Sodium salt recycling method based on vanadium slag sodium salt roasting-water leaching vanadium extraction

Abstract

The invention belongs to the technical field of vanadium extraction metallurgy, and particularly relates to a sodium salt recycling method based on vanadium slag sodium salt roasting-water leaching vanadium extraction, which comprises the following steps: a) uniformly mixing vanadium slag and sodium salt, and then carrying out sodium salt roasting to obtain sodium salt clinker; (b) the sodium modified clinker is subjected to water leaching and solid-liquid separation, and leaching residues and leaching liquid are obtained; c) removing chromium, silicon and phosphorus from the leachate, and then evaporating and concentrating to obtain high-concentration vanadium liquid; d) adding an ammonium compound into the high-concentration vanadium liquid to precipitate vanadium, and carrying out solid-liquid separation to obtain ammonium metavanadate and a vanadium precipitation supernatant; e) introducing CO2 into the vanadium precipitation supernatant to adjust the pH to crystallize sodium bicarbonate, carrying out solid-liquid separation to obtain sodium bicarbonate and crystallization mother liquor, and returning the sodium bicarbonate to the step a); and f) heating the crystallization mother liquor to decompose ammonium bicarbonate for deamination to obtain deamination liquid and ammonia-containing gas, returning the deamination liquid to the step b), and returning the ammonia-containing gas to the step d) to be absorbed by the high-concentration vanadium liquid. The method provided by the invention has good economic benefits and environmental benefits.

Description

Technical Field

[0001] The invention belongs to the technical field of vanadium extraction and metallurgy, and particularly relates to a sodium salt recycling method based on sodium roasting of vanadium slag and water leaching to extract vanadium. Background Art

[0002] Vanadium slag is the main raw material for the production of vanadium oxide. The traditional industrial production process is sodium roasting and water leaching to extract vanadium. During the sodium roasting process of vanadium slag, part of the chromium reacts with sodium salts to form water-soluble sodium chromate, which enters the solution during the sodium clinker water leaching process, forming an alkaline leachate of sodium vanadate containing chromium; after the leachate is dephosphorized with calcium chloride, sulfuric acid is used to adjust the pH of the leachate to ≈ 2 and ammonium sulfate is added to precipitate vanadium to obtain the intermediate product ammonium polyvanadate. The vanadium precipitation wastewater mainly contains Na + NH4 + 、SO4 2- 、V 5+ Cr 6+ 、SiO3 2- etc., by adding reducing agents such as sodium metabisulfite, sodium bisulfite or sodium sulfite to V 5+ Cr 6+ Restore to V 4+ Cr 3+, The solution is then adjusted to a pH of 7-9 using sodium hydroxide, allowing vanadium, chromium, and silicon to precipitate, forming a vanadium-chromium reduction sludge. After separating the vanadium, chromium, and silicon, the solution is further adjusted to a pH of approximately 12 using sodium hydroxide. Ammonia is then removed by stripping, absorbed with sulfuric acid, and recycled for vanadium precipitation. The deammoniated wastewater is evaporated, concentrated, and crystallized to produce sodium sulfate as a solid waste. Every ton of vanadium pentoxide produced generates approximately 2.5-3 tons of sodium sulfate as solid waste. In the existing sodium roasting-water leaching process for vanadium extraction, sodium roasting consumes large amounts of sodium carbonate and involves complex water treatment, resulting in high vanadium oxide production costs. Treatment of vanadium precipitation wastewater generates large amounts of sodium sulfate as solid waste, posing significant environmental risks. Recycling sodium salts, reducing vanadium oxide production costs, and minimizing sodium sulfate as solid waste from water treatment are key challenges facing the vanadium industry.

[0003] To achieve the recycling of sodium salts during the vanadium slag extraction process, Publication No. CN 116837230A provides a method for clean vanadium extraction from vanadium slag and its application. The main technical concept involves using calcium oxide to remove silicon and phosphorus from the sodium-containing leachate, then using the calcium oxide to precipitate vanadium, producing an intermediate product, calcium vanadate, and a vanadium precipitation mother liquor. The vanadium precipitation mother liquor is then passed through CO2 to crystallize sodium bicarbonate, which is then returned to the roasting process as a sodium-forming additive. The sodium bicarbonate mother liquor is separated to remove chromium and then recycled for clinker leaching. The calcium vanadate is leached with sodium bicarbonate, evaporated and concentrated to produce solid sodium metavanadate. This solid is then mixed with an aluminum additive, reduced at high temperature, and leached to separate vanadium and sodium. The sodium-containing solution is then passed through CO2 to crystallize sodium bicarbonate, which is then returned to the roasting process as a sodium-forming additive. This process achieves the recycling of sodium salts, but it also presents challenges such as a lengthy process flow and high water evaporation.

[0004] Publication No. CN 116854109A provides a method for producing sodium carbonate and ammonium sulfate from sodium sulfate. The main technical concept involves dissolving the sodium sulfate and mixing it with ammonium bicarbonate, causing a double decomposition reaction to produce solid sodium bicarbonate. The crystallization mother liquor is then evaporated and crystallized to produce a mixed salt of sodium sulfate and ammonium sulfate, which is then separated to produce ammonium sulfate. The sodium bicarbonate can be recycled as a sodium-forming additive for vanadium slag roasting. A small amount of ammonium sulfate can be recycled as an ammonium salt in the vanadium precipitation process, with the majority of the ammonium sulfate being sold externally. While this method achieves the recycling of sodium salts, it also presents challenges such as a longer vanadium oxide production process and the addition of ammonium sulfate as a byproduct.

[0005] The problem of how to achieve low-cost recycling of sodium salt and reduce the generation of sodium sulfate as solid waste from water treatment based on the existing vanadium slag sodium roasting-water leaching vanadium production process and equipment remains to be solved. At present, there are no relevant technical reports that can effectively solve the above problems and have prospects for industrial application. Summary of the Invention

[0006] In view of this, the object of the present invention is to provide a sodium salt recycling method based on sodium roasting of vanadium slag and water leaching of vanadium, which can realize the recycling of sodium salt, reduce water evaporation, reduce the production cost of preparing vanadium oxide from vanadium slag, and avoid the generation of sodium sulfate as solid waste in water treatment.

[0007] The present invention provides a sodium salt recycling method based on sodium roasting of vanadium slag and water leaching to extract vanadium, comprising the following steps:

[0008] a) mixing vanadium slag and sodium salt and then performing sodium roasting to obtain sodium clinker; wherein the sodium salt includes sodium carbonate and / or sodium bicarbonate;

[0009] b) subjecting the sodium-treated clinker to water leaching and solid-liquid separation to obtain a leaching residue and a leachate;

[0010] c) removing chromium, silicon and phosphorus from the leachate and then evaporating and concentrating it to 50-60% of its original volume to obtain a high-concentration vanadium solution; wherein the chromium removal agent is one or more of hydrazine, hydroxylamine, formaldehyde, methanol, ethanol, sucrose and starch, and the silicon and phosphorus removal agent is sodium aluminate;

[0011] d) adding ammonium compound to the high-concentration vanadium solution to precipitate vanadium, and separating the solid and liquid to obtain ammonium metavanadate solid and vanadium precipitation upper layer liquid; wherein the ammonium compound is one or more of ammonium carbonate, ammonium bicarbonate and ammonia water;

[0012] e) passing CO2 through the upper layer of the vanadium precipitate to adjust the pH and crystallize sodium bicarbonate, performing solid-liquid separation to obtain sodium bicarbonate solid and crystallization mother liquor; the sodium bicarbonate solid is returned to step a) for mixing with the vanadium slag and then sodiumization and roasting;

[0013] f) heating the crystallization mother liquor to decompose ammonium bicarbonate for deamination, thereby obtaining a deammonified liquid and an ammonia-containing gas; the deammonified liquid is returned to step b) to participate in water leaching; the ammonia-containing gas is returned to step d) to participate in vanadium precipitation after being absorbed by the high-concentration vanadium solution.

[0014] Preferably, in step a), the molar ratio of the sodium salt calculated as Na2O to the vanadium slag calculated as V2O5 is (2.5-3.5):1.

[0015] Preferably, in step a), the temperature of the sodium calcination is 750-850° C., and the time is 60-120 min.

[0016] Preferably, in step a), the vanadium slag is obtained by blowing vanadium-containing molten iron in a converter.

[0017] Preferably, in step c), the leaching solution simultaneously removes chromium, silicon and phosphorus, or the leaching solution first removes chromium and then removes silicon and phosphorus.

[0018] Preferably, in step c), the reaction temperature for removing chromium is 70-200°C, and the reaction time is 30-210 min; the reaction temperature for removing silicon and phosphorus is 60-100°C, and the reaction time is 20-60 min, and the reaction pH is 8.5-10.5.

[0019] Preferably, in step c), the molar ratio of the sodium aluminate calculated as Al to the leachate calculated as Si is (0.8-1.2):1.

[0020] Preferably, in step c), the Cr concentration of the leachate after silicon, phosphorus and chromium are removed is ≤1 g / L.

[0021] Preferably, in step d), NH4 + The molar ratio of the ammonium compound to the high-concentration vanadium solution is (2.8-3.5):1; the reaction temperature of the vanadium precipitation is 10-20° C., the reaction pH is 9.3-9.8, and the reaction time is 3-5 hours.

[0022] Preferably, in step e), the pH value of the upper layer of the vanadium precipitation liquid is adjusted to 7.5-8.5 by passing CO2.

[0023] Compared with the prior art, the present invention provides a sodium salt recycling method based on sodium roasting of vanadium slag and water leaching of vanadium, comprising the following steps: a) mixing vanadium slag with sodium salt and then performing sodium roasting to obtain sodium clinker; wherein the sodium salt comprises sodium carbonate and / or sodium bicarbonate; b) leaching the sodium clinker in water, performing solid-liquid separation to obtain leaching residue and leachate; c) removing chromium, silicon and phosphorus from the leachate and then evaporating and concentrating it to 50-60% of the original volume to obtain a high-concentration vanadium solution; wherein the chromium removal agent is one or more of hydrazine, hydroxylamine, formaldehyde, methanol, ethanol, sucrose and starch, and the silicon and phosphorus removal agent is sodium aluminate; d) adding the sodium clinker to the high-concentration vanadium solution. Concentrated vanadium liquid is added with ammonium compound to precipitate vanadium, and the solid and liquid are separated to obtain ammonium metavanadate solid and upper layer liquid of vanadium precipitation; wherein the ammonium compound is one or more of ammonium carbonate, ammonium bicarbonate and ammonia water; e) CO2 is passed through the upper layer liquid of vanadium precipitation to adjust the pH to crystallize sodium bicarbonate, and the solid and liquid are separated to obtain sodium bicarbonate solid and crystallization mother liquor; the sodium bicarbonate solid is returned to step a) to be mixed with vanadium slag and then sodiumized and roasted; f) the crystallization mother liquor is heated to decompose ammonium bicarbonate to deaminate, to obtain deaminated liquid and ammonia-containing gas; the deaminated liquid is returned to step b) to participate in water immersion; the ammonia-containing gas is returned to step d) to be absorbed by the high-concentration vanadium liquid and then participate in vanadium precipitation. The method provided by the present invention can realize the recycling of sodium salts, reduce water evaporation, reduce the production cost of preparing vanadium oxide from vanadium slag, and avoid the generation of solid waste sodium sulfate from water treatment. More specifically, it has at least the following beneficial effects:

[0024] (1) The sodium salt in the solution system is separated in the form of sodium bicarbonate and returned to the roasting process for recycling as a sodium additive, which reduces the consumption of sodium salt and avoids the generation of solid waste sodium sulfate.

[0025] (2) The upper layer of vanadium precipitation liquid is treated by heating for deammoniation and the vanadium solution directly absorbs ammonia and then recycles it, which saves the consumption of sodium hydroxide and sulfuric acid in the existing water treatment for deammoniation.

[0026] (3) After a small amount of process water is evaporated from the leachate, vanadium is precipitated, and no evaporation crystallization process is subsequently introduced, so the evaporation amount of process water is greatly reduced. BRIEF DESCRIPTION OF THE DRAWINGS

[0027] In order to more clearly illustrate the embodiments of the present invention or the technical solutions in the prior art, the following briefly introduces the drawings required for use in the embodiments or the description of the prior art. Obviously, the drawings described below are merely embodiments of the present invention. For ordinary technicians in this field, other drawings can be obtained based on the provided drawings without paying any creative work.

[0028] Figure 1 It is a process flow chart provided by an embodiment of the present invention. DETAILED DESCRIPTION

[0029] The following will clearly and completely describe the technical solutions in the embodiments of the present invention in conjunction with the accompanying drawings. 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 efforts are within the scope of protection of the present invention.

[0030] The present invention provides a sodium salt recycling method based on sodium roasting of vanadium slag and water leaching of vanadium, such as Figure 1 As shown, the following steps are included:

[0031] a) mixing vanadium slag and sodium salt and then performing sodium roasting to obtain sodium clinker; wherein the sodium salt includes sodium carbonate and / or sodium bicarbonate;

[0032] b) subjecting the sodium-treated clinker to water leaching and solid-liquid separation to obtain a leaching residue and a leachate;

[0033] c) removing chromium, silicon and phosphorus from the leachate and then evaporating and concentrating it to 50-60% of its original volume to obtain a high-concentration vanadium solution; wherein the chromium removal agent is one or more of hydrazine, hydroxylamine, formaldehyde, methanol, ethanol, sucrose and starch, and the silicon and phosphorus removal agent is sodium aluminate;

[0034] d) adding ammonium compound to the high-concentration vanadium solution to precipitate vanadium, and separating the solid and liquid to obtain ammonium metavanadate solid and vanadium precipitation upper layer liquid; wherein the ammonium compound is one or more of ammonium carbonate, ammonium bicarbonate and ammonia water;

[0035] e) passing CO2 through the upper layer of the vanadium precipitate to adjust the pH and crystallize sodium bicarbonate, performing solid-liquid separation to obtain sodium bicarbonate solid and crystallization mother liquor; the sodium bicarbonate solid is returned to step a) for mixing with the vanadium slag and then sodiumization and roasting;

[0036] f) heating the crystallization mother liquor to decompose ammonium bicarbonate for deamination, thereby obtaining a deammonified liquid and an ammonia-containing gas; the deammonified liquid is returned to step b) to participate in water leaching; the ammonia-containing gas is returned to step d) to participate in vanadium precipitation after being absorbed by the high-concentration vanadium solution.

[0037] In the method provided herein, in step a), the vanadium slag is mixed with the sodium salt and then roasted. This allows the vanadium in the vanadium slag to react with the sodium salt to form soluble sodium vanadate, facilitating the dissolution of the vanadium into the solution during the subsequent water leaching process. Furthermore, during the roasting process, a small amount of chromium in the vanadium slag reacts with the sodium to form water-soluble sodium chromate.

[0038] In the method provided herein, in step a), the vanadium slag is preferably obtained by converting vanadium-containing molten iron through vanadium extraction and blowing in a converter. The main components of the vanadium slag preferably include V2O5, FeO, SiO2, TiO2, CaO, MgO, MnO, and Cr2O3. The particle size of the vanadium slag is preferably ≤300 μm, more preferably ≤200 μm, even more preferably ≤150 μm, and most preferably ≤125 μm (120 mesh).

[0039] In the method provided by the present invention, in step a), the molar ratio of the sodium salt calculated as Na2O to the vanadium slag calculated as V2O5 is preferably (2.5-3.5):1, specifically 2.5:1, 2.6:1, 2.7:1, 2.8:1, 2.9:1, 3:1, 3.1:1, 3.2:1, 3.3:1, 3.4:1 or 3.5:1.

[0040] In the method provided by the present invention, in step a), the temperature of the sodium calcination is preferably 750-850°C, specifically 750°C, 760°C, 770°C, 780°C, 790°C, 800°C, 810°C, 820°C, 830°C, 840°C or 850°C; the time of the sodium calcination is preferably 60-120 min, specifically 60 min, 65 min, 70 min, 75 min, 80 min, 85 min, 90 min, 95 min, 100 min, 105 min, 110 min, 115 min or 120 min.

[0041] In the method provided by the present invention, in step a), after completing the sodium roasting, the obtained sodium clinker is preferably ground. The particle size of the ground sodium clinker is preferably ≤300 μm, more preferably ≤200 μm, further preferably ≤150 μm, and most preferably ≤125 μm (120 mesh).

[0042] In the method provided by the present invention, in step b), the sodiumized clinker is immersed in water so that the water-soluble sodium vanadate in the clinker is dissolved into the solution to obtain a leachate. The leachate sodium concentration has a greater impact on the vanadium precipitation effect of acidic ammonium salts. When the sodium concentration is high, the acidic ammonium salts are prone to sodium vanadate hydrolysis, resulting in a high content of sodium ammonium polyvanadate. In severe cases, a "vanadium bubble" phenomenon will occur, affecting the normal progress of the vanadium precipitation process. Therefore, the conventional vanadium oxide production generally controls the leachate vanadium concentration to be no more than 30g / L (essence is to control the sodium concentration). Since the present invention circulates the deammoniation liquid for leaching clinker, the leachate sodium concentration will increase compared with conventional processes. Therefore, subsequent vanadium precipitation selects the weakly alkaline ammonium salt vanadium precipitation mode to avoid the impact of the increased sodium concentration on the vanadium precipitation effect.

[0043] In the method provided by the present invention, in step c), the main impurity elements contained in the leachate are chromium, silicon and phosphorus. Silicon and phosphorus are removed by adding sodium aluminate to form sodium aluminosilicate and aluminum phosphate for precipitation and separation; since excess aluminum will hydrolyze and precipitate in the form of aluminum hydroxide, the addition of sodium aluminate will not introduce new impurities. The chromium removal reagent is selected from one or more of hydrazine, hydroxylamine, formaldehyde, methanol, ethanol, sucrose and starch, and the products after the redox reaction with sodium chromate are mainly gases such as N2 and NH3 and sodium carbonate and sodium hydroxide. No new impurity ions are introduced into the solution system, and the recycling of process water and sodium salt is not affected. The present invention chooses to use an organic reducing agent to remove chromium, which can avoid the problem that sulfur-containing reducing agents such as sodium metabisulfite, sodium sulfite, sodium bisulfite, sodium sulfide, and sodium hydrosulfide, which are commonly used in existing chromium removal technologies, introduce sulfur, affecting the recovery of sodium salt in the form of sodium bicarbonate or sodium carbonate and the recycling of process water.

[0044] In the method provided by the present invention, in step c), the leaching solution is simultaneously used to remove chromium, silicon and phosphorus, and the obtained impurity-removed slag is mainly a mixed slag containing chromium and silicon. The removal can also be carried out in steps, first removing chromium, and then removing silicon and phosphorus. This method is conducive to the full precipitation of chromium and silicon, while reducing the number of solid-liquid separations of the solution.

[0045] In the method provided by the present invention, in step c), the reaction temperature for removing chromium is preferably 70-200°C, specifically 70°C, 75°C, 80°C, 85°C, 90°C, 95°C, 100°C, 120°C, 140°C, 160°C, 180°C or 200°C; the reaction time is preferably 30-210 min, specifically 30 min, 45 min, 60 min, 75 min, 90 min, 105 min, 120 min, 130 min, 150 min, 180 min or 210 min.

[0046] In the method provided by the present invention, in step c), the reaction temperature for removing silicon and phosphorus again is preferably 60-100°C, specifically 60°C, 65°C, 70°C, 75°C, 80°C, 85°C, 90°C, 95°C or 100°C; the reaction time is preferably 20-60 min, specifically 20 min, 25 min, 30 min, 35 min, 40 min, 45 min, 50 min, 55 min or 60 min; and the reaction pH is preferably 8.5-10.5, specifically 8.5, 9, 9.5, 10 or 10.5.

[0047] In the method provided by the present invention, in step c), the molar ratio of the sodium aluminate calculated as Al to the leachate calculated as Si is preferably (0.8-1.2):1, specifically 0.8:1, 0.85:1, 0.9:1, 0.95:1, 1:1, 1.05:1, 1.1:1, 1.15:1 or 1.2:1.

[0048] In the method provided herein, in step c), since the solution after chromium removal undergoes vanadium precipitation with ammonium salts, crystallization and separation of sodium bicarbonate, and subsequent recycling for leaching, chromium is not enriched in the solution system. Therefore, the chromium concentration of the solution after chromium removal is controlled to not affect the quality of the vanadium product obtained by vanadium precipitation with weakly alkaline ammonium salts. In the present invention, the Cr content of the leachate after silicon, phosphorus, and chromium removal is preferably controlled to ≤1 g / L.

[0049] In the method provided by the present invention, in step d), the ammonium carbonate, ammonium bicarbonate, and ammonia water are selected as the ammonium compound for precipitating vanadium because the sodium salt in the solution needs to be crystallized and separated in the form of sodium bicarbonate. The introduction of carbonate ions is beneficial to the subsequent separation of sodium salts, and also avoids the introduction of SO4 by ammonium salts such as ammonium sulfate and ammonium chloride. 2- 、Cl - Affects the circulation of sodium salt and process water.

[0050] In the method provided by the present invention, in step d), the pH value of the vanadium precipitation reaction is preferably 9.3 to 9.8, specifically 9.3, 9.4, 9.5, 9.6, 9.7 or 9.8. The present invention preferably controls the pH value of the vanadium precipitation reaction to be 9.3 to 9.8 because the solubility of sodium bicarbonate is relatively small, and a higher pH value of vanadium precipitation can increase the CO3 in the solution. 2- concentration, and correspondingly reduce HCO3 - concentration to avoid a large amount of sodium ions crystallizing out in the form of sodium bicarbonate during the vanadium precipitation process.

[0051] In the method provided by the present invention, in step d), the reaction time, reaction temperature and amount of ammonium compound used in the vanadium precipitation can be selected and determined according to the vanadium precipitation rate index of the weak alkaline ammonium salt vanadium precipitation test. + The molar ratio of the ammonium compound to the high-concentration vanadium solution in terms of V is preferably (2.8-3.5):1, specifically 2.8:1, 2.9:1, 3:1, 3.1:1, 3.2:1, 3.3:1, 3.4:1 or 3.5:1; the reaction temperature for vanadium precipitation is preferably 10-20°C, specifically 10°C, 11°C, 12°C, 13°C, 14°C, 15°C, 16°C, 17°C, 18°C, 19°C or 20°C; the reaction time for vanadium precipitation is preferably 3-5h, specifically 3h, 3.5h, 4h, 4.5h or 5h.

[0052] In the method provided by the present invention, in step e), the upper layer of vanadium precipitation solution mainly contains Na + NH4 + 、CO3 2- 、HCO3 - VO3 - The purpose of passing CO2 into the upper layer of vanadium precipitation is to lower the pH value of the solution and make the CO3 2- Converted to HCO3 - Since the solubility of sodium bicarbonate is much lower than that of sodium carbonate, ammonium carbonate, ammonium bicarbonate and sodium vanadate, the sodium ions in the solution will crystallize out in the form of sodium bicarbonate, thus recovering the sodium salt in the solution. The recovered sodium bicarbonate is recycled as a sodium-forming additive in the sodium-forming roasting process of step a). The sodium bicarbonate decomposes into sodium carbonate during the high-temperature roasting process, which has the same vanadium conversion function as the existing vanadium slag added with sodium carbonate for roasting.

[0053] In the method provided by the present invention, in step e), the pH value of the upper layer liquid of the vanadium precipitation is preferably adjusted to 7.5-8.5 by passing CO2, specifically 7.5, 7.6, 7.7, 7.8, 7.9, 8, 8.1, 8.2, 8.3, 8.4 or 8.5.

[0054] In the method provided by the present invention, in step f), the crystallization mother liquor after separating sodium bicarbonate mainly contains sodium vanadate, sodium bicarbonate, ammonium bicarbonate, etc. High-temperature heating can decompose the ammonium bicarbonate into CO2 and NH3, which escape into the solution system, thereby reducing the ammonia concentration of the solution. Deamination of the crystallization mother liquor is intended to prevent excessive ammonia concentration in the solution recycled for sodium clinker leaching. The volatilization of ammonia during the clinker leaching process affects the operating environment, and the high ammonia concentration in the leachate leads to crystallization and precipitation of ammonium metavanadate, resulting in vanadium loss. The ammonia-containing gas generated during the deamination process is absorbed by the high-concentration vanadium solution, achieving low-cost recycling of ammonia.

[0055] In the method provided by the present invention, in step f), the heating temperature is preferably above 80°C, more preferably 80-100°C, specifically 80°C, 85°C, 90°C, 95°C or 100°C.

[0056] For the purpose of greater clarity, the present invention is described in detail with reference to the following examples.

[0057] Example 1

[0058] A sodium salt recycling method based on sodium roasting of vanadium slag and water leaching to extract vanadium comprises the following steps:

[0059] 1000g of converter vanadium slag (main chemical components: V2O5 15.35wt%, FeO 36.35wt%, SiO2 14.97wt%, TiO2 12.19wt%, CaO 2.21wt%, MgO 3.34wt%, MnO 7.18wt%, Cr2O3 3.17wt%; particle size -120 mesh, i.e., the undersize after passing a 120-mesh sieve) was added with 380g of sodium bicarbonate according to a Na2O / V2O5 molar ratio of 2.68, and calcined at 800°C for 90min to obtain 1225.6g of sodium clinker. The sodium clinker was ground to -120 mesh, added to a Büchner funnel containing filter paper, and then filled with tap water at about 90°C to cover the slag surface. The slag was soaked for 8min and then drained. The leaching process was repeated several times to obtain 2420mL of leachate and 934.1g of leached residue. The leachate had a chemical composition of 32.76 g / L V, 24.93 g / L Na, 2.48 g / L Cr, 0.63 g / L Si, and 0.03 g / L P, with a pH of 10.37. The vanadium content of the leached residue was 0.72 wt%, and the vanadium leaching rate was 92.18%. 7 g of hydrazine was added to the leachate, and the mixture was stirred in an 80°C water bath for 60 minutes to remove chromium. Then, 3.6 g of sodium aluminate was added, calculated at an Al / Si molar ratio of 0.81, and the mixture was stirred for 30 minutes to remove silicon and phosphorus. The pH of the solution after silicon and phosphorus removal was 10.46. After the impurity removal reaction is completed, the solid-liquid separation and the impurity removal residue are washed to obtain 2540mL of impurity-removed liquid with the chemical composition of V 30.65g / L, Na23.94g / L, Cr 0.66g / L, Si 0.02g / L and P 0.02g / L. The impurity-removed liquid is evaporated and concentrated to 1450mL to obtain a high-concentration vanadium solution. + Calculated at a molar ratio of vanadium / V = 3.32, 400 g of ammonium bicarbonate was added to the high-concentration vanadium solution, and the vanadium precipitation reaction was carried out by stirring at 12-15°C and pH = 9.54 for 300 minutes. After the reaction, the solid-liquid separation and washing of the solid produced 156.40 g of ammonium metavanadate and 1490 mL of the vanadium precipitation supernatant. The ammonium metavanadate contained 43.28 wt% of V and 0.03 wt% of Na; the vanadium precipitation supernatant contained 6.82 g / L of V, 40.78 g / L of Na, and NH4 + 45.13g / L. CO2 was passed through the upper layer of the vanadium precipitate to adjust the solution pH to 8.18. Sodium bicarbonate was crystallized under stirring and solid-liquid separation was performed to obtain 105.73g of solid sodium bicarbonate and crystallization mother liquor. The crystallization mother liquor was heated to 95-100℃ for deamination to obtain NH4 + Deammoniation liquid with a content of 3.86g / L.

[0060] Sodium bicarbonate is recycled as a sodium-forming additive for the next round of vanadium slag roasting; the deammoniation liquid is recycled as a leaching agent for the next round of sodium-forming clinker leaching; the ammonia-containing gas generated during the deammoniation process of the crystallization mother liquor is absorbed by the high-concentration vanadium liquid prepared in the next round and then participates in the vanadium precipitation reaction.

[0061] Example 2

[0062] A sodium salt recycling method based on sodium roasting of vanadium slag and water leaching to extract vanadium comprises the following steps:

[0063] 1000g of converter vanadium slag (main chemical components: V2O5 15.35wt%, FeO 36.35wt%, SiO2 14.97wt%, TiO2 12.19wt%, CaO 2.21wt%, MgO 3.34wt%, MnO 7.18wt%, Cr2O3 3.17wt%; particle size -120 mesh) was added with 225g of sodium carbonate according to a Na2O / V2O5 molar ratio of 2.52, and calcined at 820°C for 90 minutes to obtain 1221.3g of sodium clinker. The sodium clinker was ground to -120 mesh and added to a Buchner funnel containing filter paper. Then, tap water at about 90°C was added to cover the slag surface, soaked for 8 minutes, and then drained. Leaching was repeated several times in the above manner to obtain 2480mL of leachate and 951.3g of leach residue. The leachate had a chemical composition of 31.30 g / L V, 3.12 g / L Na, 1.96 g / L Cr, 0.56 g / L Si, and 0.03 g / L P, with a pH of 10.02. The leached residue had a V content of 0.88 wt%, and a vanadium leaching rate of 90.27%. 9 g of hydroxylamine was added to the leachate, and the mixture was stirred in a 95°C water bath for 90 minutes to remove chromium. Subsequently, 3.3 g of sodium aluminate was added, calculated at an Al / Si molar ratio of 0.81, and the mixture was stirred for 30 minutes to remove silicon and phosphorus. The pH of the solution after silicon and phosphorus removal was 10.10. After the impurity removal reaction is completed, the solid-liquid separation and the impurity removal residue are washed to obtain 2520mL of impurity-removed liquid with the chemical composition of V 30.43g / L, Na22.84g / L, Cr 0.29g / L, Si 0.02g / L and P 0.02g / L. The impurity-removed liquid is evaporated and concentrated to 1480mL to obtain a high-concentration vanadium solution. + Calculated at a molar ratio of 3.30, 260 g of ammonium bicarbonate and 80 g of ammonium carbonate were added to the high-concentration vanadium solution. The vanadium precipitation reaction was carried out at 12-15°C and pH 9.38 with stirring for 300 min. After the reaction, the solid-liquid separation and washing of the solid produced 157.10 g of ammonium metavanadate and 1530 mL of the vanadium precipitation supernatant. The ammonium metavanadate contained 3.24 wt% of V and 0.02 wt% of Na; the vanadium precipitation supernatant contained 5.72 g / L of V, 37.60 g / L of Na, and NH4 +42.66g / L. CO2 was passed through the upper layer of the vanadium precipitate to adjust the solution pH to 8.12. Sodium bicarbonate was crystallized under stirring and solid-liquid separation was performed to obtain 90.74g of solid sodium bicarbonate and crystallization mother liquor. The crystallization mother liquor was heated to 95-100℃ for deamination to obtain NH4 + Deammoniation liquid with a content of 3.62g / L.

[0064] Sodium bicarbonate is recycled as a sodium-forming additive for the next round of vanadium slag roasting; the deammoniation liquid is recycled as a leaching agent for the next round of sodium-forming clinker leaching; the ammonia-containing gas generated during the deammoniation process of the crystallization mother liquor is absorbed by the high-concentration vanadium liquid prepared in the next round and then participates in the vanadium precipitation reaction.

[0065] Example 3

[0066] A sodium salt recycling method based on sodium roasting of vanadium slag and water leaching to extract vanadium comprises the following steps:

[0067] 1000g of converter vanadium slag (main chemical components: V2O5 15.35wt%, FeO 36.35wt%, SiO2 14.97wt%, TiO2 12.19wt%, CaO 2.21wt%, MgO 3.34wt%, MnO 7.18wt%, Cr2O3 3.17wt%; particle size -120 mesh) was added with 171g of sodium bicarbonate and 132g of sodium carbonate according to a Na2O / V2O5 molar ratio of 2.68, and calcined at 800°C for 90 minutes to obtain 1225.1g of sodium clinker. The sodium clinker was ground to -120 mesh and added to a Buchner funnel containing filter paper. Then, tap water at about 90°C was added to cover the slag surface, soaked for 8 minutes, and then drained. Leaching was repeated several times in the above manner to obtain 2570mL of leachate and 933.7g of leach residue. The leachate contains 30.73 g / L V, 3.18 g / L Na, 2.34 g / L Cr, 0.58 g / L Si, and 0.03 g / L P, with a pH of 10.39. The leaching residue contains 0.75 wt% V and a vanadium extraction rate of 91.86%. 2.8 g of methanol was added to the leachate, and the mixture was stirred in an autoclave at 160°C for 180 minutes to remove chromium through reduction. After the reduction reaction, the pressure was released and the temperature was lowered to 80°C. Then, 3.6 g of sodium aluminate was added, calculated at an Al / Si molar ratio of 0.82, and the mixture was stirred for 30 minutes to remove silicon and phosphorus. The pH of the solution after silicon and phosphorus removal was 10.43. After the impurity removal reaction is completed, the solid-liquid separation and the impurity removal residue are washed to obtain 2440mL of impurity-removed liquid with the chemical composition of V 31.98g / L, Na 24.57g / L, Cr0.25g / L, Si0.02g / L and P 0.02g / L. The impurity-removed liquid is evaporated and concentrated to 1430mL to obtain a high-concentration vanadium solution. +Calculated by a molar ratio of vanadium / V = 3.10, 375g of ammonium bicarbonate was added to the high-concentration vanadium solution, and the reaction was stirred at 12-15°C and pH = 9.58 for 300 minutes. After the reaction, the solid-liquid separation and solid washing were carried out to obtain 154.66g of ammonium metavanadate and 1470mL of vanadium precipitation supernatant. Among them, the ammonium metavanadate contained 43.25wt% of V and 0.04wt% of Na; the vanadium precipitation supernatant contained 7.58g / L of V, 0.74g / L of Na4, and NH4 + 42.06g / L. CO2 was passed through the upper layer of the vanadium precipitate to adjust the solution pH to 8.26. Sodium bicarbonate was crystallized under stirring and solid-liquid separation was performed to obtain 103.91g of solid sodium bicarbonate and crystallization mother liquor. The crystallization mother liquor was heated to 95-100℃ for deamination to obtain NH4 + Deammoniation liquid with a content of 3.54g / L.

[0068] Sodium bicarbonate is recycled as a sodium-forming additive for the next round of vanadium slag roasting; the deammoniation liquid is recycled as a leaching agent for the next round of sodium-forming clinker leaching; the ammonia-containing gas generated during the deammoniation process of the crystallization mother liquor is absorbed by the high-concentration vanadium liquid prepared in the next round and then participates in the vanadium precipitation reaction.

[0069] In summary, the present invention provides a sodium salt recycling method based on sodium roasting of vanadium slag and water leaching for vanadium extraction. The sodium-treated leaching solution is subjected to alkaline conditions to remove chromium, silicon and phosphorus, and then ammonium metavanadate is precipitated to achieve separation and extraction of vanadium; the sodium in the upper layer of the vanadium precipitation liquid is crystallized and separated in the form of sodium bicarbonate; the crystallization mother liquor is deaminated and directly recycled to leach the clinker, thereby achieving the recycling of media such as sodium and process water, reducing the cost of the vanadium oxide preparation process, avoiding the generation of sodium sulfate as a solid waste from water treatment, and improving the competitiveness of the vanadium oxide preparation industry from vanadium slag.

[0070] The above is only a preferred embodiment of the present invention. It should be pointed out that for ordinary technicians in this technical field, several improvements and modifications can be made without departing from the principles of the present invention. These improvements and modifications should also be regarded as within the scope of protection of the present invention.

Claims

1. A sodium salt recycling method based on sodium roasting of vanadium slag and water leaching of vanadium, characterized in that: The following steps are involved: a) mixing vanadium slag and sodium salt and then performing sodium roasting to obtain sodium clinker; wherein the sodium salt includes sodium carbonate and / or sodium bicarbonate; b) subjecting the sodium-treated clinker to water leaching and solid-liquid separation to obtain a leaching residue and a leachate; c) removing chromium, silicon and phosphorus from the leachate and then evaporating and concentrating it to 50-60% of its original volume to obtain a high-concentration vanadium solution; wherein the chromium removal agent is one or more of hydrazine, hydroxylamine, formaldehyde, methanol, ethanol, sucrose and starch, and the silicon and phosphorus removal agent is sodium aluminate; d) adding ammonium compound to the high-concentration vanadium solution to precipitate vanadium, and separating the solid and liquid to obtain ammonium metavanadate solid and vanadium precipitation upper layer liquid; wherein the ammonium compound is one or more of ammonium carbonate, ammonium bicarbonate and ammonia water; e) passing CO2 through the upper layer of the vanadium precipitate to adjust the pH and crystallize sodium bicarbonate, performing solid-liquid separation to obtain sodium bicarbonate solid and crystallization mother liquor; the sodium bicarbonate solid is returned to step a) for mixing with the vanadium slag and then sodiumization and roasting; f) heating the crystallization mother liquor to decompose ammonium bicarbonate for deamination, thereby obtaining a deammonified liquid and an ammonia-containing gas; the deammonified liquid is returned to step b) to participate in water leaching; the ammonia-containing gas is returned to step d) to participate in vanadium precipitation after being absorbed by the high-concentration vanadium solution.

2. The sodium salt recycling method according to claim 1, characterized in that: In step a), the molar ratio of the sodium salt calculated as Na2O to the vanadium slag calculated as V2O5 is (2.5-3.5):

1.

3. The sodium salt recycling method according to claim 1, wherein In step a), the temperature of the sodium calcination is 750-850° C., and the time is 60-120 minutes.

4. The sodium salt recycling method according to claim 1, characterized in that: In step a), the vanadium slag is obtained by blowing vanadium-containing molten iron in a converter.

5. The sodium salt recycling method according to claim 1, wherein In step c), the leaching solution simultaneously removes chromium, silicon and phosphorus, or the leaching solution first removes chromium and then removes silicon and phosphorus.

6. The sodium salt recycling method according to claim 5, characterized in that: In step c), the reaction temperature for removing chromium is 70-200° C., and the reaction time is 30-210 min; the reaction temperature for removing silicon and phosphorus is 60-100° C., and the reaction time is 20-60 min. The reaction pH value is 8.5-10.

5.

7. The sodium salt recycling method according to claim 1, characterized in that: In step c), the molar ratio of the sodium aluminate calculated as Al to the leachate calculated as Si is (0.8-1.2):

1.

8. The sodium salt recycling method according to claim 1, characterized in that: In step c), the Cr concentration of the leaching solution after silicon, phosphorus and chromium are removed is ≤1 g / L.

9. The sodium salt recycling method according to claim 1, characterized in that: In step d), NH4 + The molar ratio of the ammonium compound to the high-concentration vanadium solution is (2.8-3.5):1; the reaction temperature of the vanadium precipitation is 10-20° C., the reaction pH is 9.3-9.8, and the reaction time is 3-5 hours.

10. The sodium salt recycling method according to claim 1, characterized in that: In step e), the pH value of the upper layer of the vanadium precipitation liquid is adjusted to 7.5-8.5 by passing CO2.

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