A method for the synergistic extraction of aluminum, lithium, gallium and scandium from fly ash

Abstract

The present invention belongs to the technical field of fly ash resource recycling and utilization, and relates to a method for the synergistic extraction of aluminum, lithium, gallium, and scandium from fly ash. The method comprises the following steps: (S1) pretreatment: adding fly ash to an aqueous solution of an inorganic base, subjecting it to microwave irradiation, and then pressurizing and activating it to obtain a pretreated product; (S2) graded acid leaching: subjecting the pretreated product to primary acid leaching in hydrochloric acid to obtain leachate I and a Si-rich residue I; subjecting the Si-rich residue I to secondary acid leaching in a mixed solution of hydrochloric acid and choline chloride-urea to obtain leachate II and a Si-rich residue II; and (S3) separation and purification: comprising selective adsorption of Ga / Sc / Fe by ion exchange resin, Ca removal and Al / Li separation, step-by-step resin elution, Ga / Sc extraction and separation, Ga stripping, and electrolysis. The present invention achieves efficient synergistic extraction and full resource utilization of aluminum, lithium, gallium, and scandium from fly ash, and the entire process has low energy consumption.

Description

Technical Field

[0001] The invention belongs to the technical field of fly ash resource recycling and utilization, and particularly relates to a method for the coordinated extraction of aluminum, lithium, gallium and scandium from fly ash. Background Art

[0002] Fly ash is an industrial solid waste generated by thermal power plants. With the development of the coal-fired power industry, fly ash production has increased rapidly annually, making its utilization a pressing environmental and resource-intensive issue. Fly ash, which typically contains valuable metals such as silicon, aluminum, iron, gallium, lithium, and rare earth elements, is a valuable secondary resource. Recovering these metals from fly ash is of strategic importance for reducing environmental pollution and alleviating the current shortage of metal minerals. However, due to its unique physical and chemical properties, the metal elements within fly ash are difficult to separate and leach, especially the high-value lithium and gallium. Activation of the fly ash is generally required. Experimental and practical applications have demonstrated that calcining the fly ash with an activator is an effective method for activating the fly ash. Metal ions in the calcined product can be effectively extracted under acid leaching conditions. Although extensive research has been conducted on the extraction of metal elements such as aluminum, gallium, lithium, and rare earth elements from fly ash, low recovery rates and high costs have hindered the development of a large-scale industrial production process.

[0003] CN103382531B discloses a method for enriching gallium from the mother liquor of the alumina production process using high-aluminum fly ash. This method specifically addresses the enrichment and extraction of gallium from the fly ash aluminum extraction process. Using the mother liquor containing a low gallium concentration from the high-aluminum fly ash alumina production process as raw material, ion exchange and process intensification techniques are employed to obtain a gallium-enriched solution through adsorption and elution. After further treatment, metallic gallium is obtained through electrolysis. This method achieves the synergistic extraction and utilization of aluminum, silicon, and gallium. However, it suffers from disadvantages such as high sintering temperature, high energy consumption, and large slag volumes. Furthermore, it cannot achieve the synergistic extraction of multiple metals, and the metal recovery rate needs to be improved.

[0004] CN116377225B discloses a method and intelligent equipment for recycling fly ash. The method comprises acid leaching the calcined activated product to obtain an acid leaching solution and a Si-rich residue; passing the acid leaching solution through a first ion exchange resin designed to adsorb Ga and Fe to obtain an exchange solution rich in Al and Li; removing Ca from the Al- and Li-rich exchange solution, and adjusting the pH of the Ca-removed solution to obtain an Al-rich residue and a Li-rich liquid; eluting the Ga- and Al-adsorbed first ion exchange resin to obtain an eluate rich in Ga and Fe; and removing Fe from the Ga- and Fe-rich eluate using a second ion exchange resin to obtain a Ga-rich liquid. This method, targeting solid waste fly ash, synergistically separates and enriches multiple metals in a single process flow; however, the invention utilizes separate calcination and activation steps, resulting in high energy consumption. Summary of the Invention

[0005] In view of the problem that the existing technology cannot achieve efficient synergistic extraction of multiple metals from fly ash and the production process has high energy consumption, the present invention provides a method for the synergistic extraction of aluminum, lithium, gallium and scandium from fly ash, which realizes the efficient synergistic extraction and full resource utilization of aluminum, lithium, gallium and scandium, and the energy consumption of the entire process is low.

[0006] To achieve the above object, the present invention adopts the following technical solutions:

[0007] A method for the synergistic extraction of aluminum, lithium, gallium and scandium from fly ash, comprising the following steps:

[0008] (S1) Pretreatment: Fly ash is added to an aqueous solution of an inorganic base and subjected to microwave irradiation, followed by pressure activation, and filtered to obtain a pretreated product;

[0009] (S2) graded acid leaching: the pretreated product is added to hydrochloric acid A for primary acid leaching to obtain leachate I and Si-rich residue I; the Si-rich residue I is added to a mixture of hydrochloric acid B and choline chloride-urea at 55-65°C for secondary acid leaching to obtain leachate II and Si-rich residue II;

[0010] (S3) Separation and purification: The product after graded acid leaching is separated and purified, including the following steps:

[0011] (S301) Ion exchange resin adsorption: after evaporating the choline chloride-urea from the leachate II, the leachate is mixed with the leachate I to obtain a combined leachate, and the combined leachate is passed through an ion exchange resin that selectively adsorbs Ga, Sc, and Fe to obtain a loaded resin rich in Ga / Sc / Fe and an effluent rich in Al / Li / Ca;

[0012] (S302) Ca removal and Al / Li separation: After removing Ca from the Al / Li / Ca-rich effluent, the pH is adjusted to 5.5-6.5 to obtain an Al-rich residue and a Li-rich liquid;

[0013] (S303) stepwise elution of the resin: the loaded resin rich in Ga / Sc / Fe obtained in step (S301) is eluted with an aqueous oxalic acid solution to obtain an eluate rich in iron; and then eluted with hydrochloric acid to obtain an eluate rich in Ga / Sc;

[0014] (S304) Ga / Sc extraction separation: Concentrated hydrochloric acid is added to the Ga / Sc-rich eluate until the concentration of hydrochloric acid in the system is 5-7 mol / L, and then extracted with an extractant to separate the oil and water phases to obtain a Ga-rich oil phase and a Sc-rich water phase;

[0015] (S305) Ga stripping and electrolysis: The Ga-rich oil phase is stripped using dilute acid to separate the oil-water phase to obtain a Ga-rich water phase, which is then subjected to pulse electrolysis to obtain high-purity Ga.

[0016] Furthermore, in step (S1), the particle size of the fly ash is 1~10 μm; the inorganic base is KOH or NaOH, preferably NaOH, and the concentration of the aqueous solution of the inorganic base is 10~20wt%; the solid-liquid ratio of the fly ash to the aqueous solution of the inorganic base is 1g:3~4mL; the conditions of the microwave irradiation are: frequency 2~3GHz, power 5~10kW, time 20~40min; the conditions of the pressurized activation are: pressure 0.5~1.5MPa, preferably 0.5~1MPa, temperature 150~180℃, time 1~3h.

[0017] Furthermore, in step (S2), the volume ratio of hydrochloric acid B to choline chloride-urea is 3-4:1, preferably 3:1; and the molar ratio of choline chloride to urea in choline chloride-urea is 1:0.5-1.5, preferably 1:1-1.5. Choline chloride-urea is a choline-based deep eutectic solvent and an ionic liquid. The inventors unexpectedly discovered that the use of choline chloride-urea as an auxiliary hydrochloric acid in secondary acid leaching significantly improved the recovery rate of gallium-scandium. This is likely due to the chloride ions in choline chloride-urea destroying the silicon-oxygen bonds, that is, the silicon encapsulation structure. Urea provides coordination sites to dissolve gallium-scandium, thereby improving the gallium-scandium leaching rate and, in turn, the gallium-scandium recovery rate. However, the volume ratio of hydrochloric acid B to choline chloride-urea must be controlled within the above range. If the amount of choline chloride-urea used is too much, the viscosity of the system will be high, the mass transfer rate will be reduced, thereby affecting the leaching rate, and the salt concentration will be reduced too much, which is not conducive to the destruction of the silicate structure. If the amount of choline chloride-urea used is too low, its effective auxiliary role is limited, and the recovery rate of gallium and scandium is not significantly improved.

[0018] The choline chloride-urea in the present invention can be prepared in-house. For example, the choline chloride and urea are mixed and placed in a sealed container, and stirred at 50-60° C. until the mixture becomes uniform and transparent, thereby obtaining the choline chloride-urea.

[0019] Furthermore, in step (S2), the concentration of hydrochloric acid A is 3-4 mol / L, the solid-to-liquid ratio of the pretreated product to hydrochloric acid A is 1 g:4-5 mL, and the conditions for the primary acid leaching are: temperature 80-95°C, time 2-4 hours; the concentration of hydrochloric acid B is 5-6 mol / L, the solid-to-liquid ratio of the Si-rich residue I to the mixed solution is 1 g:4-5 mL, and the secondary acid leaching time is 1-3 hours. The primary acid leaching dissolves most of the Al and Li, and a secondary acid leaching is then performed using choline chloride-urea-assisted hydrochloric acid to dissolve the remaining Al and difficult-to-dissolve Ga and Sc. Temperature control during secondary acid leaching is critical. Choline is easily decomposed and urea is easily hydrolyzed at high temperatures (>70°C) under acidic conditions, resulting in the failure of the deep eutectic solvent. When the temperature is below 55°C, the viscosity of choline chloride-urea is too high, and the mass transfer rate is reduced, thus affecting the leaching rate. Therefore, the temperature must be strictly controlled within the range of 55-65°C during secondary acid leaching. At the same time, due to the auxiliary effect of choline chloride-urea, the secondary acid leaching temperature can achieve good leaching effect even if it is lower than the primary acid leaching temperature.

[0020] Optionally, step (S2) includes collecting the Si-rich residue I and the Si-rich residue II and calcining them at 1000-1200° C. to obtain a silicon oxide product, such as a molecular sieve.

[0021] Furthermore, the conditions for removing the choline chloride-urea by distillation in step (S301) are as follows: leachate II is subjected to vacuum distillation at 50-70°C and 5-10 kPa, and then mixed with leachate I. The purpose of the vacuum distillation is to recover the choline chloride-urea for recycling.

[0022] Furthermore, the ion exchange resin in step (S301) is a macroporous aminophosphonic acid chelating resin, such as Lewatit TP260, D851 or LSC-640. The macroporous aminophosphonic acid chelating resin can efficiently and simultaneously selectively adsorb Ga, Sc and Fe.

[0023] Furthermore, in step (S302), the conditions for removing Ca are as follows: ammonium oxalate is added to the effluent rich in Al / Li / Ca, and solid-liquid separation is performed to obtain calcium oxalate and a liquid after removing Ca. After Ca is removed, the liquid after removing Ca is rich in Al and Li, and the pH is in the range of 5.5-6.5. Al 3+ Al(OH)3 precipitation is formed, and after solid-liquid separation, Al-rich residue and Li-rich liquid are obtained, thereby separating Al and Li. The control of pH is critical here. pH>6.5 is prone to Li+ Coprecipitation losses.

[0024] Optionally, step ( S302 ) includes calcining the Al-rich residue at a temperature of 900-1200° C. to obtain aluminum oxide.

[0025] Furthermore, step (S302) includes concentrating and separating the Li-rich liquid through a nanofiltration membrane to obtain a high-purity lithium solution.

[0026] Furthermore, in step (S303), the concentration of the oxalic acid aqueous solution is 0.5-1 mol / L; the concentration of the hydrochloric acid is 3-4 mol / L. When eluting with oxalic acid aqueous solution, the selectivity of oxalic acid is related to Fe 3+ The complex is formed and eluted from the ion exchange resin. Oxalic acid is used to preferentially elute Fe 3+ , which is more selective than traditional hydrochloric acid elution.

[0027] Optionally, step (S303) further comprises evaporating and concentrating the iron-rich eluate at 70-100°C to obtain a liquid with a concentration of 20-30 wt%, which can be used to prepare iron red pigment.

[0028] Furthermore, the extractant in step (S304) is tributyl phosphate (TBP) or trioctylphosphine oxide (TOPO). 3+ The complex formed with hydrochloric acid at high acidity is selectively extracted by the extractant TBP or TOPO, thereby remaining in the oil phase.

[0029] Optionally, step (S304) further comprises extracting the Sc-rich aqueous phase using an extractant P507, stripping with hydrochloric acid, evaporating, and calcining at 900-1000° C. to obtain scandium oxide.

[0030] Furthermore, the dilute acid in step (S305) is sulfuric acid with a concentration of 0.1-0.3 mol / L; the conditions of the pulse electrolysis are: current density 50-100 A / m², duty cycle 30%-40%, frequency 15 kHz-25 kHz; and the high purity is 4N purity (99.99%).

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

[0032] 1. The present invention uses a combination of microwave irradiation and pressurized alkali activation to pretreat fly ash. Microwaves can penetrate powder and micron-sized fly ash, selectively heat the aluminum-rich phase, and accelerate the destruction reaction of the mullite structure. The pressurized alkali activation conditions further promote the dissolution of aluminum / lithium. In addition, the pretreatment of the present invention is more energy-efficient and efficient than the conventional alkali roasting process.

[0033] 2. The present invention adopts a graded acid leaching process to gradually dissociate metal ions; at the same time, choline chloride-urea assisted leaching is used in the secondary acid leaching, which significantly improves the recovery rate of gallium and scandium. The possible reason is that the chloride ions in choline chloride-urea destroy the silicon-oxygen bond, that is, destroy the silicon encapsulation structure, and urea provides coordination sites to dissolve gallium and scandium, thereby improving the gallium and scandium leaching rate, and then improving the gallium and scandium recovery rate; that is, through graded acid leaching and choline chloride-urea assisted leaching, multiple metals are synergistically dissolved.

[0034] 3. The present invention comprehensively utilizes ion exchange resin, step-by-step elution, extraction, pulse electrolysis and other processes in the separation and purification stage. In particular, when separating Ga, Sc and Fe, oxalic acid is first used to remove iron, and then hydrochloric acid is eluted and extracted under high acidity to separate Ga, Sc and Fe one by one.

[0035] 4. The present invention utilizes core processes such as microwave-pressurized alkaline activation pretreatment, graded acid leaching (choline chloride-urea assisted leaching is used in the secondary acid leaching), and high-selectivity separation technology to achieve a recovery rate of over 85% for Al, Ga, and Sc, over 90% for Li and Fe, and simultaneously achieves the recovery of Ca and Si. This means that the present invention achieves efficient synergistic extraction and full resource utilization of aluminum, lithium, gallium, and scandium from fly ash, and the entire process consumes relatively little energy, which is beneficial for reducing costs and increasing profits. BRIEF DESCRIPTION OF THE DRAWINGS

[0036] Figure 1 It is a schematic diagram of the process flow of the present invention. DETAILED DESCRIPTION

[0037] To make the purpose, technical solutions and advantages of the present invention clearer, the technical solutions of the present invention will be described in detail below. The following examples are convenient for better understanding of the present invention, but are not intended to limit the present invention. The experimental methods in the following examples are all conventional methods unless otherwise specified.

[0038] The fly ash was produced by Yangquan No. 3 Power Plant, with a particle size of about 4.5 μm. X-ray fluorescence spectrometer was used to detect the Al2O3 content in the fly ash, which was 31.53 wt%, CaO was 3.10 wt%, and Fe2O3 was 4.97 wt%. Inductively coupled plasma mass spectrometry (ICP-MS) was used to detect the Li content in the fly ash, which was 246 ppm, 40.1 ppm, and 16.9 ppm.

[0039] Choline chloride-urea was prepared in-house. The specific method was as follows: choline chloride and urea were fully mixed at a molar ratio of 1:0.5, 1:1, and 1:1.5, respectively, to form mixed solution A, mixed solution B, and mixed solution C, respectively. Then, mixed solution A, mixed solution B, and mixed solution C were sealed and stirred at 50°C for 20 min to obtain transparent choline chloride-urea A (molar ratio of choline chloride to urea 1:0.5), choline chloride-urea B (molar ratio of choline chloride to urea 1:1), and choline chloride-urea C (molar ratio of choline chloride to urea 1:1.5).

[0040] The flow rate in the embodiment is in units of "BV / h", which represents the volume of liquid flowing through the resin per hour in multiples of the resin volume. For example, a flow rate of 4 BV / h means that when the resin dosage is 100 mL, the liquid flow rate is 400 mL / h.

[0041] Macroporous aminophosphonic acid chelating resin D851 was selected from Jiangsu Suqing Water Treatment Engineering Group Co., Ltd.

[0042] Macroporous aminophosphonic acid chelating resin LSC-640 was selected from Xi'an Lanxiao Technology New Materials Co., Ltd.

[0043] The nanofiltration membrane NF270 was selected from Dow Chemical and had a molecular weight cut-off of 200-250 Da.

[0044] Figure 1 It is a schematic diagram of the process flow of the present invention.

[0045] Example 1

[0046] A method for the synergistic extraction of aluminum, lithium, gallium and scandium from fly ash, comprising the following steps:

[0047] (S1) Pretreatment: Fly ash was added to a 15 wt% NaOH aqueous solution (solid-to-liquid ratio of 1 g:4 mL) and irradiated with microwaves at 2.45 GHz and 8 kW for 30 min in a microwave reactor. The mixture was then activated in an autoclave (the microwave reactor and the autoclave were connected via an airtight valve to prevent leakage of high-temperature steam; a spiral discharge port was provided at the bottom of the microwave reactor to achieve continuous production) at 150°C and 1 MPa for 2.5 h. The pretreated product was then filtered.

[0048] (S2) graded acid leaching: the pretreated product was added to 3 mol / L hydrochloric acid (solid-liquid ratio of 1 g:5 mL), and subjected to primary acid leaching at 80°C in a ceramic-lined reactor 1 for 4 h to obtain leachate I and Si-rich residue I; 5 mol / L hydrochloric acid was mixed with choline chloride-urea C in a volume ratio of 3:1 to form a mixed solution, and the Si-rich residue I was added to the mixed solution (solid-liquid ratio of 1 g:4 mL) and subjected to secondary acid leaching at 55°C in a ceramic-lined reactor 2 for 2 h to obtain leachate II and Si-rich residue II; the Si-rich residue I and the Si-rich residue II were collected and calcined at 1000°C to obtain a silicon oxide product;

[0049] (S3) Separation and purification: The product after graded acid leaching is separated and purified, including the following steps:

[0050] (S301) Ion exchange resin adsorption: Leachate II is subjected to reduced pressure distillation at 60°C and 5 kPa to recover choline chloride-urea for recycling, leachate II from which choline chloride-urea has been removed is mixed with leachate I to obtain a combined leachate, and the combined leachate is passed through a macroporous aminophosphonic acid chelating resin D851 at a flow rate of 0.5 BV / h to obtain a loaded resin rich in Ga / Sc / Fe and an effluent rich in Al / Li / Ca;

[0051] (S302) Ca removal and Al / Li separation: ammonium oxalate is dissolved in water to form an ammonium oxalate solution, which is then added dropwise to the effluent rich in Al / Li / Ca under stirring until precipitation is complete. Calcium oxalate and a liquid after Ca removal are obtained after filtration. The pH value of the liquid after Ca removal is adjusted to 6.2 with aqueous ammonia to make Al 3+ A precipitate is formed, which is filtered to obtain an Al-rich residue and a Li-rich liquid. The Al-rich residue is calcined at 1100°C to obtain alumina, and the Li-rich liquid is concentrated and separated through a nanofiltration membrane NF270 (operating pressure 1.0 MPa) to obtain a high-purity lithium solution.

[0052] (S303) stepwise elution of the resin: eluting the Ga / Sc / Fe-rich loaded resin obtained in step (S301) with a 1 mol / L oxalic acid aqueous solution (flow rate 4 BV / h), and monitoring the Ga / Sc concentration in the eluate in real time by inductively coupled plasma optical emission spectrometry (ICP-OES). When the Ga / Sc concentration exceeds 1 mg / L, eluting with oxalic acid is stopped to obtain an iron-rich eluate; then eluting with 4 mol / L hydrochloric acid (flow rate 5 BV / h) for 4 BV to obtain a Ga / Sc-rich eluate;

[0053] (S304) Ga / Sc extraction and separation: Concentrated hydrochloric acid was added to the Ga / Sc-rich eluate until the concentration of hydrochloric acid in the system was 5 mol / L. Tributyl phosphate (TBP) was then added as an extractant (the volume ratio of TBP to eluate was 1:2). The mixture was mixed and shaken for 15 minutes. The mixture was allowed to stand for 30 minutes before phase separation. The extraction was repeated three times, and the oil phases were combined to obtain a Ga-rich oil phase and a Sc-rich aqueous phase.

[0054] (S305) Ga stripping and electrolysis: 0.3 mol / L sulfuric acid stripping agent (the volume ratio of 0.3 mol / L sulfuric acid to oil phase is 1:1) was added to the Ga-rich oil phase, mixed and shaken for 15 minutes, and allowed to stand for 30 minutes before phase separation; the stripping was repeated twice, and the aqueous phases were combined to obtain a Ga-rich aqueous phase; the Ga-rich aqueous phase was then placed in an electrolytic cell and pulsed electrodeposition (current density 100 A / m², duty cycle 40%, frequency 20 kHz) was used to obtain Ga with a purity of 4N (99.99%).

[0055] Example 2

[0056] The rest is the same as in Example 1, except that choline chloride-urea B is used instead of choline chloride-urea C in step (S2).

[0057] Example 3

[0058] The rest is the same as in Example 1, except that choline chloride-urea A is used instead of choline chloride-urea C in step (S2).

[0059] Example 4

[0060] The rest is the same as in Example 1, except that the conditions for the primary acid leaching in step (S2) are different. In the secondary acid leaching, the volume ratio of 5 mol / L hydrochloric acid to choline chloride-urea C is 4:1. Accordingly, the temperature and time of the secondary acid leaching are also adjusted, specifically as follows:

[0061] (S1) Same as Example 1;

[0062] (S2) graded acid leaching: the pretreated product was added to 4 mol / L hydrochloric acid (solid-liquid ratio of 1 g:4 mL), and subjected to primary acid leaching at 95°C for 2 h in a ceramic-lined reactor 1 to obtain leachate I and Si-rich residue I; 5 mol / L hydrochloric acid was mixed with choline chloride-urea C in a volume ratio of 4:1 to form a mixed solution, and the Si-rich residue I was added to the mixed solution (solid-liquid ratio of 1 g:4 mL) and subjected to secondary acid leaching at 65°C for 3 h in a ceramic-lined reactor 2 to obtain leachate II and Si-rich residue II; the Si-rich residue I and the Si-rich residue II were collected and calcined at 1000°C to obtain a silicon oxide product;

[0063] (S3) Same as Example 1.

[0064] Example 5

[0065] The rest is the same as in embodiment 1, except that the pretreatment conditions in step (S1) are different, specifically:

[0066] (S1) Pretreatment: Fly ash was mixed with a 20 wt% NaOH aqueous solution at a ratio of 1 g:3 mL, and microwave irradiated at 2.45 GHz and 8 kW for 30 min, followed by activation at 180°C and 0.5 MPa for 2 h. The pretreated product was filtered.

[0067] (S2) Same as Example 1;

[0068] (S3) Same as Example 1.

[0069] Example 6

[0070] The rest is the same as Example 1, except that: in step (S301), the macroporous aminophosphonic acid chelating resin LSC-640 is used instead of D851; in (S304), the extractant trioctylphosphine oxide (TOPO) is used instead of TBP; and the pulse electrodeposition conditions in (S305) are adjusted. Specifically,

[0071] (S1) Same as Example 1;

[0072] (S2) Same as Example 1;

[0073] (S3) Separation and purification: The product after graded acid leaching is separated and purified, including the following steps:

[0074] (S301) Ion exchange resin adsorption: Leachate II is subjected to reduced pressure distillation at 60°C and 5 kPa to recover choline chloride-urea for recycling, leachate II from which choline chloride-urea has been removed is mixed with leachate I to obtain a combined leachate, and the combined leachate is passed through a macroporous aminophosphonic acid chelating resin LSC-640 at a flow rate of 1 BV / h to obtain a loaded resin rich in Ga / Sc / Fe and an effluent rich in Al / Li / Ca;

[0075] (S302) Ca removal and Al / Li separation: ammonium oxalate is dissolved in water to form an ammonium oxalate solution, which is then added dropwise to the effluent rich in Al / Li / Ca under stirring until precipitation is complete. Calcium oxalate and a liquid after Ca removal are obtained after filtration. The pH value of the liquid after Ca removal is adjusted to 5.8 with ammonia water to make Al 3+ A precipitate forms, which is filtered to yield an Al-rich residue and a Li-rich liquid. The Al-rich residue is calcined at 1100°C to yield alumina, while the Li-rich liquid is concentrated and separated through an NF270 nanofiltration membrane (operating pressure 1.0 MPa). The permeate is a high-purity lithium solution.

[0076] (S303) stepwise elution of the resin: the Ga / Sc / Fe-rich loaded resin obtained in step (S301) was eluted with a 1 mol / L oxalic acid aqueous solution (flow rate 5 BV / h), and the Ga / Sc concentration in the eluate was monitored in real time by inductively coupled plasma optical emission spectrometry (ICP-OES). When the Ga / Sc concentration exceeded 1 mg / L, the oxalic acid elution was stopped to obtain an iron-rich eluate. The resin was then eluted with 4 mol / L hydrochloric acid (flow rate 5 BV / h) for 4 BV to obtain a Ga / Sc-rich eluate.

[0077] (S304) Ga / Sc extraction and separation: Concentrated hydrochloric acid was added to the Ga / Sc-rich eluate until the concentration of hydrochloric acid in the system was 7 mol / L. Trioctylphosphine oxide (TOPO) was then added as an extractant (the volume ratio of TOPO to eluate was 1:2). The mixture was mixed and shaken for 15 minutes. The mixture was allowed to stand for 30 minutes before phase separation. The extraction was repeated three times, and the oil phases were combined to obtain a Ga-rich oil phase and a Sc-rich aqueous phase.

[0078] (S305) Ga stripping and electrolysis: 0.3 mol / L sulfuric acid stripping agent (the volume ratio of 0.3 mol / L sulfuric acid to oil phase is 1:1) was added to the Ga-rich oil phase, mixed and shaken for 15 minutes, and allowed to stand for 30 minutes before phase separation; the extraction was repeated twice, and the aqueous phases were combined to obtain a Ga-rich aqueous phase; the Ga-rich aqueous phase was then placed in an electrolytic cell and pulsed electrodeposition (current density 100 A / m², duty cycle 30%, frequency 15 kHz) was used to obtain Ga with a purity of 4N (99.99%).

[0079] Comparative Example 1

[0080] The rest is the same as Example 1, except that: in step (S2), choline chloride-urea is not used in the secondary acid leaching treatment, and the solid-liquid ratio of the Si-rich residue I and 5 mol / L hydrochloric acid is 1 g:4 mL.

[0081] Comparative Example 2

[0082] The rest is the same as Example 1, except that: in step (S1), microwave irradiation is not performed.

[0083] The recoveries of Al, Li, Fe, Ca, Ga and Sc in the above embodiments and comparative examples were tested, and the test results are shown in Table 1.

[0084] Recovery rate (%) = (mass of element in final product / mass of element in initial fly ash) × 100%

[0085] Table 1 Recovery rate of each metal element

[0086] .

[0087] As can be seen from Table 1, the recovery rates of Al, Ga, Sc, and Fe in the fly ash of the embodiments of the present invention are over 85%, the recovery rate of Li is over 90%, and the recovery of Ca and Si is achieved at the same time. That is, the method of the present invention achieves efficient synergistic extraction and full resource utilization of aluminum, lithium, gallium, and scandium in fly ash in the same process flow, and the energy consumption of the entire process is low, which is conducive to reducing costs and increasing profits.

Claims

1. A method for the synergistic extraction of aluminum, lithium, gallium and scandium from fly ash, characterized in that: The steps include: (S1) Pretreatment: Fly ash is added to an aqueous solution of an inorganic base and subjected to microwave irradiation, followed by pressure activation, and filtered to obtain a pretreated product; (S2) graded acid leaching: the pretreated product is added to hydrochloric acid A for primary acid leaching to obtain leachate I and Si-rich residue I; the Si-rich residue I is added to a mixture of hydrochloric acid B and choline chloride-urea at 55-65°C for secondary acid leaching to obtain leachate II and Si-rich residue II; (S3) Separation and purification: The product after graded acid leaching is separated and purified, including the following steps: (S301) evaporating the choline chloride-urea from the leachate II and mixing it with the leachate I to obtain a combined leachate, and passing the combined leachate through an ion exchange resin that selectively adsorbs Ga, Sc, and Fe to obtain a loaded resin rich in Ga / Sc / Fe and an effluent rich in Al / Li / Ca; (S302) After removing Ca from the Al / Li / Ca-rich effluent, the pH is adjusted to 5.5-6.5 to obtain an Al-rich residue and a Li-rich liquid; (S303) eluting the Ga / Sc / Fe-rich loaded resin obtained in step (S301) with an oxalic acid aqueous solution to obtain an iron-rich eluate; and then eluting with hydrochloric acid to obtain a Ga / Sc-rich eluate; (S304) adding concentrated hydrochloric acid to the Ga / Sc-rich eluate until the concentration of hydrochloric acid in the system is 5-7 mol / L, and then extracting with an extractant to separate the oil and water phases to obtain a Ga-rich oil phase and a Sc-rich water phase; (S305) The Ga-rich oil phase is stripped using dilute acid to separate the oil-water phase to obtain a Ga-rich water phase, and then high-purity Ga is obtained by pulse electrolysis.

2. The method according to claim 1, characterized in that In step (S1), the particle size of the fly ash is 1-10 μm; the inorganic base is KOH or NaOH, and the concentration of the aqueous solution of the inorganic base is 10-20 wt%; the solid-liquid ratio of the fly ash to the aqueous solution of the inorganic base is 1 g:3-4 mL; the conditions of the microwave irradiation are: frequency 2-3 GHz, power 5-10 kW, time 20-40 min; the conditions of the pressurized activation are: pressure 0.5-1.5 MPa, temperature 150-180° C., and time 1-3 h.

3. The method according to claim 1, characterized in that In step (S2), the volume ratio of hydrochloric acid B to choline chloride-urea is 3-4:1; and the molar ratio of choline chloride to urea in choline chloride-urea is 1:0.5-1.

5.

4. The method according to claim 3, characterized in that In step (S2), the volume ratio of hydrochloric acid B to choline chloride-urea is 3:1; and the molar ratio of choline chloride to urea in choline chloride-urea is 1:1-1.

5.

5. The method according to claim 1, wherein In step (S2), the concentration of hydrochloric acid A is 3-4 mol / L, the solid-liquid ratio of the pretreated product to hydrochloric acid A is 1 g:4-5 mL, and the conditions for the primary acid leaching are: temperature 80-95° C., time 2-4 h; the concentration of hydrochloric acid B is 5-6 mol / L, the solid-liquid ratio of the Si-rich residue I to the mixed solution is 1 g:4-5 mL, and the time for the secondary acid leaching is 1-3 h.

6. The method according to claim 1, characterized in that The conditions for distilling off the choline chloride-urea in step (S301) are as follows: the leachate II is subjected to reduced pressure distillation at 50-70°C and 5-10 kPa, and then mixed with the leachate I; and the ion exchange resin is a macroporous aminophosphonic acid chelating resin.

7. The method according to claim 1, characterized in that In step (S302), the conditions for removing Ca are as follows: adding ammonium oxalate to the effluent rich in Al / Li / Ca, performing solid-liquid separation, and obtaining calcium oxalate and a liquid after removing Ca; and / or Step (S302) includes concentrating and separating the Li-rich liquid through a nanofiltration membrane to obtain a high-purity lithium solution.

8. The method according to claim 1, characterized in that In step (S303), the concentration of the oxalic acid aqueous solution is 0.5-1 mol / L; the concentration of the hydrochloric acid is 3-4 mol / L.

9. The method according to claim 1, characterized in that The extractant in step (S304) is tributyl phosphate or trioctylphosphine oxide.

10. The method according to claim 1, characterized in that The dilute acid in step (S305) is sulfuric acid with a concentration of 0.1-0.3 mol / L. The conditions of the pulse electrolysis are: current density 50-100 A / m², duty cycle 30%-40%, frequency 15 kHz-25 kHz. The high purity is 4N purity.

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