Green process method for extracting alkali metal from composite ore

Abstract

The invention discloses a green process method for extracting alkali metal from composite ore, and relates to the field of metal extraction.The green process method comprises the following steps that S1, aluminosilicate composite ore containing alkali metal and ammonium carbonate are mixed and roasted, and activated clinker is obtained; s2, mixing the activated clinker with an ammonium citrate solution, carrying out a leaching reaction at 80-95 DEG C, and carrying out solid-liquid separation after the reaction to obtain a leaching solution and leaching residues; according to the green process method for extracting the alkali metal from the composite ore, through systematic coupling of a low-temperature ammonium carbonate activation-ammonium citrate complexing leaching-manganese system / ammonium phosphomolybdate specific adsorption-bipolar membrane electrodialysis closed-loop regeneration technology chain, a synergistic interaction effect is formed; the activation step significantly improves the leaching efficiency of the subsequent mild leaching agent.

Description

Technical Field

[0001] This invention relates to metal extraction technology, specifically to a green process for extracting alkali metals from composite ores. Background Technology

[0002] Lithium, rubidium, cesium, and other alkali metals are indispensable key raw materials for strategic emerging industries such as new energy, new materials, and high-end manufacturing. Currently, the mainstream processes for extracting these metals from ores (such as the sulfuric acid process, sulfate roasting process, and limestone process) generally suffer from high energy consumption (usually requiring high-temperature roasting above 800℃), heavy pollution (generating large amounts of waste gas containing fluorine, sulfur, and chlorine, high-salinity wastewater, and radioactive waste residue, which are difficult and costly to treat), poor selectivity (low comprehensive recovery rate for symbiotic rare alkali metals such as rubidium and cesium, resulting in serious resource waste), and poor economic efficiency (poor adaptability to low-grade ores with complex compositions, and large reagent consumption).

[0003] Therefore, developing a new process that can adapt to complex ores, achieve efficient and selective extraction of alkali metals under mild conditions, and is clean, low-carbon, and economically feasible throughout the entire process has become an urgent technical challenge to be solved in this field. Summary of the Invention

[0004] The purpose of this invention is to provide a green process for extracting alkali metals from complex ores, in order to solve the problems of high energy consumption, heavy pollution, poor selectivity, and insufficient adaptability to complex ores in existing processes for extracting alkali metals from complex ores.

[0005] To achieve the above objectives, the present invention provides the following technical solution: a green process method for extracting alkali metals from composite ores, comprising the following steps:

[0006] S1. Alkali metal-containing aluminosilicate composite ore is mixed with ammonium carbonate and roasted to obtain activated clinker;

[0007] S2. The activated clinker is mixed with ammonium citrate solution and leached at 80-95°C. After the reaction, solid-liquid separation is performed to obtain leachate and leaching residue.

[0008] S3. The leachate is sequentially passed through a manganese-based lithium ion sieve adsorbent and an ammonium phosphomolybdate adsorbent for selective adsorption and desorption, respectively, to obtain a purified lithium solution and a purified rubidium / cesium solution.

[0009] S4. The remaining mother liquor after adsorption and separation in step S3 is subjected to bipolar membrane electrodialysis to obtain a regenerated citric acid solution in the acid chamber and an ammonium hydroxide solution in the alkali chamber.

[0010] The citric acid solution obtained from the regeneration is returned to step S2 for the preparation of ammonium citrate solution.

[0011] Further, in step S1, the alkali metal-containing aluminosilicate composite ore contains at least one of lepidolite, petalite, cesium garnet, or rubidium-containing feldspar; the particle size of the composite ore after crushing is 100-200 mesh; the mass ratio of the composite ore to ammonium carbonate is 1:(0.05-0.15); and the roasting is carried out at 350-450℃ for 30-90 minutes.

[0012] Further, in step S2, the concentration of the ammonium citrate solution is 0.5-2.0 mol / L; the liquid-to-solid ratio of the activated clinker to the ammonium citrate solution is (3-5):1 L / kg; and the leaching reaction is carried out at a stirring speed of 300-500 rpm for 60-180 minutes.

[0013] Further, in step S3, the manganese-based lithium-ion sieve adsorbent is an HMnO type or λ-MnO2 type lithium-ion sieve; the lithium adsorption flow rate is 2-5 BV / h, and the desorbent is 0.5-1.5 mol / L hydrochloric acid; the ammonium phosphomolybdate adsorbent is (NH4)3PMo 12 O 40 The flow rate for adsorbing rubidium / cesium is 1-3 BV / h, and the desorbent is 1.0-3.0 mol / L hydrochloric acid.

[0014] Furthermore, in step S4, the current density of the bipolar membrane electrodialysis treatment is 100-300 A / m. 2 The endpoint of the operation is controlled when the pH value of the remaining mother liquor drops to 1.5-2.5.

[0015] Furthermore, the method also includes step S5: washing and drying the leaching residue obtained in step S2, and using the resulting material as a building material additive or soil conditioner.

[0016] S1. Co-activation roasting: The alkali metal-containing aluminosilicate composite ore is crushed to 100-200 mesh and mixed evenly with the activator ammonium carbonate at a mass ratio of 1:(0.05-0.15). The mixture is roasted at 350-450℃ for 30-90 minutes to loosen the ore structure and transform the alkali metal elements into a more easily leached form.

[0017] S2, Green Complex Leaching: The activated clinker obtained in step S1 is mixed with an ammonium citrate solution with a concentration of 0.5-2.0 mol / L at a liquid-to-solid ratio of (3-5):1 (L / kg), and placed in a reactor. Leaching is carried out at 80-95℃ and a stirring speed of 300-500 rpm for 60-180 minutes. Citrate ions form stable complexes with alkali metal ions, promoting their transfer from the solid phase to the liquid phase. After leaching, solid and liquid are separated to obtain a leachate rich in alkali metals and a leachate residue mainly composed of silicon and aluminum.

[0018] S3. Selective Adsorption Separation and Purification: First, the leachate obtained in step S2 is passed through an adsorption column packed with a manganese-based lithium-ion sieve adsorbent (such as HMnO or λ-MnO2 type) at a flow rate of 2-5 BV / h to selectively adsorb lithium ions. After adsorption saturation, it is desorbed with 0.5-1.5 mol / L dilute hydrochloric acid at a flow rate of 1-3 BV / h to obtain a high-purity lithium chloride solution. Subsequently, the delithiated effluent is passed through an adsorption column packed with ammonium phosphomolybdate adsorbent at a flow rate of 1-3 BV / h to selectively adsorb rubidium and cesium ions. After adsorption saturation, it is desorbed with 1.0-3.0 mol / L dilute hydrochloric acid to obtain a chloride solution rich in rubidium and cesium, which can be further separated by stepwise crystallization or extraction.

[0019] S4. Electrochemical Regeneration and Resource Recovery of Leaching Agent: The remaining mother liquor (mainly containing unused ammonium citrate and a small amount of impurity ions) after adsorption separation in step S3 is introduced into a bipolar membrane electrodialysis system. Electrodialysis is carried out at a current density of 100-300 A / m² until the pH of the mother liquor chamber drops to 1.5-2.5. During this process, the bipolar membrane decomposes water into H₂. + and OH - H + It enters the acid chamber and combines with citrate ions, regenerating to obtain a citric acid solution; OH - The leaching agent enters the alkali chamber and combines with ammonium ions to obtain an ammonium hydroxide solution. The regenerated citric acid solution can be returned to step S2 to prepare the leaching agent, and the ammonium hydroxide solution can be used for process pH adjustment or as a byproduct. This step achieves the recycling of the leaching agent and the recovery of ammonium, avoiding the generation of nitrogenous and high-salinity wastewater at the source.

[0020] S5. Resource utilization of leaching residue: The leaching residue obtained in step S2, after washing and drying, is mainly composed of amorphous silicon aluminum oxide, which has a large specific surface area and high activity. It can be used directly or after simple modification as cement admixture, ceramic raw material or soil conditioner to achieve zero discharge of solid waste.

[0021] Compared with existing technologies, the green process for extracting alkali metals from composite ores provided by this invention has the following advantages:

[0022] The systematic coupling of the low-temperature ammonium carbonate activation-ammonium citrate complexation leaching-manganese / ammonium phosphomolybdate specific adsorption-bipolar membrane electrodialysis closed-loop regeneration technology chain creates a synergistic effect. The activation step significantly improves the leaching efficiency of subsequent mild leaching agents; the adsorption separation step is highly targeted, with high purity and yield; and the membrane electrolysis regeneration step achieves greening and economic efficiency, solving the industry pain point of high cost of environmentally friendly leaching agents.

[0023] Excellent environmental friendliness: The entire process uses green chemicals such as ammonium citrate, with no large consumption or discharge of strong acids and alkalis; the leaching temperature is low and energy consumption is small; bipolar membrane electrodialysis realizes the recycling of leaching agent and water, and the amount of wastewater generated is extremely small; the leaching residue is harmless and resource-efficient. Attached Figure Description

[0024] To more clearly illustrate the technical solutions in the embodiments of this application or the prior art, the drawings used in the embodiments will be briefly described below. Obviously, the drawings described below are merely some embodiments recorded in this invention, and those skilled in the art can obtain other drawings based on these drawings.

[0025] Figure 1 This is a schematic diagram of the overall process preparation flow provided for an embodiment of the present invention. Detailed Implementation

[0026] To enable those skilled in the art to better understand the technical solution of the present invention, the present invention will be further described in detail below with reference to the accompanying drawings.

[0027] As attached Figure 1 As shown:

[0028] Example 1:

[0029] Raw material: A low-grade lithium mica ore in Jiangxi Province, composition (wt%): Li2O 1.2%, Rb2O 0.5%, Cs2O 0.15%.

[0030] Steps and parameters:

[0031] S1 synergistic activation: Crush the ore to 100 mesh, mix it with ammonium carbonate at a mass ratio of 1:0.05, place it in a muffle furnace, and calcine it at 350℃ for 90 minutes.

[0032] S2 Green Leaching: Take 100g of activated calcined material and add 300 mL of 0.5 mol / L ammonium citrate solution (liquid-solid ratio 3:1). In a 1L glass reactor, maintain the temperature at 80℃ and the stirring speed at 300 rpm for 180 minutes. After filtration, the Li content was measured. + The leaching rate was 86.5%, Rb + The leaching rate was 85.2%, Cs + The leaching rate was 84.1%.

[0033] S3 Adsorption Separation: The leachate was passed through a λ-MnO2 adsorption column (50 mL) at a flow rate of 2 BV / h to adsorb lithium, and desorbed with 0.5 mol / L HCl at a flow rate of 1 BV / h. After delithiation, the solution was passed through an ammonium phosphomolybdate column at a flow rate of 1 BV / h to adsorb rubidium and cesium, and desorbed with 1.0 mol / L HCl.

[0034] S4 membrane electrolytic regeneration: The remaining mother liquor is pumped into the bipolar membrane electrodialysis unit (BP-ED-01) and operated at a current density of 100A / m² until the pH in the feed chamber drops to 2.5, and citric acid solution is obtained through regeneration.

[0035] S5 slag resource utilization: After drying, the leaching slag was tested and found to have an activity index of 68% after 7 days, making it suitable for use as a building material auxiliary material.

[0036] Example 2:

[0037] Raw material: The same low-grade lithium mica ore from Jiangxi Province as in Example 1.

[0038] Steps and parameters:

[0039] S1 synergistic activation: The ore is crushed to 150 mesh and mixed with ammonium carbonate at a mass ratio of 1:0.1. The mixture is placed in a muffle furnace and roasted at 400°C for 60 minutes.

[0040] S2 Green Leaching: Take 100g of activated calcined material and add 400 mL of 1.2 mol / L ammonium citrate solution (liquid-solid ratio 4:1). In a 1L glass reactor, maintain the temperature at 90℃ and the stirring speed at 400 rpm for 120 minutes. After filtration, the Li content was measured. + The leaching rate was 93.8%, Rb + The leaching rate was 92.7%, Cs + The leaching rate was 91.5%.

[0041] S3 Adsorption Separation: The leachate was passed through a λ-MnO2 adsorption column at a flow rate of 3 BV / h to adsorb lithium, and then desorbed with 0.8 mol / L HCl at a flow rate of 2 BV / h to obtain a LiCl solution with a purity >99.5%. The delithiated solution was then passed through an ammonium phosphomolybdate column at a flow rate of 2 BV / h to adsorb rubidium and cesium, and then desorbed with 2.0 mol / L HCl.

[0042] S4 membrane electrolysis regeneration: The remaining mother liquor is pumped into the bipolar membrane electrodialysis unit at a current density of 200 A / m 2 The process continues until the pH in the feed chamber drops to 2.0, at which point a citric acid solution of approximately 1.0 mol / L is obtained through regeneration, which can be directly reused.

[0043] S5 slag resource utilization: After drying, the leaching slag was tested and found to have an activity index of 78% after 7 days, making it an excellent cement admixture.

[0044] Comparative experiment: Compared with Example 1, the alkali metal leaching rate under the optimal parameters increased by an average of about 7 percentage points, the leaching time was shortened by 1 / 3, the regeneration efficiency was higher, and the overall effect was significantly better.

[0045] Example 3:

[0046] Raw material: The same low-grade lithium mica ore from Jiangxi Province as in Example 1.

[0047] Steps and parameters:

[0048] S1 synergistic activation: The ore is crushed to 200 mesh and mixed with ammonium carbonate at a mass ratio of 1:0.15. The mixture is placed in a muffle furnace and roasted at 450°C for 30 minutes.

[0049] S2 Green Leaching: Take 100g of activated calcined material and add 500 mL of 2.0 mol / L ammonium citrate solution (liquid-solid ratio 5:1). In a 1L glass reactor, maintain the temperature at 95℃ and the stirring speed at 500 rpm for leaching for 60 minutes. After filtration, the Li content was measured. + The leaching rate was 90.1%, Rb + The leaching rate was 88.9%, Cs + The leaching rate was 87.3%.

[0050] S3 Adsorption Separation: The leachate was passed through a λ-MnO2 adsorption column at a flow rate of 5 BV / h to adsorb lithium, and then desorbed with 1.5 mol / L HCl at a flow rate of 3 BV / h. The delithiated solution was then passed through an ammonium phosphomolybdate column at a flow rate of 3 BV / h to adsorb rubidium and cesium, and then desorbed with 3.0 mol / L HCl.

[0051] S4 membrane electrolysis regeneration: The remaining mother liquor is pumped into the bipolar membrane electrodialysis unit at a current density of 300 A / m 2 The process is repeated until the pH in the feed chamber drops to 1.5, at which point a high concentration of citric acid solution is obtained through rapid regeneration.

[0052] S5 slag resource utilization: After drying, the leaching slag was tested and found to have an activity index of 72% after 7 days, making it suitable for use as a building material auxiliary material.

[0053] Effect analysis: Compared with Example 2, this example achieved a higher leaching rate in a shorter leaching time, but the reagent concentration and energy consumption increased, demonstrating that the process is still effective and feasible at the upper limit of the parameters.

[0054] Comparative Example 1:

[0055] One kilogram of lepidolite ore from the same batch was mixed with concentrated sulfuric acid at a mass ratio of 1:0.3 and roasted in a rotary kiln at 850°C for 2 hours. After cooling, the roasted ore was quenched in water and leached with stirring for 1 hour. After filtration, the leachate was adjusted for pH and impurities with limestone, concentrated, and then sodium carbonate was added to precipitate lithium carbonate. Results: The lithium leaching rate was 88.5%, but rubidium and cesium volatilized in large quantities into the flue gas during the roasting stage, resulting in extremely low recovery rates. The process generated a large amount of fluorinated sulfuric acid waste gas, requiring alkaline washing, and produced 8 kg of sodium sulfate slag and highly acidic wastewater, resulting in high treatment costs.

[0056] Comparative Example 2:

[0057] One kg of lepidolite from the same batch was mixed with calcium chloride at a mass ratio of 1:0.4 and roasted at 850℃ for 1.5 hours. The roasted ore was then leached with 3 mol / L hydrochloric acid at 80℃ for 2 hours at a liquid-to-solid ratio of 4:1. The results showed that the leaching rates for lithium, rubidium, and cesium were 90.2%, 78.5%, and 75.6%, respectively. However, the roasting process was energy-intensive, the leachate contained high concentrations of impurities such as calcium and aluminum, separation was difficult, and a large amount of calcium chloride waste residue was generated.

[0058] Comparative Example 3:

[0059] One kg of lepidolite ore from the same batch was leached directly under the same conditions as in Example 2 (1.2 mol / L ammonium citrate, 90°C, liquid-to-solid ratio 4:1, 120 minutes) without ammonium carbonate activation roasting. Results: The leaching rates of lithium, rubidium, and cesium were only 42.1%, 39.8%, and 38.5%, respectively. This comparison directly demonstrates that without the specific activation step of this invention, even using the same green leaching agent, efficient alkali metal extraction cannot be achieved.

[0060] Table 1: Comparison of Data Detection Between Examples and Comparative Examples

[0061]

[0062] The above description is only a preferred embodiment of the present invention, but the scope of protection of the present invention is not limited thereto. Any equivalent substitutions or modifications made by those skilled in the art within the scope of the technology disclosed in the present invention, based on the technical solution and inventive concept of the present invention, should be covered within the scope of protection of the present invention.

Claims

1. A green process for the extraction of alkali metals from complex ores, characterized in that, The method comprises the following steps: S1, mixing and roasting alkali metal-containing silicate-aluminate composite ore with ammonium carbonate to obtain activated clinker; S2, mixing the activated clinker with an ammonium citrate solution, carrying out leaching reaction at 80-95℃, and then carrying out solid-liquid separation to obtain leaching liquid and leaching residue; S3, sequentially passing the leaching liquid through manganese lithium ion sieve adsorbent and ammonium phosphomolybdate adsorbent for selective adsorption and desorption to obtain purified lithium liquid and purified rubidium / cesium liquid, respectively; S4, carrying out bipolar membrane electrodialysis treatment on the remaining mother liquor after adsorption and separation in step S3 to obtain regenerated citric acid solution in the acid chamber and ammonium hydroxide solution in the base chamber; The regenerated citric acid solution is returned to step S2 for preparing ammonium citrate solution.

2. A green process for extraction of alkali metals from complex ores as claimed in claim 1 wherein, In step S1, the alkali metal-containing silicate-aluminate composite ore comprises at least one of lepidolite, petalite, pollucite or rubidium-containing feldspar; the particle size of the crushed composite ore is 100-200 mesh; the mass ratio of the composite ore to ammonium carbonate is 1:(0.05-0.15); and the roasting is carried out at 350-450℃ for 30-90 minutes.

3. A green process for extraction of alkali metals from complex ores as claimed in claim 1 wherein, In step S2, the concentration of the ammonium citrate solution is 0.5-2.0 mol / L; the liquid-solid ratio of the activated clinker to the ammonium citrate solution is (3-5):1 L / kg; and the leaching reaction is carried out at a stirring speed of 300-500 rpm for 60-180 minutes.

4. A green process for extraction of alkali metals from complex ores as claimed in claim 1 wherein, In step S3, the manganese-based lithium ion sieve adsorbent is HMnO type or λ-MnO2 type lithium ion sieve; the flow rate of adsorbing lithium is 2-5 BV / h, the desorbent is 0.5-1.5 mol / L hydrochloric acid; the ammonium phosphomolybdate adsorbent is (NH4)3PMo 12 O 40 ; the flow rate of adsorbing rubidium / cesium is 1-3 BV / h, and the desorbent is 1.0-3.0 mol / L hydrochloric acid.

5. A green process for extraction of alkali metals from complex ores as claimed in claim 1 wherein, In step S4, the current density of the bipolar membrane electrodialysis treatment is 100-300 A / m 2 The running end control is that the pH value of the residual mother liquor is reduced to 1.5-2.

5.

6. A green process for extraction of alkali metals from complex ores as claimed in claim 1 wherein, The method further comprises step S5: washing and drying the leaching residue obtained in step S2, and using the obtained material as a building material additive or soil conditioner.

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