CN116532235A - Resource comprehensive utilization method of spodumene smelting slag - Google Patents

Abstract

The invention relates to a resource comprehensive utilization method of spodumene smelting slag, and belongs to the technical field of resource treatment of solid waste for extracting lithium from ores. The invention solves the technical problem of providing a resource comprehensive utilization method of spodumene smelting slag. The method comprises pulping, ore grinding, leaching, pulp mixing, flotation, magnetic separation, gravity separation and weak magnetic separation. The process provided by the invention can recycle lithium, tantalum and niobium in the lithium slag, is simple in process, efficient and environment-friendly, and can be used for remarkably improving the comprehensive utilization rate of lithium resources, not only realizing de-iron of slag, but also recycling tantalum and niobium lithium in slag, and further obtaining a silicon-aluminum refined powder product and additional product gypsum for glass fiber industry. In addition, the invention can reduce the water treatment capacity and save the production cost; the method can also be used for completely eating and squeezing the lithium slag, and waste slag and waste water are not generated, so that the problem of digestion of the lithium slag is fundamentally solved, the comprehensive utilization of the lithium slag is realized, and waste is changed into valuable.

Description

Comprehensive utilization method of spodumene smelting slag resources

technical field

The invention relates to a resource comprehensive utilization method of spodumene smelting slag, and belongs to the technical field of solid waste resource treatment for lithium extraction from ore.

Background technique

Spodumene smelting slag is the solid waste of ore lithium extraction, and spodumene lithium extraction process is a relatively mature ore lithium extraction process. This method first roasts natural spodumene at 950-1100 ° C to make it from a single The α-spodumene of the oblique crystal system is transformed into the β-spodumene of the tetragonal system. Due to the transformation of the crystal form, the physical and chemical properties of the mineral also change significantly with the change of the crystal structure, and the chemical activity increases. Bases undergo various reactions. In the process of extracting lithium from spodumene, it is inevitable that some lithium remains in the lithium slag. It will be of great significance to fully tap the value of secondary recycling of lithium in solid waste resources.

The main chemical components of spodumene smelting slag are SiO ₂ and Al ₂ O ₃ , mainly aluminosilicate _and quartz, followed by gypsum and a small amount of spodumene and iron minerals. 3-8%, Fe ₂ O ₃ content 0.8-1.2%, Li ₂ O content 0.3-0.6%, (TaNb) ₂ O ₅ content 120-180ppm, among which the valuable metal lithium and tantalum niobium have comprehensive recycling value .

In the process of using spodumene to produce lithium salt, about 8 to 10 tons of lithium slag are discharged for every ton of lithium salt produced. According to this discharge, a large amount of lithium slag will be produced, which not only leads to waste of land resources due to stacking, but also poor storage , The loss of slag water containing alkali and acid will harm farmland and pollute the environment. At present, the comprehensive utilization of lithium slag is mainly used in the cement building materials industry, with low added value. A small amount of lithium, tantalum, niobium metal and gypsum have not been recycled. The consumption is close to saturation, therefore, the consumption of lithium slag will become an urgent problem to be solved in the future.

The patent of invention with the publication number CN110015855A discloses a treatment method for lithium slag. By performing sulfuric acid leaching on lepidolite lithium slag, the leaching rates of lithium, rubidium, cesium, potassium, aluminum and sodium are all up to 88%. As mentioned above, the main components of the obtained acid leaching residue are quartz and gypsum, and the quartz and gypsum can be used as concrete admixtures for secondary use. Although this method recovers valuable elements such as lithium in the slag, the recovery cost is relatively high, and it is easy to generate more solid slag or waste liquid, which has certain environmental risks, and does not substantially solve the problem of lithium slag consumption. The vast majority of lithium slag can only be used as an admixture of cheap cement building materials.

The invention patent with the publication number CN114702048A discloses a lithium slag solid waste resource recovery process, through the acid-base reaction of lithium slag, potassium sulfate, sodium sulfate, lithium carbonate, cesium carbonate, rubidium carbonate and other products are obtained. The process is relatively complicated, the cost is high, and hydrofluoric acid needs to be introduced, which is likely to cause environmental pollution. Most of the acid-insoluble solids in the lithium slag are only used as building materials, and the lithium slag cannot be truly recycled.

The invention patent with the publication number CN113621811A discloses a method for recovering tantalum and niobium from spodumene slag. The prerequisite for this technology is to add a small amount of waste acid to make the pH of the slag slurry = 4-5, which is easy to cause corrosion to equipment and has certain environmental risks. .

The invention patent with the publication number CN114226413A discloses a comprehensive treatment process for lithium slag, including the use of grinding, magnetic separation, flotation, alkali conversion and other processes to obtain silicon-aluminum powder. Generally, the fineness of spodumene smelting slag is relatively fine. This process does not carry out graded grinding of lithium slag, and uses the method of adding sodium carbonate or potassium carbonate for alkali conversion to reduce the sulfur in the micropowder, resulting in high grinding costs and desulfurization costs; although this process can also obtain glass fiber Si-alumina fine powder, but the valuable metal lithium and gypsum in the lithium slag have not been recovered reasonably, and it is difficult to realize the resource disposal of lithium slag in a true sense.

The invention patent with the publication number CN113976309A discloses a method for comprehensively recovering lithium, tantalum-niobium, silicon-aluminum powder, iron concentrate and gypsum from lithium slag. After gravity separation of lithium slag, the magnetic field is weakened to obtain concentrate 1, tailings 1, concentrate 1 Coarse tantalum and niobium-rich material and coarse iron concentrate are obtained through weak magnetic separation; tailings 1 are subjected to flotation to obtain gypsum and tailings 2; after tailings 2 are crushed, weak magnetic separation is performed to obtain fine iron concentrate and Tailings 3; strong magnetic separation of tailings 3 to obtain concentrate 2 and tailings 4, drying tailings 4 to obtain silicon-aluminum powder; concentrate 2 to obtain fine-grained tantalum-niobium concentrate and high-iron lithium-rich material, and then obtain Lithium recovery from high-iron lithium-rich materials. Using this method to recover lithium in lithium slag, only part of the lithium in the high-iron lithium-rich material with a small output is recovered, and Li ₂ O in the silicon-aluminum micropowder product is not recovered. The recovery rate of lithium in the whole process is 20.5%. The rate is low, resulting in the waste of lithium resources.

Contents of the invention

In view of the above defects, the technical problem to be solved by the present invention is to provide a comprehensive utilization method for spodumene smelting slag resources.

The comprehensive utilization method of spodumene smelting slag of the present invention comprises the following steps:

a. Pulping: Spodumene slag and water are mixed evenly to prepare lithium slag slurry with a solid-liquid mass ratio of 1:1-3;

B. Grinding: The lithium slag slurry in step a is classified into particle size, divided into fine-grained slurry and coarse-grained slurry, and the coarse-grained slurry is wet-grinded. After grinding, it is merged with the fine-grained slurry to obtain fine slurry;

c. Leaching: add sulfuric acid to the fine slurry, adjust the pH to 1-1.5, heat and stir for leaching; then separate the solid and liquid to obtain acidic slag and leachate; the leachate is lithium-poor liquid;

d. Pulping: After washing the acid slag, add water or return water to stir and make pulp, adjust the concentration of the pulp to 25-35wt%, and adjust the pH of the pulp to 6-7 by adding limestone, quicklime or slaked lime;

e. Flotation: Flotation desulfurization is carried out on the slurry after step d slurry adjustment to obtain desulfurization lithium slag and flotation foam products;

f. Magnetic separation: Desulfurization lithium slag adopts medium magnetic separation to obtain iron-containing material A and medium magnetic concentrate; medium magnetic concentrate adopts strong magnetic separation to obtain iron-containing material B and strong magnetic concentrate; strong magnetic concentrate Filter and dry to obtain silica-alumina fine powder and filtrate A;

g. Gravity separation: Iron-containing material A and iron-containing material B are subjected to tantalum-niobium gravity separation to obtain tantalum-niobium rough concentrate and gravity separation tailings;

h. Weak magnetic separation: tantalum and niobium rough concentrate is removed by weak magnetic magnetic separation, concentrated and filtered to obtain tantalum and niobium concentrate and filtrate B, magnetic iron impurities and gravity separation tailings are combined, concentrated and filtered to obtain iron slag and filtrate C.

In a specific embodiment of the present invention, in step b, the fine-grained slurry is a material with a particle size ≤ 45 μm, and the coarse-grained slurry is a material with a particle size > 45 μm.

In a specific embodiment of the present invention, the grinding fineness of coarse-grained slurry wet grinding is -45 μm content≥90%.

In one embodiment of the present invention, in step c, the leaching temperature is 60-90° C., and the leaching time is 1-3 hours. In a preferred embodiment, in step c, the leaching temperature is 80-90° C., and the leaching time is 2-3 hours.

In a preferred embodiment, in step c, the leaching solution is returned to step a as a lithium-poor solution to replace part of the water, and the lithium-rich solution is obtained after cyclic leaching; the washing solution produced by washing in step d is returned to step a as a lithium-poor solution to replace part of the water ; Wherein, the Li ₂ O content in the lithium-poor solution is less than 5g/L; the Li ₂ O content in the lithium-rich solution is ≥ 5g/L; the number of cycles of leaching is preferably 2 to 4 times.

In a specific embodiment of the present invention, in step d, limestone, quicklime or slaked lime is used to adjust the pH value; the pulp concentration is adjusted to 28-32%.

In a specific embodiment of the present invention, in step e, the froth product is flotation, concentrated and filtered to obtain the gypsum product and filtrate A'.

In one embodiment of the present invention, in step f, the field strength of the medium magnetic separation is 0.2-0.6T, and the field strength of the strong magnetic separation is 1.0-1.7T. In a preferred embodiment, the field strength of the medium magnetic separation is 0.3-0.5T, and the field strength of the strong magnetic separation is 1.0-1.5T.

In one embodiment of the present invention, in step g, re-selection adopts spiral chute+shaking table mineral processing, or felt machine+shaking table mineral processing, or centrifugal mineral processing machine+shaking table mineral processing.

In one embodiment of the present invention, in step h, the field strength of the weak magnetic separation is 0.1-0.2T; preferably, the field strength of the weak magnetic separation is 0.12-0.16T.

In one embodiment of the present invention, the filtration of steps f and h is preceded by concentration.

In one embodiment of the present invention, the filtrate A' is returned to the backwater pool for recycling after water treatment, and the filtrate A, filtrate B and filtrate C are directly returned to the backwater pool for recycling; the backwater of the backwater pool is returned to Pulping, washing, pulping, flotation, magnetic separation, and re-selection links are recycled; preferably, the water treatment method includes at least one of sedimentation, adsorption, and activated sludge treatment.

Compared with the prior art, the present invention has the following beneficial effects:

1. The present invention can recycle lithium in lithium slag, so that Li ₂ O in lithium slag is reduced to below 0.15%, and the lithium leaching rate is ≥ 65%, which can significantly improve the comprehensive utilization rate of lithium resources, and the cost of lithium recovery is low .

2. The present invention can recycle tantalum and niobium in lithium slag to obtain tantalum-niobium concentrate with (TaNb) ₂ O ₅ content ≥ 30%. The tantalum and niobium in the slag are eliminated, and the green and efficient recycling of the valuable metal tantalum and niobium in the lithium slag is realized.

3. After the present invention desulfurizes and removes iron from lithium slag, a silica-alumina fine powder product used in the glass fiber industry and an additional product gypsum can be obtained. The Fe ₂ O ₃ of the silica-alumina fine powder is ≤0.4%, and SO ₃ ≤ 0.3%, to meet the requirements of high-quality silica-alumina fine powder that can be used for glass fiber; gypsum SO ₃ grade ≥ 40%, can be used in the construction industry.

4. The present invention separates the treatment of the filtered water of the flotation system from the water of the magnetic separation system, wherein the gypsum filtered liquid is treated separately, while the water of the magnetic separation system can be circulated without treatment, which has the advantage of reducing the amount of water treatment and saving production costs.

5. This patent "eats dry and squeezes out" lithium slag, no waste residue waste water is generated, fundamentally solves the problem of lithium slag consumption, realizes comprehensive utilization of lithium slag resources, and turns waste into treasure.

Description of drawings

Fig. 1 is a process flow chart of the comprehensive utilization method of spodumene smelting slag resources in Embodiment 1-3 of the present invention.

Fig. 2 is the XRD spectrum of the spodumene smelting slag in Example 1 of the present invention.

Fig. 3 is a scanning electron microscope spectrum of the spodumene smelting slag in Example 1 of the present invention, showing that a small amount of spodumene is wrapped by HAlSi ₂ O ₆ (reaction product of acid and β-spodumene).

4 is a scanning electron microscope spectrum of the spodumene smelting slag in Example 1 of the present invention, showing that Columbite, HAlSi ₂ O ₆ , and Orthoclase are co-grown.

Detailed ways

The comprehensive utilization method of spodumene smelting slag of the present invention comprises the following steps:

a. Pulping: Spodumene slag and water are mixed evenly to prepare lithium slag slurry with a solid-liquid mass ratio of 1:1-3;

B. Grinding: The lithium slag slurry in step a is classified into particle size, divided into fine-grained slurry and coarse-grained slurry, and the coarse-grained slurry is wet-grinded. After grinding, it is merged with the fine-grained slurry to obtain fine slurry;

c. Leaching: add sulfuric acid to the fine slurry, adjust the pH to 1-1.5, heat and stir for leaching; then separate the solid and liquid to obtain acidic slag and leachate; the leachate is lithium-poor liquid;

d. Slurry: After washing the acid slag, add water and limestone to mix, and adjust the pH to 6-7, and adjust the pulp concentration to 25-35wt%;

e. Flotation: Flotation desulfurization is carried out on the slurry after step d slurry adjustment to obtain desulfurization lithium slag and flotation foam products;

f. Magnetic separation: Desulfurization lithium slag adopts medium magnetic separation to obtain iron-containing material A and medium magnetic concentrate; medium magnetic concentrate adopts strong magnetic separation to obtain iron-containing material B and strong magnetic concentrate; strong magnetic concentrate Filter and dry to obtain silica-alumina fine powder and filtrate A;

g. Gravity separation: Iron-containing material A and iron-containing material B are subjected to tantalum-niobium gravity separation to obtain tantalum-niobium rough concentrate and gravity separation tailings;

h. Weak magnetic separation: the rough concentrate of tantalum and niobium is separated by weak magnetic separation to remove magnetic iron impurities, filtered to obtain tantalum and niobium concentrate and filtrate B, magnetic iron impurities and gravity separation tailings are combined, filtered to obtain iron slag and filtered Liquid C.

The method of the invention can recycle and utilize the lithium slag to the greatest extent without generating waste residue wastewater, fundamentally solves the problem of lithium slag consumption, realizes the comprehensive utilization of lithium slag resources, and turns waste into treasure.

Wherein, step a is pulping, stirring and mixing the spodumene smelting slag with water to prepare lithium slag slurry with a solid-to-liquid mass ratio of 1:1-3.

Step b is ore grinding. The lithium slag slurry prepared in step a is subjected to particle size classification and divided into fine-grained slurry and coarse-grained slurry. The materials were combined to obtain a fine slurry.

In a specific embodiment of the present invention, the particle size of the particle size classification is 45 μm, that is, the fine-grained slurry is a material with a particle size ≤ 45 μm, and the coarse-grained slurry is a material with a particle size > 45 μm.

In a specific embodiment of the present invention, the grinding fineness of coarse-grained slurry wet grinding is -45 μm content≥90%.

Step c is leaching, using sulfuric acid leaching method, adding sulfuric acid to the fine slurry, adjusting the pH to 1-1.5, heating and stirring for leaching, and then separating solid and liquid to obtain acidic residue and leachate.

_Li2O content of 0.3-0.6% remains in the spodumene smelting slag, and the content of β-lithium accounts for 70-80% after testing. Alpha Lithium with incomplete crystal transformation. Therefore, the present invention adopts the method of grinding + heating and acid leaching to recover the residual β lithium in the lithium slag.

The leaching temperature and time commonly used in the art are applicable to the present invention. In one embodiment of the present invention, the leaching temperature is 60-90° C., and the leaching time is 1-3 hours. In a preferred embodiment, the leaching temperature is 80-90° C., and the leaching time is 2-3 hours. At this time, the leaching rate of lithium is ≥65%.

The solid-liquid separation in step c can adopt common methods in the art, including but not limited to filtration or centrifugation. The solid obtained by solid-liquid separation is acid slag, and the liquid is leachate.

Since the Li ₂ O content in the spodumene smelting slag is not high, at this time, the leaching solution is a lithium-poor solution, and its Li ₂ O content is less than 5g/L. In order to enrich lithium, in a preferred embodiment, the lithium-poor solution is returned to step a to replace part of the water, and the lithium-rich solution is obtained after cyclic leaching. Preferably, the number of cycles of leaching is 2 to 4 times. At this time, the Li ₂ O content in the obtained lithium-rich solution is ≥5 g/L. The lithium-rich liquid can be returned to the leaching section of the lithium salt plant to produce lithium carbonate or lithium hydroxide products.

Step d is pulping, after washing the acid slag, mixing it with water, adjusting the pH to 6-7, and adjusting the pulp concentration to 25-35 wt%.

In order to improve the recycling rate of lithium and recycle water resources at the same time, in step d, the washing liquid generated by washing is returned to step a as a lithium-poor liquid to replace part of the water.

After the acid slag is washed, there is still a small amount of acid remaining, and alkali needs to be added to adjust the pH value. In a specific embodiment of the present invention, in step d, limestone, quicklime or slaked lime is used to adjust the pH value.

In a preferred embodiment, in step d, the pulp concentration is adjusted to be 28-32%.

Step e is flotation, and the slurry after slurry adjustment in step d is used as a feed material for flotation to carry out desulfurization by lithium slag flotation to obtain desulfurized lithium slag and flotation foam products.

Flotation desulfurization is to separate gypsum from the slag by flotation. Specifically, add flotation gypsum collector agent and regulator to the flotation tank and inflate to generate foam. Separation and extraction of gypsum; _SO3 content in the desulfurization lithium slag is ≤0.3%; the flotation foam product is slag gypsum, and in the flotation process, the foam of rough selection and sweeping is selected to obtain higher purity Gypsum foam product; the gypsum collector is an anionic collector, including but not limited to sodium methyl cocoyl taurate, sodium lauroyl amate, sodium coconut oil fatty acid alanine, cocoyl amphoteric Sodium Acetate, Lauroyl Glycine, Sodium Cocoyl Glycinate, Sodium Cocoyl Alaninate, Sodium Etheramide Acetate, Sodium Dodecylbenzene Sulfonate, Sodium Lauryl Sulfate, Laureth Phosphate, Monodecyl One or more of dialkyl phosphate potassium, polyoxyethylene monoalkyl phosphate, and the regulator includes but not limited to at least one of water glass, sodium hexametaphosphate, and CMC.

In one embodiment of the present invention, the flotation froth product is concentrated and filtered to obtain the gypsum product and the filtrate A'.

The gypsum filtered water contains gypsum collecting agent, and direct recycling will have a great impact on the desulfurization index of lithium slag. Therefore, the filtrate A' needs to be treated and can be returned to the pool for recycling. Conventional water treatment methods can be used, including but not limited to at least one of sedimentation, adsorption, and activated sludge treatment.

Step f is magnetic separation. The desulfurized lithium slag adopts medium magnetic separation to obtain iron-containing material A and medium magnetic concentrate; medium magnetic concentrate adopts strong magnetic separation to further reduce the iron content in the concentrate to obtain weak magnetic iron-containing Material B and strong magnetic concentrate; strong magnetic concentrate is filtered and dried to obtain silica-aluminum fine powder and filtrate A.

The medium magnetic concentrate is the material after the slag has been subjected to wet medium magnetic separation to remove iron. In one embodiment of the present invention, the field strength of the medium magnetic separation is 0.2-0.6T. Preferably, the field strength of the medium magnetic separation is 0.3-0.5T.

Strong magnetic concentrate is the material after the slag is removed by wet strong magnetic separation. In one embodiment of the present invention, the field strength of the strong magnetic separation is 1.0-1.7T. Preferably, the field strength of the strong magnetic separation is 1.0-1.5T.

After the strong magnetic concentrate is filtered and dried, silicon-aluminum fine powder can be obtained, wherein, the content of Fe ₂ O ₃ in the silicon-aluminum fine powder is ≤0.4%, and SO ₃ ≤0.3%. The filtrate A is directly returned to the return pool for recycling.

Preferably, the strong magnetic concentrate is concentrated before being filtered.

The g step is re-election, and the iron-containing material A and the iron-containing material B are subjected to tantalum-niobium re-election to obtain tantalum-niobium rough concentrate and gravity-election tailings. Among them, the material with high density after gravity separation is tantalum niobium rough concentrate, and the material with low density after gravity separation is gravity separation tailings. The iron-containing material A and the iron-containing material B can be combined for re-selection of tantalum and niobium, or the iron-containing material A and the iron-containing material B can be separately re-selected for tantalum and niobium.

In one embodiment of the present invention, re-selection adopts spiral chute + shaker beneficiation, or felt machine + shaker beneficiation, or centrifugal separator + shaker beneficiation.

Step h is weak magnetic separation. The tantalum and niobium rough concentrate is removed by weak magnetic separation and then filtered to obtain the tantalum and niobium concentrate and filtrate B. Preferably, concentration is performed prior to filtration.

In one embodiment of the present invention, the field strength of the weak magnetic separation is 0.1-0.2T. Preferably, the field strength of the weak magnetic separation is 0.12-0.16T.

Weak magnetic separation can remove magnetic iron impurities such as magnetite and iron filings to obtain tantalum-niobium concentrate. The (TaNb) ₂ O ₅ content of the tantalum-niobium concentrate of the present invention is ≥30%. Magnetic iron impurities can be used in cement building materials industry.

The magnetic iron impurities are combined with gravity separation tailings, and filtered to obtain iron slag and filtrate C. Preferably, concentration is performed prior to filtration.

In one embodiment of the present invention, the filtrate A, filtrate B and filtrate C are directly returned to the backwater pool for recycling; the backwater in the backwater pool is returned to the steps of pulping, washing, pulping, flotation, magnetic separation, and gravity separation recycle. In this way, water resources can be saved, zero discharge of industrial wastewater can be realized, water cost can be saved, and the environment is friendly.

The specific implementation of the present invention will be further described below in conjunction with the examples, and the present invention is not limited to the scope of the examples.

The lithium slags of Examples 1-3 come from a certain spodumene lithium extraction smelting slag in Sichuan, Jiangsu, and Jiangxi respectively, and their main chemical components are shown in Table 1.

Table 1 Lithium smelting slag chemical composition/%

Spodumene smelting slag SO _{3 Fe2O3 _} Al ₂ O ₃ SiO ₂ Li ₂ O Ta ₂ O ₅ Nb ₂ O ₅ Example 1 4.43 1.16 20.10 54.60 0.39 0.0078 0.0063 Example 2 5.39 0.89 21.33 52.58 0.53 0.0085 0.0071 Example 3 6.87 1.08 19.87 53.26 0.56 0.0098 0.0076

Example 1

As shown in Figure 1, a method for comprehensive utilization of spodumene smelting slag resources. The spodumene smelting slag comes from Sichuan, and its composition is shown in Table 1. The specific implementation steps are as follows:

1. Add water to the spodumene smelting slag and stir it to make a slurry with a solid-to-liquid ratio of 1:1; the slurry is classified by a cyclone, and the coarse-grained slurry above 45 μm enters a ball mill for grinding. The medium adopts ceramic medium, and it is ground to a fineness of -45 μm, accounting for 92%; the ground slurry is combined with the fine-grained slurry below 45 μm classified by the cyclone as the leaching raw material.

2. Put the leaching raw materials into the reaction kettle or the reaction tank, and at the same time add concentrated acid to adjust the pH to 1.0, heat to 80°C, and stir for leaching for 2 hours.

3. After the leaching is completed, use a centrifuge for solid-liquid separation to obtain leaching slag and lithium-poor leaching solution A; wash the leaching slag to obtain lithium-poor leaching solution B, combine lithium-poor leaching solution A and lithium-poor leaching solution B and return to stirring and pulping The process is to carry out cyclic leaching; the lithium-rich solution obtained after 2 times of cyclic leaching, the lithium-rich solution returns to the leaching section of the lithium carbonate factory to continue producing lithium carbonate products. The lithium content and lithium leaching rate or yield in the lithium-rich solution are shown in Table 2.

4. After washing the leached slag, add limestone to neutralize the residual acid in the leached slag, adjust the pH of the pulp to 6-7, and adjust the concentration of the pulp to 32%, then add gypsum collectors and regulators, of which the collector Select 30-50 parts of sodium lauroyl amate and 10-20 parts of sodium etheramine acetate, water glass is used as the regulator, the amount of gypsum collector is 400g/t, and the amount of water glass is 3000g/t. After one rough selection, After 3 times of sweeping and 2 times of selected flotation desulfurization process, the foam product and desulfurized lithium slag slurry are obtained. The foam product is concentrated in the concentration pool and filtered by the filter to obtain the gypsum product. After the filtered water is treated by chemical flocculation and sedimentation Return to the pool for recycling. The chemical composition of the gypsum product is shown in Table 3.

5. After flotation and desulfurization, the lithium slag slurry is magnetically separated in the magnetic machine with a magnetic field strength of 0.3T in the drum and a high-gradient strong magnetic machine with a magnetic field strength of 1.5T to obtain de-ironized slag material and iron-containing material A and Iron-containing material B.

6. Iron-containing material A and iron-containing material B are combined to obtain tantalum-niobium rough concentrate and iron slag A after being re-selected by spiral chute + shaking table; the tantalum-niobium rough concentrate is removed by a weak magnetic separator with a magnetic field strength of 0.12T After impurities, tantalum-niobium concentrate and iron slag B are obtained. The tantalum-niobium concentrate is concentrated and filtered in the concentration pool to obtain tantalum-niobium concentrate products, and the filtered water is recycled for use after returning water; iron slag A and iron slag B are combined and concentrated The pool is concentrated and filtered to obtain iron slag. The tantalum and niobium concentrate is concentrated and filtered in the concentration pool to obtain tantalum and niobium concentrate products. The filtered water enters the backwater pool for recycling. The product grade and recovery rate of tantalum-niobium concentrate are shown in Table 4.

7. After the iron removal slag is concentrated in the concentration pool and filtered by the filter, it is dried to obtain the silica-aluminum fine powder product, and the filtered water enters the return pool for recycling. The chemical composition of silica-alumina fine powder products is shown in Table 5.

Example 2

As shown in Figure 1, a comprehensive utilization method of spodumene smelting slag resources, the spodumene smelting slag is from Jiangsu, and its composition is shown in Table 1. The specific implementation steps are as follows:

1. Add water to the spodumene smelting slag and stir it to make a slurry with a solid-to-liquid ratio of 1:2; the slurry is classified by a cyclone, and the coarse-grained slurry above 45 μm enters a ball mill for grinding. The medium adopts ceramic medium, and it is ground to a fineness of -45 μm, accounting for 90%; the ground slurry is combined with the fine-grained slurry below 45 μm classified by the cyclone as the leaching raw material.

2. Put the leaching raw materials into the reaction kettle or reaction tank, add concentrated acid to adjust the pH to 1.2, heat to 85°C, and stir for leaching for 2.5 hours.

3. After the leaching is completed, use filtration for solid-liquid separation to obtain leaching slag and lithium-poor leaching solution A; wash the leaching slag to obtain lithium-poor leaching solution B, lithium-poor leaching solution A and lithium-poor leaching solution B are combined and returned to the stirring and pulping process , carry out cyclic leaching; the lithium-rich solution obtained after 3 times of cyclic leaching, the lithium-rich solution returns to the leaching section of the lithium carbonate factory to continue producing lithium carbonate products. The lithium content and lithium leaching rate or yield in the lithium-rich solution are shown in Table 2.

4. After washing the leached slag, add quicklime to neutralize the residual acid in the leached slag, adjust the pH of the pulp to 6-7, and adjust the concentration of the pulp to 28%, then add gypsum collector and regulator respectively. 50-70 parts of coconut oil fatty acid sodium alanine, 2-5 parts of sodium dodecylbenzene sulfonate, 5-10 parts of lauryl ether phosphate, the regulator uses sodium hexametaphosphate, and the dosage of gypsum collector is 600g /t, the amount of hexametaphosphoric acid is 1000g/t, after 1 roughing, 3 times of sweeping and 2 times of selective flotation desulfurization process, the foam product and desulfurized lithium slag slurry are obtained, and the foam product passes through the concentration pool Concentrate and filter to obtain gypsum products, and the filtered water is treated by activated carbon adsorption and returned to the pool for recycling. The chemical composition of the gypsum product is shown in Table 3.

5. After flotation and desulfurization, the lithium slag slurry is magnetically separated in the magnetic machine with a magnetic field strength of 0.4T in the drum and a high-gradient strong magnetic machine with a magnetic field strength of 1.0T to obtain deironing slag material and iron-containing material A and Iron-containing material B.

6. Iron-containing material A and iron-containing material B are combined to obtain tantalum-niobium rough concentrate and iron slag A after being re-selected by blanket machine + shaking table; the tantalum-niobium rough concentrate is removed by a weak magnetic separator with a magnetic field strength of 0.15T After impurities, tantalum-niobium concentrate and iron slag B are obtained; iron slag A and iron slag B are combined to obtain iron slag after being concentrated and filtered in the concentration pool; The water enters the return pool for recycling. The product grade and recovery rate of tantalum and niobium concentrate are shown in Table 4.

7. After the iron removal slag is concentrated in the concentration pool and filtered by the filter, it is dried to obtain the silica-aluminum fine powder product, and the filtered water enters the return pool for recycling. The chemical composition of silica-alumina fine powder products is shown in Table 5.

Example 3

As shown in Figure 1, a comprehensive utilization method of spodumene smelting slag resources, the spodumene smelting slag comes from Jiangxi, its composition is shown in Table 1. The specific implementation steps are as follows:

1. Add water to the spodumene smelting slag and stir it to make a slurry with a solid-to-liquid ratio of 1:3; the slurry is classified by a cyclone, and the coarse-grained slurry above 45 μm enters a ball mill for grinding. The medium adopts ceramic medium, which is ground to a fineness of -45 μm, accounting for 95%; the ground slurry is combined with the fine-grained slurry below 45 μm classified by the cyclone as the leaching raw material.

2. Put the leaching raw materials into the reaction kettle or the reaction tank, and at the same time add concentrated acid to adjust the pH to 1.5, heat to 90°C, and stir for leaching for 3 hours.

3. After the leaching is completed, use filtration for solid-liquid separation to obtain leaching slag and lithium-poor leaching solution A; wash the leaching slag to obtain lithium-poor leaching solution B, lithium-poor leaching solution A and lithium-poor leaching solution B are combined and returned to the stirring and pulping process , carry out cyclic leaching; the lithium-rich solution obtained after 4 times of cyclic leaching, the lithium-rich solution returns to the leaching section of the lithium carbonate factory to continue producing lithium carbonate products. The lithium content and lithium leaching rate or yield in the lithium-rich solution are shown in Table 2.

4. After washing the leached slag, add limestone to neutralize the residual acid in the leached slag, adjust the pH of the pulp to 6-7, and adjust the concentration of the pulp to 32%, then add gypsum collector and regulator respectively. 60-80 parts of sodium cocoyl alanine, 5-10 parts of sodium lauryl sulfate, CMC as regulator, 550g/t of gypsum collector, 200g/t of CMC, after 1 rough selection , 3 times of sweeping and 2 times of selected flotation desulfurization processes to obtain foam products and desulfurized lithium slag slurry. The foam products are concentrated in the concentration pool and filtered by the filter to obtain gypsum products. The filtered water is passed through the activated sludge process. After treatment, return to the pool for recycling. The chemical composition of the gypsum product is shown in Table 3.

5. After flotation and desulfurization, the lithium slag slurry is magnetically separated in the magnetic machine with a magnetic field strength of 0.5T in the drum and a high-gradient strong magnetic machine with a magnetic field strength of 1.3T to obtain de-ironized slag material and iron-containing material A and Iron-containing material B.

6. Iron-containing material A and iron-containing material B are combined with centrifugal concentrator + shaking table gravity separation to obtain tantalum-niobium rough concentrate and iron slag A; tantalum-niobium rough concentrate is removed by a weak magnetic separator with a magnetic field strength of 0.16T to remove magnetic iron After impurities, tantalum-niobium concentrate and iron slag B are obtained; iron slag A and iron slag B are combined and concentrated in the concentration pool, filtered by a filter to obtain iron slag, and tantalum-niobium concentrate is concentrated and filtered in the concentration pool to obtain tantalum-niobium concentrate products , filtered water enters the backwater pool for recycling. The product grade and recovery rate of tantalum-niobium concentrate are shown in Table 4.

7. After the iron removal slag is concentrated in the concentration pool and filtered by the filter, it is dried to obtain the silica-aluminum fine powder product, and the filtered water enters the return pool for recycling. The chemical composition of silica-alumina fine powder products is shown in Table 5.

Table 2 Example 1～3 Lithium slag acid leaching index

Table 3 Chemical composition/% of the gypsum product obtained in Examples 1 to 3

Implementation item SO _{3 Fe2O3 _} Example 1 40.82 0.26 Example 2 41.03 0.21 Example 3 42.76 0.18

The tantalum-niobium concentrate product index/% that table 4 embodiment 1～3 gains

Chemical composition/% of the silicon-aluminum fine powder product that table 5 embodiment 1～3 gains

Silica-alumina fine powder SO _{3 Fe2O3 _} Al ₂ O ₃ SiO ₂ Li ₂ O Example 1 0.28 0.39 24.54 67.32 0.11 Example 2 0.27 0.36 23.88 68.61 0.10 Example 3 0.25 0.38 24.76 67.17 0.11

It can be seen that by adopting the method of the present invention, elements such as lithium, tantalum, niobium, silicon, and aluminum in the lithium slag can be comprehensively recycled, and no waste water and waste residue are produced, which can solve the problem of lithium slag consumption in the future, reduce environmental pollution, and realize lithium slag resources. comprehensive utilization.

Claims (10)

1. The method for resource utilization of spodumene smelting slag is characterized in that it comprises the following steps: a. Pulping: Spodumene slag and water are mixed evenly to prepare lithium slag slurry with a solid-liquid mass ratio of 1:1-3; B. Grinding: The lithium slag slurry in step a is classified into particle size, divided into fine-grained slurry and coarse-grained slurry, and the coarse-grained slurry is wet-grinded. After grinding, it is merged with the fine-grained slurry to obtain fine slurry; c. Leaching: add sulfuric acid to the fine slurry, adjust the pH to 1-1.5, heat and stir for leaching; then separate the solid and liquid to obtain acidic slag and leachate; the leachate is lithium-poor liquid; d. Pulping: After washing the acid slag, add water or return water to stir and make pulp, adjust the concentration of the pulp to 25-35wt%, and adjust the pH of the pulp to 6-7 by adding limestone, quicklime or slaked lime; e. Flotation: Flotation desulfurization is carried out on the slurry after step d slurry adjustment to obtain desulfurization lithium slag and flotation foam products; f. Magnetic separation: Desulfurization lithium slag adopts medium magnetic separation to obtain iron-containing material A and medium magnetic concentrate; medium magnetic concentrate adopts strong magnetic separation to obtain iron-containing material B and strong magnetic concentrate; strong magnetic concentrate Filter and dry to obtain silica-alumina fine powder and filtrate A; g. Gravity separation: Iron-containing material A and iron-containing material B are subjected to tantalum-niobium gravity separation to obtain tantalum-niobium rough concentrate and gravity separation tailings; h. Weak magnetic separation: tantalum and niobium rough concentrate is removed by weak magnetic magnetic separation, concentrated and filtered to obtain tantalum and niobium concentrate and filtrate B, magnetic iron impurities and gravity separation tailings are combined, concentrated and filtered to obtain iron slag and filtrate C. 2. The comprehensive utilization method of spodumene smelting slag resources according to claim 1, characterized in that: in step b, the fine-grained slurry is a material with a particle size ≤ 45 μm, and the coarse-grained slurry is a particle Materials with a diameter > 45 μm; preferably the grinding fineness of wet grinding is -45 μm and the content is ≥ 90%. 3. The method for comprehensive utilization of spodumene smelting slag resources according to claim 1, characterized in that: in step c, the leaching temperature is 60-90°C, and the leaching time is 1-3h; preferably the leaching temperature is 80-90°C ℃, the leaching time is 2~3h. 4. The comprehensive utilization method of spodumene smelting slag resources according to claim 1, characterized in that: in step c, the leaching solution is returned to step a as a lithium-poor solution to replace part of the water, and lithium-rich solution is obtained after cyclic leaching; step d The washing liquid produced by middle washing is returned to step a as a lithium-poor liquid to replace part of the water; wherein, the Li ₂ O content in the lithium-poor liquid is less than 5g/L; the Li ₂ O content in the lithium-rich liquid is ≥ 5g/L; the preferred cycle The number of times of leaching is 2 to 4 times. 5. The comprehensive utilization method of spodumene smelting slag as a resource according to claim 1, characterized in that: in step d, limestone, quicklime or slaked lime is used to adjust the pH value; the slurry concentration is adjusted to 28-32%. 6. The comprehensive utilization method of spodumene smelting slag resources according to claim 1, characterized in that: in step e, the foam product is flotation, concentrated and filtered to obtain the gypsum product and the filtrate A'. 7. The comprehensive utilization method of spodumene smelting slag resources according to claim 1, characterized in that: in step f, the field strength of medium magnetic separation is 0.2-0.6T, and the field strength of strong magnetic separation is 0.2-0.6T. It is 1.0-1.7T; preferably, the field strength of medium magnetic separation is 0.3-0.5T, and the field strength of strong magnetic separation is 1.0-1.5T. 8. The comprehensive utilization method of spodumene smelting slag resources according to claim 1, characterized in that: in step g, re-selection is using spiral chute+shaking table mineral processing, or blanket machine+shaking table mineral processing, or centrifugal mineral processing Machine + shaker beneficiation; In step h, the field strength of the weak magnetic separation is 0.1-0.2T; preferably, the field strength of the weak magnetic separation is 0.12-0.16T. 9. The method for comprehensive utilization of spodumene smelting slag resources according to claim 1, characterized in that: the f step and h step are first concentrated before filtering. 10. The comprehensive utilization method of spodumene smelting slag resources according to claim 1, characterized in that: the filtrate A' is returned to the backwater pool for recycling after water treatment, and the filtrate A, filtrate B and filtrate C directly return to the backwater pool for recycling; the backwater in the backwater pool is returned to the pulping, washing, pulping, flotation, magnetic separation, and re-election links for recycling; the preferred water treatment methods include sedimentation, adsorption, and activated sludge treatment. at least one.