mixed salt return to roasting - lithium carbonate precipitation in the liquid after salt precipitation - lithium carbonate refining - cesium extraction in the liquid after lithium precipitation - rubidium extraction in the liquid after cesium extraction - rubidium return to leaching and impurity removal in the liquid after rubidium extraction" to treat lepidolite concentrate. While obtaining a high recovery rate of lithium, rubidium, cesium and potassium, the cost of reagents is further saved, and zero wastewater discharge is achieved, the amount of waste gas and waste residue is small, and the environmental protection benefits are good.

[0224] (1) Lithium, rubidium, cesium, and potassium are efficiently leached from lepidolite, with lithium, rubidium, and cesium leaching rates of 90% to 95%, potassium leaching rates of 70% to 90%, lithium comprehensive recovery rates of 85% to 90%, and rubidium and cesium comprehensive recovery rates greater than 80%, meeting the process indicators. At the same time, aluminum, iron, and silicon in the ore are not leached, and only trace amounts of F (20 to 30 mg / L) and Mg (10 to 20 mg / L) are leached into the solution, reducing the pressure of impurity removal.

[0225] (2) The F in the ore is fixed in the form of CaF2, avoiding corrosion of equipment and loss of lithium during roasting.

[0226] (3) Compared with the conventional potassium sulfate process system (which requires the addition of ~50% potassium salt), the new process system can leach 70% to 90% of the potassium in the ore. Taking the K2O content of 8.37% as an example, 1 t of ore can produce about 100 kg of potassium sulfate. Under the condition of potassium sulfate salt price of 6,000 yuan / ton, the self-balance of potassium sulfate in the smelting process is achieved, which can reduce the reagent cost by 600 yuan / ton of lithium concentrate.

[0227] (4) The potassium and sodium salts, calcium sulfate slag and calcium carbonate slag are returned for roasting, which reduces the loss caused by the mixed salt and impurity removal slag entrained with lithium, rubidium and cesium, and improves the comprehensive recovery rate of lithium, rubidium and cesium in the system.

[0228] (5) Part of the liquid after rubidium extraction is returned to the second stage for purification and calcium removal, which utilizes the remaining carbonate ions from the sodium carbonate precipitation of lithium and reduces the amount of reagents used.

[0229] (6) Part of the rubidium extraction liquid is returned to the wet grinding and leaching process, which simplifies the process flow and improves the comprehensive recovery rate of lithium and the comprehensive utilization rate of potassium and sodium.

[0230] (7) The system can achieve acid-base self-balance without the need for additional reagents to adjust the acid-base to meet the working pH of each process.

[0231] (8) The overall operation of this process has the advantages of low-cost roasting and high comprehensive recovery rate of lithium, rubidium and cesium. The entire process does not require acid-base adjustment of the feed solution, which reduces the consumption of reagents required for acid-base neutralization in the potassium sulfate method. In a true sense, the process innovation and improvement are utilized to save reagent costs while increasing economic benefits, forming a high-efficiency process with high comprehensive recovery rate.

[0232] Although the above embodiment provides a detailed description of the present invention, it is only a part of the embodiments of the present invention, not all of the embodiments. Other embodiments can be obtained based on this embodiment without creativity, and these embodiments all fall within the scope of protection of the present invention.

Claims

1. A method for recovering lithium, rubidium and cesium from lepidolite concentrate, characterized in that: The following steps are involved: Mixing lepidolite concentrate and a reaction agent and dry grinding them to obtain a mixed material; the reaction agent includes gypsum and CaCO3; calcining the mixed material to obtain calcine; mixing the calcine and the impregnation liquid and performing wet grinding to obtain a slurry; The impregnation liquid is water or an aqueous solution containing lithium ions; leaching the ore pulp to obtain leached ore pulp; performing a first solid-liquid separation on the leached slurry to obtain a leached precious liquid; Mixing the leached noble solution and lime slurry, performing pH adjustment and a first purification, to obtain a first purified slurry; mixing the first purified slurry and the carbonate-containing solution for a second purification to obtain a second purified slurry; separating the second purified slurry into a second solid-liquid phase to obtain a purified liquid and impurity-removed residue; evaporating and concentrating the purified liquid to obtain a concentrated liquid; Cooling the concentrated liquid and performing a third solid-liquid separation to obtain a mixed salt containing impurities potassium and sodium and a concentrated solution containing lithium; The mixed salt containing impurities of potassium and sodium is first washed with tap water or a solution of mixed salt containing potassium and sodium to obtain a purified mixed salt of potassium and sodium and a washed solution; removing calcium and magnesium from the lithium-containing concentrated solution by using an adsorption resin to obtain a calcium- and magnesium-removed solution; The calcium and magnesium-removed solution is mixed with a sodium carbonate solution to precipitate lithium, thereby obtaining a crude lithium carbonate product and a lithium precipitation mother liquor, respectively; The crude lithium carbonate is subjected to slurrying, hydrogenation, filter pressing, pyrolysis and drying to obtain battery-grade lithium carbonate; The lithium-precipitated mother liquor is mixed with the cesium-unloaded organic phase and subjected to a first countercurrent continuous extraction to obtain a cesium-loaded organic phase and a cesium-extracted liquid, respectively; the cesium-loaded organic phase is subjected to a first continuous washing to obtain a purified cesium-loaded organic phase; the purified cesium-loaded organic phase is subjected to a first continuous stripping to obtain a cesium-rich stripping solution and a cesium-unloaded organic phase; the cesium-rich stripping solution is evaporated and crystallized to obtain cesium sulfate; The cesium-extracted liquid is mixed with a rubidium-free organic phase and subjected to a second countercurrent continuous extraction to obtain a rubidium-loaded organic phase and a rubidium-extracted liquid; the rubidium-loaded organic phase is subjected to a second continuous washing to obtain a purified rubidium-loaded organic phase; the purified rubidium-loaded organic phase is subjected to a second continuous stripping to obtain a rubidium-rich stripping solution and a rubidium-free organic phase; the rubidium-rich stripping solution is evaporated and crystallized to obtain rubidium sulfate; during the first countercurrent continuous extraction and the second countercurrent continuous extraction, the extractant is t-BAMBP and the diluent is kerosene; and in the mixed system consisting of the extractant and the diluent, the volume percentage of the extractant is 10% to 30%.

2. The method according to claim 1, characterized in that The reaction reagents also include any one or more of lime, Na2SO4 and K2SO4; The mass ratio of the lepidolite concentrate to the reaction reagent is 1:(0.2-2.2); In the lepidolite concentrate, the mass percentage of lepidolite concentrate with a fineness of less than 0.153 mm is ≥50%.

3. The method according to claim 2, characterized in that The reaction reagents include gypsum, CaCO3, lime, Na2SO4 and K2SO4; The mass ratio of the gypsum measured as CaSO4·2H2O to the lepidolite concentrate is (0.1-0.6):1; The mass ratio of the CaCO3 measured as pure CaCO3 to the lepidolite concentrate is (0.1-0.7):1; The mass ratio of the lime, measured as quicklime CaO, to the lepidolite concentrate is (0-0.3):1, and is not 0; The mass ratio of the Na2SO4 to the lepidolite concentrate is (0-0.3):1, and is not 0; The mass ratio of the K2SO4 and the lepidolite concentrate is (0-0.3):1, and is not 0.

4. The method according to claim 1, wherein The calcination temperature is 850-1150° C., and the calcination holding time is 0.5-5 hours.

5. The method according to claim 1, wherein The leaching temperature is 5-95° C.; the leaching time is 0.1-5 hours; and the leaching is performed by pipeline leaching or single-stage or multi-stage stirring tank leaching.

6. The method according to claim 1, characterized in that The first solid-liquid separation obtains a filter residue phase; After the first solid-liquid separation, the further step includes: performing a second washing on the filter residue phase to obtain a washed filter cake and a washing liquid; the washed filter cake is used to make a raw material for a gelling material or a raw material for cement; The second washing includes one or more combinations of countercurrent thickening washing, belt filtration washing and filter press washing, and the number of stages of the countercurrent thickening washing is 2 to 6; the washing water used in the second washing is water and / or condensed water obtained by the evaporation and concentration; the washing liquid obtained by the washing is reused as the impregnation liquid.

7. The method according to claim 1, characterized in that The first washing includes one or more stages of plate and frame filtration washing, belt filtration washing or centrifugal filtration washing; the purified potassium and sodium mixed salt obtained by the first washing is reused as part of the reaction reagent, and the washed liquid obtained by the first washing is returned to the first purification.

8. The method according to claim 2, characterized in that When the leached precious solution and lime slurry are mixed for pH adjustment and first purification, the mass concentration of the lime slurry is 5% to 30%; The impurity-removing residue is recycled as part of the reaction reagent.

9. The method according to claim 1, characterized in that During the first continuous washing and the second continuous washing, the washing agent is a first sulfuric acid solution, and the concentration of the first sulfuric acid solution is independently 2 to 20 g / L; the first continuous washing and the second continuous washing also produce a potassium-sodium washing solution with a low cesium concentration and a potassium-sodium washing solution with a low rubidium concentration, respectively, and the potassium-sodium washing solution with a low cesium concentration and the potassium-sodium washing solution with a low rubidium concentration are reused as the impregnation solution; During the first continuous stripping and the second continuous stripping, the stripping liquid is a second sulfuric acid solution, and the mass concentration of the second sulfuric acid solution is independently 5 to 30 g / L; The mass concentration of the second sulfuric acid solution is greater than the mass concentration of the first sulfuric acid solution.

10. The method according to claim 1, characterized in that After the second countercurrent continuous extraction, the method further comprises: recycling the rubidium-extracted solution as an impregnation solution and / or recycling the rubidium-extracted solution as a carbonate-containing solution for the second purification.

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