Method for preparing vanadyl sulfate by utilizing vanadium slag calcification roasting tail gas in short process

Abstract

The invention provides a short-process method for preparing vanadyl sulfate from vanadium slag calcification roasting tail gas. The method mainly comprises the following steps: mixing vanadium slag and a calcification agent, and roasting to obtain vanadium slag clinker; the clinker enters an acid leaching process to extract vanadium, and roasting tail gas is in countercurrent contact with pre-reduction liquid for pre-reduction; the clinker and dilute sulphuric acid are subjected to countercurrent leaching, and pentavalent vanadium-rich acid leaching liquid is obtained; adding a diluent into the pentavalent vanadium-rich acid leaching solution to obtain a pre-reduction solution; roasting tail gas is introduced into the pre-reduction liquid for reduction, and pre-reduction liquid is obtained; excessive oxalic acid is added into the pre-reduction liquid for impurity removal, and meanwhile pentavalent vanadium ions which are not completely reduced are reduced to form tetravalent vanadium liquid; and separating to obtain a tetravalent vanadium extraction stock solution, and finally preparing a vanadyl sulfate solution through an extraction-washing-reverse extraction process. The method realizes efficient reduction and utilization of the roasting tail gas, has the advantages of simple process, low cost, high vanadium recovery rate, high resource utilization rate and environmental friendliness, and can greatly reduce the treatment cost of the roasting tail gas.

Description

Technical Field

[0001] This invention relates to the field of short-process preparation of high-purity vanadium oxysulfate technology, specifically a method for preparing vanadium oxysulfate using the tail gas from vanadium slag calcification roasting. Background Technology

[0002] Vanadium slag, a primary raw material for vanadium electrolytes, is mainly derived from a byproduct of steelmaking using vanadium-titanium magnetite. In vanadium slag, vanadium exists primarily as low-valence oxides (V₂O₃, VO₂), requiring roasting and oxidation to convert it into soluble pentavalent vanadium oxides (V₂O₅) before it can be used in electrolyte preparation. Current industrial vanadium extraction processes mainly include three routes: sodium roasting, blank roasting, and calcification roasting. Among these, the calcification roasting-acid leaching process demonstrates significant economic and environmental advantages in industrial applications due to its low-cost calcification agent, high vanadium leaching rate (up to 85% or more), and simple wastewater treatment.

[0003] However, the existing calcination roasting process still faces the following technical bottlenecks: (1) harmful gases such as SO2 generated during the roasting process require tail gas treatment facilities; (2) the acid leaching stage requires a large amount of sulfuric acid (usually 1.5 to 2 times the theoretical amount), which leads to a sudden increase in the temperature of the leaching system (up to 70 to 80°C), thereby inducing vanadium hydrolysis and precipitation (generating insoluble products such as VO(OH)2); (3) the reduction treatment requires the addition of additional reducing agents, which not only introduces impurity ions but also increases production costs. Therefore, developing a green vanadium extraction technology that can reduce the amount of acid and reducing agents used, realize the resource utilization of tail gas, and improve the vanadium leaching rate has important industrial value.

[0004] CN109182760A proposes a method for recovering vanadium from vanadium extraction tailings through calcification. This method involves secondary roasting and leaching of the vanadium-containing tailings to further extract the vanadium content, achieving full recovery and utilization of vanadium in the tailings. This method is simple and easy to implement, and has broad application value. However, this technology suffers from drawbacks such as high roasting energy consumption and a long process flow, and its effect on improving the vanadium leaching rate in the main process is limited.

[0005] CN107475518A proposes a multi-stage pH-controlled vanadium leaching method. By precisely controlling the pH and temperature of the seven leaching stages, a leaching rate of over 96% can be achieved. However, this process suffers from problems such as complex procedures, high acid consumption, and high energy consumption, making it difficult to meet the economic requirements of industrial production.

[0006] Existing calcination-acid leaching processes still have significant shortcomings in terms of tail gas treatment, acid consumption control, and leaching efficiency. To address these shortcomings, this invention innovatively proposes a short-process vanadium electrolyte preparation method. By selectively reusing calcination-calcination tail gas, it achieves a synergistic improvement in the efficient reduction and leaching rate of pentavalent vanadium, while simultaneously solving the problems of environmental pollution and resource waste. Summary of the Invention

[0007] The purpose of this invention is to solve the problems in the prior art and provide a method for efficiently utilizing roasting tail gas to reduce pentavalent vanadium while simultaneously improving vanadium leaching rate. This method is simple in process, low in cost, has a high vanadium recovery rate, high resource utilization rate, and is environmentally friendly, thus achieving efficient utilization of roasting tail gas.

[0008] To achieve the above-mentioned objectives, the specific technical solution of the present invention is as follows: A method for preparing vanadium oxysulfate using tail gas from vanadium slag calcination roasting in a short process, the method comprising the following steps: Step 1: The finely ground vanadium slag is evenly mixed with the calcifying agent and fed into a rotary kiln for vanadium slag calcification roasting treatment to obtain vanadium slag (roasting) clinker; the vanadium slag clinker enters the acid leaching process to extract vanadium, and the roasting (reduction) tail gas containing SO2 generated during the rotary kiln roasting process is cooled and then comes into countercurrent contact with the pre-reduction liquid (acid leaching liquid) of Step 4 for pre-reduction. Step 2: The vanadium slag clinker obtained in Step 1 is leached with dilute sulfuric acid solution using a multi-stage countercurrent process to obtain pentavalent vanadium-rich acid leaching solution and vanadium extraction tailings. The vanadium extraction tailings are washed with water and then sent to a tailings treatment company. Step 3: Add a diluent to the pentavalent vanadate leaching solution from Step 2 to adjust the concentration and obtain the solution before reduction; Step 4: Pass the roasting tail gas from Step 1 into the pre-reduction liquid from Step 3, and use the SO2 gas in the tail gas for reduction to obtain the pre-reduced liquid. Step 5: Add excess oxalic acid to the pre-reduction solution in Step 4 to allow some metal impurity ions such as iron and aluminum to combine with oxalate ions to form metal salt precipitates, thereby removing metal impurities; the pentavalent vanadium ions that were not completely reduced in Step 4 are reduced by oxalic acid to form tetravalent vanadium solution; after the reaction is completed, the tetravalent vanadium extraction stock solution and the impurity-removed filter cake are obtained by solid-liquid separation; the impurity-removed filter cake is handed over to a professional institution for recycling. Step 6: Adjust the pH value of the tetravalent vanadium extraction solution from Step 5, mix the tetravalent vanadium extraction solution with the extraction organic phase (tetravalent vanadium extractant) to obtain the loaded organic phase and raffinate. The raffinate is then recycled after being treated in the wastewater treatment process. Step 7: Wash the loaded organic phase obtained in step 6 with a low concentration of acid (detergent), and return the washing residue directly to steps 2 and 3 for recycling; Step 8: Use high-concentration acid (back-extraction agent) to back-extract the acid-washed loaded organic phase in Step 7 to obtain tetravalent vanadium solution and blank organic phase; return the blank organic phase to Step 6 for cyclic extraction; Further, the calcifying agent mentioned in step 1 is one or more of limestone or calcium sulfate.

[0009] Further, the roasting conditions described in step 1 are as follows: roasting temperature is 800-850℃, roasting time is 1.5-3h, and the molar ratio of calcium to vanadium is 2.2-2.5; the oxygen concentration during roasting is 8-15% by volume; and the particle size of the vanadium slag material is 80-200 mesh.

[0010] Further, the concentration of the dilute sulfuric acid solution in step 2 is less than 38%; and / or, the leaching conditions are: a leaching solid-liquid ratio of 1:(2~4); in order to increase mass transfer efficiency, the leaching method is to use a stirred tank for multi-stage countercurrent contact, with 3~5 leaching stages.

[0011] Further, the diluent mentioned in step 3 is one or more of pure water, process water, and recycled water from wastewater treatment processes. After being added, the vanadium concentration of the acid leaching solution is diluted to 10.0~20.0 g / L. After dilution and mixing, the vanadium content of the acid leaching solution is measured to obtain the pre-reduction solution.

[0012] Furthermore, in step 4, the acid leaching solution is pre-reduced by countercurrent contact with the calcination roasting tail gas from step 1. The main component of the roasting tail gas is sulfur dioxide gas (which plays a reducing role). The sulfur dioxide gas is generated in the roasting section of step 1 and is introduced into the pre-reduction solution after being filtered by a cyclone dust collector. And / or, the gas flow rate of the roasting tail gas is controlled at 100 mL / min·L to 500 mL / min·L. This step uses the roasting tail gas and uses the SO2 reducing gas in the tail gas for the reduction of pentavalent vanadium. After the acid leaching solution absorbs SO2 from the tail gas, there may be incomplete reduction. Subsequently, excess oxalic acid is added to completely reduce the remaining pentavalent vanadium. This measure can significantly reduce the cost of tail gas treatment and the cost of reducing agent.

[0013] Furthermore, the impurity remover mentioned in step 5 (which serves to remove impurities and perform deep reduction) is oxalic acid. Excess oxalic acid can not only effectively remove most of the impurity ions, but also ensure that the reduction reaction is thorough, deeply reducing the small amount of incompletely reduced pentavalent vanadium ions in the pre-reduction solution to tetravalent vanadium ions, while precipitating and removing some aluminum, magnesium, manganese and other metal impurities. The amount of oxalic acid added is determined according to the content of residual pentavalent vanadium ions and aluminum, magnesium and manganese impurities in the pre-reduction solution, with an excess coefficient of 20% to 50%. In addition, impurity removers such as EDTA, phosphoric acid and citric acid also have similar effects, but excess oxalic acid also acts as a reducing agent.

[0014] And / or, in step 5, the pH of the reducing solution is adjusted to 2-3 using an acid or base before the impurity removal treatment; and / or, the acid is preferably sulfuric acid; and / or, the base is preferably sodium hydroxide.

[0015] Further, in step 6, the pH of the tetravalent vanadium solution is adjusted to 1.8-2.6, preferably 2.0-2.4.

[0016] Further, the organic phase solution extracted in step 6 comprises an anionic extractant, a phase dispersant, and an organic diluent. The anionic extractant includes, but is not limited to, di(2-ethylhexyl) phosphate (P204) and trioctylphosphine oxide (P507), the phase dispersant includes, but is not limited to, TBP and sec-octanol, and the organic diluent is sulfonated kerosene. And / or, the anionic extractant is preferably P204, to achieve efficient separation of vanadium and chromium and reduce vanadium loss. The volume ratio of the extractant to the tetravalent vanadium solution is O / A = 0.2–1:1.

[0017] Further, the extraction conditions described in step 6 are: extraction temperature 25-35℃, extraction time 15-30 min, extraction method is multi-stage countercurrent extraction, and the number of extraction stages is 4-6. Furthermore, the low-concentration acid solution mentioned in step 7 refers to sulfuric acid with a concentration of 0.05 mol / L to 0.5 mol / L, which can effectively wash away impurity ions in the supported organic phase; and / or, the resulting washing residue can be directly returned to step 2 or 3 for recycling, thereby realizing wastewater recycling and saving resources. Further, in step 7, the loaded organic phase is mixed with a low-concentration acid solution at a volume ratio of O / A = 1 to 15:1; and / or, the washing conditions are: washing temperature of 20 to 30°C, washing time of 10 to 20 minutes, washing method of multi-stage countercurrent extraction, and extraction stages of 2 to 4 stages. Further, the stripping agent in step 8 is a high-concentration acid; and / or, the high-concentration acid refers to sulfuric acid with a concentration of 3~4 mol / L, which can efficiently strip vanadium from the supported organic phase to the aqueous phase; Further, in step 8, the back-extraction agent is mixed with the supported organic phase at a volume ratio of O / A = 1 to 15:1; and / or, the back-extraction conditions are: back-extraction temperature of 20 to 30°C, back-extraction time of 1 to 2 hours, extraction method of multi-stage countercurrent extraction, and extraction stages of 4 to 6 stages.

[0018] Furthermore, the blank organic phase described in step 8 can be directly returned to step 6 for recycling, which can effectively avoid the loss of organic phase.

[0019] Compared with the prior art, the positive effects of the present invention are reflected in: 1. This invention significantly improves the leaching efficiency of vanadium by optimizing the calcination roasting and acid leaching processes; it innovatively integrates the reduction characteristics of roasting tail gas with a short-process extraction process to successfully produce high-purity vanadium oxysulfate, achieving efficient resource utilization and reduced production costs. 2. This invention employs a dynamic countercurrent leaching system, which effectively improves the vanadium leaching rate while making full use of the leaching acid solution, reducing resource consumption and lowering production costs; 3. This invention employs a roasting tail gas (SO2) + oxalic acid synergistic reduction system, which can effectively achieve efficient reduction of pentavalent vanadium, reduce the cost of tail gas alkaline neutralization and absorption, and reduce additional reducing agent consumption and production costs while being green and environmentally friendly. 4. This invention achieves high vanadium extraction efficiency through pre-purification treatment, further improving the quality and purity of vanadium electrolyte; 5. By recycling the wastewater generated in the extraction section, water resources are reused, environmental pollution is reduced, and green environmental protection requirements are met. Attached Figure Description

[0020] Figure 1 This is a schematic diagram of the process for preparing vanadium oxysulfate using the tail gas from the calcination roasting of vanadium slag, as provided by the present invention. Detailed Implementation

[0021] All features disclosed in this specification, or steps in all methods or processes disclosed herein, may be combined in any way, except for mutually exclusive features and / or steps.

[0022] Any feature disclosed in this specification (including the claims and abstract) may be replaced by other equivalent or similar features, unless specifically stated otherwise. That is, unless specifically stated otherwise, each feature is merely one example of a series of equivalent or similar features.

[0023] The features and performance of the present invention will be further described in detail below with reference to embodiments.

[0024] Unless otherwise specified, the process methods used in the embodiments are conventional methods; unless otherwise specified, the materials used are commercially available.

[0025] In this application, unless otherwise specified, % refers to mass percentage content, and all ratios refer to mass ratios; the liquid-to-solid ratio is the ratio of liquid volume (mL) to solid mass (g).

[0026] The main components and percentage content (%) of vanadium slag in the following examples are as follows:

[0027] Example 1 This example provides a short-process method for preparing vanadium oxysulfate using tail gas from vanadium slag calcification roasting. Vanadium slag is used as raw material for calcification roasting and continuous countercurrent acid leaching to improve the quality of vanadium electrolyte. This example describes a short-process method for preparing vanadium oxysulfate using the tail gas from the calcination roasting of vanadium slag, which mainly includes the following steps: Step 1: Ball mill 4 kg of vanadium slag, then pass the treated vanadium slag through an 80-mesh standard sieve. Mix the material passing through the 80-mesh sieve with limestone powder passing through the 80-mesh standard sieve, with a mass ratio of vanadium slag to limestone powder of 1:1.2. After mixing evenly, add the mixture to a rotary kiln. Under an atmosphere with an O2 content of 10% (gas volume percentage), a coke oven gas content of 16%, and an air content of 74%, control the roasting temperature at 820℃ and the roasting time at 2 hours to obtain roasted clinker (vanadium slag clinker). At the same time, the generated gas is passed into the pre-reduction liquid. Step 2: After the temperature of the roasted clinker obtained in Step 1 is reduced to room temperature, it is added to a stirring tank filled with acid solution for three-stage dynamic countercurrent leaching. The inlet concentration of sulfuric acid used is 38 wt%, and the liquid-solid ratio is controlled at 3:1. After leaching, pentavalent vanadium-rich acid leaching solution and vanadium tailings are obtained. Step 3: Add an equal volume of pure water to the 10L pentavalent vanadate leaching solution obtained in Step 2, stir well, and adjust the pH of the solution to 2.0 using sulfuric acid or liquid alkali to obtain the solution before reduction. Step 4: Pass the gas (sulfur dioxide) generated in Step 1 into 20L of the pre-reduction liquid; control the gas flow rate to 500mL / min·L, which is the sum of the unit total volume of the aqueous phase and the oil phase, and stir at room temperature for 0.5 hours to obtain the pre-reduction liquid. Step 5: Add 810g of oxalic acid (9mol) to the untreated solution obtained in Step 4, stir for 1 hour, and then filter to obtain the original extract. Step 6: Adjust the pH of the extraction solution to 2.0, and at room temperature and with a phase O / A ratio of 1:1.5, perform a 6-stage countercurrent extraction of the tetravalent vanadium solution using P204 extractant to obtain the supported organic phase; Step 7: At room temperature and with a ratio of O / A = 10:1, the loaded tetravalent vanadium organic phase is washed in a 4-stage countercurrent manner using a 0.2 mol / L sulfuric acid solution. The washing residue is directly returned to step 2 or 3 for recycling. Step 8: At room temperature and with a ratio of O / A = 6:1, perform a 6-stage countercurrent back-extraction with 3 mol / L sulfuric acid to obtain 2.2 L of tetravalent vanadium solution; The analysis revealed the following concentrations in the back-extraction solution: vanadium 91.4 g / L, sodium 11 mg / L, iron 22 mg / L, calcium 15 mg / L, aluminum 34 mg / L, chromium 0 mg / L, silicon 0.3 mg / L, magnesium 0.9 mg / L, and manganese 0 mg / L. Example 2 This example describes a short-process method for preparing vanadium oxysulfate using the tail gas from the calcination roasting of vanadium slag, which mainly includes the following steps: Step 1: Ball mill 4 kg of vanadium slag, then pass the treated vanadium slag through a 200-mesh standard sieve. Mix the 200-mesh sieve underflow with limestone powder (also 200-mesh standard sieve underflow) at a mass ratio of 1:1.2. Add the mixture to a rotary kiln. Under an atmosphere of 10% O2, 16% coke oven gas, and 74% air, control the roasting temperature at 800℃ and the roasting time at 2.5 hours to obtain roasted clinker. Simultaneously, pass the generated gas into the pre-reduction liquid. Step 2: After the clinker temperature drops to room temperature, it is added to a stirring tank filled with acid solution for three-stage dynamic countercurrent leaching. The inlet concentration of sulfuric acid used is 24%, and the liquid-solid ratio is controlled at 4.75:1. After leaching, pentavalent vanadium-rich acid leaching solution and vanadium tailings are obtained. Step 3: Add 9L of pure water to 31.6L of pentavalent vanadate leaching solution, stir well, and adjust the pH of the solution to 2.0 using sulfuric acid or liquid alkali to obtain the solution before reduction; Step 4: Pass the gas (sulfur dioxide) generated in step 1 of roasting into the pre-reduction liquid; control the gas flow rate to 300 mL / min·L, stir at room temperature for 40 min to obtain the pre-purification liquid; Step 5: Add 756g of oxalic acid to the liquid before impurity removal, stir for 1 hour, and then filter to obtain the original extract. Step 6: Adjust the pH of the pre-extraction solution to 2.1, and at room temperature and with a phase O / A ratio of 1:1.4, perform a 6-stage countercurrent extraction of the tetravalent vanadium solution using P204 extractant to obtain the supported organic phase; Step 7: Wash the loaded tetravalent vanadium organic phase with 0.2 mol / L sulfuric acid solution in a 4-stage countercurrent manner at room temperature and a ratio of O / A = 10:1. Return the washing liquid to step 2 or 3 for recycling. Step 8: At room temperature and with a ratio of O / A = 8:1, 1.6 L of tetravalent vanadium solution was obtained by 6-stage countercurrent back-extraction using 3 mol / L sulfuric acid.

[0028] The analysis revealed that the vanadium concentration in the back-extraction solution was 124.6 g / L, and / or sodium concentration was 3.6 mg / L, and / or iron concentration was 20 mg / L, and / or calcium concentration was 8.7 mg / L, and / or aluminum concentration was 12 mg / L, and / or chromium concentration was 0 mg / L, silicon concentration was 0.2 mg / L, and / or magnesium concentration was 0.7 mg / L, and / or manganese concentration was 0 mg / L. Comparative Example 1 This comparative example provides a short-process method for preparing vanadium oxysulfate using the tail gas from the calcination roasting of vanadium slag, mainly including the following steps: Step 1: Ball mill 4 kg of vanadium slag, then pass the treated vanadium slag through an 80-mesh standard sieve. Mix the material passing through the 80-mesh sieve with limestone powder passing through the 80-mesh standard sieve, with a mass ratio of vanadium slag to limestone powder of 1:1.2. After mixing evenly, add the mixture to a rotary kiln. Under an atmosphere with an O2 content of 10% (gas volume percentage), a coke oven gas content of 16%, and an air content of 74%, control the roasting temperature at 820℃ and the roasting time at 2 hours to obtain roasted clinker (vanadium slag clinker). At the same time, the generated gas is passed into the pre-reduction liquid. Step 2: After the temperature of the roasted clinker obtained in Step 1 is reduced to room temperature, it is added to a stirring tank filled with acid solution for three-stage dynamic countercurrent leaching. The inlet concentration of sulfuric acid used is 38 wt%, and the liquid-solid ratio is controlled at 3:1. After leaching, pentavalent vanadium-rich acid leaching solution and vanadium tailings are obtained. Step 3: Add an equal volume of pure water to the 10L pentavalent vanadate leaching solution obtained in Step 2, stir well, and adjust the pH of the solution to 2.0 using sulfuric acid or liquid alkali to obtain the solution before reduction. Step 4: Pass the gas (sulfur dioxide) generated in Step 1 into 20L of the pre-reduction liquid; control the gas flow rate to 500mL / min·L, which is the sum of the unit total volume of the aqueous phase and the oil phase, and stir at room temperature for 0.5 hours to obtain the pre-reduction liquid. Step 5: Add 810g of oxalic acid (9mol) to the untreated solution obtained in Step 4, stir for 1 hour, and then filter to obtain the original extract. Step 6: Adjust the pH of the extraction solution to 1.4, and at room temperature and with a phase O / A ratio of 1:1.5, perform a 6-stage countercurrent extraction of the tetravalent vanadium solution using P204 extractant to obtain the supported organic phase; Step 7: At room temperature and with a ratio of O / A = 10:1, the loaded tetravalent vanadium organic phase is washed in a 4-stage countercurrent manner using a 0.2 mol / L sulfuric acid solution. The washing residue is directly returned to step 2 or 3 for recycling. Step 8: At room temperature and with a ratio of O / A = 6:1, perform a 6-stage countercurrent back-extraction with 3 mol / L sulfuric acid to obtain 2.2 L of tetravalent vanadium solution; The analysis revealed the following concentrations in the back-extraction solution: vanadium 76 g / L, sodium 89 mg / L, iron 65 mg / L, calcium 23 mg / L, aluminum 67 mg / L, chromium 7.8 mg / L, silicon 0.8 mg / L, magnesium 1.3 mg / L, and manganese 0 mg / L.

[0029] The experimental conditions for this comparative example were the same as those for Example 1, except that in step 6, the pH of the pre-extraction solution was adjusted to 1.4, while the other steps were the same.

[0030] The final analysis of the back-extraction solution showed that the vanadium concentration was 76 g / L, and the vanadium content in the raffinate produced in step 6 was greater than 2 g / L, indicating a significant reduction in the vanadium extraction efficiency.

[0031] Comparative Example 2 This example provides a short-process method for preparing vanadium oxysulfate using the tail gas from the calcination roasting of vanadium slag, mainly including the following steps: Step 1: Ball mill 4 kg of vanadium slag, then pass the treated vanadium slag through an 80-mesh standard sieve. Mix the material passing through the 80-mesh sieve with limestone powder passing through the 80-mesh standard sieve, with a mass ratio of vanadium slag to limestone powder of 1:1.2. After mixing evenly, add the mixture to a rotary kiln. Under an atmosphere with an O2 content of 10% (gas volume percentage), a coke oven gas content of 16%, and an air content of 74%, control the roasting temperature at 820℃ and the roasting time at 2 hours to obtain roasted clinker (vanadium slag clinker). At the same time, the generated gas is passed into the pre-reduction liquid. Step 2: After the temperature of the roasted clinker obtained in Step 1 is reduced to room temperature, it is added to a stirring tank filled with acid solution for three-stage dynamic countercurrent leaching. The inlet concentration of sulfuric acid used is 38 wt%, and the liquid-solid ratio is controlled at 3:1. After leaching, pentavalent vanadium-rich acid leaching solution and vanadium tailings are obtained. Step 3: Add an equal volume of pure water to the 10L pentavalent vanadate leaching solution obtained in Step 2, stir well, and adjust the pH of the solution to 2.0 using sulfuric acid or liquid alkali to obtain the solution before reduction. Step 4: Pass the gas (sulfur dioxide) generated in Step 1 into 20L of the pre-reduction liquid; control the gas flow rate to 500mL / min·L, which is the sum of the unit total volume of the aqueous phase and the oil phase, and stir at room temperature for 0.5 hours to obtain the pre-reduction liquid. Step 5: Add 486g of oxalic acid to the pre-extraction solution obtained in Step 4, stir for 1 hour, and then filter to obtain the original extract. Step 6: Adjust the pH of the extraction solution to 2.0, and at room temperature and with a phase O / A ratio of 1:1.5, perform a 6-stage countercurrent extraction of the tetravalent vanadium solution using P204 extractant to obtain the supported organic phase; Step 7: At room temperature and with a ratio of O / A = 10:1, the loaded tetravalent vanadium organic phase is washed in a 4-stage countercurrent manner using a 0.2 mol / L sulfuric acid solution. The washing residue is directly returned to step 2 or 3 for recycling. Step 8: At room temperature and with a ratio of O / A = 6:1, perform a 6-stage countercurrent back-extraction with 3 mol / L sulfuric acid to obtain 2.2 L of tetravalent vanadium solution; The analysis revealed the following concentrations in the back-extraction solution: vanadium 82 g / L, sodium 64 mg / L, iron 52 mg / L, calcium 24 mg / L, aluminum 57 mg / L, chromium 3.4 mg / L, silicon 0 mg / L, magnesium 1.5 mg / L, and manganese 0 mg / L.

[0032] The experimental conditions in this case were the same as in Example 1, except that in step 5, the amount of oxalic acid added was 486g.

[0033] The final back-extraction solution test results showed that the vanadium concentration was 82 g / L. Pentavalent vanadium was detected in the raffinate produced in step 6, with a content greater than 500 mg / L, proving that insufficient oxalic acid led to incomplete reduction of pentavalent vanadium.

[0034] The embodiments described above merely illustrate specific implementation methods of this application, and while the descriptions are detailed and specific, they should not be construed as limiting the scope of protection of this application. It should be noted that those skilled in the art can make various modifications and improvements without departing from the concept of the technical solution of this application, and these modifications and improvements all fall within the scope of protection of this application.

[0035] This background section is provided to generally present the context of the invention. The work of the currently named inventors, the work to the extent described in this background section, and aspects of this section that did not constitute prior art at the time of application are neither expressly nor impliedly acknowledged as prior art to the invention.

Claims

1. A method for preparing vanadyl sulfate by short process of vanadium slag calcification roasting tail gas, characterized in that, The method comprises the following steps: Step 1: uniformly mix the ground vanadium slag and a calcifying agent, and then put them into a rotary kiln to perform a calcification roasting treatment on the vanadium slag to obtain vanadium slag clinker; put the vanadium slag clinker into an acid leaching process to extract vanadium, and perform countercurrent contact between the SO2-containing roasting tail gas generated in the roasting process in the rotary kiln and the pre-reduction liquid in step 4 after the roasting tail gas is cooled to perform pre-reduction; Step 2: use a multi-stage countercurrent process to leach the vanadium slag clinker in step 1 with a dilute sulfuric acid solution to obtain a pentavalent vanadium-rich acid leaching liquid and a vanadium extraction tailing; after the vanadium extraction tailing is washed with water, it is handed over to a tailing treatment enterprise for treatment; Step 3: add a dilution liquid to the pentavalent vanadium-rich acid leaching liquid in step 2, and after the concentration is adjusted, a pre-reduction liquid is obtained; Step 4: introduce the roasting tail gas in step 1 into the pre-reduction liquid in step 3 to absorb SO2 gas in the tail gas to perform a reduction reaction, and a pre-reduction liquid is obtained; Step 5: add an excess amount of oxalic acid to the pre-reduction liquid in step 4, part of metal impurity ions are combined with oxalate to generate metal salt precipitates to remove the metal impurities, and at the same time, the pentavalent vanadium ions that are not completely reduced in step 4 are reduced by the oxalic acid to form a tetravalent vanadium liquid; After the reaction is completed, a tetravalent vanadium extraction original liquid and a removal impurity filter cake are obtained through solid-liquid separation, and the filter cake is recycled by a professional institution; Step 6: adjust the pH value of the tetravalent vanadium extraction original liquid in step 5, mix the tetravalent vanadium extraction original liquid with an extraction organic phase to perform extraction, and a loaded organic phase and a raffinate liquid are obtained, and the raffinate liquid is treated in a wastewater treatment process and then recycled; Step 7: use a low-concentration acid to wash the loaded organic phase obtained in step 6, and the obtained acid washing liquid is directly returned to steps 2 and 3 for recycling; Step 8: use a high-concentration acid to back-extract the loaded organic phase washed in step 7 to obtain a tetravalent vanadium liquid and a blank organic phase, the blank organic phase is returned to step 6 for cyclic extraction, and the tetravalent vanadium liquid is a vanadyl sulfate solution.

2. A process for the preparation of vanadyl sulphate using vanadium slag calcification roasting tail gas short process as claimed in claim 1 wherein, In step 1, the calcifying agent is one or more of limestone or calcium sulfate; and / or, the temperature during the roasting treatment is 800-850 DEG C, the roasting time is 1.5-3 h, the molar ratio of calcium to vanadium is 2.2-2.5, and the oxygen concentration during the roasting treatment is 8-15% in terms of volume percentage; and the particle size of the vanadium slag is 80-200 mesh. ​ 3. A process for the preparation of vanadyl sulphate using vanadium slag calcification roasting tail gas short process as claimed in claim 1 wherein, In step 2, the concentration of the dilute sulfuric acid solution is less than 38 wt%; and / or, the solid-liquid ratio during the leaching is 1:2-4; and / or, the leaching temperature is 20-35 DEG C. ​ 4. A process for the preparation of vanadyl sulphate using vanadium slag calcification roasting tail gas short process as claimed in claim 1 wherein, In step 3, the dilution liquid is one or more of pure water, production process water and recycled water in the wastewater treatment process; after the dilution liquid is added, the vanadium concentration in the acid leaching liquid is diluted to 10.0-20.0 g / L, the vanadium content in the acid leaching liquid is determined after the dilution liquid is uniformly mixed, and then a pre-reduction liquid is obtained. ​ 5. A process for the preparation of vanadyl sulphate using short flow process of vanadium slag calcination flue gas as claimed in claim 1 wherein, In step 4, the pre-reduction liquid in step 3 is countercurrently contacted with the roasting tail gas in step 1 to perform pre-reduction on the acid leaching liquid; and / or, the flow rate of the roasting tail gas introduced is 100 mL / min·L-500 mL / min·L. ​ 6. A process for the preparation of vanadyl sulphate using short flow process of vanadium slag calcination flue gas as claimed in claim 1 wherein, In step 5, the amount of oxalic acid added is determined according to the residual pentavalent vanadium ions and the contents of aluminum, magnesium and manganese impurities in the pre-reduction liquid; and the excess coefficient of the oxalic acid is 20%-50%. ​ 7. A process for the preparation of vanadyl sulphate using short flow process of vanadium slag calcination flue gas as claimed in claim 1 wherein, The extraction organic phase in step 6 is a mixture of anion extractant, phase dispersant and organic diluent; wherein the anion extractant comprises di(2-ethylhexyl)phosphate and / or tri-octylphosphine oxide; the phase dispersant comprises TBP and / or secondary octanol; the organic diluent is sulfonated kerosene; and / or, the volume ratio of the extractant to the tetravalent vanadium solution is O / A=0.2-1:1; ​ The anion extractant is selected from one or more of P507, P204; and / or, the volume ratio of the extraction organic phase to the tetravalent vanadium extraction stock solution is O / A=0.2-1:1; and / or, the extraction mode is multi-stage countercurrent extraction.

8. A process for the preparation of vanadyl sulphate using short flow process of vanadium slag calcination flue gas as claimed in claim 1 wherein, The extraction conditions in step 6 are: extraction temperature 25-35℃, extraction time 15-30min. ​ 9. A process for the preparation of vanadyl sulphate using short flow process of vanadium slag calcination flue gas as claimed in claim 1 wherein, The low concentration acid in step 7 refers to sulfuric acid with a concentration of 0.05mol / L-0.5mol / L; and / or, the volume ratio of the low concentration acid to the loaded organic phase is O / A=1-15:1; and / or, the washing mode is multi-stage countercurrent extraction. ​ 10. A process for the preparation of vanadyl sulphate using short flow process of vanadium slag calcination flue gas as claimed in claim 1 wherein, The high concentration acid used in stripping in step 8 refers to sulfuric acid with a concentration of 3mol / L-5mol / L; and / or, the volume ratio of the high concentration acid to the loaded organic phase is O / A=1-15:1; and / or, the stripping mode is multi-stage countercurrent extraction. ​

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