A method for recovering metal elements from lepidolite and impurity removal residue

Abstract

This invention discloses a comprehensive method for recovering valuable metal elements from the neutralized and impurity-removed slag of lepidolite. The method involves obtaining clinker through high-temperature roasting of the slag; obtaining a lithium-containing leachate and aluminum-, calcium-, and iron-containing waste residue through alkaline leaching; extracting lithium from the lithium-containing leachate through resin purification, concentration, and precipitation; leaching the aluminum-, calcium-, and iron-containing waste residue with sulfuric acid and hydrogen peroxide to obtain calcium sulfate and an iron-aluminum-containing aqueous leachate; and recovering aluminum through extraction and ammonium-aluminum crystallization reaction to obtain iron oxide red and ammonium alum products. This method achieves comprehensive recovery of valuable elements such as lithium, iron, aluminum, and calcium from the neutralized and impurity-removed slag of lepidolite. The waste residue and wastewater generated throughout the process are properly utilized, effectively solving the problem of secondary resource utilization of lepidolite neutralized and impurity-removed slag. This achieves the goals of resource utilization, harmlessness, and volume reduction of lepidolite neutralized and impurity-removed slag, aligning with my country's goals of developing a circular economy and energy conservation and emission reduction.

Description

Technical Field

[0001] This invention relates to the field of resources and environmental technology, and in particular to a method for recovering metal elements from lepidolite and impurity removal residue. Background Technology

[0002] Lithium possesses unique electrochemical properties and some distinctive characteristics, making it widely used in ceramics, glass, lithium batteries, and other fields. In recent years, with the rapid development of new energy vehicles and energy storage industries, lithium has become a crucial mineral resource for countries worldwide. While lithium content in the Earth's crust is scarce, global lithium resources are abundant, mainly concentrated in a few countries such as Bolivia, Chile, Argentina, China, Australia, and the United States, primarily in the form of salt lakes. Currently, my country's salt lakes have a high magnesium-to-lithium ratio, and large-scale industrial production has not yet been achieved. Domestic lithium production still mainly relies on lithium extraction from lepidolite and spodumene.

[0003] Currently, the lithium extraction process using sulfate roasting of lepidolite mainly includes: roasting a mixture of lepidolite and sulfate, followed by water leaching. Quicklime is then added to the resulting leachate for primary purification, producing a neutralization residue and a primary purified solution. This primary purified solution undergoes further purification and lithium precipitation to obtain lithium carbonate. During this process, the roasting product from the mixed roasting of lepidolite and sulfate dissolves impurities such as iron and aluminum during water leaching. The primary purification of the leachate with quicklime produces a neutralization residue, which is an amorphous precipitate of iron and aluminum hydroxides. Simultaneously, due to selective precipitation, some lithium in the leachate co-precipitates in the residue, resulting in lithium loss. To recover lithium from the neutralization residue, the industry commonly uses a roasting and water leaching process. However, this method suffers from the redissolution of elements such as aluminum, iron, calcium, and manganese during water leaching, resulting in a high impurity content in the leachate. To ensure the purity of the recovered lithium, a secondary neutralization and purification process is required. However, this secondary neutralization and purification process inevitably leads to further lithium loss. In addition, the products roasted after water immersion can only be stored in piles, which leads to problems such as the failure to fully utilize valuable elements such as iron and aluminum in the slag. Summary of the Invention

[0004] Based on the technical problems existing in the background art, the present invention proposes a method for recovering metal elements from lepidolite and impurity removal slag.

[0005] This invention proposes a method for recovering metal elements from lepidolite and impurity removal residue, comprising the following steps:

[0006] S1. The lepidolite neutralized and impurity-removed residue is roasted and crushed to obtain clinker; the clinker is leached with alkaline solution to obtain lithium-containing leaching solution and aluminum-calcium-iron-containing waste residue.

[0007] S2. The lithium-containing eluent is purified by ion exchange resin to obtain a purified mother liquor; the purified mother liquor is concentrated, and sodium carbonate solution is added to the concentrated solution to precipitate it to obtain battery-grade lithium carbonate.

[0008] S3. The aluminum-calcium-iron waste residue is leached with hydrogen peroxide and concentrated sulfuric acid to obtain an iron-aluminum leachate and calcium sulfate; the iron-aluminum leachate is extracted with an extractant to obtain an iron-containing organic phase and raffinate; the iron-containing organic phase is back-extracted with sulfuric acid, neutralized with ammonia, and calcined to obtain iron oxide red.

[0009] S4. The raffinate is subjected to an ammonium-aluminum reaction, and then cooled to crystallize, yielding ammonium alum.

[0010] Preferably, the lepidolite neutralization and impurity removal slag is the waste residue generated by leaching the calcined product of lepidolite mixed with sulfate with water, and then adding quicklime to the resulting water leachate for primary purification; the main components of the lepidolite neutralization and impurity removal slag include: Li2SO4 1%~4%, Fe(OH)3 5%~10%, Al(OH)3 25%~40%, and CaSO4 30%~45%.

[0011] Preferably, in S1, the calcination temperature is 400℃~700℃, and the calcination time is 20min~240min.

[0012] Preferably, in S1, the pH of the alkaline solution is 8 to 11, and the alkaline solution is at least one of caustic soda solution, sodium carbonate solution, and ammonia water; in S1, the liquid-solid ratio of the alkaline solution to the clinker is 1:1 to 3:1.

[0013] In S1, the main reactions that occur are:

[0014] 2Al(OH)3→Al2O3+3H2O

[0015] 2Fe(OH)3→Fe2O3+3H2O

[0016] Ca 2+ +CO3 2- →CaCO3↓

[0017] Al 3+ +3OH - →Al(OH)3↓.

[0018] High-temperature calcination decomposes amorphous hydroxides into oxides with no adsorption capacity, disrupting the chemical composition of lithium-aluminum co-precipitates and converting lithium in the impurity residue into soluble components. Alkaline rinsing further inhibits the decomposition of aluminum, iron, and manganese oxides, as well as the leaching of soluble calcium and aluminum salts from the impurity residue.

[0019] Preferably, in step S2, the ion exchange resin is a macroporous weakly acidic cation exchange resin; in step S2, the concentration of the sodium carbonate solution is 300 g / L to 360 g / L, the concentration of lithium ions in the concentrate is 15 g / L to 40 g / L, and the amount of sodium carbonate used is 0.5 to 0.8 times the total molar amount of lithium in the concentrate.

[0020] Preferably, the ion exchange resin is one of HP-8 (Jiangsu Haipu), D403 (Shengquan Chemical), or CH-93 (Kehaisi).

[0021] In S2, the main reactions that occur are:

[0022] R2R-COOLi+Ca 2+ →R2R-COOCa+Li +

[0023] Na2CO3+Li2SO4→Na2SO4+Li2CO3↓

[0024] Ion exchange is used to deeply remove impurities such as calcium and magnesium from the rinsing solution, while enriching the lithium concentration in the brine and then evaporating and concentrating it to meet the precipitation conditions.

[0025] Preferably, in S2, the flow rate of the lithium-containing rinsing solution used for impurity removal by ion exchange resin is ≤30m / h.

[0026] Preferably, in step S3, the aluminum-calcium-iron waste residue is mixed evenly with water to obtain a slurry with a solid content of 30-50%, then hydrogen peroxide equivalent to 1%-5% of the slurry volume is added, and then concentrated sulfuric acid is slowly added and stirred for 60-200 minutes to leach, controlling the final pH value to be greater than 0 and less than or equal to 3, to obtain an iron-aluminum leachate and calcium sulfate.

[0027] Preferably, the hydrogen peroxide is industrial hydrogen peroxide with a mass concentration of 30-35%.

[0028] Preferably, in S3, the mass concentration of the concentrated sulfuric acid is 92-98%.

[0029] Preferably, in step S3, the method for extracting the iron-aluminum leaching solution using a composite extractant containing N235 is as follows: the iron-aluminum leaching solution is mixed with the extractant for single-stage or multi-stage extraction, the extraction time is 2 min to 15 min, and the oil-water ratio O / A is 1:1 to 1:5. The extractant comprises the following components by mass percentage: extractant N235 20% to 30%, TBP 5% to 10%, and the balance is sulfonated kerosene. In step S3, the iron-containing organic phase is mixed with a sulfuric acid solution with a concentration of 20 g / L to 50 g / L for single-stage or multi-stage back-extraction, the back-extraction time is 5 min to 20 min, and the oil-water ratio O / A is 5:1 to 1:1. Then, ammonia water is added to neutralize to a pH of 5 to 6. After filtration, the precipitate is calcined at 400℃ to 600℃ to obtain iron oxide red.

[0030] In S3, the main reactions that occur are:

[0031] Ca4Al2O6SO4·11H2O+H2SO4→Al2(SO4)3+CaSO4↓+H2O

[0032] Fe2O3+2FeO+6H2SO4++H2O2→2Fe2(SO4)3+7H2O

[0033] xFe 3+ +x(3-z) / 2SO4 2- +xy / 2(R3NH)2SO4+xzH2O→[(R3NH) y Fe(OH) z (SO4) (3+y-z) / 2 ] x

[0034] +xzH +

[0035] In the formula: x is the extractability of the extract, y is the number of tertiary amine molecules, and z is the number of hydroxyl groups.

[0036] Fe2(SO4)3+6NH3·H2O→2Fe(OH)3↓+3(NH4)2SO4

[0037] 2Fe(OH)3→Fe2O3+3H2O

[0038] Preferably, in step S4, a saturated ammonium sulfate solution is added to the raffinate, and an ammonium-aluminum reaction is carried out at 50°C to 90°C for 20 min to 240 min. The resulting reaction solution is then cooled and crystallized at 0°C to 10°C for 20 min to 80 min to obtain ammonium alum. The amount of ammonium sulfate used is 0.5 to 0.8 times the total molar amount of aluminum in the raffinate. The addition rate of the saturated ammonium sulfate solution is 2 mL / min to 10 mL / min.

[0039] Preferably, the method for recovering metal elements from lepidolite and impurity removal residue further includes: reusing the calcium sulfate obtained in S3 for mixed roasting of lepidolite and sulfate.

[0040] Preferably, the method for recovering metal elements from lepidolite and impurity residue further includes: adjusting the supernatant from cooling and crystallizing in S4 to neutral or weakly acidic using concentrated sulfuric acid, and then reusing it in the leaching process in S3. Since the supernatant from cooling and crystallizing in S4 contains ammonium ions and unrecovered aluminum ions, returning it to the leaching process in S3 allows for the recycling of aluminum and ammonium ions.

[0041] The beneficial effects of this invention are as follows:

[0042] 1) This invention uses alkaline solution to rinse the calcined and neutralized residue clinker, transferring water-soluble lithium into the solution. At the same time, under alkaline conditions, oxides such as aluminum, iron, calcium, and manganese are difficult to enter the rinsing solution, thus achieving efficient extraction of lithium from the neutralized residue of lepidolite.

[0043] 2) For eluents with low impurity content, this invention uses resin for impurity removal, which eliminates the shortcomings of traditional precipitation impurity removal processes, effectively reduces lithium loss, and uses sodium carbonate to obtain battery-grade lithium carbonate products with a purity of over 99.6% and a lithium recovery rate of over 96%.

[0044] 3) This invention uses sulfuric acid and hydrogen peroxide to leach the waste residue generated after rinsing, transferring valuable components such as iron and aluminum from the neutralized lepidolite residue to the solution. Then, through extraction separation and ammonium aluminum reaction, the valuable elements such as iron and aluminum in the residue are extracted simultaneously. The iron recovery rate can reach more than 95%, the aluminum recovery rate can reach more than 95%, and products such as iron oxide red and ammonium alum are obtained. The calcium sulfate produced after leaching can be directly returned to the lepidolite roasting process after crushing, and the supernatant after ammonium aluminum reaction crystallization can be directly returned to the leaching process after acid adjustment.

[0045] In summary, this invention relates to a comprehensive method for recovering valuable metal elements from lepidolite neutralization and impurity removal slag. The method involves obtaining clinker through high-temperature roasting to neutralize the slag; obtaining a lithium-containing leachate and aluminum-, calcium-, and iron-containing waste residue through alkaline leaching; extracting lithium from the lithium-containing leachate through resin purification, concentration, and precipitation; leaching the aluminum-, calcium-, and iron-containing waste residue with sulfuric acid and hydrogen peroxide to obtain calcium sulfate and an iron-aluminum-containing aqueous leachate; and recovering aluminum through extraction and ammonium-aluminum crystallization reaction to obtain iron oxide red and ammonium alum products. This method achieves comprehensive recovery of valuable elements such as lithium, iron, aluminum, and calcium from lepidolite neutralization and impurity removal slag. All waste residue and wastewater generated during the process are properly utilized, effectively solving the problem of secondary resource utilization of lepidolite neutralization and impurity removal slag. This achieves the goals of resource utilization, harmlessness, and volume reduction of lepidolite neutralization and impurity removal slag, aligning with my country's goals of developing a circular economy and energy conservation and emission reduction. Attached Figure Description

[0046] Figure 1 This is a schematic diagram of the process for recovering metal elements from lepidolite and impurity removal slag proposed in this invention. Detailed Implementation

[0047] The technical solution of the present invention will now be described in detail through specific embodiments.

[0048] Example 1

[0049] Metal elements are recovered from the neutralized residue of lepidolite, whose main components include: Li₂SO₄ 3%, Fe(OH)₃ 7%, Al(OH)₃ 32%, and CaSO₄ 41%. The recovery method is as follows:

[0050] S1. The neutralized and impurity-removed slag of lithium mica is fed into a rotary kiln and roasted at 550°C for 20 minutes. Then it is crushed to 30 mesh to obtain clinker. The clinker is leached with an alkaline solution with a pH of 10 to obtain a lithium-containing leaching solution and aluminum-calcium-iron-containing waste residue. The alkaline solution is a mixed solution of caustic soda and sodium carbonate, and the liquid-solid ratio of the alkaline solution to the clinker is 1.5:1.

[0051] S2. The lithium-containing eluent was purified using CH-93 resin at a feed rate of 20 m / h to obtain a purified mother liquor. The mother liquor was concentrated to a lithium ion concentration of 36 g / L using MVR evaporation technology. A sodium carbonate solution with a concentration of 360 g / L was added to the concentrated solution, and precipitation was carried out at 95°C for 1 h. The precipitate was washed three times with water to obtain battery-grade lithium carbonate. The amount of sodium carbonate used was 0.55 times the total molar amount of lithium in the concentrated solution.

[0052] S3. Mix aluminum-calcium-iron waste residue with water to obtain a slurry with a solid content of 40%. Then add industrial hydrogen peroxide equivalent to 2% of the slurry volume, and slowly add concentrated sulfuric acid with a mass concentration of 98% while stirring and leaching for 100 minutes. Control the final pH to 1.5 to obtain an iron-aluminum leaching solution and calcium sulfate. The obtained calcium sulfate is recycled for calcination of lepidolite and sulfate.

[0053] The iron-aluminum leachate was mixed with a composite extractant containing N235 and subjected to four-stage cross-flow extraction for 10 min at an oil-water ratio of O / A = 1:1. The resulting product was an iron-containing organic phase and a raffinate. The composite extractant containing N235 consisted of the following components by mass percentage: N235 extractant 20%, TBP 10%, and the balance being sulfonated kerosene.

[0054] The iron-containing organic phase was mixed with a sulfuric acid solution with a concentration of 25 g / L and subjected to two-stage countercurrent back-extraction for 15 min. The oil-to-water ratio O / A = 2:1. Ammonia water was then added to neutralize the solution to pH 5-6. After filtration, the precipitate was calcined at 550 °C to obtain iron red.

[0055] S4. Add a saturated ammonium sulfate solution to the raffinate at a rate of 5 mL / min, and carry out the ammonium-aluminum reaction at 60°C for 100 min. Then, cool the resulting reaction solution at 5°C for 40 min to crystallize and obtain ammonium alum. The amount of ammonium sulfate used is 0.55 times the total molar amount of aluminum in the raffinate. The addition rate of the saturated ammonium sulfate solution is 5 mL / min. Adjust the supernatant of the cooled crystallized solution to neutral or weakly acidic using concentrated sulfuric acid, and then reuse it in the leaching process in S3.

[0056] The recovery rates for lithium were 95%, iron 92%, aluminum 93%, and calcium 96%.

[0057] Example 2

[0058] Metal elements are recovered from the neutralized residue of lepidolite, whose main components are: Li₂SO₄ 3%, Fe(OH)₃ 7%, Al(OH)₃ 32%, and CaSO₄ 41%. The recovery method is as follows:

[0059] S1. The neutralized and impurity-removed slag of lithium mica is fed into a rotary kiln and roasted at 650°C for 40 minutes. Then it is crushed to 50 mesh to obtain clinker. The clinker is leached with an alkaline solution with a pH of 10 to obtain a lithium-containing leaching solution and aluminum-calcium-iron-containing waste residue. The alkaline solution is a mixed solution of caustic soda and sodium carbonate, and the liquid-solid ratio of the alkaline solution to the clinker is 2:1.

[0060] S2. The lithium-containing eluent was purified using CH-93 resin at a feed rate of 20 m / h to obtain a purified mother liquor. The mother liquor was concentrated to a lithium ion concentration of 36 g / L using MVR evaporation technology. A sodium carbonate solution with a concentration of 360 g / L was added to the concentrated solution, and precipitation was carried out at 95°C for 1 h. The precipitate was washed three times with water to obtain battery-grade lithium carbonate. The amount of sodium carbonate used was 0.6 times the total molar amount of lithium in the concentrated solution.

[0061] S3. Mix aluminum-calcium-iron waste residue with water to obtain a slurry with a solid content of 30%. Then add industrial hydrogen peroxide equivalent to 4% of the slurry volume, and slowly add concentrated sulfuric acid with a mass concentration of 98% while stirring and leaching for 140 minutes. Control the final pH to 1.0 to obtain an iron-aluminum leaching solution and calcium sulfate. The obtained calcium sulfate is recycled for calcination of lepidolite and sulfate.

[0062] The iron-aluminum leachate was mixed with a composite extractant containing N235 and subjected to four-stage cross-flow extraction for 10 min at an oil-water ratio of O / A = 1:1. The resulting product was an iron-containing organic phase and a raffinate. The composite extractant containing N235 consisted of the following components by mass percentage: N235 extractant 20%, TBP 10%, and the balance being sulfonated kerosene.

[0063] The iron-containing organic phase was mixed with a sulfuric acid solution with a concentration of 20 g / L to 50 g / L and subjected to two-stage countercurrent back-extraction for 15 min. The oil-to-water ratio O / A was 2:1. Ammonia water was then added to neutralize the solution to pH 5 to 6. After filtration, the precipitate was calcined at 550 °C to obtain iron red.

[0064] S4. Add a saturated ammonium sulfate solution to the raffinate and carry out the ammonium-aluminum reaction at 60°C for 100 min. Then, cool the resulting reaction solution at 5°C for 40 min to crystallize and obtain ammonium alum. The amount of ammonium sulfate used is 0.55 times the total molar amount of aluminum in the raffinate. The addition rate of the saturated ammonium sulfate solution is 5 mL / min. Adjust the supernatant of the cooled crystallized solution to neutral or weakly acidic using concentrated sulfuric acid, and then reuse it in the leaching process in S3.

[0065] The recovery rates for lithium were 96%, iron 94%, aluminum 95%, and calcium 97%.

[0066] Example 3

[0067] Metal elements are recovered from the neutralized residue of lepidolite, whose main components are: Li₂SO₄ 3%, Fe(OH)₃ 7%, Al(OH)₃ 32%, and CaSO₄ 41%. The recovery method is as follows:

[0068] S1. The neutralized and impurity-removed slag of lithium mica is fed into a rotary kiln and roasted at 650°C for 40 minutes. Then it is crushed to 50 mesh to obtain clinker. The clinker is leached with an alkaline solution with a pH of 10 to obtain a lithium-containing leaching solution and aluminum-calcium-iron-containing waste residue. The alkaline solution is a mixed solution of caustic soda and sodium carbonate, and the liquid-solid ratio of the alkaline solution to the clinker is 2.5:1.

[0069] S2. The lithium-containing eluent was purified using CH-93 resin at a feed rate of 20 m / h to obtain a purified mother liquor. The mother liquor was concentrated to a lithium ion concentration of 36 g / L using MVR evaporation technology. A sodium carbonate solution with a concentration of 360 g / L was added to the concentrated solution, and precipitation was carried out at 95°C for 1 h. The precipitate was washed three times with water to obtain battery-grade lithium carbonate. The amount of sodium carbonate used was 0.6 times the total molar amount of lithium in the concentrated solution.

[0070] S3. Mix aluminum-calcium-iron waste residue with water to obtain a slurry with a solid content of 50%. Then add industrial hydrogen peroxide equivalent to 5% of the slurry volume, and slowly add concentrated sulfuric acid with a mass concentration of 98% while stirring and leaching for 140 minutes. Control the final pH to 0.8 to obtain an iron-aluminum leaching solution and calcium sulfate. The obtained calcium sulfate is recycled for calcination of lepidolite and sulfate.

[0071] The iron-aluminum leachate was mixed with a composite extractant containing N235 and subjected to four-stage cross-flow extraction for 15 min at an oil-water ratio of O / A = 1:1. The resulting product was an iron-containing organic phase and a raffinate. The composite extractant containing N235 consisted of the following components by mass percentage: N235 extractant 25%, TBP 5%, and the balance being sulfonated kerosene.

[0072] The iron-containing organic phase was mixed with a sulfuric acid solution with a concentration of 30 g / L and subjected to two-stage countercurrent back-extraction for 20 min. The oil-to-water ratio O / A = 2:1. Ammonia water was then added to neutralize the solution to pH 5-6. After filtration, the precipitate was calcined at 550 °C to obtain iron red.

[0073] S4. Add a saturated ammonium sulfate solution to the raffinate and carry out the ammonium-aluminum reaction at 60°C for 120 min. Then, cool the resulting reaction solution at 5°C for 60 min to crystallize and obtain ammonium alum. The amount of ammonium sulfate used is 0.55 times the total molar amount of aluminum in the raffinate. The addition rate of the saturated ammonium sulfate solution is 5 mL / min. Adjust the supernatant of the cooled crystallized solution to neutral or weakly acidic using concentrated sulfuric acid, and then reuse it in the leaching process in S3.

[0074] The recovery rates for lithium, iron, aluminum, and calcium were 96%, 95%, 95%, and 98%, respectively.

[0075] Example 4

[0076] Metal elements are recovered from the neutralized residue of lepidolite, whose main components are: Li₂SO₄ 3%, Fe(OH)₃ 7%, Al(OH)₃ 32%, and CaSO₄ 41%. The recovery method is as follows:

[0077] S1. The neutralized and impurity-removed slag of lithium mica is fed into a rotary kiln and roasted at 650°C for 40 minutes. Then it is crushed to 50 mesh to obtain clinker. The clinker is leached with an alkaline solution with a pH of 10 to obtain a lithium-containing leaching solution and aluminum-calcium-iron-containing waste residue. The alkaline solution is a mixed solution of caustic soda and sodium carbonate, and the liquid-solid ratio of the alkaline solution to the clinker is 2.5:1.

[0078] S2. The lithium-containing eluent was purified using ion exchange resin at a feed rate of 20 m / h to obtain a purified mother liquor. The mother liquor was concentrated to a lithium ion concentration of 36 g / L using MVR evaporation technology. A sodium carbonate solution with a concentration of 360 g / L was added to the concentrated solution, and precipitation was carried out at 95°C for 1 h. The precipitate was washed three times with water to obtain battery-grade lithium carbonate. The amount of sodium carbonate used was 0.6 times the total molar amount of lithium in the concentrated solution.

[0079] S3. Mix aluminum-calcium-iron waste residue with water to obtain a slurry with a solid content of 50%. Then add industrial hydrogen peroxide equivalent to 5% of the slurry volume, and slowly add concentrated sulfuric acid with a mass concentration of 98% while stirring and leaching for 140 minutes. Control the final pH to 0.5 to obtain an iron-aluminum leaching solution and calcium sulfate. The obtained calcium sulfate is recycled for calcination of lepidolite and sulfate.

[0080] The iron-aluminum leachate was mixed with a composite extractant containing N235 and subjected to four-stage cross-flow extraction for 15 min at an oil-water ratio of O / A = 1:1. The resulting product was an iron-containing organic phase and a raffinate. The composite extractant containing N235 consisted of the following components by mass percentage: N235 extractant 25%, TBP 5%, and the balance being sulfonated kerosene.

[0081] The iron-containing organic phase was mixed with a sulfuric acid solution with a concentration of 30 g / L and subjected to two-stage countercurrent back-extraction for 20 min. The oil-to-water ratio O / A = 2:1. Ammonia water was then added to neutralize the solution to pH 5-6. After filtration, the precipitate was calcined at 550 °C to obtain iron red.

[0082] S4. Add a saturated ammonium sulfate solution to the raffinate and carry out the ammonium-aluminum reaction at 65°C for 180 min. Then, cool the resulting reaction solution at 5°C for 80 min to crystallize and obtain ammonium alum. The amount of ammonium sulfate used is 0.6 times the total molar amount of aluminum in the raffinate. The addition rate of the saturated ammonium sulfate solution is 3 mL / min. Adjust the supernatant of the cooled crystallized solution to neutral or weakly acidic using concentrated sulfuric acid, and then reuse it in the leaching process in S3.

[0083] The recovery rates for lithium, iron, aluminum, and calcium were 96%, 95%, 96%, and 98%, respectively.

[0084] The above description is only a preferred embodiment of the present invention, but the scope of protection of the present invention is not limited thereto. Any equivalent substitutions or modifications made by those skilled in the art within the scope of the technology disclosed in the present invention, based on the technical solution and inventive concept of the present invention, should be covered within the scope of protection of the present invention.

Claims

1. A method for recovering metal elements from lepidolite and impurity removal residue, characterized in that, Includes the following steps: S1. The lepidolite neutralized and impurity-removed residue is roasted and crushed to obtain clinker; the clinker is leached with an alkaline solution to obtain a lithium-containing leaching solution and aluminum-calcium-iron-containing waste residue, wherein the pH of the alkaline solution is 8-11, and the alkaline solution is at least one of caustic soda solution, sodium carbonate solution, and ammonia water; the lepidolite neutralized and impurity-removed residue is the waste residue generated by leaching the roasted product of lepidolite mixed with sulfate after water leaching, and then adding quicklime to the resulting water leaching solution for primary purification. S2. The lithium-containing eluent is purified by ion exchange resin to obtain a purified mother liquor; the purified mother liquor is concentrated, and sodium carbonate solution is added to the concentrated solution to precipitate it to obtain battery-grade lithium carbonate. S3. The aluminum-calcium-iron waste residue is leached with hydrogen peroxide and concentrated sulfuric acid to obtain an iron-aluminum leachate and calcium sulfate; the iron-aluminum leachate is extracted with an extractant to obtain an iron-containing organic phase and raffinate; the iron-containing organic phase is back-extracted with sulfuric acid, neutralized with ammonia, and calcined to obtain iron oxide red. S4. The raffinate is subjected to an ammonium-aluminum reaction, and then cooled to crystallize, yielding ammonium alum.

2. The method for recovering metal elements from lepidolite and impurity removal residue according to claim 1, characterized in that, The main components of the lepidolite neutralization and impurity removal residue include: Li2SO4 1%~4%, Fe(OH)3 5%~10%, Al(OH)3 25%~40%, and CaSO4 30%~45%.

3. The method for recovering metal elements from lepidolite and impurity removal residue according to claim 1, characterized in that, In S1, the calcination temperature is 400℃~700℃, and the calcination time is 20min~240min.

4. The method for recovering metal elements from lepidolite and impurity removal residue according to claim 1, characterized in that, In S1, the liquid-to-solid ratio of the alkali solution to the clinker is 1:1 to 3:

1.

5. The method for recovering metal elements from lepidolite and impurity removal residue according to claim 1, characterized in that, In S2, the ion exchange resin is a macroporous weakly acidic cation exchange resin; in S2, the concentration of the sodium carbonate solution is 300 g / L to 360 g / L, the concentration of lithium ions in the concentrate is 15 g / L to 40 g / L, and the amount of sodium carbonate used is 0.5 to 0.8 times the total molar amount of lithium in the concentrate.

6. The method for recovering metal elements from lepidolite and impurity removal residue according to claim 1, characterized in that, In step S3, the aluminum-calcium-iron waste residue is mixed evenly with water to obtain a slurry with a solid content of 30-50%. Then, hydrogen peroxide equivalent to 1%-5% of the slurry volume is added, followed by slow addition of concentrated sulfuric acid and stirring for 60-200 minutes. The final pH value is controlled to be greater than 0 and less than or equal to 3 to obtain an iron-aluminum leachate and calcium sulfate.

7. The method for recovering metal elements from lepidolite and impurity removal residue according to claim 1, characterized in that, In S3, the method for extracting the iron-aluminum leaching solution using an extractant is as follows: the iron-aluminum leaching solution is mixed with the extractant for single-stage or multi-stage extraction, the extraction time is 2 min to 15 min, and the oil-water ratio O / A is 1:1 to 1:

5. The extractant comprises the following components by mass percentage: extractant N235 20% to 30%, TBP 5% to 10%, and the balance is sulfonated kerosene. In S3, the iron-containing organic phase is mixed with a sulfuric acid solution with a concentration of 20 g / L to 50 g / L for single-stage or multi-stage back-extraction, the back-extraction time is 5 min to 20 min, and the oil-water ratio O / A is 5:1 to 1:

1. Then, ammonia water is added to neutralize to a pH of 5 to 6. After filtration, the obtained precipitate is calcined at 400℃ to 600℃ to obtain iron oxide red.

8. The method for recovering metal elements from lepidolite and impurity removal residue according to claim 1, characterized in that, In step S4, a saturated ammonium sulfate solution is added to the raffinate, and an ammonium-aluminum reaction is carried out at 50℃ to 90℃ for 20 min to 240 min. The resulting reaction solution is then cooled and crystallized at 0℃ to 10℃ for 20 min to 80 min to obtain ammonium alum. The amount of ammonium sulfate used is 0.5 to 0.8 times the total molar amount of aluminum in the raffinate. The addition rate of the saturated ammonium sulfate solution is 2 mL / min to 10 mL / min.

9. The method for recovering metal elements from lepidolite and impurity removal residue according to claim 1, characterized in that, Also includes: The calcium sulfate obtained from S3 was recycled for roasting a mixture of lepidolite and sulfate.

10. The method for recovering metal elements from lepidolite and impurity removal residue according to claim 1, characterized in that, Also includes: The supernatant from the cooling crystallization in S4 is adjusted to acidity with concentrated sulfuric acid and then reused in the leaching process in S3.

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