Method for comprehensive utilization of lithium-aluminum-silicate smelting slag

Abstract

The present application relates to a lithium spodumene smelting slag resource comprehensive utilization method, and belongs to the technical field of solid waste resource processing of ore lithium extraction.The technical problem solved by the present application is to provide a lithium spodumene smelting slag resource comprehensive utilization method.The method comprises pulp preparation, ore grinding, leaching, pulp conditioning, flotation, magnetic separation, gravity separation and low-intensity magnetic separation.The present application can recycle lithium, tantalum and niobium in lithium slag, and the process is simple, efficient and environmentally friendly, significantly improves the comprehensive utilization rate of lithium resources, realizes iron removal of slag, recovers tantalum, niobium and lithium in the slag, and can obtain a silicon-aluminum fine powder product for the glass fiber industry and an additional product gypsum.In addition, the present application can reduce water treatment capacity and save production cost;the lithium slag can be fully utilized, no waste slag and waste water are generated, the problem of lithium slag disposal is fundamentally solved, the lithium slag is comprehensively utilized, and waste is turned into treasure.

Description

Technical Field

[0001] This invention relates to a method for the comprehensive utilization of lithium spodumene smelting slag, belonging to the field of solid waste resource utilization technology in lithium extraction from ore. Background Technology

[0002] Spodumene smelting slag is a solid waste product from lithium extraction from ore. Spodumene extraction is a relatively mature process for lithium extraction from ore. This method first roasts natural spodumene at 950–1100℃, transforming it from monoclinic α-spodumene to tetragonal β-spodumene. Due to this crystal transformation, the physicochemical properties of the mineral change significantly, with increased chemical activity, enabling it to react with various acids and alkalis. During spodumene extraction, some lithium inevitably remains in the lithium slag. Therefore, fully exploring the secondary recycling value of lithium from solid waste resources is of great significance.

[0003] The main chemical components of spodumene smelting slag are SiO2 and Al2O3, mainly aluminosilicates and quartz, and also contain gypsum and small amounts of spodumene and iron minerals. According to the test, the content of SO3 in the lithium slag raw material is 3-8%, Fe2O3 content is 0.8-1.2%, Li2O content is 0.3-0.6%, and (TaNb)2O5 content is 120-180 ppm. Among them, the valuable metals lithium, tantalum and niobium have comprehensive recycling value.

[0004] In the process of producing lithium salts using spodumene, approximately 8 to 10 tons of lithium slag are generated for every ton of lithium salt produced. This level of discharge results in a large amount of lithium slag, which not only wastes land resources due to stockpiling but also leads to the loss of alkaline and acidic slag water if not properly stored, harming farmland and polluting the environment. Currently, the comprehensive utilization of lithium slag is mainly applied in the cement and building materials industry, resulting in low added value. Small amounts of lithium, tantalum, niobium metals, and gypsum are not recycled. Furthermore, with the explosive growth in lithium slag volume, the cement and building materials industry's capacity for lithium slag disposal is nearing saturation. Therefore, the issue of lithium slag disposal will become an urgent problem to be solved in the future.

[0005] Chinese patent CN110015855A discloses a method for treating lithium slag. This method involves sulfuric acid leaching of lithium slag from lepidolite, achieving leaching rates of over 88% for lithium, rubidium, cesium, potassium, aluminum, and sodium. The resulting acid-leached residue primarily consists of quartz and gypsum, which can then be reused as concrete admixtures. While this method recovers valuable elements like lithium from the slag, the recovery cost is relatively high, and it easily generates more solid slag or waste liquid, posing certain environmental risks. Furthermore, it does not substantially solve the problem of lithium slag disposal; the vast majority of lithium slag can only be used as inexpensive cement admixtures.

[0006] The invention patent with publication number CN114702048A discloses a process for the resource-based recycling of lithium slag solid waste. This process involves acid-base reactions of the lithium slag to obtain products such as potassium sulfate, sodium sulfate, lithium carbonate, cesium carbonate, and rubidium carbonate. However, this process is complex, costly, and requires the introduction of hydrofluoric acid, which can easily cause environmental pollution. Furthermore, most of the acid-insoluble solids in the lithium slag are only used as building materials, failing to achieve true resource-based disposal of the lithium slag.

[0007] The invention patent with publication number CN113621811A discloses a method for recovering tantalum and niobium from spodumene slag. The prerequisite for this technology is that a small amount of waste acid needs to be added to make the pH of the slag slurry 4 to 5, which can easily cause corrosion to the equipment and poses certain environmental risks.

[0008] The invention patent with publication number CN114226413A discloses a comprehensive lithium slag treatment process, which includes processes such as grinding, magnetic separation, flotation, and alkali conversion to obtain silicon-aluminum micro powder. Generally, spodumene smelting slag has a fine particle size. This process does not perform graded grinding of the lithium slag and uses the addition of sodium carbonate or potassium carbonate for alkali conversion to reduce the sulfur in the micro powder, resulting in high grinding and desulfurization costs. Although this process can also obtain silicon-aluminum micro powder for glass fiber, it does not reasonably recover valuable lithium metal and gypsum from the lithium slag, making it difficult to truly achieve the resource-based disposal of lithium slag.

[0009] The invention patent with publication number CN113976309A discloses a method for the comprehensive recovery of lithium, tantalum-niobium, silica-alumina powder, iron concentrate, and gypsum from lithium slag. The method involves gravity separation of the lithium slag followed by weak magnetic separation to obtain concentrate 1 and tailings 1. Concentrate 1 is then separated using weak magnetic separation to obtain coarse-grained tantalum-niobium rich material and coarse-grained iron concentrate. Tailings 1 is then floated to obtain gypsum and tailings 2. Tailings 2 is pulverized and then separated using weak magnetic separation to obtain fine-grained iron concentrate and tailings 3. Tailings 3 is then separated using strong magnetic separation to obtain concentrate 2 and tailings 4. Tailings 4 is dried to obtain silica-alumina powder. Concentrate 2 is gravity separated to obtain fine-grained tantalum-niobium concentrate and high-iron lithium-rich material, from which lithium is then recovered. However, this method only recovers a portion of the lithium from the relatively small-volume high-iron lithium-rich material. Li₂O in the silica-alumina powder product is not recovered, resulting in a total lithium recovery rate of only 20.5%, which is low and leads to a waste of lithium resources. Summary of the Invention

[0010] To address the above deficiencies, the technical problem solved by this invention is to provide a method for the comprehensive utilization of spodumene smelting slag resources.

[0011] The present invention provides a method for the comprehensive utilization of lithium spodumene smelting slag resources, comprising the following steps:

[0012] a. Pulping: Mix spodumene smelting slag and water to prepare a lithium slag slurry with a solid-liquid mass ratio of 1:1 to 3;

[0013] b. Grinding: The lithium slag slurry from step a is classified by particle size into fine slurry and coarse slurry. The coarse slurry is wet-ground and then combined with the fine slurry to obtain fine slurry.

[0014] c. Leaching: Sulfuric acid is added to the fine slurry to adjust the pH to 1-1.5, and the mixture is heated and stirred for leaching; then solid-liquid separation is performed to obtain acidic slag and leachate; the leachate is a lithium-poor solution;

[0015] d. Slurry preparation: After washing the acidic slag, add water or recycled water and stir to prepare slurry. Adjust the slurry concentration to 25-35 wt%, and adjust the pH of the slurry to 6-7 by adding limestone, quicklime or hydrated lime.

[0016] e. Flotation: The slurry after preparation in step d is subjected to flotation desulfurization to obtain desulfurized lithium slag and flotation foam products;

[0017] f. Magnetic separation: The desulfurized lithium slag is separated by medium magnetic separation to obtain iron-containing material A and medium magnetic concentrate; the medium magnetic concentrate is separated by strong magnetic separation to obtain iron-containing material B and strong magnetic concentrate; the strong magnetic concentrate is filtered and dried to obtain silicon-aluminum concentrate and filtrate A.

[0018] 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;

[0019] h. Weak magnetic separation: The tantalum-niobium rough concentrate is separated by weak magnetic separation to remove magnetic iron impurities. After concentration and filtration, tantalum-niobium concentrate and filtrate B are obtained. The magnetic iron impurities are combined with the gravity separation tailings and concentrated and filtered to obtain iron slag and filtrate C.

[0020] In one 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.

[0021] In one specific embodiment of the present invention, the grinding fineness of the coarse-grained slurry wet grinding is ≥90% -45μm content.

[0022] 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.

[0023] In a preferred embodiment, in step c, the leachate is returned to step a as a lithium-poor solution to replace part of the water, and after cyclic leaching, a lithium-rich solution is obtained; the washing liquid generated during washing in step d is returned to step a as a lithium-poor solution to replace part of the water; wherein, the Li2O content in the lithium-poor solution is <5g / L; the Li2O content in the lithium-rich solution is ≥5g / L; preferably, the number of cyclic leaching cycles is 2 to 4.

[0024] In one specific embodiment of the present invention, in step d, limestone, quicklime, or hydrated lime is used to adjust the pH value; the slurry concentration is adjusted to 28-32%.

[0025] In one specific embodiment of the present invention, in step e, the flotation foam product is concentrated and filtered to obtain gypsum product and filtrate A′.

[0026] In one embodiment of the present invention, in step f, the field strength of the medium magnetic separation is 0.2–0.6 T, and the field strength of the strong magnetic separation is 1.0–1.7 T. In a preferred embodiment, the field strength of the medium magnetic separation is 0.3–0.5 T, and the field strength of the strong magnetic separation is 1.0–1.5 T.

[0027] In one embodiment of the present invention, in step g, the re-selection is carried out by spiral chute + shaking table beneficiation, or blanket mill + shaking table beneficiation, or centrifugal concentrator + shaking table beneficiation.

[0028] In one embodiment of the present invention, in step h, the field strength of the weak magnetic separation is 0.1 to 0.2 T; preferably, the field strength of the weak magnetic separation is 0.12 to 0.16 T.

[0029] In one embodiment of the invention, concentration is performed before filtration in steps f and h.

[0030] In one embodiment of the present invention, filtrate A′ is returned to the return water tank for recycling after water treatment, and filtrate A, filtrate B and filtrate C are directly returned to the return water tank for recycling; the return water in the return water tank is returned to the pulping, washing, pulping, flotation, magnetic separation and gravity separation stages for recycling; preferably, the water treatment method includes at least one of sedimentation, adsorption and activated sludge treatment.

[0031] Compared with the prior art, the present invention has the following beneficial effects:

[0032] 1. This invention can recover and utilize lithium from lithium slag, reducing the Li2O content in the lithium slag to below 0.15% and achieving a lithium leaching rate of ≥65%, which can significantly improve the comprehensive utilization rate of lithium resources and reduce lithium recovery costs.

[0033] 2. This invention can recover tantalum and niobium from lithium slag to obtain tantalum and niobium concentrate with (TaNb)2O5 content ≥30%. The process is simple, efficient and environmentally friendly. It not only removes iron from the slag but also recovers tantalum and niobium from the slag, realizing the green and efficient recycling of valuable metals tantalum and niobium in lithium slag.

[0034] 3. After desulfurization and deironization of lithium slag, this invention can obtain a silicon-aluminum concentrate product for the glass fiber industry and a by-product gypsum. The silicon-aluminum concentrate has Fe2O3≤0.4% and SO3≤0.3%, which meets the requirements for high-quality silicon-aluminum concentrate that can be used in glass fiber; the gypsum has an SO3 content ≥40%, which can be used in the construction industry.

[0035] 4. This invention separates the water treatment of the flotation system filtration system from the water of the magnetic separation system. The gypsum filtration liquid is treated separately, while the water of the magnetic separation system can be circulated without treatment. Its advantage is that it reduces the amount of water to be treated and saves production costs.

[0036] 5. This patent makes full use of lithium slag, generating no waste residue or wastewater, fundamentally solving the problem of lithium slag disposal, realizing the comprehensive utilization of lithium slag resources, and turning waste into treasure. Attached Figure Description

[0037] Figure 1 This is a process flow diagram of the comprehensive utilization method of spodumene smelting slag resource in Examples 1-3 of the present invention.

[0038] Figure 2 The image shows the XRD pattern of the spodumene smelting slag in Example 1 of this invention.

[0039] Figure 3 The scanning electron microscope image of the spodumene smelting slag in Example 1 of the present invention shows that a small amount of spodumene is encapsulated by HAlSi2O6 (a product of the reaction between acid and β-spodumene).

[0040] Figure 4 The scanning electron microscope image of the spodumene smelting slag in Example 1 of the present invention shows that columbite is intergrown with HAlSi2O6 and orthoclase. Detailed Implementation

[0041] The present invention provides a method for the comprehensive utilization of lithium spodumene smelting slag resources, comprising the following steps:

[0042] a. Pulping: Mix spodumene smelting slag and water to prepare a lithium slag slurry with a solid-liquid mass ratio of 1:1 to 3;

[0043] b. Grinding: The lithium slag slurry from step a is classified by particle size into fine slurry and coarse slurry. The coarse slurry is wet-ground and then combined with the fine slurry to obtain fine slurry.

[0044] c. Leaching: Sulfuric acid is added to the fine slurry to adjust the pH to 1-1.5, and the mixture is heated and stirred for leaching; then solid-liquid separation is performed to obtain acidic slag and leachate; the leachate is a lithium-poor solution;

[0045] d. Slurry preparation: After washing the acidic slag, add water and limestone to mix, and adjust the pH to 6-7, and adjust the slurry concentration to 25-35 wt%.

[0046] e. Flotation: The slurry after preparation in step d is subjected to flotation desulfurization to obtain desulfurized lithium slag and flotation foam products;

[0047] f. Magnetic separation: The desulfurized lithium slag is separated by medium magnetic separation to obtain iron-containing material A and medium magnetic concentrate; the medium magnetic concentrate is separated by strong magnetic separation to obtain iron-containing material B and strong magnetic concentrate; the strong magnetic concentrate is filtered and dried to obtain silicon-aluminum concentrate and filtrate A.

[0048] 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;

[0049] h. Weak magnetic separation: The tantalum-niobium rough concentrate is separated by weak magnetic separation to remove magnetic iron impurities. After filtration, tantalum-niobium concentrate and filtrate B are obtained. The magnetic iron impurities are combined with the gravity separation tailings and filtered to obtain iron slag and filtrate C.

[0050] The method of this invention can maximize the recycling and utilization of lithium slag, without generating waste slag or wastewater, fundamentally solving the problem of lithium slag disposal, realizing the comprehensive utilization of lithium slag resources, and turning waste into treasure.

[0051] Step a is pulping, in which spodumene smelting slag is mixed with water to prepare a lithium slag slurry with a solid-liquid mass ratio of 1:1 to 3.

[0052] Step b is grinding. The lithium slag slurry prepared in step a is classified by particle size into fine-particle slurry and coarse-particle slurry. The coarse-particle slurry is wet-ground and then combined with the fine-particle slurry to obtain fine slurry.

[0053] In one specific embodiment of the present invention, the particle size for particle size classification is 45 μm, that is, fine-grained slurry is material with a particle size ≤ 45 μm, and coarse-grained slurry is material with a particle size > 45 μm.

[0054] In one specific embodiment of the present invention, the grinding fineness of the coarse-grained slurry wet grinding is ≥90% -45μm content.

[0055] Step c is leaching, which uses sulfuric acid leaching. Sulfuric acid is added to the fine slurry to adjust the pH to 1-1.5. The leaching is carried out by heating and stirring, followed by solid-liquid separation to obtain acidic slag and leachate.

[0056] Lithium spodumene smelting slag contains 0.3-0.6% residual Li₂O, of which β-lithium accounts for 70-80%. This portion of lithium consists of unreacted lithium from the ore leaching process, while the remaining lithium is α-lithium that has not undergone complete crystal transformation. Therefore, this invention employs a grinding + heated acid leaching method to recover residual β-lithium from the lithium slag.

[0057] The leaching temperatures and times commonly used in this art are applicable to this invention. In one embodiment of this 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. In this case, the lithium leaching rate is ≥65%.

[0058] The solid-liquid separation in step c can be performed using methods commonly used in the art, including but not limited to filtration or centrifugation. The solid obtained from the solid-liquid separation is an acidic slag, and the liquid is a leachate.

[0059] Because the Li₂O content in the spodumene smelting slag is low, the leachate is lithium-poor, with a Li₂O content < 5 g / L. To enrich lithium, in a preferred embodiment, this lithium-poor solution is returned to step a to replace part of the water, and after repeated leaching, a lithium-rich solution is obtained. Preferably, the leaching cycle is 2 to 4 times. At this point, the resulting lithium-rich solution has a Li₂O content ≥ 5 g / L. This lithium-rich solution can be returned to the leaching section of a lithium salt plant to produce lithium carbonate or lithium hydroxide products.

[0060] Step d is slurry preparation. After washing the acidic slag, it is mixed with water, and the pH is adjusted to 6-7, and the slurry concentration is adjusted to 25-35 wt%.

[0061] To improve the lithium recovery rate and recycle water resources, in step d, the washing liquid generated from washing is returned to step a as a lithium-poor solution to replace part of the water.

[0062] After washing, acidic slag still retains a small amount of acid, requiring the addition of alkali to adjust the pH value. In one specific embodiment of the present invention, in step d, limestone, quicklime, or hydrated lime is used to adjust the pH value.

[0063] In a preferred embodiment, in step d, the pulp concentration is adjusted to 28-32%.

[0064] Step e is flotation. The slurry prepared in step d is used as the feed for flotation to desulfurize lithium slag, resulting in desulfurized lithium slag and flotation foam products.

[0065] Flotation desulfurization involves separating gypsum from flotation slag. Specifically, a flotation gypsum collector and modifier are added to the flotation cell, and aeration is used to generate foam. After the gypsum reacts with the collector, it is adsorbed onto the foam and scraped off, thus achieving the separation and removal of gypsum. The SO3 content in the desulfurized lithium slag is ≤0.3%. The flotation foam product is slag gypsum. During the flotation process, the foam from the roughing and scavenging processes is further refined to obtain a higher purity gypsum foam product. The gypsum collector is an anionic collector, including but not limited to one or more of the following: sodium methyl cocoyl taurate, sodium lauroyl taurine, sodium coconut oil fatty acid alanine, sodium cocoamphoacetate, lauroyl glycine, sodium cocoyl glycine, sodium cocoyl alanine, sodium etheramine acetate, sodium dodecylbenzene sulfonate, sodium dodecyl sulfate, lauryl ether phosphate, potassium monododecyl phosphate, and polyoxyethylene monoalkyl phosphate. The modifier includes but is not limited to at least one of water glass, sodium hexametaphosphate, and CMC.

[0066] In one embodiment of the present invention, flotation foam product is concentrated and filtered to obtain gypsum product and filtrate A′.

[0067] The gypsum filtration water contains gypsum collecting agents, and direct recycling would significantly impact the desulfurization performance of lithium slag. Therefore, filtrate A′ needs to undergo water treatment before being entirely returned to the reclaimed water tank for reuse. Conventional water treatment methods can be used, including but not limited to at least one of sedimentation, adsorption, and activated sludge treatment.

[0068] Step f is magnetic separation. The desulfurized lithium slag is separated by medium magnetic separation to obtain iron-containing material A and medium magnetic concentrate. The medium magnetic concentrate is further separated by strong magnetic separation to reduce the iron content in the concentrate to obtain weakly magnetic iron-containing material B and strong magnetic concentrate. The strong magnetic concentrate is filtered and dried to obtain silicon-aluminum concentrate and filtrate A.

[0069] The medium magnetic concentrate is the material obtained after removing iron from slag through wet medium magnetic separation. In one embodiment of the present invention, the magnetic field strength of the medium magnetic separation is 0.2 to 0.6 T. Preferably, the magnetic field strength of the medium magnetic separation is 0.3 to 0.5 T.

[0070] The high-intensity magnetic concentrate is the material obtained after removing iron from slag through wet high-intensity magnetic separation. In one embodiment of the present invention, the magnetic field strength of the high-intensity magnetic separation is 1.0 to 1.7 T. Preferably, the magnetic field strength of the high-intensity magnetic separation is 1.0 to 1.5 T.

[0071] After filtration and drying, the strong magnetic concentrate yields a silica-alumina concentrate, wherein the Fe2O3 content of the silica-alumina concentrate is ≤0.4% and the SO3 content is ≤0.3%. The filtrate A is directly returned to the recycling pool for reuse.

[0072] Preferably, the strong magnetic concentrate is concentrated before filtration.

[0073] Step g is gravity separation. Iron-containing materials A and B undergo tantalum-niobium gravity separation to obtain tantalum-niobium rough concentrate and gravity separation tailings. The material with the higher density after gravity separation is the tantalum-niobium rough concentrate, and the material with the lower density is the gravity separation tailings. Iron-containing materials A and B can be combined for tantalum-niobium gravity separation, or they can be separated separately.

[0074] In one embodiment of the present invention, the re-selection is carried out by spiral chute + shaking table beneficiation, or blanket mill + shaking table beneficiation, or centrifugal concentrator + shaking table beneficiation.

[0075] Step h involves weak magnetic separation. The tantalum-niobium rough concentrate is separated using weak magnetic separation to remove magnetic iron impurities, then filtered to obtain tantalum-niobium concentrate and filtrate B. Preferably, concentration is performed before filtration.

[0076] In one embodiment of the present invention, the field strength of the weak magnetic separation is 0.1 to 0.2 T. Preferably, the field strength of the weak magnetic separation is 0.12 to 0.16 T.

[0077] Weak magnetic separation can remove magnetic iron impurities such as magnetite and iron filings to obtain tantalum-niobium concentrate. The tantalum-niobium concentrate of this invention has a (TaNb)₂O₅ content ≥30%. Magnetic iron impurities can be used in the cement and building materials industry.

[0078] Magnetic iron impurities are combined with gravity separation tailings and filtered to obtain iron slag and filtrate C. Preferably, concentration is performed before filtration.

[0079] In one embodiment of the present invention, filtrate A, filtrate B, and filtrate C are directly returned to the recycling tank for reuse; the recycled water from the recycling tank is then returned to the pulping, washing, pulp conditioning, flotation, magnetic separation, and gravity separation processes for reuse. This saves water resources, achieves zero discharge of industrial wastewater, reduces water costs, and is environmentally friendly.

[0080] The specific embodiments of the present invention will be further described below with reference to examples, but the present invention is not limited to the scope of the embodiments described herein.

[0081] The lithium slags in Examples 1-3 were respectively obtained from lithium spodumene extraction smelting slags in Sichuan, Jiangsu and Jiangxi provinces, and their main chemical components are shown in Table 1.

[0082] Table 1 Chemical composition of lithium smelting slag / %

[0083] spodumene smelting slag <![CDATA[SO3]]> <![CDATA[Fe2O3]]> <![CDATA[Al2O3]]> <![CDATA[SiO2]]> <![CDATA[Li2O]]> <![CDATA[Ta2O5]]> <![CDATA[Nb2O5]]> 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

[0084] Example 1

[0085] like Figure 1 As shown, a method for the comprehensive utilization of spodumene smelting slag is presented. The spodumene smelting slag originates from Sichuan Province, and its composition is shown in Table 1. The specific implementation steps are as follows:

[0086] 1. Water is added to slag for stirring and slurry preparation to form a slurry with a solid-liquid ratio of 1:1. The slurry is classified by hydrocyclone. The coarse slurry with a particle size of 45μm or larger is fed into a ball mill for grinding. Ceramic media are used for grinding. The slurry is ground to a fineness of -45μm, which accounts for 92%. The finely ground slurry is combined with the fine slurry with a particle size of less than 45μm from the hydrocyclone classification as leaching feedstock.

[0087] 2. Add the leaching material to the reaction vessel or reaction tank, and add concentrated acid to adjust the pH to 1.0. Heat to 80°C and stir for leaching for 2 hours.

[0088] 3. After leaching, a centrifuge is used for solid-liquid separation to obtain leaching residue and lithium-poor leachate A. The leaching residue is washed to obtain lithium-poor leachate B. Lithium-poor leachate A and lithium-poor leachate B are combined and returned to the stirring and pulping process for cyclic leaching. The lithium-rich solution obtained after two cycles of cyclic leaching is returned to the leaching section of the lithium carbonate plant to continue the production of lithium carbonate products. The lithium content and lithium leaching rate or yield in this lithium-rich solution are shown in Table 2.

[0089] 4. After washing, limestone is added to neutralize the residual acid in the leaching residue, adjusting the slurry pH to 6-7 and the slurry concentration to 32%. Then, gypsum collector and modifier are added separately. The collector consists of 30-50 parts sodium lauroyl urate and 10-20 parts sodium etheramine acetate, while the modifier is water glass. The dosage of gypsum collector is 400 g / t, and the dosage of water glass is 3000 g / t. After one roughing, three scavenging, and two cleaning flotation desulfurization processes, a froth product and desulfurized lithium slag slurry are obtained. The froth product is concentrated in a thickener and filtered to obtain the gypsum product. The filtered water is treated by chemical flocculation and sedimentation and then returned to the recycling tank for reuse. The chemical composition of this gypsum product is shown in Table 3.

[0090] 5. The lithium slag slurry after flotation desulfurization is subjected to magnetic separation in stages using a 0.3T magnetic drum and a 1.5T high-gradient magnetic separator to remove iron, resulting in de-ironized slag, iron-containing material A, and iron-containing material B.

[0091] 6. Iron-containing materials A and B are combined and subjected to gravity separation via a spiral chute and shaking table to obtain tantalum-niobium rough concentrate and iron slag A. The tantalum-niobium rough concentrate is then subjected to a weak magnetic separator with a magnetic field strength of 0.12T to remove magnetic iron impurities, resulting in tantalum-niobium concentrate and iron slag B. The tantalum-niobium concentrate is concentrated and filtered in a thickener to obtain the tantalum-niobium concentrate product, and the filtered water is recycled after treatment. Iron slag A and iron slag B are combined and concentrated in a thickener, then filtered to obtain iron slag. The tantalum-niobium concentrate is concentrated and filtered in a thickener to obtain the tantalum-niobium concentrate product, and the filtered water is recycled in a recycling tank. The grade and recovery rate of the tantalum-niobium concentrate product are shown in Table 4.

[0092] 7. After the de-iron slag is concentrated in a thickener and filtered by a filter, it is dried to obtain silicon-aluminum concentrate. The filtered water is recycled in a return water tank. The chemical composition of the silicon-aluminum concentrate is shown in Table 5.

[0093] Example 2

[0094] like Figure 1 As shown, a method for the comprehensive utilization of spodumene smelting slag is presented. The spodumene smelting slag originates from Jiangsu Province, and its composition is shown in Table 1. The specific implementation steps are as follows:

[0095] 1. Add water to slag for smelting and stir to prepare a slurry with a solid-liquid ratio of 1:2. The slurry is classified by hydrocyclone. The coarse slurry with a particle size of 45μm or larger is fed into a ball mill for grinding. Ceramic media are used for grinding. The slurry is ground until 90% of the particles are -45μm. The finely ground slurry is combined with the fine slurry with a particle size of less than 45μm from the hydrocyclone classification and used as leaching feedstock.

[0096] 2. Add the leaching material to the reactor or reaction tank, and add concentrated acid to adjust the pH to 1.2. Heat to 85°C and stir for leaching for 2.5 hours.

[0097] 3. After leaching, solid-liquid separation is performed by filtration to obtain leaching residue and lithium-poor leachate A. The leaching residue is washed to obtain lithium-poor leachate B. Lithium-poor leachate A and lithium-poor leachate B are combined and returned to the stirring and pulping process for cyclic leaching. After 3 cycles of cyclic leaching, a lithium-rich solution is obtained and returned to the leaching section of the lithium carbonate plant to continue the production of lithium carbonate products. The lithium content and lithium leaching rate or yield in this lithium-rich solution are shown in Table 2.

[0098] 4. After washing, quicklime is added to the leaching residue to neutralize the residual acid, adjust the slurry pH to 6-7, and adjust the slurry concentration to 28%. Then, gypsum collector and modifier are added. The collector consists of 50-70 parts of sodium alanine from coconut oil fatty acids, 2-5 parts of sodium dodecylbenzene sulfonate, and 5-10 parts of lauryl ether phosphate. The modifier is sodium hexametaphosphate. The dosage of gypsum collector is 600 g / t, and the dosage of hexametaphosphate is 1000 g / t. After one roughing, three scavenging, and two cleaning flotation desulfurization processes, a foam product and desulfurized lithium slag slurry are obtained. The foam product is concentrated in a thickener and filtered to obtain the gypsum product. The filtered water is treated by activated carbon adsorption and then returned to the recycling tank for reuse. The chemical composition of this gypsum product is shown in Table 3.

[0099] 5. The lithium slag slurry after flotation desulfurization is subjected to magnetic separation in stages using a 0.4T magnetic drum and a 1.0T high-gradient magnetic separator to remove iron, resulting in de-ironized slag, iron-containing material A, and iron-containing material B.

[0100] 6. Iron-containing materials A and B are combined and subjected to gravity separation using a blanket separator and shaking table to obtain tantalum-niobium rough concentrate and iron slag A. The tantalum-niobium rough concentrate is then subjected to a weak magnetic separator with a magnetic field strength of 0.15T to remove magnetic iron impurities, yielding tantalum-niobium concentrate and iron slag B. Iron slag A and iron slag B are combined, concentrated in a thickener, and filtered to obtain iron slag. The tantalum-niobium concentrate is then concentrated in a thickener and filtered to obtain tantalum-niobium concentrate product. The filtered water is recycled into a return water tank. The grade and recovery rate of the tantalum-niobium concentrate product are shown in Table 4.

[0101] 7. After the de-iron slag is concentrated in a thickener and filtered by a filter, it is dried to obtain silicon-aluminum concentrate. The filtered water is recycled in a return water tank. The chemical composition of the silicon-aluminum concentrate is shown in Table 5.

[0102] Example 3

[0103] like Figure 1 As shown, a method for the comprehensive utilization of spodumene smelting slag is presented. The spodumene smelting slag originates from Jiangxi Province, and its composition is shown in Table 1. The specific implementation steps are as follows:

[0104] 1. Add water to slag for smelting and stir to prepare a slurry with a solid-liquid ratio of 1:3. The slurry is classified by hydrocyclone. The coarse slurry with a particle size of 45μm or larger is fed into a ball mill for grinding. Ceramic media are used for grinding. The slurry is ground to a fineness of -45μm, which accounts for 95%. The finely ground slurry is combined with the fine slurry with a particle size of less than 45μm from the hydrocyclone classification and used as leaching raw material.

[0105] 2. Add the leaching material to the reactor or reaction tank, and add concentrated acid to adjust the pH to 1.5. Heat to 90°C and stir for leaching for 3 hours.

[0106] 3. After leaching, solid-liquid separation is performed by filtration to obtain leaching residue and lithium-poor leachate A. The leaching residue is washed to obtain lithium-poor leachate B. Lithium-poor leachate A and lithium-poor leachate B are combined and returned to the stirring and pulping process for cyclic leaching. After 4 cycles of cyclic leaching, a lithium-rich solution is obtained and returned to the leaching section of the lithium carbonate plant to continue the production of lithium carbonate products. The lithium content and lithium leaching rate or yield in this lithium-rich solution are shown in Table 2.

[0107] 4. After washing, limestone is added to the leaching residue to neutralize the residual acid, adjust the slurry pH to 6-7, and adjust the slurry concentration to 32%. Then, gypsum collector and modifier are added. The collector consists of 60-80 parts sodium cocoyl alanine and 5-10 parts sodium dodecyl sulfate. The modifier is CMC. The dosage of gypsum collector is 550 g / t, and the dosage of CMC is 200 g / t. After one roughing, three scavenging, and two cleaning flotation desulfurization processes, foam product and desulfurized lithium slag slurry are obtained. The foam product is concentrated in a thickener and filtered to obtain gypsum product. The filtered water is treated by activated sludge and returned to the recycling tank for reuse. The chemical composition of this gypsum product is shown in Table 3.

[0108] 5. The lithium slag slurry after flotation desulfurization is subjected to magnetic separation in stages using a 0.5T magnetic drum and a 1.3T high-gradient magnetic separator to remove iron, resulting in de-ironized slag, iron-containing material A, and iron-containing material B.

[0109] 6. Iron-containing materials A and B are combined and then separated by centrifugal concentrator + shaking table gravity separation to obtain tantalum-niobium rough concentrate and iron slag A. The tantalum-niobium rough concentrate is then subjected to a weak magnetic separator with a magnetic field strength of 0.16T to remove magnetic iron impurities, yielding tantalum-niobium concentrate and iron slag B. Iron slag A and iron slag B are combined and concentrated in a thickener, then filtered to obtain iron slag. The tantalum-niobium concentrate is further concentrated and filtered in a thickener to obtain tantalum-niobium concentrate product. The filtered water is recycled into a return water tank. The grade and recovery rate of the tantalum-niobium concentrate product are shown in Table 4.

[0110] 7. After the de-iron slag is concentrated in a thickener and filtered by a filter, it is dried to obtain silicon-aluminum concentrate. The filtered water is recycled in a return water tank. The chemical composition of the silicon-aluminum concentrate is shown in Table 5.

[0111] Table 2. Acid leaching indexes of lithium slag in Examples 1-3

[0112]

[0113] Table 3 Chemical composition (%) of gypsum products obtained in Examples 1-3

[0114] Implementation Items <![CDATA[SO3]]> <![CDATA[Fe2O3]]> Example 1 40.82 0.26 Example 2 41.03 0.21 Example 3 42.76 0.18

[0115] Table 4. Indicators of tantalum-niobium concentrate products obtained in Examples 1-3 / %

[0116]

[0117] Table 5 Chemical composition (%) of the silicon-aluminum concentrate products obtained in Examples 1-3

[0118] Silicon-aluminum concentrate <![CDATA[SO3]]> <![CDATA[Fe2O3]]> <![CDATA[Al2O3]]> <![CDATA[SiO2]]> <![CDATA[Li2O]]> 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

[0119] It is evident that the method of this invention can comprehensively recover and utilize elements such as lithium, tantalum, niobium, silicon, and aluminum in lithium slag, without generating wastewater or waste residue. This can solve the future problem of lithium slag disposal, reduce environmental pollution, and realize the comprehensive utilization of lithium slag resources.

Claims

1. A method for the comprehensive utilization of spodumene smelting slag resources, characterized in that, Includes the following steps: a. Pulping: Mix spodumene smelting slag and water to prepare a lithium slag slurry with a solid-liquid mass ratio of 1:1 to 3; b. Grinding: The lithium slag slurry from step a is classified by particle size into fine slurry and coarse slurry. The coarse slurry is wet-ground and then combined with the fine slurry to obtain fine slurry. c. Leaching: Sulfuric acid is added to the fine slurry to adjust the pH to 1-1.5, and the mixture is heated and stirred for leaching; then solid-liquid separation is performed to obtain acidic slag and leachate; the leachate is a lithium-poor solution; the leaching temperature is 60-90℃, and the leaching time is 1-3 h; d. Slurry preparation: After washing the acidic slag, add water or recycled water and stir to prepare slurry. Adjust the slurry concentration to 25-35 wt%, and adjust the pH of the slurry to 6-7 by adding limestone, quicklime or hydrated lime. e. Flotation: The slurry after preparation in step d is subjected to flotation desulfurization to obtain desulfurized lithium slag and flotation foam products; f. Magnetic separation: The desulfurized lithium slag is separated by medium magnetic separation to obtain iron-containing material A and medium magnetic concentrate; the medium magnetic concentrate is separated by strong magnetic separation to obtain iron-containing material B and strong magnetic concentrate; the strong magnetic concentrate is filtered and dried to obtain silicon-aluminum concentrate 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 tantalum-niobium rough concentrate is separated by weak magnetic separation to remove magnetic iron impurities, then concentrated and filtered to obtain tantalum-niobium concentrate and filtrate B. The magnetic iron impurities are combined with the gravity separation tailings, concentrated and filtered to obtain iron slag and filtrate C. In step c, the leachate is returned to step a as a lithium-poor solution to replace part of the water, and after repeated leaching, a lithium-rich solution is obtained; the washing liquid generated during washing in step d is returned to step a as a lithium-poor solution to replace part of the water; wherein, the Li2O content in the lithium-poor solution is <5g / L; the Li2O content in the lithium-rich solution is ≥5g / L; and the number of leaching cycles is 2 to 4.

2. The method for comprehensive utilization 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 material with a particle size > 45 μm.

3. The method for comprehensive utilization of spodumene smelting slag resources according to claim 2, characterized in that: In step b, the grinding fineness of wet grinding is -45μm with a content ≥90%.

4. The method for comprehensive utilization of spodumene smelting slag resources according to claim 1, characterized in that: In step c, the leaching temperature is 80–90℃ and the leaching time is 2–3 h.

5. The method for comprehensive utilization of spodumene smelting slag resources according to claim 1, characterized in that: In step d, the pulp concentration is adjusted to 28-32%.

6. The method for comprehensive utilization of spodumene smelting slag resources according to claim 1, characterized in that: In step e, the flotation of the foam product is followed by concentration and filtration to obtain the gypsum product and the filtrate A′.

7. The method for comprehensive utilization of spodumene smelting slag resources according to claim 1, characterized in that: In step f, the field strength of the medium magnetic separation is 0.2 to 0.6 T, and the field strength of the strong magnetic separation is 1.0 to 1.7 T.

8. The method for comprehensive utilization of spodumene smelting slag resources according to claim 7, characterized in that: In step f, the field strength of the medium magnetic separation is 0.3 to 0.5 T, and the field strength of the strong magnetic separation is 1.0 to 1.5 T.

9. The method for comprehensive utilization of spodumene smelting slag resources according to claim 1, characterized in that: In step g, the re-selection is carried out by spiral chute + shaking table beneficiation, or blanket mill + shaking table beneficiation, or centrifugal concentrator + shaking table beneficiation; In step h, the field strength of the weak magnetic separation is 0.1 to 0.2 T.

10. The method for comprehensive utilization of spodumene smelting slag resources according to claim 9, characterized in that: In step h, the field strength of the weak magnetic separation is 0.12 to 0.16 T.

11. The method for comprehensive utilization of spodumene smelting slag resources according to claim 1, characterized in that: Concentration is performed before filtration in steps f and h.

12. The method for comprehensive utilization of spodumene smelting slag resources according to claim 6, characterized in that: After water treatment, filtrate A′ is returned to the return water tank for recycling. Filtrates A, B, and C are directly returned to the return water tank for recycling. The return water from the return water tank is recycled to the pulping, washing, pulp conditioning, flotation, magnetic separation, and gravity separation stages.

13. The method for comprehensive utilization of spodumene smelting slag resources according to claim 12, characterized in that: Water treatment methods include at least one of sedimentation, adsorption, and activated sludge treatment.

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