Method for recovering valuable components in fused salt chlorination slag through dressing and smelting combination

Abstract

The invention discloses a method for recovering valuable components in fused salt chlorination slag through dressing and smelting combination, which comprises the following steps: (1) after the fused salt chlorination slag is discharged from a furnace, adding an additive to carry out mineral phase conversion, and cooling to room temperature to obtain a phase conversion product; (2) crushing a phase conversion product, leaching under the condition of an aqueous solution, and carrying out solid-liquid separation on a water leaching product to obtain a sodium chloride leaching solution and leaching residues; (3) the leaching residues are subjected to reduction-magnetic separation, magnetic separation concentrate and magnetic separation tailings are obtained, and the magnetic separation concentrate is used for preparing manganese-iron alloy; and (4) magnetic separation tailings are subjected to leaching, enrichment, oxalic acid precipitation and roasting treatment, and a rare earth enriched product is obtained. According to the method, a dressing and smelting combined technical route is provided, various valuable elements in the fused salt chlorination residues are comprehensively recycled, meanwhile, cyclic recycling of sodium chloride in the fused salt chlorination residues can be achieved, economic benefits are remarkable, and the problem that a large amount of fused salt chlorination residues are piled up and pollute the environment is solved.

Description

Technical Field

[0001] The invention belongs to the technical field of titanium metallurgical solid waste recycling, and in particular relates to a method for recovering valuable components in molten salt chloride slag through combined selection and smelting. Background Art

[0002] Titanium is a rising "fourth-generation metal material" after copper, iron, and aluminum. Its high- and low-temperature resistance, non-magnetic properties, and strong corrosion resistance make it widely used in aerospace and marine engineering. Titanium also has a wide range of applications in the automotive, medical, and chemical industries, as well as in daily life. Demand for titanium is rapidly growing both domestically and internationally.

[0003] Currently, my country's titanium industry primarily produces low- and medium-end sulfuric acid-processed titanium dioxide. High-end titanium products, including chloride-processed titanium dioxide and titanium sponge, account for approximately 10% of the total, resulting in a severe supply shortage. Furthermore, the quality of high-end titanium products lags significantly behind that of the United States, Russia, and Japan. There is an urgent need to accelerate the development of a high-end titanium industry by leveraging my country's existing titanium resources.

[0004] The production of existing high-end titanium products all uses titanium tetrachloride (TiCl4) as an intermediate. TiCl4 is reduced to produce titanium sponge, which is then oxidized to produce chloride-based titanium dioxide. Currently, the main industrial methods for producing TiCl4 are boiling chlorination and molten salt chlorination. The boiling chlorination method requires high-quality titanium-rich materials with TiO2 > 92%, CaO + MgO < 1.5%, and a particle size distribution that ensures a controllable fluidization state. The molten salt chlorination method can use titanium-rich materials with high levels of impurities such as calcium and magnesium, and has no stringent requirements for particle size. The reaction temperature is low, the chlorination process has minimal blowdown, and the resulting crude titanium tetrachloride has a low impurity content. my country has the world's largest titanium resource reserves, primarily located in the Panzhihua region. High levels of calcium and magnesium impurities make the molten salt chlorination method more suitable for TiCl4 production. Currently, to maintain the stability of the molten salt chlorination reaction system, the molten salt chlorination method requires regular replenishment of new salt and discharge of waste salt, with slag discharge reaching up to 400 kg per ton of TiCl4. Since the molten salt chlorination process was introduced and put into production in the last century, a large amount of molten salt chlorination slag has been stored, and there is no effective treatment method, which poses a great threat to the environment.

[0005] Molten salt chloride slag is a complex product, containing more than ten types of metal chlorides, unreacted oxides, and petroleum coke. It contains high levels of rare earth resources such as iron, manganese, calcium, and magnesium, as well as niobium, gallium, scandium, yttrium, and cerium, and possesses extremely high economic value. However, metal chlorides are highly soluble and hazardous, and their direct storage or landfilling can lead to serious problems such as groundwater pollution, soil salinization, and resource waste. Currently, hydrometallurgy or pyrometallurgy are the primary methods used to recycle molten salt chloride slag. The main recovered products include sodium chloride, magnesium-manganese alloy, pure iron, and scandium trioxide. Sodium chloride is the primary component of molten salt chloride slag, exceeding 40%, making it the primary target for recycling. Currently, the main method used for recovery is water dissolution and impurity removal. Although this method can recycle the NaCl in molten salt chloride slag, it has problems such as a long impurity ion removal process, complex technology, and low economic value of process products and by-products, and its industrial application prospects are limited. Magnesium, manganese, and iron are valuable metal elements with high content in molten salt chloride slag. Electrolysis is mainly used to recover and prepare magnesium-manganese alloy or pure iron. This method can effectively recover valuable metal elements such as magnesium, manganese, and iron from molten salt chloride slag, but the sorting and remelting of the molten salt chloride slag before electrolysis is inefficient and the process is long. The NaCl-CaCl2 molten salt and other valuable metal elements in the electrolytic cell have not yet been recovered. The recovery of scandium from molten salt chloride slag mainly uses water dissolution, impurity removal, and extraction to prepare a scandium-rich solution. The scandium-rich solution is precipitated with oxalic acid and calcined to obtain high-purity scandium trioxide. However, due to the large variety and complex composition of impurity ions in the molten salt chloride slag leachate, the impurity removal process before leachate extraction is long, the extractant is severely emulsified in the multi-stage countercurrent extraction process, and the scandium loss rate is high.

[0006] In summary, current treatment methods for molten salt chloride slag, both domestically and internationally, are complex, resulting in difficult treatment of chlorine-containing wastewater. A significant amount of soluble chloride salts remain ineffectively recovered, and selective extraction of other valuable components, including rare metals and rare earth elements, remains incomplete. If the valuable elements in molten salt chloride slag can be economically and efficiently recovered, fundamentally resolving the issue of large-scale molten salt chloride slag stockpiling, the molten salt chlorination process will retain its advantages over the boiling chlorination process. Summary of the Invention

[0007] In view of the shortcomings of the existing technology, the purpose of the present invention is to provide a method for recovering valuable components in molten salt chloride slag by combined metallurgy and smelting with good economic benefits and high processing efficiency, so as to realize the recycling of sodium chloride in molten salt chloride slag and solve the problem of large-scale storage of molten salt chloride slag.

[0008] In order to achieve the above object, the present invention provides a method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting, comprising the following steps:

[0009] (1) Adding additives to the molten salt chloride slag after it is discharged from the furnace to carry out mineral phase transformation, and cooling it to room temperature to obtain a phase transformation product;

[0010] (2) crushing the phase transformation product and leaching it under aqueous solution conditions, separating the solid and liquid of the leaching product to obtain sodium chloride leachate and leach residue;

[0011] (3) The leached residue is subjected to reduction-magnetic separation to obtain magnetic concentrate and magnetic tailings, and the magnetic concentrate is used to prepare manganese iron alloy;

[0012] (4) The magnetic separation tailings are treated by leaching-enrichment-oxalic acid precipitation-roasting to obtain rare earth enriched products.

[0013] Preferably, in step (1), the additive is one or more of carbonates, bicarbonates, and silicates, including but not limited to sodium carbonate, potassium carbonate, sodium bicarbonate, potassium bicarbonate, sodium silicate, and potassium silicate; more preferably, sodium carbonate.

[0014] More preferably, the amount of the additive added is 0 to 2 times, preferably 1.1 to 1.5 times, the total molar amount of calcium chloride, magnesium chloride, manganese chloride, ferrous chloride and ferric chloride in the molten salt chloride slag.

[0015] Preferably, in step (1), the molten salt chloride slag is fully reacted after adding the additive, the reaction temperature is 500°C to 700°C, the reaction time is 30min to 180min, and the reaction is cooled to room temperature after completion of the reaction. The reaction time depends on the reaction rate of the molten salt chloride slag and the additive. The longer the reaction time, the more complete the mineral phase conversion, the better the conversion effect, the higher the sodium chloride content in the leachate, and the lower the content of impurity elements such as calcium, magnesium, manganese, and iron.

[0016] More preferably, the reaction temperature is 700° C. and the reaction time is 120 min.

[0017] Preferably, in step (2), the phase transformation product is crushed to less than 1 mm to ensure that the sodium chloride, metal carbonates, and metal oxides in the phase transformation product are fully dissociated. This is beneficial for the subsequent water leaching process, in which the sodium chloride is fully dissolved into the leachate, and the metal carbonates and metal oxides are absorbed into the leached residue.

[0018] Preferably, the amount of water used in the water leaching process should be 1 to 1.3 times the saturated solubility of sodium chloride. The amount of water used in the leaching process should be minimized to reduce the energy consumption of subsequent evaporation crystallization or sodium chloride concentration.

[0019] More preferably, the amount of water used in the water leaching process should be 1.1 times the saturated solubility of sodium chloride.

[0020] Preferably, in the step (2), the equipment used for crushing the phase transformation product is one of a jaw crusher, a ball mill, and a roller crusher.

[0021] Preferably, in the step (2), the solid-liquid separation method is one of pressure filtration, vacuum filtration, and centrifugal filtration.

[0022] Preferably, in the step (3), during the leaching residue reduction process, the reducing agent is one of a carbon-based reducing agent, an aluminum-based reducing agent, and a silicon-based reducing agent.

[0023] More preferably, carbon-based reducing agents include graphite powder, coke powder, coal powder, metallurgical coke, etc.; aluminum-based reducing agents include aluminum powder, scrap aluminum and scrap aluminum alloy, etc.; silicon-based reducing agents include silicon powder, ferrosilicon powder, etc.

[0024] Preferably, in step (3), the reduction temperature of the leaching residue reduction process is 1400-1500° C., and the reduction time is 30-180 minutes. Iron and manganese compounds are easily reduced to metallic iron and metallic manganese under the action of a reducing agent, and the leaching residue reduction products are mainly metallic iron and manganese and other weakly magnetic or non-magnetic phases.

[0025] During the leaching residue reduction process, iron and manganese compounds are reduced to metallic iron and manganese under suitable reducing agents and temperature conditions, which are highly magnetic. Calcium, magnesium, and rare earth oxides are difficult to reduce and are weakly magnetic or non-magnetic. During the magnetic separation process, the magnetic field intensity is controlled, and metallic iron and manganese enter the magnetic concentrate, while the remaining components enter the magnetic tailings.

[0026] Preferably, in the step (3), the magnetic concentrate obtained after grinding and magnetic separation of the reduction product has a main phase of metallic iron and manganese, and the grinding method includes but is not limited to conventional grinding equipment such as ball milling and rod milling. The grinding time is 30 to 60 minutes, and the pulp ratio is 1 to 3.

[0027] Preferably, in the step (3), the magnetic separation process is a weak magnetic field separation of ferromagnetic metals iron and manganese, and the magnetic field strength is 700 to 2000 GS.

[0028] Preferably, in step (4), the leaching agent includes but is not limited to acidic leaching agents such as hydrochloric acid and sulfuric acid; the rare earth ion enrichment method includes but is not limited to solution extraction, liquid membrane extraction, resin ion exchange, etc.; the precipitant used for rare earth ion precipitation includes but is not limited to oxalic acid; and the roasting system is 800-900°C for 2-4h.

[0029] The mechanism and effects of the present invention are as follows:

[0030] Chloride salts such as CaCl2, MgCl2, MnCl2, FeCl2, FeCl3, and ScCl3 in the molten salt chloride slag undergo mineral phase transformation with the additives sodium carbonate or potassium carbonate, generating the corresponding metal carbonates or metal oxides and sodium chloride. After cooling, the reaction product undergoes crushing, water dissolution, and solid-liquid separation to produce a sodium chloride leachate and leach residue, successfully achieving dechlorination, desodiumization, and volume reduction of the molten salt chloride slag, as well as initial enrichment of valuable metal elements.

[0031] The main components of the leached residue are metal carbonates and metal oxides. Iron and manganese oxides are easily reduced. Under a reducing atmosphere, iron and manganese oxides are reduced to metallic iron and manganese. Metallic iron and manganese are strongly magnetic, while the remaining phases are weakly magnetic or non-magnetic. Under the influence of a weak magnetic field, the strongly magnetic metallic iron and manganese are separated from the remaining components and enter the magnetic concentrate, while calcium, magnesium, and rare earth oxides enter the magnetic tailings. This achieves the selective separation of iron and manganese from calcium, magnesium, and rare earth elements.

[0032] After phase conversion separation pre-enrichment and secondary enrichment through reduction grinding, the rare earth content in the magnetic separation tailings is significantly increased several times compared to molten salt chloride slag. During the acid leaching process, all rare earth elements are dissolved into the leachate, where they are further enriched using ion exchange or solvent extraction. Finally, oxalic acid is used as a precipitant to selectively precipitate the rare earth ions in the rare earth enrichment solution. After calcination, the initial rare earth oxides are obtained. BRIEF DESCRIPTION OF THE DRAWINGS

[0033] Figure 1 This is a process flow chart for the combined recovery of valuable components from molten salt chloride slag. DETAILED DESCRIPTION

[0034] The technical solutions in the embodiments of the present invention will be described clearly and completely below. Obviously, the described embodiments are only part of the embodiments of the present invention, not all of the embodiments. Based on the embodiments of the present invention, all other embodiments obtained by ordinary technicians in this field without making creative work are within the scope of protection of the present invention.

[0035] The present invention is further described below with reference to specific embodiments and accompanying drawings:

[0036] Example 1

[0037] Such as process flow Figure 1 As shown, the additive sodium carbonate is added to the molten salt chloride slag (NaCl 37.46wt%, CaCl211.39wt%, MgCl217.07wt%, MnCl29.39wt%, FeCl25.53wt%, Sc 220ppm, Nb 300ppm) discharged from the furnace, and the amount of the additive is controlled to be 1.2 times the total molar amount of calcium chloride, magnesium chloride, manganese chloride, ferrous chloride and ferric chloride in the molten salt chloride slag. The phase inversion temperature is 700℃, the reaction time is 120min, and the reaction is gradually cooled to room temperature after the reaction is completed. After the phase inversion product is crushed and leached with water, the water consumption in the water leaching process should be 1.1 times the saturated solubility of sodium chloride, the NaCl content in the leachate reaches 99.65%, and the Ca in the leachate is 1.1 times the saturated solubility of sodium chloride. 2+ The content is 0.36ppm, Mg 2+The content is 0.11ppm, and after evaporation and crystallization, it is returned to the molten salt chlorination furnace for recycling. Graphite powder is added to the leached slag, the reduction temperature is 1450℃, the reduction time is 60min, and the iron and manganese oxides are selectively reduced to metallic iron and metallic manganese, while the remaining oxides are not reduced. When ball milling is used for 30min, the slurry ratio is 2, and the magnetic field is 1000GS, the main phases of the recovered magnetic concentrate are metallic iron and metallic manganese, which can be used to prepare manganese-iron alloy. The magnetic tailings are leached with sulfuric acid, the sulfuric acid concentration is 15%, the liquid-solid ratio is 3, the leaching temperature is 50℃, and the Sc 3+ The leaching rate is 96.31%. P507 is used as the extractant. After two-stage extraction, a scandium-rich solution is obtained. After sodium oxalate precipitation, it is roasted at 850℃ for 3h to obtain a rare earth-enriched product with a Sc2O3 content of 99.11%.

[0038] Example 2

[0039] The molten salt chloride slag (NaCl 46.24wt%, CaCl212.63wt%, MgCl219.71wt%, MnCl211.32wt%, FeCl25.99wt%, Sc 220ppm, Nb 300ppm) was added with the additive sodium carbonate. The amount of the additive was controlled to be 1.1 times the total molar amount of calcium chloride, magnesium chloride, manganese chloride and ferric chloride in the molten salt chloride slag. The phase inversion temperature was 650℃, the reaction time was 180min, and the slag was gradually cooled to room temperature. After the phase inversion product was crushed and leached with water, the water consumption in the leaching process should be 1.3 times the saturated solubility of sodium chloride. The NaCl content in the leaching solution reached 99.98%, and the Ca content in the leaching solution was 99.98%. 2+ The content is 0.57ppm, Mg 2+ The content is 0.15ppm, and after evaporation and crystallization, it is returned to the molten salt chlorination furnace for recycling. Metal aluminum powder is added to the leached slag, the reduction temperature is 1500℃, the reduction time is 60min, and the iron-manganese oxide is reduced to metallic iron and metallic manganese. When the ball mill is used for 60min, the pulp ratio is 1, and the magnetic field is 1500GS, the main components of the magnetic concentrate are metallic iron and metallic manganese, which can be used to prepare manganese-iron alloy. The magnetic tailings are leached with hydrochloric acid, the hydrochloric acid concentration is 20%, the liquid-solid ratio is 3, the leaching temperature is 65℃, and the Sc 3+ The leaching rate is 99.21%. Cyanex925 is used as the extractant. After two-stage extraction, a scandium-rich solution is obtained. After sodium oxalate precipitation, it is roasted at 850℃ for 3h to obtain a rare earth-enriched product with a Sc2O3 content of 99.94%.

[0040] Example 3

[0041] The molten salt chloride slag (NaCl 33.09wt%, CaCl210.77wt%, MgCl215.79wt%, MnCl28.11wt%, FeCl35.98wt%, Sc 200ppm, Nb 310ppm) was added with the additive sodium carbonate. The amount of the additive was controlled to be 1.5 times the total molar amount of calcium chloride, magnesium chloride, manganese chloride and ferric chloride in the molten salt chloride slag. The phase inversion temperature was 600℃ and the reaction time was 60min. After the reaction was completed, the slag was gradually cooled to room temperature. After the phase inversion product was crushed and leached with water, the water consumption in the leaching process should be 1.2 times the saturated solubility of sodium chloride. The NaCl content in the leachate reached 99.52%, and the Ca in the leachate reached 99.52%. 2+ The content is 0.88ppm, Mg 2+ The content is 0.38ppm. After evaporation and crystallization, it is returned to the molten salt chlorination furnace for recycling. Silicon powder is added to the leached slag, the reduction temperature is 1500℃, the reduction time is 120min, and the iron-manganese oxides are reduced to metallic iron and metallic manganese. When the ball mill is used for 100min, the pulp ratio is 1, and the magnetic field strength is 1300GS, the main components of the magnetic concentrate are metallic iron and metallic manganese, which can be used to prepare manganese-iron alloy. The magnetic tailings are leached with hydrochloric acid, the hydrochloric acid concentration is 30%, the liquid-solid ratio is 3, the leaching temperature is 40℃, and the Sc 3+ The leaching rate is 93.89%. P204 is used as the extractant. After two-stage extraction, a scandium-rich solution is obtained. After sodium oxalate precipitation, it is roasted at 850℃ for 3h to obtain a rare earth-enriched product with a Sc2O3 content of 97.44%.

[0042] Example 4

[0043] The molten salt chloride slag (NaCl 46.24wt%, CaCl212.63wt%, MgCl219.71wt%, MnCl211.32wt%, FeCl25.99wt%, Sc 220ppm, Nb 300ppm) was added with potassium carbonate as an additive. The amount of the additive was controlled to be 1.2 times the total molar amount of calcium chloride, magnesium chloride, manganese chloride and ferric chloride in the molten salt chloride slag. The phase inversion temperature was 700℃, the reaction time was 180min, and the mixture was gradually cooled to room temperature. After the phase inversion product was crushed and leached with water, the water consumption in the leaching process should be 1.2 times the saturated solubility of sodium chloride. The NaCl+KCl content in the leaching solution reached 99.82%, and the Ca in the leaching solution was 1.2 times the saturated solubility of sodium chloride. 2+ The content is 0.34ppm, Mg 2+The content is 0.07ppm, and after evaporation and crystallization, it is returned to the molten salt chlorination furnace for recycling. Graphite powder is added to the leached slag, the reduction temperature is 1600℃, the reduction time is 180min, and the iron-manganese oxides are reduced to metallic iron and metallic manganese. When the ball mill is used for 60min, the pulp ratio is 1, and the magnetic field is 2000GS, the main components of the magnetic concentrate are metallic iron and manganese, which can be used to prepare manganese-iron alloy. The magnetic tailings are leached with hydrochloric acid, the hydrochloric acid concentration is 30%, the liquid-solid ratio is 3, the leaching temperature is 50℃, and the Sc 3+ The leaching rate is 99.28%. P204 is used as the extractant. After two-stage extraction, a scandium-rich solution is obtained. After sodium oxalate precipitation, it is roasted at 850℃ for 3h to obtain a rare earth enriched product with a Sc2O3 content of 99.91%.

[0044] Example 5

[0045] The molten salt chloride slag (NaCl 46.24wt%, CaCl212.63wt%, MgCl219.71wt%, MnCl211.32wt%, FeCl25.99wt%, Sc 220ppm, Nb 300ppm) was added with the additive sodium bicarbonate. The amount of the additive was controlled to be 1.2 times the total molar amount of calcium chloride, magnesium chloride, manganese chloride and ferric chloride in the molten salt chloride slag. The phase inversion temperature was 750℃, the reaction time was 180min, and the slag was gradually cooled to room temperature. After the phase inversion product was crushed and leached with water, the water consumption in the leaching process should be 1.1 times the saturated solubility of sodium chloride. The NaCl content in the leachate reached 99.99%, and the Ca in the leachate reached 99.99%. 2+ The content is 0.46ppm, Mg 2+ The content is 0.08ppm, and after evaporation and crystallization, it is returned to the molten salt chlorination furnace for recycling. Graphite powder is added to the leached slag, the reduction temperature is 1600℃, the reduction time is 180min, and the iron-manganese oxide is reduced to metallic iron and metallic manganese. When the ball mill is used for 60min, the pulp ratio is 2, and the magnetic field is 700GS, the main components of the magnetic concentrate are metallic iron and metallic manganese, which can be used to prepare manganese-iron alloy. The magnetic tailings are leached with hydrochloric acid, the hydrochloric acid concentration is 10%, the liquid-solid ratio is 3, the leaching temperature is 50℃, and the Sc 3+ The leaching rate is 96.14%. P507 is used as the extractant. After two-stage extraction, a scandium-rich solution is obtained. After sodium oxalate precipitation, it is roasted at 850℃ for 3h to obtain a rare earth-enriched product with a Sc2O3 content of 99.25%.

[0046] Comparative Example 1

[0047] The molten salt chloride slag (NaCl 46.24wt%, CaCl2 12.63wt%, MgCl2 19.71wt%, MnCl2 11.32wt%, FeCl2 5.99wt%, Sc 220ppm, Nb 300ppm) was directly held at 650°C for 180 minutes and then gradually cooled to room temperature. After the product was crushed and leached in water, the water consumption during the leaching process should be 1.3 times the saturated solubility of sodium chloride. The NaCl content in the leachate was 62.72%, and other metal chlorides including CaCl2, MgCl2, MnCl2, and FeCl2 all entered the leachate, making it difficult to return the molten salt chlorination furnace for recycling after evaporation and crystallization. The main phases in the leached slag are silica, alumina, and titania, and the enrichment of iron, manganese, and rare earth elements is not achieved, making it difficult to recover iron-manganese alloys and rare earth oxides using a subsequent reduction grinding-leaching extraction method.

[0048] Comparative Example 2

[0049] The additive sodium carbonate is added to the molten salt chloride slag (NaCl 46.24wt%, CaCl2 12.63wt%, MgCl2 19.71wt%, MnCl2 11.32wt%, FeCl2 5.99wt%, Sc 220ppm, Nb 300ppm) discharged from the furnace. The additive dosage is controlled to 0.5 times the total molar amount of calcium chloride, magnesium chloride, manganese chloride, and ferric chloride in the molten salt chloride slag. The phase inversion temperature is 750℃, the reaction time is 180min, and the product is gradually cooled to room temperature. After the product is crushed and water-leached, the water consumption in the water leaching process should be 1.3 times the saturated solubility of sodium chloride. Due to the small amount of additive, the NaCl content in the leachate is only 71.93%, and other metal chlorides including CaCl2, MgCl2, MnCl2, FeCl2, etc. enter the leachate, making it difficult to return them to the molten salt chlorination furnace for recycling after evaporation and crystallization. A large amount of iron, manganese, and rare earth chloride salts enter the leachate without reacting with the additives, resulting in ineffective enrichment of iron, manganese, and rare earth elements in the leaching residue. This makes it difficult to recover the iron-manganese alloy and rare earth oxides using subsequent reduction grinding and leaching extraction.

[0050] Comparative Example 3

[0051] Sodium carbonate was added to the molten salt chloride slag (NaCl 46.24wt%, CaCl2 12.63wt%, MgCl2 19.71wt%, MnCl2 11.32wt%, FeCl2 5.99wt%, Sc 220ppm, Nb 300ppm) at a concentration of 1.2 times the total molar amount of calcium chloride, magnesium chloride, manganese chloride, and ferric chloride in the molten salt chloride slag. The phase inversion temperature was 400°C, the reaction time was 30 minutes, and the mixture was gradually cooled to room temperature. The product was crushed and leached in water, with the water consumption for the leaching process being 1.3 times the saturated solubility of sodium chloride. Due to the low phase inversion temperature and short phase inversion time, the solid-solid reaction rate between the additive and the molten salt chloride slag is slow, leaving a large amount of metal chloride salts unreacted with the additive. The NaCl content in the leachate is only 63.22%. The remaining unreacted metal chloride salts, including CaCl2, MgCl2, MnCl2, and FeCl2, enter the leachate and are difficult to return to the molten salt chlorination furnace for recycling after evaporation and crystallization. A large amount of iron, manganese, and rare earth chloride salts enter the leachate without reacting with the additive, resulting in ineffective enrichment of iron, manganese, and rare earth elements in the leachate slag. This makes it difficult to recover the iron-manganese alloy and rare earth oxides using subsequent reduction grinding and leaching extraction.

[0052] The above is only a preferred embodiment of the present invention, and the scope of protection of the present invention is not limited to the above embodiment. For those skilled in the art, improvements and modifications obtained without departing from the technical concept of the present invention should also be considered as the scope of protection of the present invention.

Claims

1. A method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting, characterized in that: The following steps are involved: (1) Adding additives to the molten salt chloride slag after it is discharged from the furnace to carry out mineral phase transformation, and cooling it to room temperature to obtain a phase transformation product; (2) crushing the phase transformation product and leaching it under aqueous solution conditions, separating the solid and liquid of the leaching product to obtain sodium chloride leachate and leach residue; (3) The leached residue is subjected to reduction-magnetic separation to obtain magnetic concentrate and magnetic tailings, and the magnetic concentrate is used to prepare manganese iron alloy; (4) The magnetic separation tailings are treated by leaching-enrichment-oxalic acid precipitation-roasting to obtain rare earth enriched products.

2. The method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting according to claim 1, characterized in that: In the step (1), the additive is one or more of carbonates, bicarbonates, and silicates, including but not limited to sodium carbonate, potassium carbonate, sodium bicarbonate, potassium bicarbonate, sodium silicate, and potassium silicate.

3. The method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting according to claim 2, characterized in that: The additive amount is 0 to 2 times, preferably 1.1 to 1.5 times, the total molar amount of calcium chloride, magnesium chloride, manganese chloride, ferrous chloride and ferric chloride in the molten salt chloride slag.

4. The method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting according to claim 1, characterized in that: In the step (1), the molten salt chloride slag is fully reacted after adding the additive, the reaction temperature is 500° C. to 700° C., and the reaction time is 30 min to 180 min.

5. The method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting according to claim 1, characterized in that: In the step (2), the phase transformation product is crushed to less than 1 mm.

6. The method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting according to claim 1, characterized in that: The amount of water used in the water leaching process should be 1 to 1.3 times the saturated solubility of sodium chloride.

7. The method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting according to claim 1, characterized in that: In the step (2), the equipment used for crushing the phase transformation product is a jaw crusher, a ball mill, or a roller crusher; The solid-liquid separation method is one of pressure filtration, vacuum filtration and centrifugal filtration.

8. The method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting according to claim 1, characterized in that: In the step (3), during the leaching residue reduction process, the reducing agent is one of a carbon-based reducing agent, an aluminum-based reducing agent, and a silicon-based reducing agent.

9. The method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting according to claim 1, characterized in that: In the step (3), the reduction temperature of the leaching residue reduction process is 1400-1500° C., and the reduction time is 30-180 minutes.

10. The method for recovering valuable components from molten salt chloride slag by combined metallurgy and smelting according to claim 1, characterized in that: In the step (3), the magnetic concentrate obtained after grinding and magnetic separation of the reduction product has a main phase of metallic iron and manganese, and the grinding method includes but is not limited to conventional grinding equipment such as ball milling and rod milling, the grinding time is 30 to 60 minutes, and the pulp ratio is 1 to 3; The magnetic separation process is to separate the strongly magnetic metals iron and manganese in a weak magnetic field, and the magnetic field strength is 700-2000 GS.

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