A method for separating lithium from zinnwaldite

Abstract

The present invention relates to the technical field of lithium extraction from lithium ores, and specifically discloses a method for separating lithium from zinnwaldite. The method for separating lithium from zinnwaldite provided by the present invention includes: mixing zinnwaldite powder with an acid solution and adding the mixture into a high-pressure reaction kettle for oxidative leaching reaction to obtain a lithium-containing leaching solution; adding a calcium salt to the lithium-containing leaching solution until the pH of the system is neutral or alkaline, filtering to obtain a filtrate; adding a carbonate solution to the filtrate to remove the excess calcium ions, filtering, concentrating the filtrate, adding a saturated carbonate solution to the concentrated solution for precipitation reaction, filtering, and drying to obtain battery-grade lithium carbonate. Through the method for separating lithium from zinnwaldite provided by the present invention, battery-grade lithium carbonate can be obtained, and the lithium recovery rate is high, the operation is simple, the energy consumption is low, the consumption of acids and alkalis is small, and the amount of waste liquid generated is small. It is suitable for industrial batch extraction of lithium from lithium ores, has high practical value, and is of great significance for the sustainable development of the lithium battery field.

Description

Technical Field

[0001] The present invention relates to the technical field of lithium extraction from lithium ores, and particularly to a method for separating lithium from zinnwaldite. Background Art

[0002] Lithium is considered an important strategic resource in this century and to a certain extent determines the development lifeline of the electric vehicle industry. It can be predicted that for a long time in the future, the global market demand for lithium will steadily increase. Therefore, the development and utilization of lithium resources have attracted great attention and emphasis from various countries. There are three main types of lithium resources in nature: salt lakes (sulfates, carbonates, chlorides, nitrates, etc.), pegmatites (spodumene, lepidolite, etc.), and sedimentary rocks (bauxite, coal, kaolin, etc.). Lithium resources in China mainly exist in the form of salt lake lithium and lepidolite. Among them, salt lake lithium resources are relatively rich, but there are problems such as high magnesium-lithium ratio in salt lakes, difficult separation, and high cost. Although lepidolite resources are also relatively rich, the ore grade is low, the lithium extraction cost is high, and the environmental pollution is large.

[0003] The chemical general formula of zinnwaldite is KLiFeAl(AlSi 3 O 10 )(F,OH) 2 , which is a gray metallic layered silicate mineral belonging to the mica group. Among them, the Li 2 O content ranges from 1.1 to 5.0 wt%. The methods for separating lithium from zinnwaldite mainly include dilute acid leaching method, sulfate roasting method, chlorination roasting method, hydrofluoric acid method, autoclaving method, etc. However, the roasting method and autoclaving method have high energy consumption, require more reagents to be added during the extraction process, and generate a large amount of slag; the acid leaching method cannot avoid impurity ions from entering the leaching solution, and it is necessary to further remove the impurity ions in the leaching solution. Inevitably, lithium loss will be involved during the impurity removal process, resulting in a low lithium recovery rate. At the same time, the acid corrodes the extraction equipment greatly. Therefore, there is an urgent need to develop a lithium extraction method for lithium ores with simple operation, low energy consumption, low impurity removal difficulty, and high recovery rate. Summary of the Invention

[0004] Aiming at the problems existing in the existing methods for extracting lithium from lepidolite ores, such as high energy consumption, complex operation, large amount of slag, difficult impurity removal, low recovery rate, and difficulty in obtaining battery-grade lithium salts, the present invention provides a method for separating lithium from zinnwaldite. The method uses high-pressure low-concentration acid combined with oxidation reaction to extract lithium, without performing high-temperature roasting and high-concentration acid reaction, so that most of the leached impurity ions precipitate in the form of ions, realizing the separation of lithium and slag, reducing the difficulty of impurity removal, increasing the lithium recovery rate, reducing the dosage of neutralizing reagents, reducing the amount of slag and the corrosion of equipment, facilitating industrial application, and having important significance for promoting the sustainable development of lithium batteries.

[0005] To solve the above technical problems, the technical solution provided by the present invention is as follows:

[0006] A method for separating lithium from zinnwaldite, comprising the following steps:

[0007] Step a: Crushing the zinnwaldite ore to obtain zinnwaldite powder;

[0008] Step b: Mixing the zinnwaldite powder evenly with a low-concentration acid solution to obtain a mixed material; the acid solution includes sulfuric acid;

[0009] Step c: Performing an oxidative leaching reaction on the mixed material and an oxidant under high-pressure conditions, followed by solid-liquid separation to obtain a lithium-containing leaching solution;

[0010] Step d: Adding a calcium salt to the lithium-containing leaching solution until the pH of the system is neutral or alkaline, followed by solid-liquid separation to obtain a filtrate;

[0011] Step e: Adding a carbonate solution to the filtrate to remove excess calcium ions, followed by solid-liquid separation. After concentrating the obtained filtrate, a concentrated solution is obtained;

[0012] Step f: Adding a saturated carbonate solution to the concentrated solution to carry out a precipitation reaction, followed by solid-liquid separation and drying to obtain battery-grade lithium carbonate.

[0013] In the existing acid leaching method for lithium, since it is necessary to add alkaline substances to the lithium-containing acid leaching solution to remove iron ions and aluminum ions, and a large amount of Li will be entrained during the precipitation process, especially when the precipitation amount is large, more Li will be entrained. For example, generally, calcium hydroxide or calcium carbonate is added to the lithium-containing acid solution to precipitate aluminum hydroxide. During the process, lithium in the lithium-containing acid leaching solution and aluminum hydroxide will form a stable lithium sulfate-aluminum hydroxide double salt compound. Moreover, Fe will form iron oxyhydroxide in an alkaline environment, entraining a part of lithium, resulting in a large loss of lithium.

[0014] Compared with the prior art, in the method for separating lithium from zinnwaldite provided by the present invention, first, during the leaching of lithium under acidic conditions, Fe 2+ is oxidized, so that the oxidized Fe 3+ and Al 3+ precipitate out in the form of alum and iron vanadate, and Si precipitates out in the form of silicon dioxide, effectively reducing the leaching of Al and Fe impurity ions. This not only reduces the large amount of precipitation generated in the subsequent alkali addition for impurity removal process, but also reduces the entrainment of Li in the precipitation, thus effectively improving the recovery rate of Li; at the same time, since most of the Al and Fe impurities have precipitated and fixed out in the process of acid leaching of lithium, the content of impurity ions in the lithium-containing extraction solution is less. Subsequently, only by simply adding a calcium salt to remove F - , un-precipitated Al 3+ , Fe3+ , and adding carbonate to remove the excess Ca 2+ A high-purity lithium-containing extraction solution can be obtained, and finally battery-grade lithium carbonate can be obtained through a simple lithium precipitation reaction, greatly simplifying the impurity removal process; in addition, when extracting lithium under high-pressure oxidation conditions, a low-concentration acid solution can be used to fully extract lithium from iron-lithium mica, not only reducing the consumption of acid, but also avoiding the problem of introducing impurity salts in the traditional high-temperature roasting method.

[0015] Through the method for separating lithium from iron-lithium mica provided by the present invention, battery-grade lithium carbonate can be obtained, with high lithium recovery rate, simple operation, low energy consumption, small consumption of acids and alkalis, and small amount of waste liquid generated. It is suitable for industrial batch extraction of lithium from lithium ores, has high practical value, and is of great significance for the sustainable development of the lithium battery field.

[0016] Furthermore, in step a, the content of the powder with a particle size not less than 100 mesh in the iron-lithium mica powder is ≥ 90%.

[0017] Furthermore, step a in the present invention specifically includes: drying, crushing, fine grinding and sieving the iron-lithium mica ore to obtain iron-lithium ore powder with the number of powders with a particle size not less than 100 mesh ≥ 90%.

[0018] The preferred particle size of the iron-lithium ore powder is beneficial to promoting the full leaching and extraction of Li.

[0019] Furthermore, in step b, the concentration of the acid solution is 0.5 mol / L to 3 mol / L.

[0020] Furthermore, in step b, the acid solution is a sulfuric acid solution, or an acid solution mixed with at least one of hydrochloric acid and nitric acid.

[0021] Specifically, in step b, the acid solution is a sulfuric acid solution; or the acid solution is a mixed solution of sulfuric acid and hydrochloric acid; or the acid solution is a mixed solution of sulfuric acid and nitric acid.

[0022] Furthermore, in step b, the liquid-solid ratio of the acid solution to the iron-lithium mica powder is (3 - 6) mL:1 g.

[0023] Oxidative leaching is carried out under high-pressure oxidation conditions, which effectively promotes the leaching of Li. A low-concentration acid solution can be used to achieve the leaching and extraction of Li from iron-lithium mica. Compared with the traditional acid leaching method, the method provided by the present invention effectively reduces the dosage and concentration of the acid solution, not only reducing the subsequent addition amount of alkali, reducing the generation amount of waste residue and waste liquid, but also effectively reducing the corrosion of equipment during the extraction process, and is more suitable for industrial application.

[0024] Under the above acid solution concentration and addition ratio, it can be ensured that step c is completed under the condition of pH < 1, so that Fe in lepidolite is oxidized to form iron vanadate, and Al forms alum.

[0025] Further, in step c, the oxidant is at least one of oxygen, air, ozone, potassium persulfate, sodium persulfate or ammonium persulfate.

[0026] Preferably, in step c, the oxidant is oxygen.

[0027] Specifically, when the oxidant is oxygen, the main chemical reaction occurring in step c is:

[0028] 3K(AlFeLi)(Si 3 Al)O 10 (OH)F + 3 / 4O 2 + 15 / 2H 2 SO 4 + 3 / 2H 2 O = 2K(Al 3 (SO 4 ) 2 (OH) 6 )↓ + K(Fe 3 (SO 4 ) 2 (OH) 6 )↓ + 3 / 2Li 2 SO 4 + 9SiO 2 + 3HF↑

[0029] Specifically, when the oxidant is persulfate, the main chemical reaction occurring in step c is:

[0030] 6K(AlFeLi)(Si 3 Al)O 10 (OH)F + 3Me 2 S 2 O 8 + 12H 2 SO 4 + 6H 2 O = 4K(Al 3 (SO 4 ) 2 (OH) 6 )↓ + 2K(Fe 3 (SO 4 ) 2 (OH) 6 )↓ + 3Li 2 SO 4 + 18SiO 2 + 6HF↑ + 3Me 2 SO4 (Me = K + , Na + , NH 4 + ).

[0031] During the oxidative leaching process, Fe in the zinnwaldite is oxidized to form iron vanadate, Al forms alum, and Si is precipitated in the form of silica sand (silicon dioxide), effectively reducing the content of Fe and Al impurity ions in the lithium-containing leaching solution. Moreover, the iron vanadate and alum formed in the acidic solution do not entrap Li. At the same time, since most of the Fe and Al have been precipitated and separated out in this step, the amount of slag formed during the subsequent removal of Fe, Al, and F ions is significantly reduced, thereby significantly reducing the entrainment of Li in the precipitate, effectively improving the recovery rate of Li, simplifying the steps of lithium extraction, and providing a new, efficient, and environmentally friendly method for lithium extraction from zinnwaldite, which is of great significance for promoting the sustainable development of lithium batteries.

[0032] Furthermore, in step c, the molar ratio of the oxidant to Fe in the zinnwaldite powder is 0.5:1 to 1:1.

[0033] The present invention creatively utilizes the chemical components in zinnwaldite to carry out a chemical reaction with SO 4 2- in sulfuric acid under oxidative conditions to form iron vanadate and aluminum vanadate, so that Fe and Al impurity ions are cleverly precipitated out, effectively reducing the content of Fe and Al impurity ions in the lithium-containing leaching solution, achieving the removal of most Fe and Al impurities simultaneously during the lithium extraction process, effectively simplifying the subsequent impurity removal process, significantly reducing the amount of slag generated by subsequent impurity removal, and further reducing the entrainment loss of Li.

[0034] Furthermore, in step c, the temperature of the oxidative leaching reaction is 150°C to 300°C, the pressure is 1 MPa to 6 MPa, and the time is 1 h to 10 h.

[0035] Under the above high-pressure leaching conditions, the leaching rate of lithium ions can be increased.

[0036] Furthermore, in step d, the calcium salt is at least one of calcium oxide, calcium carbonate, calcium hydroxide, calcium chloride, calcium nitrate, or calcium oxalate, or a mixed salt thereof with at least one of sodium hydroxide, potassium hydroxide, sodium carbonate, or potassium carbonate.

[0037] Furthermore, in step d, the pH of the system is adjusted to 7 to 11.

[0038] Preferably, in step d, the pH of the system is adjusted to 8 to 9.

[0039] In the above preferred pH range, the removal rates of F, Al, and Fe ions can be increased, and the content of impurity ions in the extract can be reduced.

[0040] Further, in step e, the carbonate solution is at least one of sodium carbonate solution, potassium carbonate solution, sodium bicarbonate solution, or potassium bicarbonate solution.

[0041] Further, in step e, the concentration of the carbonate solution is 150 g / L to 200 g / L.

[0042] Further, in step e, the molar ratio of carbonate ions in the carbonate solution to calcium ions in the filtrate is 1:1 to 1.1:1.

[0043] Removing the excess Ca in step d 2+ can improve the purity of the lithium-containing extract.

[0044] Further, in step e, the concentration of lithium ions in the concentrated solution is 15 g / L to 25 g / L.

[0045] Further, in step f, the saturated carbonate solution is at least one of saturated sodium carbonate solution, saturated potassium carbonate solution, saturated sodium bicarbonate solution, or saturated potassium bicarbonate solution.

[0046] Further, in step f, the molar ratio of carbonate ions in the saturated carbonate solution to lithium ions in the concentrated solution is 1:2 to 1.1:2.

[0047] Further, in step f, the temperature of the precipitation reaction is 80°C to 95°C, and the reaction time is 1 h to 2 h.

[0048] Compared with the prior art, the present invention has the following advantages:

[0049] (1) The present invention provides a novel method for extracting lithium from zinnwaldite. This method can achieve sufficient leaching and extraction of lithium from zinnwaldite using a low-concentration sulfuric acid-containing solution, effectively reducing the dosage and concentration of acid, reducing the corrosion of equipment and the generation amount of waste liquid;

[0050] (2) The present invention leaches under oxidation conditions, and precipitates of Fe and Al can be formed during the leaching process, reducing the leaching of Fe and Al impurity ions, realizing the simultaneous extraction of lithium and impurity removal. At the same time, it also reduces the generation amount of precipitates during the alkali addition for impurity removal, reduces the entrainment loss of Li, and improves the recovery rate of Li;

[0051] (3) The present invention has the advantages of short process flow, small consumption of acids and alkalis, and small generation amounts of waste liquid and waste residue. The reagent raw materials used in the extraction process are cheap and easily available, taking into account both environmental protection and economic benefits. It is a green and efficient low-cost lithium extraction process that can be scaled up for engineering, suitable for industrial batch lithium extraction from lepidolite, and has important social, economic, and environmental values. BRIEF DESCRIPTION OF THE DRAWINGS

[0052] In order to more clearly illustrate the technical solutions in the embodiments of the present invention, the following will briefly introduce the drawings required for use in the embodiments. Obviously, the drawings described below are only some embodiments of the present invention. For those of ordinary skill in the art, without creative efforts, other drawings can also be obtained based on these drawings.

[0053] Figure 1 It is a schematic process flow diagram for lithium extraction from lepidolite ore in the embodiments of the present invention. DETAILED DESCRIPTION OF THE EMBODIMENTS

[0054] In order to make the objectives, technical solutions, and advantages of the present invention clearer, the following further details the present invention in conjunction with embodiments. It should be understood that the specific embodiments described herein are only used to explain the present invention and are not used to limit the present invention.

[0055] Embodiment 1

[0056] A method for separating lithium from lepidolite:

[0057] Step a: Dry, crush, and finely grind the lepidolite ore, pass it through a 100-mesh sieve to make the passing ratio reach more than 90%, and obtain lepidolite powder, wherein the lithium oxide content is 2.01%.

[0058] Step b: Mix the lepidolite powder and 3 mol / L sulfuric acid solution evenly according to a solid-liquid ratio of 1 g:3 mL, add them into a high-pressure reaction kettle for leaching reaction, introduce oxygen during the reaction, and the molar ratio of the introduced oxygen amount to Fe in the lepidolite powder is 0.8:1. The reaction temperature is 250 °C, the reaction pressure is 5 MPa, and the reaction time is 10 h. After the reaction, filter to obtain lithium-containing leaching solution A and filter residue B; the main phases in filter residue B are iron vanadate, aluminum vanadate, and silica sand. Sampling and analysis are carried out on the lithium-containing leaching solution A, and the leaching rate of Li is 98.27%.

[0059] Step c: Add calcium oxide to the lithium-containing leaching solution A until the solution pH reaches 7.6, filter to obtain lithium-containing leaching solution C and filter residue D; the main phase of filter residue D is calcium fluoride and calcium sulfate. Sampling and analysis are carried out on the lithium-containing leaching solution C, and the calcium ion content therein is 0.5 g / L.

[0060] Step d: Add a 150 g / L sodium carbonate solution with an equimolar amount of calcium ions to the lithium-containing leaching solution C, filter to obtain a lithium-containing leaching solution E and a filter residue F. The main phase of the filter residue F is calcium carbonate. Sampling and analysis of the lithium-containing leaching solution E shows that the removal rate of calcium ions is 99.9% and the loss rate of lithium is less than 1%.

[0061] Step e: Concentrate the lithium-containing leaching solution E to obtain a concentrated solution with a lithium ion concentration of 16.97 g / L. Add a saturated sodium carbonate solution to the concentrated solution, with the molar ratio of carbonate ions in the saturated sodium carbonate solution to lithium ions in the concentrated solution being 1.1:2. Stir and react at 90 °C for 2 h, filter while it is hot, wash with hot water at 100 °C, and dry in an oven at 120 °C to obtain battery-grade lithium carbonate. Finally, the recovery rate of lithium is measured to be 92.03%, and the purity of the lithium carbonate is detected to be 99.66%.

[0062] Example 2

[0063] A method for separating lithium from lepidolite:

[0064] Step a: Dry, crush, and finely grind the lepidolite ore, pass it through a 100-mesh sieve, and make the passing ratio reach over 90% to obtain lepidolite powder, with a lithium oxide content of 2.01%.

[0065] Step b: Mix the lepidolite powder and a 2 mol / L sulfuric acid solution evenly according to a solid-liquid ratio of 1 g:3 mL, add them to a high-pressure reaction kettle for leaching reaction. During the reaction, introduce oxygen, with the molar ratio of the introduced oxygen to Fe in the lepidolite powder being 0.8:1. The reaction temperature is 200 °C, the reaction pressure is 5 MPa, and the reaction time is 2 h. After the reaction, filter to obtain a lithium-containing leaching solution A and a filter residue B. The main phases in the filter residue B are iron vanadium, aluminum vanadium, and silica sand. Sampling and analysis of the lithium-containing leaching solution A shows that the leaching rate of Li is 95.25%.

[0066] Step c: Add calcium oxide to the lithium-containing leaching solution A until the solution pH reaches 8.1, filter to obtain a lithium-containing leaching solution C and a filter residue D. The main phase of the filter residue D is calcium fluoride and calcium sulfate. Sampling and analysis of the lithium-containing leaching solution C shows that the calcium ion content is 0.45 g / L.

[0067] Step d: Add a 150 g / L sodium carbonate solution with an equimolar amount of calcium ions to the lithium-containing leaching solution C, filter to obtain a lithium-containing leaching solution E and a filter residue F. The main phase of the filter residue F is calcium carbonate. Sampling and analysis of the lithium-containing leaching solution E shows that the removal rate of calcium ions is 99.9% and the loss rate of lithium is less than 1%.

[0068] Step e: Concentrate the lithium-containing leaching solution E to obtain a concentrated solution with a lithium ion concentration of 15.78 g / L. Add saturated sodium carbonate solution to the concentrated solution. The molar ratio of carbonate ions in the saturated sodium carbonate solution to lithium ions in the concentrated solution is 1.1:2. Stir and react at 95 °C for 1 h, filter while it is hot, wash with hot water at 100 °C, and dry in an oven at 120 °C to obtain battery-grade lithium carbonate. Finally, the recovery rate of lithium is measured to be 92.24%, and the purity of the lithium carbonate is detected to be 99.61%.

[0069] Example 3

[0070] A method for separating lithium from lepidolite:

[0071] Step a: Dry, crush, and finely grind the lepidolite ore, pass through a 100-mesh sieve, and make the passing ratio reach more than 90% to obtain lepidolite powder, in which the lithium oxide content is 2.03%.

[0072] Step b: Mix the lepidolite powder and 3 mol / L sulfuric acid solution evenly according to a solid-liquid ratio of 1 g:3 mL, add them to a high-pressure reaction kettle for leaching reaction, introduce oxygen during the reaction, and the molar ratio of the introduced oxygen to Fe in the lepidolite powder is 0.8:1. The reaction temperature is 150 °C, the reaction pressure is 5 MPa, and the reaction time is 3 h. After the reaction, filter to obtain a lithium-containing leaching solution A and a filter residue B; the main phases in the filter residue B are iron vanadate, aluminum vanadate, and silica sand. Sampling and analysis are carried out on the lithium-containing leaching solution A, and the leaching rate of Li is 94.71%.

[0073] Step c: Add calcium oxide to the lithium-containing leaching solution A until the solution pH reaches 9.3, filter to obtain a lithium-containing leaching solution C and a filter residue D; the phase of the filter residue D is mainly calcium fluoride and calcium sulfate. Sampling and analysis are carried out on the lithium-containing leaching solution C, and the calcium ion content therein is 0.49 g / L.

[0074] Step d: Add 200 g / L sodium carbonate solution with an equimolar amount of calcium ions to the lithium-containing leaching solution C, filter to obtain a lithium-containing leaching solution E and a filter residue F; the phase of the filter residue F is mainly calcium carbonate. Sampling and analysis are carried out on the lithium-containing leaching solution E, and the calcium ion removal rate is 99.8%, and the lithium loss rate is less than 1%.

[0075] Step e: Concentrate the lithium-containing leaching solution E to obtain a concentrated solution with a lithium ion concentration of 20.25 g / L. Add saturated sodium carbonate solution to the concentrated solution. The molar ratio of carbonate ions in the saturated sodium carbonate solution to lithium ions in the concentrated solution is 1:2. Stir and react at 80 °C for 2 h, filter while it is hot, wash with hot water at 100 °C, and dry in an oven at 120 °C to obtain battery-grade lithium carbonate. Finally, the recovery rate of lithium is measured to be 90.81%, and the purity of the lithium carbonate is detected to be 99.53%.

[0076] Example 4

[0077] A method for separating lithium from zinnwaldite:

[0078] Step a: Dry, crush and finely grind the zinnwaldite ore, pass it through a 100-mesh sieve, and make the passing ratio reach more than 90% to obtain zinnwaldite powder, in which the lithium oxide content is 2.03%;

[0079] Step b: Mix the zinnwaldite powder and 0.5 mol / L sulfuric acid solution evenly according to a solid-liquid ratio of 1 g:6 mL, add them into a high-pressure reactor for leaching reaction, and introduce oxygen during the reaction. The molar ratio of the introduced oxygen amount to the Fe in the zinnwaldite powder is 0.8:1. The reaction temperature is 300 °C, the reaction pressure is 5 MPa, and the reaction time is 8 h. After the reaction, filter to obtain lithium-containing leaching solution A and filter residue B; the main phases in the filter residue B are iron vanadate, aluminum vanadate and silica sand. Sampling and analysis are carried out on the lithium-containing leaching solution A, and the leaching rate of Li is 96.38%;

[0080] Step c: Add calcium oxide to the lithium-containing leaching solution A until the solution pH reaches 8.6, filter to obtain lithium-containing leaching solution C and filter residue D; the phases of the filter residue D are mainly calcium fluoride and calcium sulfate. Sampling and analysis are carried out on the lithium-containing leaching solution C, and the calcium ion content therein is 0.56 g / L;

[0081] Step d: Add a 170 g / L sodium carbonate solution with an equimolar amount of calcium ions to the lithium-containing leaching solution C, filter to obtain lithium-containing leaching solution E and filter residue F; the phase of the filter residue F is mainly calcium carbonate. Sampling and analysis are carried out on the lithium-containing leaching solution E, and the calcium ion removal rate is 99.9%, and the lithium loss rate is less than 1%;

[0082] Step e: Concentrate the lithium-containing leaching solution E to obtain a concentrated solution with a lithium ion concentration of 24.14 g / L. Add a saturated sodium carbonate solution to the concentrated solution. The molar ratio of carbonate ions in the saturated sodium carbonate solution to lithium ions in the concentrated solution is 1.1:2. Stir and react at 85 °C for 2 h, filter while it is hot, wash with hot water at 100 °C, and dry in an oven at 120 °C to obtain battery-grade lithium carbonate. Finally, the lithium recovery rate is measured to be 91.21%, and the purity of the lithium carbonate is detected to be 99.58%.

[0083] Example 5

[0084] A method for separating lithium from zinnwaldite:

[0085] Step a: Dry, crush and finely grind the zinnwaldite ore, pass it through a 100-mesh sieve, and make the passing ratio reach more than 90% to obtain zinnwaldite powder, in which the lithium oxide content is 1.98%;

[0086] Step b: Mix lepidolite powder and 1 mol / L sulfuric acid solution evenly at a solid-liquid ratio of 1 g:4 mL, add them into a high-pressure reactor for leaching reaction, introduce oxygen during the reaction, the molar ratio of the introduced oxygen to Fe in the lepidolite powder is 0.8:1, the reaction temperature is 200 °C, the reaction pressure is 1 MPa, and the reaction time is 3 h. After the reaction, filter to obtain lithium-containing leaching solution A and filter residue B; the main phases in the filter residue B are iron vanadate, aluminum vanadate and silica sand. Sampling and analysis are carried out on the lithium-containing leaching solution A, and the leaching rate of Li is 97.18%.

[0087] Step c: Add calcium oxide to the lithium-containing leaching solution A until the solution pH reaches 9.3, filter to obtain lithium-containing leaching solution C and filter residue D; the main phase of the filter residue D is calcium fluoride and calcium sulfate. Sampling and analysis are carried out on the lithium-containing leaching solution C, and the calcium ion content therein is 0.55 g / L.

[0088] Step d: Add 200 g / L sodium carbonate solution with an equimolar amount of calcium ions to the lithium-containing leaching solution C, filter to obtain lithium-containing leaching solution E and filter residue F; the main phase of the filter residue F is calcium carbonate. Sampling and analysis are carried out on the lithium-containing leaching solution E, the calcium ion removal rate is 99.8%, and the lithium loss rate is less than 1%.

[0089] Step e: Concentrate the lithium-containing leaching solution E to obtain a concentrated solution with a lithium ion concentration of 21.58 g / L. Add saturated sodium carbonate solution to the concentrated solution, and the molar ratio of carbonate ions in the saturated sodium carbonate solution to lithium ions in the concentrated solution is 1:2. Stir and react at 90 °C for 2 h, filter while it is hot, wash with hot water at 100 °C, and dry in an oven at 120 °C to obtain battery-grade lithium carbonate. Finally, the lithium recovery rate is measured to be 90.91%, and the purity of the lithium carbonate is detected to be 99.63%.

[0090] Example 6

[0091] A method for separating lithium from lepidolite:

[0092] Step a: Dry, crush and finely grind the lepidolite ore, pass through a 100-mesh sieve so that the passing ratio reaches more than 90% to obtain lepidolite powder, wherein the lithium oxide content is 2.01%.

[0093] Step b: Mix lepidolite powder and 2 mol / L sulfuric acid solution evenly at a solid-liquid ratio of 1 g:5 mL, add them into a high-pressure reactor for leaching reaction, introduce oxygen during the reaction, the molar ratio of the introduced oxygen to Fe in the lepidolite powder is 0.5:1, the reaction temperature is 250 °C, the reaction pressure is 5 MPa, and the reaction time is 8 h. After the reaction, filter to obtain lithium-containing leaching solution A and filter residue B; the main phases in the filter residue B are iron vanadate, aluminum vanadate and silica sand. Sampling and analysis are carried out on the lithium-containing leaching solution A, and the leaching rate of Li is 98.86%.

[0094] Step c: Add calcium oxide to the lithium-containing leaching solution A until the pH of the solution reaches 8.7, and then filter to obtain the lithium-containing leaching solution C and the filter residue D. The main phases of the filter residue D are calcium fluoride and calcium sulfate. Sampling and analysis are carried out on the lithium-containing leaching solution C, and the calcium ion content therein is 0.53 g / L.

[0095] Step d: Add a 170 g / L sodium carbonate solution with an equimolar amount to the calcium ions to the lithium-containing leaching solution C, and then filter to obtain the lithium-containing leaching solution E and the filter residue F. The main phase of the filter residue F is calcium carbonate. Sampling and analysis are carried out on the lithium-containing leaching solution E, and the calcium ion removal rate is 99.9%, and the lithium loss rate is less than 1%.

[0096] Step e: Concentrate the lithium-containing leaching solution E to obtain a concentrated solution with a lithium ion concentration of 18.25 g / L. Add a saturated sodium carbonate solution to the concentrated solution, and the molar ratio of carbonate ions in the saturated sodium carbonate solution to lithium ions in the concentrated solution is 1.1:2. Stir and react at 85 °C for 2 h, filter while it is hot, wash with hot water at 100 °C, and dry in an oven at 120 °C to obtain battery-grade lithium carbonate. Finally, the lithium recovery rate is measured to be 91.12%, and the purity of the lithium carbonate is detected to be 99.58%.

[0097] Example 7

[0098] A method for separating lithium from zinnwaldite:

[0099] Step a: Dry, crush, and finely grind the zinnwaldite ore, and pass through a 100-mesh sieve so that the passing ratio reaches more than 90% to obtain zinnwaldite powder, wherein the lithium oxide content is 2.11%.

[0100] Step b: Mix the zinnwaldite powder and a 2 mol / L sulfuric acid solution evenly according to a solid-liquid ratio of 1 g:5 mL, add them to a high-pressure reaction kettle for leaching reaction, and introduce oxygen during the reaction. The molar ratio of the introduced oxygen to Fe in the zinnwaldite powder is 1:1, the reaction temperature is 200 °C, the reaction pressure is 6 MPa, and the reaction time is 6 h. After the reaction, filter to obtain the lithium-containing leaching solution A and the filter residue B. The main phases in the filter residue B are iron vanadate, aluminum vanadate, and silica sand. Sampling and analysis are carried out on the lithium-containing leaching solution A, and the Li leaching rate is 98.34%.

[0101] Step c: Add calcium oxide to the lithium-containing leaching solution A until the pH of the solution reaches 9.8, and then filter to obtain the lithium-containing leaching solution C and the filter residue D. The main phases of the filter residue D are calcium fluoride and calcium sulfate. Sampling and analysis are carried out on the lithium-containing leaching solution C, and the calcium ion content therein is 0.61 g / L.

[0102] Step d: Add a 160 g / L sodium carbonate solution with an equimolar amount of calcium ions to the lithium-containing leaching solution C, filter to obtain the lithium-containing leaching solution E and the filter residue F. The main phase of the filter residue F is calcium carbonate. Sampling and analysis from the lithium-containing leaching solution E show that the removal rate of calcium ions is 99.9%, and the loss rate of lithium is less than 1%.

[0103] Step e: Concentrate the lithium-containing leaching solution E to obtain a concentrated solution with a lithium ion concentration of 17.69 g / L. Add a saturated sodium carbonate solution to the concentrated solution, with the molar ratio of carbonate ions in the saturated sodium carbonate solution to lithium ions in the concentrated solution being 1.1:2. Stir and react at 90 °C for 2 h, filter while it is hot, wash with hot water at 100 °C, and dry in an oven at 120 °C to obtain battery-grade lithium carbonate. Finally, the recovery rate of lithium is measured to be 92.31%, and the purity of the lithium carbonate is detected to be 99.72%.

[0104] Comparative Example 1

[0105] This comparative example provides a method for separating lithium from lepidolite. The only difference from Example 1 is that no oxidant is added during the leaching process, and the other steps are the same. The specific steps are as follows:

[0106] Step a: Dry, crush, and finely grind the lepidolite ore, pass through a 100-mesh sieve to make the passing ratio reach more than 90% to obtain lepidolite powder, with a lithium oxide content of 2.01%.

[0107] Step b: Mix the lepidolite powder and a 3 mol / L sulfuric acid solution evenly according to a solid-liquid ratio of 1 g:3 mL, add them to a high-pressure reactor for leaching reaction. The reaction temperature is 250 °C, the reaction pressure is 5 MPa, and the reaction time is 10 h. After the reaction, filter to obtain the lithium-containing leaching solution A and the filter residue B. The main phases in the filter residue B are aluminum vanadium and silica sand. Sampling and analysis from the lithium-containing leaching solution A show that the leaching rate of Li is 97.02%.

[0108] Step c: Add calcium oxide to the lithium-containing leaching solution A until the solution pH reaches 8.5, filter to obtain the lithium-containing leaching solution C and the filter residue D. The main phase of the filter residue D is calcium fluoride, calcium sulfate, aluminum hydroxide, and iron hydroxide. Sampling and analysis from the lithium-containing leaching solution C show that the loss rate of lithium is 10.2%.

[0109] Step d: Add a 150 g / L sodium carbonate solution with an equimolar amount of calcium ions to the lithium-containing leaching solution C, filter to obtain the lithium-containing leaching solution E and the filter residue F. The main phase of the filter residue F is calcium carbonate. Sampling and analysis from the lithium-containing leaching solution E show that the removal rate of calcium ions is 99.9%.

[0110] Step e: Concentrate the lithium-containing leaching solution E to obtain a concentrated solution with a lithium ion concentration of 17.25 g / L. Add saturated sodium carbonate solution to the concentrated solution, with the molar ratio of carbonate ions in the saturated sodium carbonate solution to lithium ions in the concentrated solution being 1.1:2. Stir and react at 90 °C for 2 h, filter while it is hot, wash with hot water at 100 °C, and dry in an oven at 120 °C to obtain battery-grade lithium carbonate. Finally, the lithium recovery rate is measured to be 75.32%, and the purity of the lithium carbonate is detected to be 99.03%.

[0111] Comparative Example 2

[0112] This comparative example provides a method for separating lithium from zinnwaldite. The only difference from Example 1 is that no oxidant is added during the leaching process, but it is added simultaneously with calcium oxide in step c, and the remaining steps are the same. The specific steps are as follows:

[0113] Step a: Dry, crush, and finely grind the zinnwaldite ore, pass through a 100-mesh sieve to make the passing ratio reach more than 90% to obtain zinnwaldite powder, in which the lithium oxide content is 2.01%.

[0114] Step b: Mix the zinnwaldite powder and 3 mol / L sulfuric acid solution evenly according to a solid-liquid ratio of 1 g:3 mL, add them to a high-pressure reactor for leaching reaction. The reaction temperature is 250 °C, the reaction pressure is 5 MPa, and the reaction time is 10 h. After the reaction, filter to obtain a lithium-containing leaching solution A and a filter residue B. The main phases in the filter residue B are aluminum vanadium and silica sand. Sampling and analysis are carried out on the lithium-containing leaching solution A, and the leaching rate of Li is 96.85%.

[0115] Step c: Add calcium oxide to the lithium-containing leaching solution A until the solution pH reaches 8.5, and then add hydrogen peroxide. The molar ratio of hydrogen peroxide to Fe in the zinnwaldite is 0.8:1. Filter to obtain a lithium-containing leaching solution C and a filter residue D. The phases of the filter residue D are mainly calcium fluoride, calcium sulfate, aluminum hydroxide, and iron hydroxide. Sampling and analysis are carried out on the lithium-containing leaching solution C, and the lithium loss rate is 15.2%.

[0116] Step d: Add 150 g / L sodium carbonate solution with an equimolar amount of calcium ions to the lithium-containing leaching solution C, filter to obtain a lithium-containing leaching solution E and a filter residue F. The phases of the filter residue F are mainly calcium carbonate. Sampling and analysis are carried out on the lithium-containing leaching solution E, and the calcium ion removal rate is 99.9%.

[0117] Step e: Concentrate the lithium-containing leaching solution E to obtain a concentrated solution with a lithium ion concentration of 17.36 g / L. Add saturated sodium carbonate solution to the concentrated solution, with the molar ratio of carbonate ions in the saturated sodium carbonate solution to lithium ions in the concentrated solution being 1.1:2. Stir and react at 90 °C for 2 h, filter while it is hot, wash with hot water at 100 °C, and dry in an oven at 120 °C to obtain battery-grade lithium carbonate. Finally, the lithium recovery rate is measured to be 80.34%, and the purity of the lithium carbonate is detected to be 98.23%.

[0118] In Examples 1 to 7, other oxidants defined in the present invention may also be used as the oxidant, such as air, ozone, persulfate, etc. The acid solution for acid leaching may also be a mixed acid solution of hydrochloric acid or nitric acid and sulfuric acid. As long as it is within the addition amount defined in the present invention, the same technical effects as those of the corresponding examples can be achieved. Similarly, carbonates or saturated carbonates, calcium salts, etc. may also be other substances defined in the present invention, and the same effects as those of the examples can be achieved as long as they are within the scope defined in the present invention.

[0119] The above are only the preferred embodiments of the present invention and are not intended to limit the present invention. Any modifications, equivalent replacements, or improvements made within the spirit and principles of the present invention shall be included in the protection scope of the present invention.

Claims

1. A method for separating lithium from iron lithium mica, characterized in that: The following steps are involved: Step a, crushing the iron lithium mica ore to obtain the iron lithium mica powder; Step b, mixing the iron lithium mica powder and a low concentration acid solution uniformly to obtain a mixed material; wherein the acid solution includes a sulfuric acid solution; and the concentration of the acid solution is 0.5 mol / L to 3 mol / L; Step c, subjecting the mixed material and the oxidant to an oxidation leaching reaction under high pressure conditions, and performing solid-liquid separation to obtain a lithium-containing leachate; the oxidant is at least one of oxygen, air, ozone, potassium persulfate, sodium persulfate or ammonium persulfate; Step d, adding calcium salt to the lithium-containing leachate until the pH of the system is neutral or alkaline, and separating the solid and liquid to obtain a filtrate; Step e, adding a carbonate solution to the filtrate to remove excess calcium ions, separating the solid and the liquid, and concentrating the obtained filtrate to obtain a concentrated solution; Step f, adding a saturated carbonate solution to the concentrated solution, performing a precipitation reaction, solid-liquid separation, and drying to obtain battery-grade lithium carbonate.

2. The method for separating lithium from iron lithium mica according to claim 1, characterized in that: In step a, the content of powder with a particle size not less than 100 mesh in the iron lithium mica powder is ≥90%.

3. The method for separating lithium from iron lithium mica according to claim 1, characterized in that: In step b, the acid solution is a sulfuric acid solution, or a mixture of sulfuric acid and at least one of hydrochloric acid and nitric acid; and / or In step b, the liquid-to-solid ratio of the acid solution to the iron lithium mica powder is (3-6) mL:1 g.

4. The method for separating lithium from iron lithium mica according to claim 1, characterized in that: In step c, the molar ratio of the oxidant to Fe in the iron lithium mica powder is 0.5:1 to 1:

1.

5. The method for separating lithium from iron lithium mica according to claim 1, characterized in that: In step c, the temperature of the oxidation leaching reaction is 150° C. to 300° C., the pressure is 1 MPa to 6 MPa, and the time is 1 h to 10 h.

6. The method for separating lithium from iron lithium mica according to claim 1, characterized in that: In step d, the calcium salt is at least one of calcium oxide, calcium carbonate, calcium hydroxide, calcium chloride, calcium nitrate or calcium oxalate, or a mixed salt thereof with at least one of sodium hydroxide, potassium hydroxide, sodium carbonate or potassium carbonate; and / or In step d, the pH of the system is adjusted to 7-11.

7. The method for separating lithium from iron lithium mica according to claim 1, characterized in that: In step e, the carbonate solution is at least one of a sodium carbonate solution, a potassium carbonate solution, a sodium bicarbonate solution or a potassium bicarbonate solution; and / or In step e, the molar ratio of carbonate ions in the carbonate solution to calcium ions in the filtrate is 1:1 to 1.1:1; and / or In step e, the concentration of lithium ions in the concentrated solution is 15 g / L to 25 g / L.

8. The method for separating lithium from iron lithium mica according to claim 1, characterized in that: In step f, the saturated carbonate solution is at least one of a saturated sodium carbonate solution, a saturated potassium carbonate solution, a saturated sodium bicarbonate solution or a saturated potassium bicarbonate solution; and / or In step f, the molar ratio of carbonate ions in the saturated carbonate solution to lithium ions in the concentrated solution is 1:2 to 1.1:2; and / or In step f, the precipitation reaction temperature is 80° C. to 95° C., and the reaction time is 1 h to 2 h.

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