Method for resourceful treatment of spodumene concentrate

Abstract

The invention discloses a spodumene concentrate resourceful treatment method which comprises the following steps: carrying out wet treatment on spodumene acid clinker to prepare lithium-removed filter residues and leachate; carrying out neutralization treatment and desulfurization treatment on the leachate to obtain a lithium-containing purified solution, gypsum and iron oxide red, and carrying out lithium extraction treatment on the lithium-containing purified solution to obtain a lithium salt and mirabilite; the lithium-removed filter residues are treated through a process combining a beneficiation method and a chemical method, and water glass, nano kaolinite, ceramsite aggregate and tantalum-niobium concentrate are obtained. According to the method, whole-process slag-free production of the spodumene concentrate is achieved, and meanwhile the lithium-containing compound with the high purity and the high quality is obtained; and various valuable products can be obtained.

Description

TECHNICAL FIELD

[0001] The present application relates to the technical field of spodumene processing, in particular to a spodumene concentrate resource processing method. BACKGROUND

[0002] Spodumene is the main raw material for producing battery-grade lithium carbonate, and its processing can be divided into calcination method and hydrothermal pressure cooking method. The current main industrialized production process for lithium extraction from spodumene is sulfuric acid calcination method. The process flow is to convert spodumene into β-spodumene by high-temperature calcination at 950℃~1200℃, then acidize the β-spodumene with excess concentrated sulfuric acid at 250~300℃, and then obtain lithium carbonate and lithium slag through leaching, neutralization, solid-liquid separation, and impurity removal.

[0003] Patent CN 117756487B discloses a lithium slag cementitious material and a preparation method thereof. In the cementitious material, lithium slag replaces part of the cement, and the silicon and aluminum components in the lithium slag and the calcium oxide or calcium hydroxide in the cement undergo a chemical reaction to produce calcium silicate hydrate or calcium aluminate hydrate gel, thereby improving the mechanical properties of concrete. Although this method can use lithium slag, it requires ball milling the lithium slag to 4~6μm, which has high particle size requirements and high processing cost. Patent CN 115286419B discloses a method for preparing and applying lightweight ceramsite aggregate from sulfuric acid lithium extraction slag. The method is to granulate and agglomerate the sulfuric acid lithium extraction slag, calcium carbonate, magnesium carbonate, and albite / potassium feldspar, and then calcine the pellets to obtain lightweight ceramsite aggregate. The lightweight ceramsite aggregate prepared by this method has a 1h water absorption rate of <5%, a cylinder compressive strength of 3.5~8.8MPa, and a bulk density of 480~650kg / m 3However, the method requires high-temperature calcination (1100-1200℃), which has high energy consumption. Patent CN108273826A discloses a full-phase high-value recycling method of lithium residue. The patent obtains a leaf talc raw material for glass fiber through slurry preparation, carbonate reaction, and magnetic separation treatment. This method needs to recover sulfate products by crystallization, has the disadvantages of high cost and low alkali conversion efficiency, and is difficult to realize industrialization application. Patent CN108147658A obtains a leaf talc raw material for glass fiber through slurry preparation, physical beneficiation desulfurization, and magnetic separation iron removal. The leaf talc for glass fiber produced by this method has the characteristics of high iron and high sulfur, and its application is limited to a certain extent. Patent CN115072749B provides a method for lithium extraction from lithium spodumene without residue. First, the ball-milled lithium spodumene concentrate is mixed with an alkaline substance, then the product is obtained after calcination, and the slurry is prepared by mixing with water. The slurry is heated and reacted, then filtered to obtain a water leaching solution and nepheline. The water leaching solution is added with calcium oxide for hydrothermal reaction, filtration, washing, drying, and concentration to obtain lithium carbonate and calcium silicate hydrate byproducts. Finally, the nepheline is hydrothermally reacted under acidification conditions to prepare nano-high kaolinite. Although this method realizes lithium extraction from lithium spodumene without residue and preparation of high-value nano-high kaolinite and calcium silicate hydrate products, it requires more raw materials, the byproducts are affected by the properties of the raw materials, and the process is complex.

[0004] As can be seen from the above, the problem of treating lithium residue generated in the lithium extraction process of lithium spodumene has not been well solved. According to the new edition of the General Portland Cement Standard (GB175-2023), the existing disposal channel of lithium residue as a cement mixing material will not be applicable. How to realize scientific disposal and efficient utilization of lithium residue has become an important problem restricting its sustainable development; at the same time, the price of lithium carbonate continues to fall, and through lithium extraction from lithium spodumene concentrate without residue, high-value development of lithium residue has good economic benefits. Therefore, it is of great significance to develop a lithium spodumene concentrate resource treatment method. SUMMARY

[0005] The purpose of the present application is to overcome the defects of the prior art and provide a lithium spodumene concentrate resource treatment method.

[0006] The technical problem of the present application is solved by adopting the following technical scheme.

[0007] The present application provides a lithium spodumene concentrate resource treatment method, which comprises: S1: wet treatment of lithium spodumene acid clinker to obtain lithium-removed filter residue and leaching solution; S2: sequentially neutralizing and treating the leaching solution, and desulfurizing to obtain a lithium-containing purified solution; S3: lithium extraction treatment of the lithium-containing purified solution to obtain a lithium salt.

[0008] The present application has the following beneficial effects: The present invention provides a method for resource processing of spodumene concentrate. Spodumene acid clinker is subjected to leaching treatment, and the leachate is sequentially subjected to neutralization treatment and desulfurization treatment to obtain a lithium-containing purified liquid, which is then subjected to lithium extraction treatment. The above method reduces the problem of impurity interference in the leachate through segmented purification and impurity removal treatment, and the lithium carbonate product produced has higher purity and more stable quality. The lithium extraction rate exceeds 95%, and it can maximize the extraction of lithium elements in lithium concentrate, reduce the waste of lithium resources, and obtain a variety of valuable products while producing lithium-containing products. BRIEF DESCRIPTION OF THE DRAWINGS

[0009] In order to more clearly illustrate the technical solutions of the embodiments of the present invention, the following briefly introduces the drawings required for use in the embodiments. It should be understood that the following drawings only illustrate certain embodiments of the present invention and therefore should not be regarded as limiting the scope. For ordinary technicians in this field, other relevant drawings can be obtained based on these drawings without paying any creative work.

[0010] Figure 1 This is a process flow chart for extracting lithium from spodumene concentrate without slag to prepare lithium carbonate, nano-kaolinite, ceramsite aggregate, tantalum-niobium concentrate, water glass, iron oxide, gypsum and sodium sulfate; Figure 2 The X-ray powder diffraction pattern of lithium carbonate prepared in an embodiment of the present invention; Figure 3 This is the X-ray powder diffraction pattern of the tantalum-niobium concentrate prepared in the embodiment of the present invention. DETAILED DESCRIPTION

[0011] To make the purpose, technical solutions and advantages of the embodiments of the present invention clearer, the technical solutions in the embodiments of the present invention are described clearly and completely below. Where specific conditions are not specified in the embodiments, conventional conditions or conditions recommended by the manufacturer are used. Where the manufacturer of the reagents or instruments is not specified, all are conventional products that can be purchased commercially.

[0012] The following is a detailed description of a spodumene concentrate resource processing method provided in an embodiment of the present invention.

[0013] The embodiment of the present invention provides a method for resource processing of spodumene concentrate, which includes: S1: wet-processing spodumene acid clinker to obtain lithium-removed residue and leachate; S2: The leachate is subjected to a neutralization treatment and a desulfurization treatment in sequence to obtain a lithium-containing purified liquid; S3: extracting lithium from the purified lithium-containing liquid to obtain lithium salt.

[0014] The current main industrial production process for extracting lithium from spodumene is sulfuric acid roasting method. The process is that spodumene is calcined by crystal transformation, and then is acidized with excess concentrated sulfuric acid. Lithium-containing purified solution is prepared through leaching, neutralization, solid-liquid separation and other processes, and then lithium is extracted from the lithium-containing purified solution. In the above method, calcium carbonate, calcium hydroxide, calcium oxide and the like are added in the neutralization process to generate gypsum, and the lithium-containing purified solution is obtained after the gypsum is separated. Although this method can precipitate a large amount of sulfate ions in the leaching solution and generate gypsum, the influence of other impurity ions contained in the solution is not considered. In addition, there is a method that sodium hydroxide, potassium hydroxide and the like are added in the water leaching solution to make the impurity ions form impurity precipitates, and the lithium-containing purified solution is obtained after the impurity precipitates are separated. Although this method can precipitate and remove the impurity ions in the solution, the influence of a large amount of sulfate in the solution is not considered, which will increase the cost of flue gas treatment.

[0015] In order to overcome the above defects, the embodiment of the present application provides a spodumene concentrate resource processing method, which comprises: sequentially performing neutralization treatment and desulfurization treatment on the leaching solution of the spodumene acid clinker subjected to wet treatment to obtain lithium-containing purified solution, and extracting lithium from the lithium-containing purified solution. In the preparation process of the lithium-containing purified solution, the leaching solution is subjected to neutralization treatment, so that a small amount of Fe 3+ precipitates is dissolved in the roasting, reducing the impurity ions in the water leaching solution, and then the filtrate after the precipitation is filtered out is subjected to desulfurization treatment, so that a large amount of sulfate in the water leaching solution is converted into gypsum. The gypsum and other minerals are pre-separated in the lithium extraction process, the separation effect is good, which is beneficial to the high-value recovery of the lithium-depleted filter residue in the subsequent process, and can also reduce the cost of flue gas treatment. The lithium-containing purified solution prepared has a small amount of impurity ions and high purity, and a lithium-containing product with higher purity and better quality can be obtained after lithium extraction treatment.

[0016] In some optional embodiments, the following steps are included: S1: wet-treating spodumene acid clinker to obtain lithium-depleted filter residue and leaching solution; S2: after the leaching solution is subjected to neutralization treatment, iron-removing filtrate and filter residue are obtained through filtration separation, and the filter residue is calcined to obtain iron red; the iron-removing filtrate is subjected to desulfurization treatment, and gypsum and lithium-containing purified solution are obtained through filtration separation; S3: the lithium-containing purified solution is subjected to lithium extraction treatment, and lithium salt and precipitation mother liquor are obtained through filtration separation, and then the precipitation mother liquor is subjected to separation treatment to obtain mirabilite and sodium precipitation mother liquor.

[0017] The embodiment of the present application provides a spodumene concentrate resource processing method, which comprises the following steps: wet processing of spodumene acid clinker to obtain delithium filter residue and leaching solution; purification and impurity removal processing of the leaching solution to obtain lithium-containing purified solution and two products of iron red and gypsum; lithium extraction processing of the lithium-containing purified solution to obtain lithium salt products, mirabilite and sodium precipitation mother liquor; and the sodium precipitation mother liquor can be returned to the lithium extraction processing step for reuse.

[0018] It can be seen that, in the lithium extraction processing for preparing lithium salt, the iron red, gypsum and mirabilite products are obtained through segmented processing, which can reduce the impurity interference problem in the leaching solution, so that the purity of the produced lithium salt product is higher and the quality is more stable. At the same time, the leaching rate of lithium is more than 95%, which can maximize the extraction of lithium elements in the lithium concentrate and reduce the waste of lithium resources. Compared with the traditional lithium extraction method, the mineral resources can be more fully utilized.

[0019] In some optional embodiments, the delithium filter residue is processed through the following steps: S4: magnetic separation processing of the delithium filter residue to obtain magnetic material and non-magnetic material; S5: mineral processing of the magnetic material to obtain tantalum-niobium concentrate and ceramic aggregate; S6: hydrothermal treatment of the non-magnetic material to obtain water glass solution and nano kaolinite.

[0020] The embodiment of the present application provides a spodumene concentrate resource processing method, which comprises the following steps: wet processing of spodumene acid clinker to obtain delithium filter residue and leaching solution; in the delithium filter residue processing process, a combined process of mineral processing method and chemical method is used to reduce the influence of iron on the preparation of water glass and nano kaolinite products from the delithium filter residue. Through resource recycling of the delithium filter residue, high-value tantalum-niobium concentrate, water glass solution, nano kaolinite and ceramic aggregate product obtained by processing of tailings can be obtained.

[0021] In some optional embodiments, the spodumene acid clinker is prepared by crystal transformation roasting and acid roasting of spodumene concentrate; Preferably, the Li2O content in the spodumene concentrate is greater than or equal to 4.0%, and the content of -100 mesh in the spodumene concentrate is greater than or equal to 70%; Preferably, the crystal transformation roasting temperature is 1100-1300 DEG C, and the time is 0.5-2 h; the acid roasting temperature is 200-300 DEG C, and the time is 30-60 min; the concentration of concentrated sulfuric acid used in the acid roasting is 95-98%, and the amount of concentrated sulfuric acid is 1.3-1.5 times the theoretical amount.

[0022] In some optional embodiments, in step S1, the wet processing comprises: mixing the spodumene acid clinker with water for leaching treatment, and performing pressure filtration, washing and filtration treatment on the leaching slurry to obtain delithium filter residue and leaching solution; Preferably, the liquid-solid ratio of the leaching treatment is 0.8:1-3:1, the temperature is 30-80℃, and the time is 0.5-2.0h; the washing times of the leaching slurry is 2-3 times; preferably, the Li2O content in the lithium-removed filter residue is ≤0.2%, and the SO3 content is ≤0.3%; the pH of the leaching solution is 0.5-2.5.

[0023] In some alternative embodiments, in step S2, the neutralization treatment comprises: slurry preparation, aging, filtration, drying and calcination treatment of the leaching solution to obtain an iron-removed filtrate and iron red; Preferably, the pH of the leaching solution after slurry preparation with NaOH is 3.5-4.5, the NaOH concentration is 2-10mol / L, the aging time is 0.5-2.0h, the filtration uses a vacuum treatment machine or a filter press, the drying temperature is 100-200℃, the drying time is 2-10h, the calcination temperature is 450-550℃, and the calcination time is 0.5-4h; preferably, the Fe2O3 content in the iron red is ≥98%.

[0024] In some alternative embodiments, in step S2, the desulfurization treatment comprises: adding quicklime to the iron-removed filtrate for purification and impurity removal, aging, filtration and drying treatment to obtain gypsum and purified liquid; Preferably, the pH of the iron-removed filtrate after adding quicklime for purification and impurity removal is 10-12, and the aging time is 0.2-2.0h; preferably, the SO3 content in the gypsum is ≥43%, and the Li2O content is ≤0.2%.

[0025] In some alternative embodiments, in step S3, the lithium extraction treatment comprises: evaporation concentration, carbonization, centrifugation and drying treatment of the lithium-containing purified liquid to prepare lithium carbonate and a precipitation mother liquor, and the precipitation mother liquor is separated by freezing to prepare mirabilite and a sodium precipitation mother liquor; Preferably, the Li2O concentration of the lithium-containing purified liquid after evaporation concentration is 50-80g / L, the soda concentration is 10-30%, and the freezing separation temperature is -5--10℃; preferably, the lithium carbonate content is ≥99.5%, the Na2SO4 content in the mirabilite is ≥44.0%, and the sodium precipitation mother liquor is returned to the evaporation concentration process and the pure soda solution preparation process.

[0026] In some alternative embodiments, in steps S4 and S5, the magnetic separation and gravity separation treatment comprises: performing a weak magnetic+strong magnetic separation treatment process on the lithium-removed filter residue to obtain magnetic material and non-magnetic material, and performing gravity separation beneficiation treatment on the magnetic material to prepare tantalum-niobium concentrate and gravity separation tailings, and performing filtration and drying treatment on the gravity separation tailings to prepare ceramic aggregate; Preferably, the strength of the weak magnetic separation is 3000-6000Gs, the strength of the strong magnetic separation is 1.0-1.8T, and the concentration of the magnetic separation slurry is 10-30%; the gravity separation equipment includes one or more combinations of a spiral chute, a shaking table and a blanket machine.

[0027] In some alternative embodiments, in step S6, the hydrothermal treatment comprises: controlling the hydrothermal reaction temperature to be 200-300℃, the pressure to be 1.0-5.0 MPa, the hydrothermal reaction time to be 2-12 h, and the amount of sodium hydroxide to be 1.1-1.3 times the theoretical amount.

[0028] In some alternative embodiments, the method comprises the following steps: S1: preparing lithium spodumene acid clinker by converting crystallization roasting and acidizing roasting of lithium spodumene concentrate; S2: mixing the lithium spodumene acid clinker with water for leaching, and performing pressure filtration, stirring washing, backwashing and separation treatment on the leaching slurry to obtain lithium-removed filter residue and leaching liquor; S3: adding the leaching liquor into NaOH solution for slurry adjustment, aging, filtration, drying and calcination treatment to obtain iron-removed filtrate and iron red; S4: adding quicklime to the iron-removed filtrate for purification and impurity removal, aging, filtration and drying treatment to obtain gypsum and purified liquor; S5: performing evaporation concentration, carbonization, centrifugation and drying treatment on the purified liquor to prepare lithium carbonate and centrifugal filtrate, and performing refrigeration separation on the centrifugal filtrate to prepare mirabilite; S6: performing magnetic separation treatment on the lithium-removed filter residue to obtain magnetic material and non-magnetic material; S7: preparing tantalum-niobium concentrate, gravity separation tailings and tailings from the magnetic material by gravity separation treatment, and preparing ceramic aggregate from the tailings by filtration and drying processes; S8: performing hydrothermal reaction on the non-magnetic material to prepare water glass solution and kaolinite raw material, and preparing nano-kaolinite from the kaolinite raw material by drying and grinding.

[0029] The main chemical reactions involved in the above treatment process are as follows: (1) Conversion roasting

[0030] (2) Acidizing roasting

[0031] (3) Neutralization of residual acid

[0032] (3) Iron removal from filter residue

[0033] (4) Preparation of iron red

[0034] (5) Desulfurization treatment ,

[0035] (6) Lithium precipitation

[0036] (7) Hydrothermal reaction

[0037] As can be seen from the above, the embodiment of the present invention provides a method for resource processing of spodumene concentrate, which makes full use of the lithium, silicon, aluminum, tantalum niobium, iron and other components in spodumene, can achieve 100% resource utilization and slag-free production in the entire process, and produces high-value-added nano-kaolinite, tantalum niobium concentrate, iron red, ceramsite aggregate, water glass, gypsum, and Glauber's salt products while preparing lithium carbonate. It solves the problem that the solid waste smelting slag in the spodumene processing industry is difficult to utilize and limits its sustainable development. The traditional lithium extraction method will produce a large amount of waste slag. The storage of these waste slags not only occupies land resources, but also may pollute the soil and water quality due to the leakage of heavy metals and harmful substances therein. The present invention greatly reduces the risk of environmental pollution and helps to protect the ecological environment and biodiversity.

[0038] The present invention will be further described below with reference to the embodiments.

[0039] Example 1 according to Figure 1 The process shown in FIG. 1 is shown in FIG. 2 . The spodumene used in this embodiment is spodumene concentrate from a lithium carbonate processing enterprise in Sichuan. The chemical composition analysis results are shown in Table 1.

[0040] Table 1 Chemical composition analysis results of spodumene (wt%)

[0041] 500g of spodumene was weighed and placed in a muffle furnace, calcined at 1150°C for 1h, cooled to 250°C, and then 700g of 98% concentrated sulfuric acid was added for acidification and calcination for 1h to obtain spodumene acid clinker. Then, clean water was added to the cooled spodumene acid clinker at a liquid-to-solid mass ratio of 1.5:1 and leached for 1h at a leaching temperature of 50-60°C to obtain a leached slurry. The leached slurry was filtered, washed three times, and filtered to obtain a filter residue and a leachate. The leachate was a mixture of lithium sulfate and sulfuric acid. The lithium content in the leachate was measured, and the lithium leaching rate was calculated to be 98.9%.

[0042] The filtrate was adjusted to pH 4.0 by adding 2 mol / L NaOH solution, and then aged for 2 h. The filtrate was filtered, and the filter residue was dried in an oven at 105 ℃ for 8 h. The dried material was calcined in a muffle furnace at 450 ℃ for 2 h to obtain iron red, and the Fe2O3 content in the iron red was 98.74%. The pH of the filtrate after iron removal was adjusted to 10 by adding quicklime, and the filtrate was aged for 1.0 h to obtain gypsum filter residue and purified solution. The gypsum filter residue was dried at 110 ℃ for 8 h to obtain gypsum product, and the SO3 content in the gypsum was 44.86%, and the Li2O content was 13%. The purified solution was pumped into a lithium precipitation reactor, heated to boiling, and evaporated to remove part of the water. The purified solution was concentrated to a Li2O concentration of 60-70 g / L, and then subjected to carbonization reaction with 20% NaOH solution. The precipitate was separated by centrifugation to obtain lithium carbonate crude product and lithium precipitation mother liquor. The lithium carbonate crude product was dried to obtain lithium carbonate, and the X-ray powder diffraction pattern of the lithium carbonate is shown in Figure 2 Fig. 1. The lithium carbonate content was 99.8% as determined. The lithium precipitation mother liquor was frozen at -5 ℃ to separate mirabilite and sodium precipitation mother liquor, and the sodium precipitation mother liquor was used to prepare NaOH solution and return to the evaporation and concentration process.

[0043] The filter residue was slurried to a pulp concentration of 30%, and then subjected to weak magnetic separation and strong magnetic separation, respectively. The magnetic field strength of the weak magnetic separation was 4000 Gs, and the magnetic field strength of the strong magnetic separation was 1.5 T. Magnetic material and non-magnetic material were obtained, and the magnetic material was subjected to spiral chute-shaking table reselection process to obtain tantalum-niobium concentrate. The X-ray powder diffraction pattern of the tantalum-niobium concentrate is shown in Figure 3 Fig. 2. The Ta2O5 and Nb2O5 contents in the tantalum-niobium concentrate were 7.12% and 5.41%, respectively. The shaking table tailings were concentrated, filtered, and dried to obtain ceramic aggregate.

[0044] The non-magnetic material and 10 mol / L NaOH solution were added to the reactor for hydrothermal reaction. The hydrothermal reaction temperature was 250 ℃, the pressure was 3 MPa, and the hydrothermal reaction time was 6.0 h. After the hydrothermal reaction, the filter cake was dried and ground to obtain nano-kaolinite.

[0045] Example 2 The spodumene used in this example was still the spodumene concentrate from a lithium carbonate processing enterprise in Sichuan used in Example 1.

[0046] 500 g of spodumene was weighed into a muffle furnace and calcined at 1200 °C for 40 min. After cooling to 270 °C, 750 g of 95% concentrated sulfuric acid was added to perform acid calcination for 0.5 h to obtain spodumene acid clinker. Then, the cooled spodumene acid clinker was added with clean water at a liquid-solid mass ratio of 1.5:1 to leach for 2 h at a temperature of 30-40 °C to obtain leaching slurry. The leaching slurry was separated by pressure filtration, 3 times of washing and filtration to obtain leaching residue and leaching solution. The leaching solution was a mixture of lithium sulfate and sulfuric acid. The content of lithium in the leaching solution was determined to calculate the leaching rate of lithium, which was 99.1%.

[0047] 6 mol / L NaOH solution was added to the leaching solution to adjust the pH of the filtrate to about 4.0 and age for 1 h. Then, the filtrate was filtered and the filter residue was dried in an oven at 105 °C for 6 h. The dried material was calcined in a muffle furnace at 500 °C for 2 h to obtain iron red. The content of Fe2O3 in the iron red was 99.21% as determined. Lime was added to the iron removal filtrate to adjust the pH of the filtrate to 10.5 and age for 0.5 h. The gypsum filter residue and purified solution were obtained by filtration. The gypsum filter residue was dried at 110 °C for 8 h to obtain gypsum product. The SO3 content in the gypsum was 44.6% and the Li2O content was 0.12%. The purified solution was pumped into a lithium precipitation reactor and heated to boiling to evaporate part of the water to concentrate the purified solution to a Li2O concentration of 70-80 g / L. Then, carbonization reaction was performed using 20% pure caustic solution. Lithium carbonate crude and lithium precipitation mother liquor were obtained by centrifugal separation of the precipitate. The lithium carbonate crude was dried to obtain lithium carbonate. The X-ray powder diffraction pattern of the lithium carbonate was similar to that of Figure 2 As determined, the lithium carbonate content was 99.6%. Mirabilite and sodium precipitation mother liquor were obtained by freezing separation of the lithium precipitation mother liquor at -5 °C. The sodium precipitation mother liquor was used to prepare pure caustic solution and returned to the evaporation and concentration process.

[0048] The filter residue was thickened to a slurry concentration of 30%, and then subjected to weak magnetic separation and strong magnetic separation, respectively. The magnetic field strength of the weak magnetic separation was 3000 Gs, and the magnetic field strength of the strong magnetic separation was 1.6 T. Magnetic material and non-magnetic material were obtained. The magnetic material was subjected to a cloth machine-shaking table gravity separation process to obtain tantalum-niobium concentrate. The X-ray powder diffraction pattern of the tantalum-niobium concentrate was similar to that of Figure 3 As determined, the Ta2O5 and Nb2O5 contents in the tantalum-niobium concentrate were 6.83% and 5.21%, respectively. The shaking table tailings were treated by concentration, filtration and drying to obtain ceramic aggregate.

[0049] Non-magnetic material and 8 mol / L NaOH solution were added to the reaction kettle to perform hydrothermal reaction at a temperature of 220 °C and a pressure of 3 MPa for 4.0 h. After the hydrothermal reaction, pressure filtration and washing were performed. The filtrate was water glass solution, and the filter residue was dried and ground to obtain nano-kaolinite.

[0050] Example 3 The spodumene used in the present embodiment is a spodumene concentrate produced in a certain place in Jiangxi Province, and the results of chemical composition analysis thereof are shown in Table 2.

[0051] Table 2 Results of chemical composition analysis of spodumene (wt%)

[0052] The spodumene concentrate was weighed at 500 g and placed in a muffle furnace for roasting at 1100°C for 2.0 h. After cooling to 200°C, 650 g of 95% concentrated sulfuric acid was added for acid roasting for 1.0 h to obtain spodumene acid clinker. Then, the cooled spodumene acid clinker was added with clean water at a liquid-to-solid mass ratio of 3:1 for leaching for 0.5 h at a leaching temperature of 30-40°C to obtain leaching slurry. The leaching slurry was separated by pressure filtration, 2 times of washing and filtration to obtain leaching residue and leaching solution. The leaching solution was a mixture of lithium sulfate and sulfuric acid. The content of lithium in the leaching solution was determined, and the leaching rate of lithium was calculated to be 98.8%.

[0053] NaOH solution was added to the leaching solution to adjust the pH value of the filtrate to about 4.5, and the solution was aged for 0.5 h. Then, the solution was filtered, and the filter residue was dried in an oven at 120°C for 8 h. The dried material was calcined in a muffle furnace at 450°C for 4 h to obtain iron red. The content of Fe2O3 in the iron red was determined to be 98.46%. Lime was added to the filtrate to adjust the pH value of the filtrate to 12, and the solution was aged for 0.5 h. The solution was filtered to obtain gypsum filter residue and purified solution. The gypsum filter residue was dried at 100°C for 6 h to obtain gypsum product. The content of SO3 in the gypsum was 43.8%, and the content of Li2O was 0.14%. The purified solution was pumped into a lithium precipitation reactor, heated to boiling, and evaporated to remove part of the water. The concentration of Li2O in the purified solution was 50-60 g / L. Then, carbonization reaction was carried out using 10% pure caustic solution. The precipitate was separated by centrifugation to obtain crude lithium carbonate and lithium precipitation mother liquor. The crude lithium carbonate was dried to obtain lithium carbonate. The X-ray powder diffraction pattern of the lithium carbonate was similar to that of Li2CO3, and the content of lithium carbonate was determined to be 99.7%. The lithium precipitation mother liquor was frozen at -10°C to separate mirabilite and sodium precipitation mother liquor. The sodium precipitation mother liquor was used to prepare pure caustic solution and return to the evaporation and concentration process. Figure 2

[0054] The filter residue was slurried to a slurry concentration of 10%, and then subjected to weak magnetic separation and strong magnetic separation, respectively. The magnetic field strength of the weak magnetic separation was 3000 Gs, and the magnetic field strength of the strong magnetic separation was 1.0 T. Magnetic material and non-magnetic material were obtained. The magnetic material was subjected to shaking table reselection process to obtain tantalum-niobium concentrate. The X-ray powder diffraction pattern of the tantalum-niobium concentrate was similar to that of 3. The content of Ta2O5 and Nb2O5 in the tantalum-niobium concentrate was 5.82% and 4.38%, respectively. The shaking table tailings were treated by concentration, filtration and drying to obtain ceramic aggregate.

[0055] ​The non-magnetic material and 4 mol / L NaOH solution were added into the reactor for hydrothermal reaction, the hydrothermal reaction temperature was 200℃, the pressure was 5 MPa, and the hydrothermal reaction time was 10.0 h. After the hydrothermal reaction, pressure filtration and washing were performed, the filtrate was a water glass solution, and the filter residue was dried and ground to obtain nano kaolinite.

[0056] Example 4 The spodumene used in this example was still the spodumene concentrate from a certain place in Jiangxi used in Example 3.

[0057] 500 g of spodumene was placed in a muffle furnace and calcined at 1300℃ for 0.5 h, and then 750 g of 98% concentrated sulfuric acid was added to the cooled spodumene for acid roasting for 0.5 h to obtain spodumene acid clinker. Then, 0.8:1 of water was added to the cooled spodumene acid clinker for leaching for 0.5 h, and the leaching temperature was 70-80℃, to obtain leaching slurry. The leaching slurry was separated by pressure filtration, twice washing and filtration to obtain filter residue and leaching solution. The leaching solution was a mixture of lithium sulfate and sulfuric acid, and the content of lithium in the leaching solution was determined to calculate the leaching rate of lithium, which was 98.6%.

[0058] 10 mol / L NaOH solution was added to the leaching solution, the pH value of the filtrate was adjusted to about 4.5, and the solution was aged for 1.0 h, and then it was filtered. The filter residue was dried in an oven at 200℃ for 3 h, and then it was calcined in a muffle furnace at 550℃ for 0.5 h to obtain iron red, and the content of Fe2O3 in the iron red was 98.36% by determination. Lime was added to the iron removal filtrate to adjust the pH value of the filtrate to 10, and the solution was aged for 2.0 h, and then it was filtered to obtain gypsum filter residue and purified solution. The gypsum filter residue was dried at 150℃ for 4 h to obtain gypsum product, and the SO3 content in the gypsum was 44.2%, and the Li2O content was 0.14%. The purified solution was pumped into a lithium precipitation reactor, heated to boiling, and evaporated to remove part of the water to concentrate the purified solution to a Li2O concentration of 70-80 g / L. Then, 30% pure caustic solution was used for carbonization reaction, and lithium carbonate crude product and lithium precipitation mother liquor were obtained by centrifugal separation. The lithium carbonate crude product was dried to obtain lithium carbonate, and the X-ray powder diffraction pattern thereof was similar to that of Figure 2 The lithium carbonate crude product was dried to obtain lithium carbonate, and the X-ray powder diffraction pattern thereof was similar to that of

[0059] The filter residue was slurried to a slurry concentration of 30%, and then weak magnetic separation and strong magnetic separation were performed, the magnetic field strength of the weak magnetic separation was 6000Gs, and the magnetic field strength of the strong magnetic separation was 1.8T, to obtain magnetic material and non-magnetic material. The magnetic material was subjected to a cloth table-shaking table gravity separation process to obtain tantalum-niobium concentrate, and the X-ray powder diffraction pattern thereof was similar to that of Figure 3Similarly, the Ta2O5 and Nb2O5 contents in the tantalum-niobium concentrate were determined to be 8.32% and 6.15%, respectively. The ceramic aggregate was obtained by concentrating, filtering and drying the tailings of the shaking table.

[0060] The non-magnetic material and the NaOH solution with a concentration of 8 mol / L were added into the reaction kettle for hydrothermal reaction, the hydrothermal reaction temperature was 300°C, the pressure was 3 MPa, and the hydrothermal reaction time was 6.0 h. After the hydrothermal reaction, pressure filtration and washing were performed, the filtrate was a water glass solution, and the filter residue was dried and ground to obtain nano-kaolinite.

[0061] Comparative Example 1 Compared with Example 1, the difference lies in that the iron removal process is omitted in the preparation of sodium carbonate from the leaching solution, and the leaching solution is directly purified and impurities are removed, and the other steps and parameters are the same as those in Example 1.

[0062] Comparative Example 2 Compared with Example 1, the difference lies in that the filter residue is directly filtered and dried to obtain the ceramic aggregate after magnetic separation, and the other steps and parameters are the same as those in Example 1.

[0063] Comparative Example 3 Similar to the steps of Example 1, the only difference is that the liquid-solid ratio of the leaching treatment during the treatment process is 5:1, the obtained leaching solution is a mixture of lithium sulfate and sulfuric acid, the pH of the leaching solution is 2.8, the content of lithium in the leaching solution is determined, and the leaching rate of lithium is calculated. At the same time, lithium carbonate, nano-kaolinite, iron red, gypsum, ceramic aggregate and mirabilite are prepared according to the steps in Example 1.

[0064] Comparative Example 4 Similar to the steps of Example 1, the only difference is that the amount of 98% concentrated sulfuric acid used during acidification roasting is 900 g, the pH of the obtained leaching solution is determined to be 1.2, and the leaching rate of lithium in the leaching solution is calculated. At the same time, lithium carbonate, nano-kaolinite, iron red, gypsum, ceramic aggregate and mirabilite are prepared according to the steps in Example 1.

[0065] Comparative Example 5 Similar to the steps of Example 1, the only difference is that the lithium ore used has a Li2O content of 3.5%, the Li2O content of the obtained leaching solution and the Li2O content of the leaching residue are determined, and the leaching rate of Li2O in the leaching solution is calculated. At the same time, lithium carbonate, nano-kaolinite, iron red, gypsum, ceramic aggregate and mirabilite are prepared according to the steps in Example 1.

[0066] Comparative Example 6 Similar to the procedure of Example 1, the only difference is that the used lithium concentrate contains only 50% of -100 mesh content. The obtained leaching solution is measured for Li2O content and the leaching rate of Li2O in the leaching solution is calculated. Meanwhile, lithium carbonate, nano-kaolin, iron red, gypsum, ceramsite aggregate and mirabilite are prepared according to the procedure of Example 1.

[0067] Comparative Example 7 Similar to the procedure of Example 1, the only difference is that the concentration of NaOH added in the leaching solution is 20 mol / L and the aging time is 20 min. Lithium carbonate, nano-kaolin, iron red, gypsum, ceramsite aggregate and mirabilite are prepared according to the procedure of Example 1.

[0068] The Li2O content of the leaching solution of Comparative Examples 1-7 is measured, and the results are shown in Table 3. The lithium carbonate, nano-kaolin, iron red, gypsum, ceramsite aggregate and mirabilite products prepared in the comparative examples are tested by XRD, SEM, chemical multi-element analysis and other tests to analyze the differences with Example 1.

[0069] Table 3 Li2O leaching rate of the leaching solution in each comparative example

[0070] The lithium carbonate, nano-kaolin, ceramsite, tantalum-niobium concentrate and mirabilite products prepared in Comparative Example 1 are similar to the examples, and only the SO3 content in the gypsum product is 42.21%, and the color is yellowish. It is indicated that the free Fe 3+Fe(OH)3precipitate is formed and adsorbed on the gypsum, affecting the purity of the gypsum. In Comparative Example 2, the prepared lithium carbonate, nanometer highland clay, iron red, gypsum, ceramsite aggregate and mirabilite products are similar to those of Example 1, but lack the tantalum-niobium concentrate product, and the ceramsite aggregate has low added value and fails to achieve high value and resource utilization. In Comparative Example 3, the liquid-solid ratio of the leaching treatment is 5:1, and the prepared lithium carbonate, nanometer highland clay, iron red, gypsum, ceramsite aggregate, tantalum-niobium concentrate and mirabilite products are similar to those of Example 1. Although the liquid-solid ratio of the leaching treatment has little effect on the product performance, it has some effect on the leaching rate of lithium, but causes difficulties in the subsequent lithium extraction process, especially the evaporation and concentration process, which increases the workload and cost. In Comparative Example 4, the prepared lithium carbonate, nanometer highland clay, iron red, gypsum, ceramsite aggregate, tantalum-niobium concentrate and mirabilite products are similar to those of Example 1, but the amount of sulfuric acid is increased, resulting in an increase in the amount of NaOH and the yield of gypsum products. In Comparative Example 5, the prepared lithium carbonate, iron red, gypsum, ceramsite aggregate, tantalum-niobium concentrate and mirabilite products are similar to those of Example 1, but the purity of the prepared nanometer highland clay is decreased. From XRD analysis, in addition to the nanometer highland clay phase, there are also orthoclase, albite and quartz phases. This is because the lithium spodumene raw material contains orthoclase, albite and quartz phases, which cannot be converted by high-temperature calcination and leaching reaction. In Comparative Example 6, the prepared lithium carbonate, iron red, gypsum, ceramsite aggregate, tantalum-niobium concentrate and mirabilite products are similar to those of Example 1, but the purity of the prepared nanometer highland clay is decreased. From XRD analysis, in addition to the nanometer highland clay phase, there is also a lithium spodumene phase. This is because the particle size is too coarse, and the calcination is not complete, so part of the α-lithium spodumene is not converted to β-lithium spodumene, resulting in a decrease in the leaching rate of lithium, and the purity of the prepared nanometer highland clay is not high. In Comparative Example 7, the prepared lithium carbonate, nanometer highland clay, ceramsite aggregate, tantalum-niobium concentrate and mirabilite products are similar to those of Example 1, but the purity of the iron red and gypsum is not high. This is mainly because the high concentration of NaOH in the leaching solution causes uneven reaction, resulting in serious inclusion of Fe(OH)3precipitate, and part of the precipitate is precipitated into the prepared gypsum, causing the gypsum to become dark in color and low in purity.

[0071] As can be seen from the above, the example of the present application provides a lithium spodumene concentrate resource treatment method, which comprises: high-temperature calcination and acid roasting of lithium spodumene, mixing the obtained lithium spodumene acid clinker with water for water leaching treatment, and filtering to obtain lithium-removed filter residue and leaching solution; the leaching solution is sequentially subjected to iron removal and desulfurization treatment to obtain iron red, gypsum and lithium-containing purified solution, the lithium-containing purified solution is treated to obtain lithium carbonate, mirabilite and sodium precipitation mother liquor, the sodium precipitation mother liquor is returned to the process for reuse, and the lithium-removed filter residue is subjected to resource treatment to obtain water glass, nanometer highland clay, ceramsite aggregate and tantalum-niobium concentrate. At the same time, during the treatment process, the purity and yield of the products are as high as possible by optimizing the process and dosage, realizing lithium slag-free treatment process, and increasing the added value of the products.

[0072] The above merely describes the preferred embodiments of the present application, and is not used to limit the present application. Any modification, equivalent replacement, improvement, etc. made within the spirit and principle of the present application shall be included in the protection scope of the present application.

Claims

1. A method for recycling spodumene concentrate, characterized in that: It includes: S1: wet-processing spodumene acid clinker to obtain lithium-removed residue and leachate; S2: performing a neutralization treatment and a desulfurization treatment on the leachate in sequence to obtain a lithium-containing purified liquid; S3: subjecting the lithium-containing purified liquid to lithium extraction treatment to obtain lithium salt.

2. The method according to claim 1, characterized in that The leachate is processed by the following steps: S1: wet-processing spodumene acid clinker to obtain lithium-removed residue and leachate; S2: After neutralizing the leachate, filtering and separating to obtain a deironing filtrate and a filter residue, and calcining the filter residue to obtain iron oxide red; desulfurizing the deironing filtrate, and filtering and separating to obtain gypsum and lithium-containing purified liquid; S3: The lithium-containing purified liquid is subjected to lithium extraction treatment, and lithium salt and precipitation mother liquor are obtained by filtration and separation, and the precipitation mother liquor is then separated to obtain Glauber's salt and sodium precipitation mother liquor.

3. The method according to claim 1 or 2, characterized in that The delithiation filter residue is subjected to the following steps: S4: performing magnetic separation on the delithiation filter residue to obtain magnetic material and non-magnetic material; S5: performing gravity separation on the magnetic material to obtain tantalum-niobium concentrate and ceramsite aggregate; S6: subjecting the non-magnetic material to hydrothermal treatment to obtain a water glass solution and nano-kaolinite.

4. The method according to claim 1 or 2, characterized in that The spodumene acid clinker is prepared by subjecting spodumene concentrate to crystallization roasting and acidification roasting; Preferably, the Li2O content in the spodumene concentrate is ≥4.0%, and the -100 mesh content in the spodumene concentrate is ≥70%; Preferably, the crystallization roasting temperature is 1100-1300° C., and the time is 0.5-2 h; the acidification roasting temperature is 200-300° C., and the time is 30-60 min; the concentration of concentrated sulfuric acid used in the acidification roasting is 95-98%, and the amount of concentrated sulfuric acid used is 1.3-1.5 times the theoretical amount.

5. The method according to claim 1 or 2, characterized in that In step S1, the wet treatment comprises: mixing the spodumene acid clinker with water for leaching treatment, and subjecting the leached slurry to filter pressing, washing and filtering to obtain a delithiation filter residue and a leachate; Preferably, the liquid-to-solid ratio of the leaching treatment is 0.8:1~3:1, the temperature is 30~80°C, and the time is 0.5~2.0h; the leaching slurry is washed 2-3 times; preferably, the Li2O content in the delithiation filter residue is ≤0.2%, and the SO3 content is ≤0.3%; the pH of the leachate is 0.7~2.

5.

6. The method according to claim 1 or 2, characterized in that In step S2, the neutralization treatment includes: slurrying, aging, filtering, drying and calcining the leachate to obtain iron removal filtrate and iron oxide red; Preferably, after adding NaOH to the leachate for slurry adjustment, the pH value is 3.5-4.5, the NaOH concentration is 2-10 mol / L, the aging time is 0.5-2.0 h, the filtration is performed using a vacuum processor or a filter press, the drying temperature is 100-200° C., the drying time is 2-10 h, the calcination temperature is 450-550° C., and the calcination time is 0.5-4 h; preferably, the Fe2O3 content in the iron red is ≥98%.

7. The method according to claim 1 or 2, characterized in that In step S2, the desulfurization treatment includes: adding quicklime to the iron removal filtrate to perform purification and impurity removal, aging, filtering, and drying to obtain gypsum and purified liquid; Preferably, the pH of the iron removal filtrate after adding quicklime for purification and impurity removal is 10-12, and the aging time is 0.2-2.0 h. Preferably, the SO3 content in the gypsum is ≥43%, and the Li2O content is ≤0.2%.

8. The method according to claim 1 or 2, characterized in that In step S3, the lithium extraction process includes: evaporating and concentrating the lithium-containing purified liquid, carbonizing, centrifuging, and drying to obtain lithium carbonate and a precipitation mother liquor, and freezing and separating the precipitation mother liquor to obtain sodium sulfate and a sodium precipitation mother liquor; Preferably, the lithium-containing purified liquid has a Li2O concentration of 50-80 g / L after evaporation and concentration, a soda ash concentration of 10-30%, a freezing separation temperature of -5--10°C, a lithium carbonate content of ≥99.5%, a Na2SO4 content in the sodium sulfate of ≥44.0%, and the sodium precipitation mother liquor is returned to the evaporation and concentration and the soda ash solution preparation process.

9. The method according to claim 3, characterized in that In steps S4 and S5, the magnetic separation and gravity separation treatment includes: subjecting the delithiation filter residue to a weak magnetic + strong magnetic magnetic separation process to obtain magnetic material and non-magnetic material, subjecting the magnetic material to gravity separation to prepare tantalum-niobium concentrate and gravity separation tailings, and filtering and drying the gravity separation tailings to obtain ceramsite aggregate; Preferably, the intensity of the weak magnetic separation is 3000~6000Gs, the intensity of the strong magnetic separation is 1.0~1.8T, and the concentration of the magnetic separation pulp is 3~30%; the gravity separation equipment includes one or more combinations of spiral chute, shaking table, and felt machine.

10. The method according to claim 3, characterized in that In step S6, the hydrothermal treatment includes: controlling the hydrothermal reaction temperature to 200-300° C., the pressure to 1.0-5.0 MPa, the hydrothermal reaction time to 2-12 h, and the amount of sodium hydroxide used to be 1.1-1.3 times the theoretical amount.

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