Resource recycling method for tungsten slag

Abstract

The invention relates to the technical field of resource recycling of tungsten slag, in particular to a resource recycling method of tungsten slag. The resource recycling method for the tungsten slag is easy to operate, high in organic metal recycling rate, good in purification effect, low in energy consumption, high in medium complete recycling resource utilization rate and wide in application prospect.

Description

Technical Field

[0001] This invention relates to the field of resource recovery technology of tungsten slag, and more particularly to a method for resource recovery of tungsten slag. Background Technology

[0002] With tungsten resources becoming increasingly scarce, tungsten smelters are compelled to upgrade their technology to improve their economic efficiency. Production techniques are constantly evolving, typically employing either caustic soda high-temperature, high-pressure boiling or soda ash high-temperature, high-pressure boiling to decompose tungsten slag. The extract is then combined with the tungsten leaching solution obtained from boiling the concentrate, followed by impurity removal. However, the treatment of tungsten slag is limited to tungsten recovery. Therefore, researching and improving processes for the resource recovery of tungsten slag and the reduction of waste emissions is crucial for reducing overall costs, minimizing pollution, improving economic efficiency, and promoting the green and low-carbon development of the industry. Summary of the Invention

[0003] In view of this, the purpose of this invention is to provide a resource recovery method for tungsten slag. The resource recovery method is simple to operate, has a high organic metal recovery rate, good purification effect, low energy consumption, and high resource utilization rate due to complete recycling of the medium, and has broad application prospects.

[0004] To achieve the above-mentioned objectives, the present invention provides the following technical solution: This invention provides a method for the resource recovery of tungsten slag, comprising the following steps: The tungsten slag is acid-leached to obtain acid leaching solution and acid leaching residue; The acid leaching solution is extracted to obtain an extract and a raffinate; The extract was back-extracted to obtain a back-extract containing rare earth mixed solids; The back-extraction solution containing the rare earth mixed solids is subjected to a first solid-liquid separation to obtain a first sodium tungstate solution and rare earth hydroxides. The tungsten trioxide in the first sodium tungstate solution and the first tungsten precipitation agent are subjected to a first tungsten precipitation reaction to obtain the first artificial scheelite and the first tungsten precipitation liquid. The acid leaching residue is mixed with water and subjected to a second solid-liquid separation to obtain tungsten-tin concentrate and washing water. The tungsten-tin concentrate is subjected to sodium roasting, water leaching and third solid-liquid separation in sequence to obtain a second sodium tungstate solution and tin concentrate. The second sodium tungstate solution is subjected to impurity removal to obtain a purified solution and impurity removal residue; The purified liquid and the second tungsten precipitation agent are subjected to a second tungsten precipitation reaction to obtain the second artificial scheelite and the second tungsten precipitation liquid. The raffinate was subjected to iron precipitation with alkali to obtain ferric hydroxide and the precipitate solution after iron precipitation. The iron-precipitated liquid and soluble carbonates are used to precipitate manganese to obtain manganese carbonate and manganese-precipitated liquid. The manganese precipitation solution, the first tungsten precipitation solution, and the second tungsten precipitation solution are mixed and evaporated to obtain a composite salt.

[0005] Preferably, the acid solution used in the acid leaching is one or more of hydrochloric acid, sulfuric acid, acetic acid, nitric acid, and phosphoric acid; The ratio of the amount of tungsten slag to acid solution is 1g:(1~5)mL; The pH value after the acid leaching is completed is 0.5~3, and the acid leaching time is 1~4 hours.

[0006] Preferably, the extractant used in the extraction includes one or more of P204, TBP, P507, N1923, N263, N235, LK-N21, MIBK, C902, kerosene, and solvent oil; The back-extraction agent used includes one or more of sodium hydroxide solution, potassium hydroxide solution, and sodium carbonate solution; the concentration of the back-extraction agent is 1~3 mol / L.

[0007] Preferably, the pH value of the precipitated iron is 2.8~3.5, the temperature is 80~100℃, and the time is 1~2h.

[0008] Preferably, the soluble carbonate includes one or more of sodium carbonate, sodium bicarbonate, potassium carbonate, and ammonium bicarbonate.

[0009] Preferably, the mass ratio of the acid leaching residue to water is 1:(1~5).

[0010] Preferably, the sodium source used in the sodium roasting is one or more of sodium carbonate, sodium bicarbonate, sodium chloride, and sodium sulfate; The mass ratio of the sodium source to the tungsten-tin concentrate is (20~50):100.

[0011] Preferably, the sodium calcination temperature is 700~1000℃ and the time is 1.5~2.0h; The solid-liquid mass ratio of the water immersion is 1:(1~5), the water immersion temperature is 40~80℃, and the time is 1~4h.

[0012] Preferably, the purifying agent used for impurity removal includes one or more of magnesium chloride, magnesium sulfate, polyaluminum chloride, and polyferric chloride, and the mass percentage of the purifying agent in the second sodium tungstate solution is 1-5%. The purification temperature is 40~80℃, and the time is 1~22h.

[0013] Preferably, the first tungsten precipitation agent and the second tungsten precipitation agent independently include one or more of calcium oxide, calcium hydroxide and calcium chloride; The tungsten precipitation reaction is carried out at a temperature of 80-100℃ for 1-3 hours.

[0014] This invention provides a method for the resource recovery of tungsten slag, comprising the following steps: acid leaching of tungsten slag to obtain acid leaching solution and acid leaching residue; extraction of the acid leaching solution to obtain extract and raffinate; back-extraction of the extract to obtain back-extraction solution containing rare earth mixed solids; first solid-liquid separation of the back-extraction solution containing rare earth mixed solids to obtain a first sodium tungstate solution and rare earth hydroxide; first tungsten trioxide in the first sodium tungstate solution and a first tungsten precipitation agent to undergo a first tungsten precipitation reaction to obtain a first artificial scheelite and a first tungsten precipitation post-liquid; mixing the acid leaching residue with water and performing a second solid-liquid separation to obtain... Tungsten-tin concentrate and washing water; the tungsten-tin concentrate is sequentially subjected to sodium roasting, water leaching and a third solid-liquid separation to obtain a second sodium tungstate solution and tin concentrate; the second sodium tungstate solution is purified to obtain a purified liquid and a purified residue; the purified liquid and a second tungsten precipitation agent are subjected to a second tungsten precipitation reaction to obtain a second artificial scheelite and a second tungsten precipitation post-liquid; the raffinate and alkali are used to precipitate iron to obtain ferric hydroxide and an iron precipitation post-liquid; the iron precipitation post-liquid and soluble carbonate are used to precipitate manganese to obtain manganese carbonate and a manganese precipitation post-liquid; the manganese precipitation post-liquid, the first tungsten precipitation post-liquid and the second tungsten precipitation post-liquid are mixed and evaporated to obtain a composite salt.

[0015] Compared with the prior art, the present invention has at least the following beneficial effects: 1) The resource recycling method described in this invention can effectively solve the problem that currently only tungsten is recovered from tungsten slag while other valuable elements are wasted in landfills; 2) The resource recycling method described in this invention has a short reaction process, low energy consumption loss, low cost of complete medium recycling operation, is clean and environmentally friendly, and is more beneficial to environmental protection, and can realize industrial production; 3) The resource recycling method described in this invention extracts multiple metals in steps through a specific metal recycling sequence. After extraction, the wastewater problem is directly solved by evaporation, without generating secondary pollution, while effectively recovering valuable metals. Attached Figure Description

[0016] Figure 1 This is a schematic flowchart of the resource recovery method for tungsten slag described in this invention. Detailed Implementation

[0017] like Figure 1 As shown, the present invention provides a method for the resource recovery of tungsten slag, comprising the following steps: The tungsten slag is acid-leached to obtain acid leaching solution and acid leaching residue; The acid leaching solution is extracted to obtain an extract and a raffinate; The extract was back-extracted to obtain a back-extract containing rare earth mixed solids; The back-extraction solution containing the rare earth mixed solids is subjected to a first solid-liquid separation to obtain a first sodium tungstate solution and rare earth hydroxides. The tungsten trioxide in the first sodium tungstate solution and the first tungsten precipitation agent are subjected to a first tungsten precipitation reaction to obtain the first artificial scheelite and the first tungsten precipitation liquid. The acid leaching residue is mixed with water and subjected to a second solid-liquid separation to obtain tungsten-tin concentrate and washing water. The tungsten-tin concentrate is subjected to sodium roasting, water leaching and third solid-liquid separation in sequence to obtain a second sodium tungstate solution and tin concentrate. The second sodium tungstate solution is subjected to impurity removal to obtain a purified solution and impurity removal residue; The purified liquid and the second tungsten precipitation agent are subjected to a second tungsten precipitation reaction to obtain the second artificial scheelite and the second tungsten precipitation liquid. The raffinate was subjected to iron precipitation with alkali to obtain ferric hydroxide and the precipitate solution after iron precipitation. The iron-precipitated liquid and soluble carbonates are used to precipitate manganese to obtain manganese carbonate and manganese-precipitated liquid. The manganese precipitation solution, the first tungsten precipitation solution, and the second tungsten precipitation solution are mixed and evaporated to obtain a composite salt.

[0018] In this invention, unless otherwise specified, all raw materials used in the preparation are commercially available products well known to those skilled in the art.

[0019] This invention involves acid leaching tungsten slag to obtain an acid leaching solution and acid leaching residue.

[0020] This invention does not impose any special limitations on the source and composition of the tungsten slag; sources and compositions well known to those skilled in the art can be used. In the embodiments of this invention, the tungsten slag comprises 2.0 wt% tungsten trioxide, 4.96 wt% tin, 20.5 wt% iron, 15 wt% manganese, and 0.5 wt% total rare earth elements.

[0021] Prior to the acid leaching, the present invention preferably includes pulping; the present invention does not impose any special limitations on the pulping process, and any process well known to those skilled in the art can be used. In the embodiments of the present invention, the pulping is specifically carried out at a mass ratio of tungsten slag to water of 1:2 or 1:3.

[0022] In this invention, the acid solution used for acid leaching is preferably one or more of hydrochloric acid, sulfuric acid, acetic acid, nitric acid, and phosphoric acid, more preferably hydrochloric acid. When the acid solution is two or more of the above-mentioned specific selections, this invention does not impose any special limitation on the ratio of the above-mentioned specific substances, and they can be mixed in any ratio. In the embodiments of this invention, the acid solution can be hydrochloric acid.

[0023] In this invention, the preferred ratio of tungsten slag to acid solution is 1g:(1~5)mL, more preferably 1g:1mL, 1g:2mL, 1g:3mL, 1g:4mL, or 1g:5mL. In an embodiment of this invention, the ratio of tungsten slag to acid solution can be 1g:2mL.

[0024] The present invention does not impose any special limitations on the concentration of the acid solution. A concentration well known to those skilled in the art can be used, and the pH value after adding the above-mentioned amount of acid solution and performing acid leaching can be within the range specified below.

[0025] In this invention, the pH value after acid leaching is preferably 0.5-3, more preferably 0.5, 1, 1.5, 2, 2.5, or 3; the acid leaching time is preferably 1-4 hours, more preferably 1 hour, 2 hours, 3 hours, or 4 hours; and the acid leaching temperature is preferably room temperature. In embodiments of this invention, the final pH value after adding acid can be 0.5 or 1, and the acid leaching time can be 2 hours. In this invention, room temperature can be understood as no additional heating or cooling.

[0026] After the acid leaching is completed, the present invention preferably includes solid-liquid separation. The present invention does not impose any special limitations on the solid-liquid separation process, and any process known to those skilled in the art can be used.

[0027] After obtaining the acid leaching solution and acid leaching residue, the present invention extracts the acid leaching solution to obtain an extract and a raffinate.

[0028] In this invention, the extractant used for extraction preferably includes one or more of P204, TBP, P507, N1923, N263, N235, LK-N21, MIBK, C902, kerosene, and solvent oil, more preferably including P204, TBP, N235, and kerosene; the volume ratio of P204, TBP, N235, and kerosene is preferably 10:5:10:75.

[0029] In this invention, the volume ratio of the extractant to the acid leaching solution is preferably (0.5~3):1, more preferably 0.5:1, 1:1, 1.5:1, 2:1, 2.5:1 or 3:1. In an embodiment of this invention, the volume ratio of the extractant to the acid leaching solution can be 1:1.

[0030] The present invention does not impose any special limitations on the extraction process; any process known to those skilled in the art can be used.

[0031] In this invention, the extract preferably comprises rare earth ions and soluble tungstic acid. The raffinate is an iron-manganese solution.

[0032] After obtaining the extract and raffinate, the present invention back-extracts the extract to obtain a back-extract containing rare earth mixed solids.

[0033] In this invention, the back-extraction agent preferably includes one or more of sodium hydroxide solution, potassium hydroxide solution, and sodium carbonate solution. When the back-extraction agent is two or more of the above-mentioned specific selections, this invention does not impose any special limitation on the ratio of the above-mentioned specific substances, and any ratio can be used. In an embodiment of this invention, the back-extraction agent can be sodium hydroxide solution.

[0034] In this invention, the concentration of the stripping agent is preferably 1-3 mol / L, more preferably 1 mol / L, 2 mol / L, or 3 mol / L. In an embodiment of this invention, the concentration of the stripping agent can be 3 mol / L.

[0035] In this invention, the volume ratio of the extractant to the stripping agent is preferably 1:(3~5), more preferably 1:3, 1:3.5, 1:4, 1:4.5 or 1:5. In an embodiment of this invention, the volume ratio of the extractant to the stripping agent can be 1:3.

[0036] The present invention does not impose any special limitations on the back-extraction process; any process known to those skilled in the art can be used.

[0037] After obtaining the back-extraction solution containing the rare earth mixed solids, the present invention performs a first solid-liquid separation on the back-extraction solution containing the rare earth mixed solids to obtain a first sodium tungstate solution and rare earth oxides.

[0038] The present invention does not impose any special limitations on the first solid-liquid separation process; any process known to those skilled in the art can be used.

[0039] After the first solid-liquid separation is completed, the resulting filter cake is a rare earth oxide, and the filtrate is a first sodium tungstate solution. The rare earth oxide is sold, and the first sodium tungstate solution undergoes further post-processing.

[0040] After obtaining the first sodium tungstate solution and rare earth oxides, the present invention performs a first tungsten precipitation reaction on tungsten trioxide in the first sodium tungstate solution and a first tungsten precipitation agent to obtain the first artificial scheelite and the first tungsten precipitation liquid.

[0041] In this invention, the first tungsten precipitation agent preferably includes one or more of calcium oxide, calcium hydroxide, and calcium chloride. When the first tungsten precipitation agent is two or more of the above-mentioned specific selections, this invention does not have any special limitations on the ligands of the above-mentioned specific substances, and they can be mixed in any proportion. In an embodiment of this invention, the first tungsten precipitation agent can be calcium chloride.

[0042] In this invention, the mass ratio of tungsten trioxide to the first tungsten precipitation agent in the first sodium tungstate solution is preferably 1:(1~3), more preferably 1:1, 1:1.5, 1:2, 1:2.5 or 1:3. In an embodiment of this invention, the mass ratio of tungsten trioxide to the first tungsten precipitation agent in the first sodium tungstate solution can be 1:1.5.

[0043] In this invention, the temperature of the first tungsten deposition reaction is preferably 80~100℃, more preferably 80℃, 90℃ or 100℃; the time is preferably 1~3h, more preferably 1h, 2h or 3h.

[0044] In this invention, the first artificial scheelite is commercially available.

[0045] The present invention also includes mixing the acid leaching residue with water to perform a second solid-liquid separation to obtain tungsten-tin concentrate and washing water.

[0046] In this invention, the preferred ratio of the acid leaching residue to water is 1g:(1~5)mL, more preferably 1g:1mL, 1g:2mL, 1g:3mL, 1g:4mL, or 1g:5mL. In an embodiment of this invention, the mass ratio of the acid leaching residue to water can be 1g:2mL.

[0047] In this invention, the purpose of mixing the acid leaching residue with water is to wash the acid leaching residue. The washing temperature is preferably room temperature, and the washing time is preferably 0.5 to 4 hours, more preferably 0.5 hours, 1 hour, 1.5 hours, 2 hours, 2.5 hours, 3 hours, 3.5 hours, or 4 hours. In an embodiment of this invention, the washing time can be 2 hours. In this invention, room temperature can be understood as no additional heating or cooling is performed.

[0048] The present invention does not impose any special limitations on the second solid-liquid separation process; any process known to those skilled in the art can be used.

[0049] In this invention, the washing water is preferably reused in the above-mentioned pulping process.

[0050] After obtaining the tungsten-tin concentrate, the present invention sequentially performs sodium roasting, water leaching and third solid-liquid separation on the tungsten-tin concentrate to obtain a second sodium tungstate solution and tin concentrate.

[0051] In this invention, the sodium source used in the sodium roasting is preferably one or more of sodium carbonate, sodium bicarbonate, sodium chloride, and sodium sulfate, more preferably sodium carbonate, sodium chloride, and sodium sulfate; the mass ratio of sodium carbonate, sodium chloride, and sodium sulfate is preferably (15~30):(1~15):(1~10), more preferably 15:(1~15):(1~10), 20:(1~15):(1~10), 25:(1~15):(1~10), 30:(1~15):(1~10), (15~30):1:(1~10), (15~30):5:(1~10), (15~30):10:(1~10), (15~30):15:(1~10), (15~30):(1~15):1, (15~30):(1~15):5, or (15~30):(1~15):10. In embodiments of the present invention, the sodium source may be sodium carbonate, sodium chloride and sodium sulfate in a mass ratio of 25:1:4; or the sodium source may be sodium carbonate, sodium chloride and sodium sulfate in a mass ratio of 15:8:7.

[0052] In this invention, the preferred mass ratio of the sodium source to the tungsten-tin concentrate is (20-50):100, more preferably 20:100, 25:100, 30:100, 35:100, 40:100, 45:100, or 50:100. In an embodiment of this invention, the mass ratio of the sodium source to the tungsten-tin concentrate can be 30:100.

[0053] In this invention, the sodium calcination temperature is preferably 700~1000℃, more preferably 700℃, 750℃, 800℃, 850℃, 900℃, 950℃, or 100℃; the time is preferably 1.5~2.0h, more preferably 1.5h, 1.6h, 1.8h, or 2h. In an embodiment of this invention, the sodium calcination temperature can be 960℃ or 900℃, and the time can be 2h.

[0054] In this invention, the solid-liquid mass ratio of the water immersion is preferably 1:(1~5), more preferably 1:1, 1:2, 1:3, 1:4 or 1:5; the water immersion temperature is preferably 40~80℃, more preferably 40℃, 50℃, 60℃, 70℃ or 80℃; the immersion time is preferably 1~4h, more preferably 1h, 2h, 3h or 4h. In an embodiment of this invention, the solid-liquid mass ratio of the water immersion can be 1:3, the water immersion temperature can be 80℃, and the immersion time can be 2h.

[0055] The present invention does not impose any special limitations on the third solid-liquid separation process; any process well known to those skilled in the art can be used.

[0056] In this invention, the tin concentrate is preferably sold commercially.

[0057] After obtaining the second sodium tungstate solution, the present invention removes impurities from the second sodium tungstate solution to obtain a purified solution and impurity-removed residue.

[0058] In this invention, the purifying agent used for impurity removal preferably includes one or more of magnesium chloride, magnesium sulfate, polyaluminum chloride, and polyferric chloride. When the purifying agent is two or more of the above-mentioned specific selections, this invention does not impose any special limitation on the ratio of the above-mentioned specific substances, and they can be mixed in any ratio. In an embodiment of this invention, the purifying agent can be magnesium chloride.

[0059] In this invention, the mass percentage of the purifying agent in the second sodium tungstate solution is preferably 1-5%, more preferably 1%, 2%, 3%, 4%, or 5%. In an embodiment of this invention, the mass percentage of the purifying agent in the second sodium tungstate solution can be 3%.

[0060] In this invention, the temperature for impurity removal is preferably 40~80℃, more preferably 40℃, 50℃, 60℃, 70℃ or 80℃; the time is preferably 1~22h, more preferably 1h, 2h, 3h, 4h, 5h, 6h, 7h, 8h or 2h. In an embodiment of this invention, the temperature for impurity removal can be 65℃ and the time can be 2h.

[0061] After the impurity removal is completed, the present invention preferably includes solid-liquid separation. The present invention does not impose any special limitations on the solid-liquid separation process, and any process known to those skilled in the art can be used.

[0062] In this invention, the molybdenum-containing residue is preferably sold as a raw material to downstream manufacturers.

[0063] After obtaining the purified liquid, the present invention performs a second tungsten precipitation reaction on the purified liquid and the second tungsten precipitation agent to obtain a second artificial scheelite and a second tungsten precipitation liquid.

[0064] In this invention, the second tungsten precipitation agent preferably comprises one or more of calcium oxide, calcium hydroxide, and calcium chloride. When the second tungsten precipitation agent is two or more of the above-mentioned specific selections, this invention does not impose any special limitations on the ligands of the above-mentioned specific substances, and they can be mixed in any proportion. In an embodiment of this invention, the second tungsten precipitation agent can be calcium chloride.

[0065] In this invention, the mass ratio of tungsten trioxide and the second tungsten precipitation agent in the purification liquid is preferably 1:(1~2), more preferably 1:1, 1:1.5 or 1:2. In an embodiment of this invention, the mass ratio of the purification liquid and the second tungsten precipitation agent can be 1:1.2.

[0066] In this invention, the temperature of the second tungsten deposition reaction is preferably 80~100℃, more preferably 80℃, 90℃ or 100℃; the time is preferably 1~3h, more preferably 1h, 2h or 3h. In embodiments of this invention, the temperature of the second tungsten deposition reaction can be 90℃ or 95℃; the time can be 1h or 2h.

[0067] In this invention, the second artificial scheelite is sold.

[0068] After the tungsten precipitation reaction is completed, the present invention preferably includes solid-liquid separation. The present invention does not have any special limitations on the solid-liquid separation process, and any process known to those skilled in the art can be used.

[0069] The present invention also includes precipitating iron with the raffinate and alkali to obtain ferric hydroxide and iron precipitation solution.

[0070] This invention does not impose any special limitations on the type and amount of alkali used; any type and amount familiar to those skilled in the art can be employed. In an embodiment of this invention, the alkali may be a 3 mol / L sodium hydroxide solution.

[0071] In this invention, the pH value of the precipitated iron is preferably 2.8-3.5, more preferably 2.8, 2.9, 3.0, 3.1, 3.2, 3.3, 3.4, or 3.5; the temperature is preferably 80-100℃, more preferably 80℃, 90℃, or 100℃; and the time is preferably 1-2 hours, more preferably 1 hour, 1.2 hours, 1.4 hours, 1.6 hours, 1.8 hours, or 2 hours. In embodiments of this invention, the pH value of the precipitated iron can be 3.0 or 2.8, the temperature can be 80℃, and the time can be 2 hours.

[0072] After the iron deposition is completed, the present invention preferably includes solid-liquid separation. The present invention does not have any special limitations on the solid-liquid separation process, and any process known to those skilled in the art can be used.

[0073] After obtaining the iron precipitation liquid, the present invention further precipitates manganese with soluble carbonates to obtain manganese carbonate and manganese precipitation liquid.

[0074] In this invention, the soluble carbonate preferably includes one or more of sodium carbonate, sodium bicarbonate, potassium carbonate, and ammonium bicarbonate. When the soluble carbonate is two or more of the above-mentioned specific selections, this invention does not impose any special limitation on the ratio of the above-mentioned specific substances, and they can be mixed in any ratio. In an embodiment of this invention, the soluble carbonate can be sodium carbonate.

[0075] In this invention, the mass ratio of the iron-precipitated liquid to the soluble carbonate is preferably 1:(1~1.5), more preferably 1:1, 1:1.1, 1:1.2, 1:1.3, 1:1.4 or 1:1.5. In an embodiment of this invention, the mass ratio of the iron-precipitated liquid to the soluble carbonate can be 1:1.2.

[0076] In this invention, the temperature for manganese precipitation is preferably 60-80°C, more preferably 60°C, 65°C, 70°C, 75°C, or 80°C; the time is preferably 1-2 hours, more preferably 1 hour, 1.5 hours, or 2 hours. In embodiments of this invention, the temperature for manganese precipitation can be 60°C or 65°C, and the time can be 1 hour or 1.5 hours.

[0077] After the manganese precipitation is completed, the present invention preferably includes solid-liquid separation. The present invention does not have any special limitations on the solid-liquid separation process, and any process known to those skilled in the art can be used.

[0078] In this invention, the manganese carbonate is preferably industrial-grade manganese carbonate.

[0079] After obtaining the manganese precipitation solution, the first sodium tungstate solution, and the second sodium tungstate solution, the present invention mixes the manganese precipitation solution, the first tungsten precipitation solution, and the second tungsten precipitation solution and evaporates them to obtain a composite salt.

[0080] In this invention, the evaporation is preferably MVR evaporation. This invention does not impose any special limitations on the MVR evaporation process, and any process well known to those skilled in the art can be used.

[0081] In this invention, the composite salt is preferably sold as a snow removal agent.

[0082] The technical solutions of this invention will be clearly and completely described below with reference to the embodiments thereof. Obviously, the described embodiments are only a part of the embodiments of this invention, and not all of them. All other embodiments obtained by those skilled in the art based on the embodiments of this invention without creative effort are within the scope of protection of this invention.

[0083] Example 1 Tungsten slag: 2.00 wt% tungsten trioxide, 4.96 wt% tin, 20.5 wt% iron, 15 wt% manganese, and 0.5 wt% total rare earth elements; Using water or the wash water obtained after washing the acid leaching residue at a mass ratio of 1:2, 1000 kg of tungsten slag was slurried, and 2000 L of hydrochloric acid was added to adjust the pH value to 0.5. The reaction was carried out at room temperature for 2 hours, and the solid and liquid were separated to obtain acid leaching solution and acid leaching residue (700 kg, with a total leaching rate of 30.00%, tungsten trioxide 2.32 wt% (leaching rate of 18.80%), tin 6.96% (enriched), iron 1.36 wt% (leaching rate of 95.36%), manganese 0.3 wt% (leaching rate of 97.68%) and total rare earth 0.01 wt% (leaching rate of 97.19%)). At a volume ratio of 1:1, a composite extractant (P204 accounting for 10% of the mass of the acid leaching solution, TBP accounting for 5% of the mass of the acid leaching solution, N235 accounting for 10% of the mass of the acid leaching solution, and kerosene accounting for 75% of the mass of the acid leaching solution) was directly added to the acid leaching solution for extraction, resulting in an extract (extracting ions are rare earth ions and soluble tungstic acid) and a raffinate (iron-manganese solution). At a volume ratio of 1:3 (oil to water), a 3 mol / L sodium hydroxide solution was added to the extract for back-extraction, followed by solid-liquid separation to obtain rare earth oxides (for sale) and sodium tungstate solution (first sodium tungstate solution). According to the mass ratio of sodium tungstate solution to calcium chloride of 1:1.5, calcium chloride was added to the sodium tungstate solution, and the reaction was carried out at 85°C for 2 hours. After solid-liquid separation, artificial scheelite (for sale, with a tungsten trioxide content of more than 55 wt%, meeting the commercial standard) and the liquid after the first tungsten precipitation were obtained. After heating the raffinate to 80°C, add 3 mol / L sodium hydroxide solution to adjust the pH to 3.0 and perform iron precipitation for 2 hours. Separate the solid and liquid to obtain ferric hydroxide (with an iron content of more than 60 wt%, meeting the standard for desulfurization catalyst raw materials) and the liquid after iron precipitation. After heating the precipitated liquid to 60°C, sodium carbonate (molar ratio of sodium carbonate to manganese ions in the precipitated liquid 1:1.2) was added and precipitated for 1 hour. Solid-liquid separation was then performed to obtain industrial-grade manganese carbonate (manganese content of more than 43wt%, meeting industrial-grade standards) and manganese-precipitated liquid (sodium salt wastewater). The acid leaching residue and water were mixed at a mass ratio of 1:2 and washed at room temperature for 1 hour. Solid-liquid separation was performed to obtain tungsten-tin concentrate and washing water (the washing water was reused in the above pulping process). The tungsten-tin concentrate, sodium carbonate (15% by mass relative to the tungsten-tin concentrate), sodium chloride (8% by mass relative to the tungsten-tin concentrate), and sodium sulfate (7% by mass relative to the tungsten-tin concentrate) are mixed and subjected to sodium roasting (at a temperature of 960°C for 2 hours) to obtain the roasted material. The roasted material and water were mixed and leached at a solid-liquid ratio of 1g:3mL (temperature 80℃, time 2h). Solid-liquid separation was performed to obtain a soluble sodium tungstate solution (second sodium tungstate solution) and tin concentrate (for sale, with a tin content of 6-7wt%, meeting the sales standard for low-grade tin ore). After heating the soluble sodium tungstate solution to 60°C, magnesium chloride at a mass percentage of 3% relative to the soluble sodium tungstate solution is added for 2 hours to remove impurities. Solid-liquid separation is then performed to obtain impurity-removed residue (containing molybdenum, which is sold as raw material to downstream manufacturers) and purified liquid. According to the mass ratio of tungsten trioxide to calcium chloride in sodium tungstate solution of 1:1.2, the purified solution was heated to 90°C, calcium chloride was added, and the reaction was carried out at 90°C for 1 hour. Solid-liquid separation was performed to obtain artificial scheelite (for sale) and tungsten-precipitated liquid. The first tungsten precipitation liquid, the second tungsten precipitation liquid, and the manganese precipitation liquid are combined and subjected to MVR evaporation to obtain a composite salt (sold as a snow removal agent). In the calcium tungstate produced throughout the entire process, tungsten trioxide content is >55.0%, and tin enrichment reaches 7%. The overall recovery rates are 95% for tungsten, 98% for tin, 95% for iron, 96% for manganese, and 99% for total rare earth elements.

[0084] Example 2 Tungsten slag: 2.00 wt% tungsten trioxide, 4.96 wt% tin, 20.5 wt% iron, 15 wt% manganese, and 0.5 wt% total rare earth elements; Using water or the wash water obtained after washing the acid leaching residue at a mass ratio of 1:3, 15,000 kg of tungsten slag was slurried, and 3,000 L of hydrochloric acid was added to adjust the pH value to 1. The reaction was carried out at room temperature for 2 hours, and the solid and liquid were separated to obtain acid leaching solution and acid leaching residue (1,000 kg, with a total leaching rate of 33.00%, tungsten trioxide 2.22 wt% (leaching rate of 26.00%), tin 6.76% (enriched), iron 1.45 wt% (leaching rate of 95.28%), manganese 0.5 wt% (leaching rate of 96.30%) and total rare earth 0.015 wt% (leaching rate of 95.91%)). At a volume ratio of 1:1, a composite extractant (P204 accounting for 10% of the mass of the acid leaching solution, TBP accounting for 5% of the mass of the acid leaching solution, N235 accounting for 10% of the mass of the acid leaching solution, and kerosene accounting for 75% of the mass of the acid leaching solution) was directly added to the acid leaching solution for extraction, resulting in an extract (extracting ions are rare earth ions and soluble tungstic acid) and a raffinate (iron-manganese solution). At a volume ratio of 1:3 (oil to water), a 3 mol / L sodium hydroxide solution was added to the extract for back-extraction, followed by solid-liquid separation to obtain rare earth oxides (for sale) and sodium tungstate solution (first sodium tungstate solution). According to the mass ratio of tungsten trioxide to calcium chloride in sodium tungstate solution of 1:1.5, calcium chloride is added to the sodium tungstate solution, and the reaction is carried out at 88°C for 2 hours. After solid-liquid separation, artificial scheelite (for sale, with a tungsten trioxide content of more than 55wt%, meeting the commercial standard) and the liquid after the first tungsten precipitation are obtained. After heating the raffinate to 80°C, add 3 mol / L sodium hydroxide solution to adjust the pH to 2.8 and perform iron precipitation for 2 hours. After solid-liquid separation, obtain ferric hydroxide (with an iron content of more than 60 wt%, meeting the standard for desulfurization catalyst raw materials) and the liquid after iron precipitation. After heating the precipitated liquid to 65°C, sodium carbonate (molar ratio of sodium carbonate to manganese ions in the precipitated liquid 1:1.2) was added to precipitate for 1.5 hours. Solid-liquid separation was then performed to obtain industrial-grade manganese carbonate (manganese content of more than 43wt%, meeting industrial-grade standards) and manganese-precipitated liquid (sodium salt wastewater). The acid leaching residue and water were mixed at a mass ratio of 1:2 and washed at room temperature for 2 hours. Solid-liquid separation was performed to obtain tungsten-tin concentrate and washing water (the washing water was reused in the above pulping process). The tungsten-tin concentrate, sodium carbonate (25% by mass relative to the tungsten-tin concentrate), sodium chloride (1% by mass relative to the tungsten-tin concentrate), and sodium sulfate (4% by mass relative to the tungsten-tin concentrate) are mixed and subjected to sodium roasting (at 900°C for 2 hours) to obtain the roasted material. The roasted material and water were mixed and leached at a solid-liquid ratio of 1g:3mL (temperature 80℃, time 2h). Solid-liquid separation was performed to obtain a soluble sodium tungstate solution (second sodium tungstate solution) and tin concentrate (for sale, with a tin content of 6-7wt%, meeting the sales standard for low-grade tin ore). After heating the soluble sodium tungstate solution to 65°C, magnesium chloride at a mass percentage of 3% relative to the soluble sodium tungstate solution is added for 2 hours to remove impurities. Solid-liquid separation is then performed to obtain impurity-removed residue (containing molybdenum, which is sold as raw material to downstream manufacturers) and purified liquid. According to the mass ratio of tungsten trioxide to calcium chloride in sodium tungstate solution of 1:1.2, the purified solution was heated to 95°C, calcium chloride was added, and the reaction was carried out at 90°C for 2 hours. Solid-liquid separation was performed to obtain artificial scheelite (for sale) and tungsten-precipitated liquid. The first tungsten precipitation liquid, the second tungsten precipitation liquid, and the manganese precipitation liquid are combined and subjected to MVR evaporation to obtain a composite salt (sold as a snow removal agent). In the calcium tungstate produced throughout the entire process, tungsten trioxide content is >55.0%, and tin enrichment reaches 6.76%. The overall recovery rates are 95% for tungsten, 99% for tin, 96% for iron, 96% for manganese, and 99.5% for total rare earth elements.

[0085] Comparative Example 1 Referring to Example 1, the difference is that, according to a 1:3 mass ratio, water or the washing water obtained after washing the acid leaching residue was used to slurry 1500 kg of tungsten slag, and 2000 L of hydrochloric acid was added to adjust the pH value to 3. The reaction was carried out at room temperature for 2 hours, and the solid and liquid were separated to obtain acid leaching solution and acid leaching residue (1200 kg, with a total leaching rate of 20.00%, tungsten trioxide 2.5 wt% (enriched), tin 6.76% (enriched), iron 22 wt% (enriched), manganese 14 wt% (leaching rate of 1.33%) and total rare earth 0.215 wt% (leaching rate of 34.4%)). It is evident that the low acidity of the acid leaching process results in poor leaching of various elements, making it impossible to leach the metals into the solution. Most of the metals remain in the acid leaching residue and cannot be leached out, which is far from the intended process route. Therefore, there is no need to conduct further comparative experiments on extraction and separation.

[0086] The above description is merely a preferred embodiment of the present invention and is not intended to limit the present invention in any way. It should be noted that those skilled in the art can make various improvements and modifications without departing from the principles of the present invention, and these improvements and modifications should also be considered within the scope of protection of the present invention.

Claims

1. A method for the resource recovery of tungsten slag, characterized in that, Includes the following steps: The tungsten slag is acid-leached to obtain an acid leaching solution and acid leaching residue; The acid leaching solution is extracted to obtain an extract and a raffinate; The extract was back-extracted to obtain a back-extract containing rare earth mixed solids; The back-extraction solution containing the rare earth mixed solids is subjected to a first solid-liquid separation to obtain a first sodium tungstate solution and rare earth hydroxides. The tungsten trioxide in the first sodium tungstate solution and the first tungsten precipitation agent are subjected to a first tungsten precipitation reaction to obtain the first artificial scheelite and the first tungsten precipitation liquid. The acid leaching residue is mixed with water and subjected to a second solid-liquid separation to obtain tungsten-tin concentrate and washing water. The tungsten-tin concentrate was subjected to sodium roasting, water leaching and third solid-liquid separation in sequence to obtain a second sodium tungstate solution and tin concentrate. The second sodium tungstate solution is purified to obtain a purified solution and a purified residue. The purified liquid and the second tungsten precipitation agent are subjected to a second tungsten precipitation reaction to obtain the second artificial scheelite and the second tungsten precipitation liquid. The raffinate was subjected to iron precipitation with alkali to obtain ferric hydroxide and the precipitate solution after iron precipitation. The iron precipitation solution and soluble carbonate are used to precipitate manganese to obtain manganese carbonate and manganese precipitation solution. The manganese precipitation solution, the first tungsten precipitation solution, and the second tungsten precipitation solution are mixed and evaporated to obtain a composite salt.

2. The resource recycling method as described in claim 1, characterized in that, The acid solution used for the acid leaching is one or more of hydrochloric acid, sulfuric acid, acetic acid, nitric acid, and phosphoric acid; The ratio of tungsten slag to acid solution is 1g:(1~5)mL; The pH value after the acid leaching is completed is 0.5~3, and the acid leaching time is 1~4 hours.

3. The resource recycling method as described in claim 1, characterized in that, The extraction solvent used includes one or more of the following: P204, TBP, P507, N1923, N263, N235, LK-N21, MIBK, C902, kerosene, and solvent oil. The back-extraction agent used includes one or more of sodium hydroxide solution, potassium hydroxide solution and sodium carbonate solution; the concentration of the back-extraction agent is 1~3 mol / L.

4. The resource recycling method as described in claim 1, characterized in that, The pH value of the precipitated iron is 2.8~3.5, the temperature is 80~100℃, and the time is 1~2h.

5. The resource recycling method as described in claim 1, characterized in that, The soluble carbonates include one or more of sodium carbonate, sodium bicarbonate, potassium carbonate, and ammonium bicarbonate.

6. The resource recycling method as described in claim 1, characterized in that, The mass ratio of the acid leaching residue to water is 1:(1~5).

7. The resource recycling method as described in claim 1, characterized in that, The sodium source used in the sodium roasting is one or more of sodium carbonate, sodium bicarbonate, sodium chloride, and sodium sulfate. The mass ratio of the sodium source to the tungsten-tin concentrate is (20~50):

100.

8. The resource recycling method as described in claim 7, characterized in that, The sodium calcination temperature is 700~1000℃, and the time is 1.5~2.0h; The solid-liquid mass ratio of the water immersion is 1:(1~5), the water immersion temperature is 40~80℃, and the time is 1~4h.

9. The resource recycling method as described in claim 1, characterized in that, The purification agent used for impurity removal includes one or more of magnesium chloride, magnesium sulfate, polyaluminum chloride, and polyferric chloride, and the mass percentage of the purification agent in the second sodium tungstate solution is 1-5%. The purification temperature is 40~80℃, and the time is 1~22h.

10. The resource recycling method as described in claim 1, characterized in that, The first tungsten precipitation agent and the second tungsten precipitation agent independently include one or more of calcium oxide, calcium hydroxide and calcium chloride; The tungsten precipitation reaction is carried out at a temperature of 80-100℃ for 1-3 hours.

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