Improved method for chemically extracting lithium from lepidolite fluorine

Abstract

The invention belongs to the technical field of extraction of lithium from lepidolite, and discloses an improved fluorine chemical extraction method of lithium from lepidolite. The method comprises the following steps: S1, crushing lepidolite and fluoride together; adding a solid material obtained after crushing into a sulfuric acid solution with the concentration of 50-90wt%, heating to 80-95 DEG C, and stirring to react for 1-3 hours to obtain viscous slurry; s2, the viscous slurry is introduced into a rotary kiln to be roasted for 1-6 h at the temperature of 800-900 DEG C; s3, crushing a solid material obtained after roasting, adding water for water leaching treatment, and filtering to obtain silicon slag and brine; s4, adding calcium oxide into the brine to adjust the pH value of the brine to 10-14, and filtering to realize primary impurity removal; adding water-soluble carbonate into the brine obtained after the first-time impurity removal, and filtering to realize second-time impurity removal, so as to obtain purified liquid; and S5, adding a lithium precipitation agent into the purified liquid to carry out lithium precipitation reaction, and filtering to obtain a lithium salt product. Roasting of the rotary kiln not only solves the problem of aluminum impurities in the water leaching process and eradicates the problem of treatment of the byproduct alum, but also reduces the problem of subsequent brine impurity removal, and further improves the lithium recovery rate.

Description

Technical Field

[0001] This invention belongs to the field of lithium extraction technology from lepidolite, specifically relating to an improved method for lithium extraction from lepidolite using fluorine chemistry. Background Technology

[0002] Lepidolite is a TOT-structured silicate mineral. In its mineral structure, the Li-O bond energy is increased due to the constraint of the Si-O bond, thus reducing the Li-O reactivity. Therefore, under conventional leaching techniques, the lithium leaching efficiency from lepidolite is extremely low. Considering that lepidolite contains 2%–4% fluorine, and that HF ​​can break the Si-O bonds in lepidolite minerals, the fluorine chemical method is considered an improved acid method to enhance the lithium extraction efficiency from lepidolite. The fluorine chemical method utilizes the strong complexation of fluorine with metal ions to break down the dense crystal structure in the ore, and then reacts with sulfuric acid to generate the corresponding soluble lithium and aluminum sulfates. Compared with traditional methods, this method requires lower reaction temperatures and consumes less energy, making it highly valuable for application.

[0003] Currently, there is a large body of related research. Patent CN101885496A, "A Fluorochemical Lithium Extraction Process from Lithium Mica," proposes a fluorochemical lithium extraction process that involves fluorine chemical leaching of lepidolite at 150-350℃, followed by impurity removal from the leachate using calcium oxide. This process achieves good lithium leaching efficiency; however, at this reaction temperature, a large amount of impurities such as potassium, aluminum, and iron from the mica enter the water leaching solution, which is a complex sulfate solution with an aluminum concentration exceeding 10 g / L. Traditional calcium oxide impurity removal inevitably leads to increased lithium loss. To address this, patent CN101974684A, "Impurity Removal Process for Lithium Mica Leaching Solution," proposes an impurity removal process for lepidolite fluorine chemical water leaching solution. This process employs two alum-forming steps to remove most of the aluminum from the water leaching solution, followed by two pH adjustments for deep impurity removal to obtain a purified solution, yielding corresponding salts of alkali metal elements such as lithium, potassium, and aluminum. However, using the above technology, processing 1 ton of lepidolite ore will generate 1.5-3 tons of alum. Currently, alum has low value and limited utilization methods, resulting in a large accumulation of alum by-products and restricting the application of fluorine chemical methods. Summary of the Invention

[0004] To address the problem of generating large amounts of alum in existing fluorine-chemical lithium extraction methods from lepidolite, the present invention aims to provide an improved fluorine-chemical lithium extraction method from lepidolite.

[0005] To achieve the above objectives, the technical solution adopted by the present invention is as follows: An improved method for fluorine-chemical lithium extraction from lepidolite, comprising the following steps: S1. Stirring and acid leaching: Lithium mica and fluoride are mixed and crushed together; the crushed solid material is added to a sulfuric acid solution with a concentration of 50~90wt%, heated to 80~95℃ and stirred for 1~3h to obtain a viscous slurry; wherein, the fluoride is calculated based on the fluorine element, and the raw material ratio is lithium mica: fluoride = 1kg: (0.1~4)mol; the mass ratio of crushed solid powder to sulfuric acid solution is 2: (1~4). S2, Low-temperature calcination: The viscous slurry obtained in S1 is fed into a rotary kiln and calcined at 800~900℃ for 1~6 hours; S3, water immersion: The solid material obtained after S2 roasting is crushed and then water is added for water immersion treatment. After water immersion, it is filtered to obtain silicon slag and brine. S4. Stepwise impurity removal: Add calcium oxide to the brine obtained in S3 to adjust the pH value to 10-14, stir for 10 min to 2 h, and then filter to achieve the first impurity removal; add water-soluble carbonate to the brine obtained after the first impurity removal, stir for 0.5-2 h, and then filter to achieve the second impurity removal, and obtain purified liquid; wherein, the molar amount of carbonate is at least 1 times the molar amount of calcium in the brine obtained after the first impurity removal; S5, Lithium precipitation: Add lithium precipitation agent to the purified liquid obtained in S4 to carry out lithium precipitation reaction, filter, and obtain lithium salt product.

[0006] Ideally, in S1, the material is broken down to 150~250µm.

[0007] Preferably, in S1, the fluoride is one or a mixture of two of sodium fluoride, calcium fluoride, or potassium fluoride.

[0008] Ideally, in S2, the rotational speed of the rotary kiln is 4~20 rpm.

[0009] Preferably, in S2, the tail gas generated during the roasting process is absorbed by water and mixed with concentrated sulfuric acid to form a sulfuric acid solution with a concentration of 50~90wt%, which is then returned to S1 for reuse.

[0010] Preferably, in S3, during water immersion, the liquid-to-solid mass ratio is (1~4):1, the water immersion temperature is 50~90℃, and the immersion time is 0.5~2h.

[0011] Preferably, in S4, the water-soluble carbonate is sodium carbonate or potassium carbonate.

[0012] Preferably, in S5, the lithium precipitation agent is sodium carbonate or potassium carbonate, and the molar amount of the lithium precipitation agent is at least 1 times the molar amount of lithium oxide in the purified liquid obtained in S4; the lithium precipitation temperature is 85~99℃ and the time is 0.5~3h.

[0013] Beneficial effects: This invention utilizes the extremely strong electronegativity of fluoride ions to destroy the structure of lepidolite minerals at low temperatures, promoting lithium leaching. However, it is necessary to control the amount of fluoride and sulfuric acid to obtain a viscous slurry. The viscous slurry is roasted in a rotary kiln, where the aluminum salts (aluminum sulfate) are converted and decomposed into water-insoluble alumina, sulfur dioxide, and sulfur trioxide. The sulfur dioxide and sulfur trioxide are absorbed by water and converted into sulfuric acid, which is then returned to the acid leaching step, improving the utilization efficiency of sulfuric acid. After water leaching, the water-insoluble alumina can be removed by filtration. The brine obtained from water leaching is then treated for impurity removal and lithium precipitation to obtain lithium salts. Rotary kiln roasting not only solves the problem of aluminum impurities in the water leaching process and eliminates the problem of treating the by-product alum, but also reduces the problem of subsequent brine impurity removal, further improving the lithium recovery rate. Attached Figure Description

[0014] Figure 1 : Schematic diagram of the process of this invention. Detailed Implementation

[0015] To enable those skilled in the art to better understand the technical solutions of the present invention, the present invention will be further described below in conjunction with specific embodiments. Obviously, the described embodiments are only some, not all, of the embodiments of the present invention. Based on the described embodiments, all other embodiments obtained by those skilled in the art without creative effort are within the protection scope of the present invention.

[0016] In the following examples, the lithium oxide grade of the lepidolite concentrate is 2.15 wt%.

[0017] Example 1

[0018] An improved fluorine-chemical lithium extraction method from lepidolite, the process flow diagram is shown below. Figure 1 As shown, the specific steps are as follows: S1. Stirring and acid leaching: Mix 1 kg of lithium mica concentrate with 40 g of sodium fluoride evenly, crush it to 200 µm after ball milling, and place it in a closed reaction vessel. Slowly add 60 wt% sulfuric acid solution at a liquid-to-solid mass ratio of 2:1, heat to 80 °C, and acid leaching reaction at a stirring speed of 200 r / min for 2 h to obtain a viscous slurry. S2, Low-temperature roasting: The viscous slurry obtained in S1 is fed into a rotary kiln for roasting. The kiln temperature is set at 800℃, the kiln speed is 4rpm, and the roasting time is 2h. The tail gas generated during the roasting process is mixed with concentrated sulfuric acid through a three-stage water circulation absorption system to prepare a sulfuric acid solution of the required concentration for S1 and then returned to S1 for reuse. S3, water immersion: The solid material obtained after S2 roasting is crushed, water is added at a liquid-solid mass ratio of 1:1, water is immersed at 60℃ for 0.5h, filtered, and silicon slag and brine are obtained. The silicon slag is discarded. S4. Stepwise impurity removal: Add calcium oxide to the brine obtained in S3 to adjust the pH value of the brine to 10, stir for 0.5 h and then filter to achieve the first impurity removal; add sodium carbonate to the brine obtained after the first impurity removal, stir for 0.5 h and then filter to achieve the second impurity removal and obtain purified liquid; wherein, the molar amount of sodium carbonate is 1.0 times the molar amount of calcium in the brine obtained after the first impurity removal. S5, Lithium precipitation: Add sodium carbonate solution (concentration of 300 g / L) to the purified solution obtained in S4, react at 95°C for 2 h, filter, and obtain lithium carbonate; wherein, the amount of sodium carbonate solution added is such that the molar amount of sodium carbonate is 1.1 times the molar amount of lithium oxide.

[0019] Example 2

[0020] The difference from Example 1 is that in S1, the amount of sodium fluoride used is 50g and the acid leaching temperature is 85℃; everything else is the same as in Example 1.

[0021] Example 3

[0022] The difference from Example 1 is that in S1, the amount of sodium fluoride used is 60g, the concentration of sulfuric acid solution is 80wt%, and the acid leaching temperature is 90℃; all other aspects are the same as in Example 1.

[0023] Example 4

[0024] The difference from Example 1 is that in S1, the amount of sodium fluoride used is 70g, the concentration of sulfuric acid solution is 80wt%, and the acid leaching temperature is 95℃; everything else is the same as in Example 1.

[0025] Example 5

[0026] The difference from Example 1 is that in S1, the amount of sodium fluoride used is 80g, the concentration of sulfuric acid solution is 80wt%, and the acid leaching temperature is 95℃; everything else is the same as in Example 1.

[0027] Comparative Example 1 The traditional method for lithium extraction from lepidolite using fluorine chemistry involves the following steps: S1. Acid leaching reaction: Add 1 kg of lithium mica concentrate, 100 g of hydrofluoric acid (concentration 30 wt%) and 800 g of sulfuric acid solution (concentration 80 wt%) to a preheating reactor, control the reaction temperature at 50℃ and the preheating time at 0.1 h; then transfer the preheated material into the reactor and react at 350℃ for 4 h. S2, Clinker Water Immersion: Crush the clinker obtained from the S1 reaction, add water at a liquid-solid mass ratio of 1:1, immerse in water at 60℃ for 0.5h, filter, and obtain silicon slag and brine, discarding the silicon slag. S3, Two-stage alum formation: Cool the brine obtained in S2 to 5°C, stir for 0.5 h, filter, and complete the first alum formation and dealuminization; add potassium sulfate to the filtrate to adjust the potassium-aluminum molar ratio of the filtrate to 1:1, and after the potassium sulfate dissolves, cool to 5°C again, stir for 0.5 h, filter, and complete the second alum formation and dealuminization. S4. Deep purification: First, use calcium oxide to adjust the pH of the filtrate obtained from the secondary alum formation in S3 to 12.5. After stirring for 0.5 hours, filter. Add sodium carbonate in an equal molar amount to the calcium in the filtrate, stir for 0.5 hours, and filter to remove calcium. S5, Lithium precipitation: Add sodium carbonate solution (concentration of 300 g / L) to the filtrate obtained in S4, react at 95°C for 2 h, filter, and obtain lithium carbonate; wherein, the amount of sodium carbonate solution added is such that the molar amount of sodium carbonate is 1.1 times the molar amount of lithium oxide.

[0028] No alum was produced in Examples 1-5 above, while Comparative Example 1 produced 2.13 kg of alum and contained 0.01 wt% lithium.

[0029] The composition of the brine, lithium leaching rate, and lithium recovery rate after water immersion in the above embodiments and Comparative Example 1 are shown in Table 1. The calculation formulas for lithium leaching rate and lithium recovery rate are as follows.

[0030] The formula for calculating lithium leaching rate is: ; In the formula: m denoted as ρ, where ρ is the mass of lithium in the silica slag obtained after water leaching, in g; M is the mass of lithium in the lepidolite concentrate, in g.

[0031] The formula for calculating the overall lithium recovery rate is: ; In the formula: m' represents the mass of lithium remaining in the silica slag and other forms (including the impurity removal slag produced during the impurity removal process; in Comparative Example 1, the traditional lithium extraction method from lepidolite via fluorine chemistry includes alum produced during two alum formation processes in addition to the impurity removal slag) after water leaching, in g; M is the mass of lithium in the lepidolite concentrate, in g.

[0032]

[0033] As shown in Table 1: 1. The present invention has a better impurity removal effect and can effectively remove aluminum ions and fluoride ions; 2. Compared with the traditional lithium extraction method of lepidolite fluoride chemical extraction in Comparative Example 1, the present invention can achieve a significant increase in leaching rate by adding a small amount of fluoride; 3. As the quality of fluoride added increases, the leaching rate of lithium can be effectively improved in the present invention.

Claims

1. An improved method for fluorine-chemical lithium extraction from lepidolite, characterized in that, The steps are as follows: S1. Stirring and acid leaching: Lithium mica and fluoride are mixed and crushed together; the crushed solid material is added to a sulfuric acid solution with a concentration of 50~90wt%, heated to 80~95℃ and stirred for 1~3h to obtain a viscous slurry; wherein, the fluoride is calculated based on the fluorine element, and the raw material ratio is lithium mica: fluoride = 1kg: (0.1~4)mol; the mass ratio of crushed solid powder to sulfuric acid solution is 2: (1~4). S2, Low-temperature calcination: The viscous slurry obtained in S1 is fed into a rotary kiln and calcined at 800~900℃ for 1~6 hours; S3, water immersion: The solid material obtained after S2 roasting is crushed and then water is added for water immersion treatment. After water immersion, it is filtered to obtain silicon slag and brine. S4. Stepwise impurity removal: Add calcium oxide to the brine obtained in S3 to adjust the pH value to 10-14, stir for 10 min to 2 h, and then filter to achieve the first impurity removal; add water-soluble carbonate to the brine obtained after the first impurity removal, stir for 0.5-2 h, and then filter to achieve the second impurity removal, and obtain purified liquid; wherein, the molar amount of carbonate is at least 1 times the molar amount of calcium in the brine obtained after the first impurity removal; S5, Lithium precipitation: Add lithium precipitation agent to the purified liquid obtained in S4 to carry out lithium precipitation reaction, filter, and obtain lithium salt product.

2. The improved fluorine-chemical lithium extraction method from lepidolite as described in claim 1, characterized in that: In S1, the material is crushed to 150~250µm.

3. The improved fluorine-chemical lithium extraction method from lepidolite as described in claim 1, characterized in that: In S1, the fluoride is one or a mixture of two of sodium fluoride, calcium fluoride, or potassium fluoride.

4. The improved fluorine-chemical lithium extraction method from lepidolite as described in claim 1, characterized in that: In S2, the rotary kiln rotates at a speed of 4~20 rpm.

5. The improved fluorine-chemical lithium extraction method from lepidolite as described in claim 1, characterized in that: In S2, the tail gas generated during the roasting process is absorbed by water and mixed with concentrated sulfuric acid to form a sulfuric acid solution with a concentration of 50~90wt%, which is then returned to S1 for reuse.

6. The improved fluorine-chemical lithium extraction method from lepidolite as described in claim 1, characterized in that: In S3, during water immersion, the liquid-to-solid mass ratio is (1~4):1, the water immersion temperature is 50~90℃, and the time is 0.5~2h.

7. The improved fluorine-chemical lithium extraction method from lepidolite as described in claim 1, characterized in that: In S4, the water-soluble carbonate is sodium carbonate or potassium carbonate.

8. The improved fluorine-chemical lithium extraction method from lepidolite as described in claim 1, characterized in that: In S5, the lithium precipitation agent is sodium carbonate or potassium carbonate, and the molar amount of the lithium precipitation agent is at least 1 times the molar amount of lithium oxide in the purified liquid obtained in S4; the lithium precipitation temperature is 85~99℃ and the time is 0.5~3h.

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