Method for cleanly extracting vanadium from high-calcium vanadium slag of converter

Abstract

The invention discloses a method for cleanly extracting vanadium from high-calcium vanadium slag of a converter. The method comprises the following steps: firstly, uniformly mixing the high-calcium vanadium slag with a magnesium sulfate solid and / or a solution containing magnesium sulfate, oxidizing and roasting at 500-700 DEG C for 30-60 minutes, heating to 850-900 DEG C, oxidizing and roasting for 30-90 minutes, cooling, and grinding to obtain a roasted material; carrying out primary acid leaching on the roasted material, and carrying out solid-liquid separation to obtain primary leachate and primary residues; ammonium sulfate is added into the primary leachate for high-temperature vanadium precipitation, ammonium polyvanadate and vanadium precipitation supernatant are obtained, and a vanadium oxide product is obtained after ammonium polyvanadate is calcined; the vanadium precipitation supernatant is mixed with the primary residues, secondary leaching is conducted, solid-liquid separation is conducted, and secondary leaching liquid and secondary residues are obtained; and finally, the secondary leaching solution can be used as an additive to be mixed and roasted with the high-calcium vanadium slag again for recycling. The process is simple and easy to implement, the leaching rate of vanadium and cleaning treatment of wastewater are improved by reasonably controlling the roasting conditions and recycling the secondary leachate, and the method has good environmental benefits and industrial application prospects.

Description

Technical Field

[0001] This invention belongs to the field of hydrometallurgical technology of vanadium, and specifically relates to a method for cleanly extracting vanadium from high-calcium vanadium slag in a converter. Background Technology

[0002] The main industrial processes for vanadium extraction from vanadium slag include sodium roasting-water leaching and calcification roasting-acid leaching. In the sodium roasting process, calcium has a significant impact on the conversion rate of vanadium. Calcium is considered a harmful element because it readily reacts with vanadium pentoxide during oxidative roasting to form insoluble calcium vanadate or calcium-containing vanadium bronze. Every 1% increase in the mass fraction of calcium oxide in the vanadium slag results in a loss of 4.7%-9.0% of vanadium pentoxide. For the calcification roasting-acid leaching process, an appropriate amount of CaO in the vanadium slag is beneficial for vanadium extraction because CaO reacts with various vanadium oxides during oxidation to form calcium vanadate, which is insoluble in water but soluble in acid. However, excessive calcium can also have adverse effects, such as low vanadium conversion rate during the roasting and acid leaching process of high-calcium vanadium slag, and the large amount of calcium sulfate produced during the leaching process leading to a viscous slurry and difficult filtration. Combining sulfuric acid with high-calcium vanadium slag for vanadium extraction during roasting can solve the problems of low vanadium transfer leaching rate and viscous slurry. However, the mixing process of sulfuric acid and high-calcium vanadium slag generates a large amount of heat, produces acid mist, and requires equipment with high corrosion resistance. When ferric sulfate is added, partial decomposition occurs at temperatures of 500℃ and above, leading to sulfate loss, necessitating an increase in the amount added.

[0003] In the treatment of vanadium precipitation wastewater, the calcification vanadium extraction process mainly uses the lime slurry neutralization method, which is low-cost and simple, but produces a large amount of neutralized gypsum slag. The sodium vanadium extraction process mainly uses the evaporation concentration method, which is costly, difficult to maintain equipment, and produces a large amount of unusable ammonium sulfate and sodium sulfate double salts. To solve the problems in wastewater treatment, patent CN104058523A first removes impurities from the wastewater in stages to remove vanadium, phosphorus, and iron, and then performs variable current electrolysis to obtain metallic manganese products that meet national standards. The treated wastewater is then recycled. Patent CN109930008A mixes and roasts low-calcium vanadium slag with a CaO / V2O3 mass ratio of approximately 0.2 with magnesium oxide, then leaches it under high acidity (pH=0.5-1.0), then precipitates vanadium oxide from the leachate to prepare vanadium oxide products. Finally, the pH of the vanadium precipitation supernatant is adjusted to obtain manganese hydroxide and magnesium precipitates, achieving resource utilization of the wastewater. This patented method for treating wastewater involves adjusting the pH by first adding ammonia water or introducing ammonia gas, thereby precipitating manganese and magnesium in the upper layer of the vanadium precipitation solution in stages. The magnesium-containing precipitate is then roasted to prepare magnesium oxide, which is then returned for further roasting. This process is relatively complex. It is worth noting that because the leaching process has a low pH, a large number of impurities from the residue enter the solution. Consequently, the V₂O₅ content of the product obtained from the subsequent vanadium precipitation is only 93%-93.7%, far below the 98% V₂O₅ content required by the YB / T 5304-2017 standard implemented in 2018.

[0004] This patent, through extensive experimental research, discovered that adding magnesium sulfate during roasting significantly increases the vanadium leaching rate from high-calcium vanadium slag. Therefore, a method for cleanly extracting vanadium from converter high-calcium vanadium slag is proposed. First, the high-calcium vanadium slag is mixed with solid magnesium sulfate and / or a solution containing magnesium sulfate, followed by high-temperature oxidative roasting. The resulting clinker is then acid-leached and vanadium-precipitated to obtain vanadium oxide. The supernatant from the vanadium precipitation is returned to the residue for deep leaching and then used as an additive for direct return to roasting for recycling. This achieves vanadium recovery from converter high-calcium vanadium slag and clean treatment of wastewater. This patent improves the total vanadium leaching rate without introducing new impurities, and the obtained product meets standard requirements. It achieves efficient vanadium extraction and wastewater recycling, and the process is simple and easy to implement. Summary of the Invention

[0005] In view of this, one of the objectives of the present invention is to provide a method for cleanly extracting vanadium from high-calcium vanadium slag in a converter, comprising the following steps:

[0006] 1) Mix the high-calcium vanadium slag with the additives evenly, oxidize and roast at 500-700 ℃ for 30-60 min, then raise the temperature to 850-900 ℃ and oxidize and roast for another 30-90 min. After cooling, grind finely to obtain the roasted material.

[0007] 2) Add the roasted material to water, add sulfuric acid for the first acid leaching, stir the reaction, separate the solid and liquid, and obtain the first leachate and the first residue;

[0008] 3) Add ammonium sulfate to the primary leachate for high-temperature vanadium precipitation to obtain ammonium polyvanadate and vanadium precipitation supernatant. Calcining the ammonium polyvanadate yields vanadium pentoxide.

[0009] 4) Mix the vanadium precipitate supernatant with the primary residue, add sulfuric acid for a second acid leaching, stir the reaction, and separate the solid and liquid to obtain a secondary leachate and a secondary residue.

[0010] The additive mentioned in step 1) is solid magnesium sulfate and / or a solution containing magnesium sulfate; the molar ratio of sulfate ions in the additive to calcium in the high-calcium vanadium slag is 0.2-0.8:1; the secondary leachate mentioned in step 4) is returned to step 1) for recycling as the solution containing magnesium sulfate.

[0011] Furthermore, the mass ratio of CaO / V2O5 in the high-calcium vanadium slag described in step 1) is 0.96-2.23.

[0012] Furthermore, the ratio of water to calcined material in step 2) is 1.8-2.5 mL:1 g.

[0013] Furthermore, the pH value of the first acid leaching described in step 2) is between 2.5 and 3.5.

[0014] Furthermore, the reaction time described in step 2) is 45-90 min.

[0015] Further, in step 4), the ratio of the vanadium precipitation supernatant to the primary residue is 1-2 mL:1 g.

[0016] Furthermore, the pH value of the second acid leaching described in step 4) is between 0.5 and 0.8.

[0017] Furthermore, the reaction time described in step 4) is 5-20 min.

[0018] This invention first mixes high-calcium vanadium slag with solid magnesium sulfate and / or a solution containing magnesium sulfate, then performs high-temperature oxidation roasting, followed by a first acid leaching to precipitate vanadium and obtain vanadium oxide product. The supernatant of the vanadium precipitate is then returned to the residue for a second acid leaching to obtain a second leachate. The second leachate is then used as an additive and returned to roasting. This invention achieves the recycling of magnesium and the recovery of vanadium in high-calcium vanadium slag from converters without introducing new impurities. Detailed Implementation

[0019] The present invention will be described in detail below with reference to embodiments. These embodiments are for illustrative purposes only and are not intended to limit the scope of application of the present invention. The present invention is not limited to the following embodiments or examples. Any modifications and variations made without departing from the spirit of the present invention should be included within the scope of the present invention. Unless otherwise specified, the experimental materials or reagents used in the following embodiments are commercially available.

[0020] Example 1:

[0021] Take 200 g of high-calcium vanadium slag (CaO=13.85%, TV=8.1%, Mn=6.3%, Mg=1.1%, CaO / V2O5 mass ratio=0.96), add 60 g of magnesium sulfate heptahydrate, the molar ratio of sulfate to calcium in the high-calcium vanadium slag is about 0.49:1, mix evenly, oxidize and roast at 500 ℃ for 60 min, then raise the temperature to 900 ℃ and continue roasting for 30 min. After roasting, cool and grind to obtain 226 g of roasted material (TV=7.17%, Mg=3.56%).

[0022] 200 g of the roasted material was added to 400 mL of water, and sulfuric acid was added to control the pH to 2.8-3.0 for the first acid leaching. The mixture was stirred for 60 min, and the solid and liquid were separated. The residue was washed to obtain a leachate (V=31.6 g / L, volume 375 mL), a primary residue (TV=0.65%, Mg=2.12%), and a washing filtrate (V=6.0 g / L, volume 210 mL). The primary leaching rate of vanadium was 91.39%, and the primary leaching rate of magnesium was 43.43%. 360 mL of the leachate was taken, 15 g of ammonium sulfate was added, and the pH was adjusted to 1.8 with sulfuric acid. Vanadium was precipitated under boiling conditions to obtain ammonium polyvanadate and a vanadium precipitation supernatant (V=0.19 g / L). The vanadium precipitation rate was 99.4%.

[0023] The vanadium supernatant was mixed with the primary residue at a liquid-to-solid ratio of 1:1 (mL:g). Sulfuric acid was added to adjust the pH of the system to 0.6 for a second acid leaching. The mixture was stirred at room temperature for 10 min. Solid-liquid separation was performed to obtain a secondary leachate (Mg=11.82 g / L, TV=3.25 g / L, Mn=10.46 g / L, Fe=2.16 g / L, sulfate 93.1 g / L) and a secondary residue (TV=0.32%, Mg=1.31%, Mn=4.22%). The total vanadium leaching rate was 96.4%, and the total magnesium leaching rate was 66.88%.

[0024] Example 2:

[0025] Take 200 g of high-calcium vanadium slag (CaO=13.85%, TV=8.1%, Mn=6.3%, Mg=1.1%, CaO / V2O5 mass ratio=0.96), add 120 mL of the secondary leaching solution from Example 1 (Mg=11.82 g / L, TV=3.25 g / L, Mn=10.46 g / L, Fe=2.16 g / L, sulfate 93.1 g / L), the molar ratio of sulfate to calcium in the high-calcium vanadium slag is approximately 0.24:1. After mixing evenly, oxidize and roast at 600 ℃ for 30 min, then raise the temperature to 850 ℃ and continue roasting for 90 min. After roasting, cool and grind to obtain 222.5 g of roasted material (TV=7.28%, Mg=1.63%).

[0026] 200 g of roasted material was added to 500 mL of water, and sulfuric acid was added to control the pH to 2.5-2.8 for the first acid leaching. The mixture was stirred for 45 min, and the solid and liquid were separated. The residue was washed to obtain a leachate (V=26.1 g / L, volume 458 mL), a primary residue (TV=0.70%, Mg=1.02%), and a washing filtrate (V=6.5 g / L, volume 200 mL). The primary leaching rate of vanadium was 90.87%, and the primary leaching rate of magnesium was 40.55%. 400 mL of the leachate was taken, 15 g of ammonium sulfate was added, and the pH was adjusted to 1.8 with sulfuric acid. Vanadium was precipitated under boiling conditions to obtain ammonium polyvanadate and a vanadium precipitation supernatant (V=0.21 g / L). The vanadium precipitation rate was 99.2%.

[0027] The vanadium precipitate supernatant was mixed with the primary residue at a liquid-to-solid ratio of 1.5:1 (mL:g). Sulfuric acid was added to adjust the pH of the system to 0.8 for a second acid leaching. The mixture was stirred at room temperature for 20 min. Solid-liquid separation was performed to obtain a secondary leaching solution (Mg=8.75 g / L, TV=1.93 g / L, Mn=8.92 g / L, Fe=1.84 g / L, sulfate 81.6 g / L) and secondary residue (TV=0.41%, Mg=0.41%, Mn=4.81%). The total vanadium leaching rate was 94.8%, and the total magnesium leaching rate was 77.36%.

[0028] Example 3:

[0029] Take 200 g of high-calcium vanadium slag (CaO=18.76%, TV=8.06%, Mn=6.4%, Mg=1.2%, CaO / V2O5 mass ratio=1.3), add 80 mL of the secondary leaching solution from Example 1 (Mg=11.82 g / L, TV=3.25 g / L, Mn=10.46 g / L, Fe=2.16 g / L, sulfate 93.1 g / L) and 49 g of magnesium sulfate heptahydrate. The molar ratio of sulfate to calcium in the high-calcium vanadium slag is approximately 0.56:1. Mix evenly, oxidize and roast at 700 ℃ for 30 min, then raise the temperature to 870 ℃ and continue oxidizing and roasting for 60 min. After roasting, cool and grind to obtain 250 g of roasted material (TV=6.45%, Mg=3.25%).

[0030] 200 g of roasted material was added to 360 mL of water, and sulfuric acid was added to control the pH to 3.0-3.5 for the first acid leaching. The mixture was stirred for 90 min, and the solid and liquid were separated. The residue was washed to obtain a leachate (V=30.5 g / L, volume 340 mL), a primary residue (TV=0.54%, Mg=2.16%), and a washing filtrate (V=7.6 g / L, volume 200 mL). The primary leaching rate of vanadium was 92.05%, and the primary leaching rate of magnesium was 36.86%. 330 mL of the leachate was taken, and 12 g of ammonium sulfate was added. The pH was adjusted to 1.8 with sulfuric acid, and vanadium was precipitated under boiling conditions to obtain ammonium polyvanadate and a vanadium precipitation supernatant (V=0.2 g / L). The vanadium precipitation rate was 99.3%.

[0031] The vanadium precipitation supernatant was mixed with the primary residue at a liquid-to-solid ratio of 1.5:1 (mL:g). Sulfuric acid was added to adjust the pH of the system to 0.5 for a second acid leaching. The mixture was stirred for 5 min, and the solid and liquid phases were separated to obtain a secondary leachate (Mg=15.5 g / L, TV=1.53 g / L, Mn=10.2 g / L, Fe=2.45 g / L, sulfate 110 g / L) and a secondary residue (TV=0.31%, Mg=1.1%, Mn=3.95%). The total vanadium leaching rate was 95.6%, and the total magnesium leaching rate was 69.54%.

[0032] Example 4:

[0033] Take 200 g of high-calcium vanadium slag (CaO=25.07%, TV=6.31%, Mn=4.05%, Mg=1.2%, CaO / V2O5 mass ratio=2.23), add 150 g of magnesium sulfate heptahydrate, the molar ratio of sulfate to calcium in the high-calcium vanadium slag is about 0.68:1, mix evenly, oxidize and roast at 500 ℃ for 60 min, then raise the temperature to 850 ℃ and continue roasting for 90 min. After roasting, cool and grind to obtain 268 g of roasted material (TV=4.71%, Mg=6.34%).

[0034] 200 g of roasted material was added to 400 mL of water, and sulfuric acid was added to control the pH to 2.8-3.0 for the first acid leaching. The mixture was stirred for 60 min, and the solid and liquid were separated. The residue was washed to obtain a leachate (V=19.61 g / L, volume 365 mL), a primary residue (TV=0.52%, Mg=4.04%), and a washing filtrate (V=6.05 g / L, volume 210 mL). The primary leaching rate of vanadium was 89.51%, and the primary leaching rate of magnesium was 39.46%. 360 mL of the leachate was taken, 15 g of ammonium sulfate was added, and the pH was adjusted to 1.8 with sulfuric acid. Vanadium was precipitated under boiling conditions to obtain ammonium polyvanadate and a vanadium precipitation supernatant (V=0.15 g / L). The vanadium precipitation rate was 99.2%.

[0035] The vanadium supernatant was mixed with the primary residue at a liquid-to-solid ratio of 1.8:1 (mL:g). Sulfuric acid was added to adjust the pH of the system to 0.6 for a second acid leaching. The mixture was stirred at room temperature for 10 min. Solid-liquid separation was performed to obtain a secondary leachate (Mg=27.67 g / L, TV=1.27 g / L, Mn=6.51 g / L, Fe=2.02 g / L, sulfate 141 g / L) and a secondary residue (TV=0.28%, Mg=2.51%, Mn=1.8%). The total vanadium leaching rate was 94.65%, and the total magnesium leaching rate was 64.37%.

[0036] Comparative Example 1: (Direct roasting, wastewater neutralization treatment)

[0037] Take 200 g of high-calcium vanadium slag (CaO=18.76%, TV=8.06%, Mn=6.4%, Mg=1.2%, CaO / V2O5 mass ratio=1.3), oxidize and roast at 600 ℃ for 30 min, then raise the temperature to 900 ℃ and continue oxidizing and roasting for 60 min. After roasting, cool and grind to obtain 213.8 g of roasted material (TV=7.54%, Mn=5.99%, Mg=1.12%).

[0038] 200 g of roasted material was added to 600 mL of water, and sulfuric acid was added to control the pH to 3.0-3.5 for the first acid leaching. The mixture was stirred for 90 min, and the solid and liquid were separated. The residue was washed to obtain a leachate (V=17.75 g / L, volume 580 mL), a primary residue (TV=2.19%, Mn=5.01%, Mg=1.06%), and a washing filtrate (V=3.55 g / L, volume 200 mL). The primary leaching rates of vanadium, manganese, and magnesium were 72.41%, 20.54%, and 10.09%, respectively. 500 mL of the leachate was taken, and 12 g of ammonium sulfate was added. The pH was adjusted to 1.8 with sulfuric acid, and vanadium was precipitated under boiling conditions to obtain ammonium polyvanadate and a vanadium precipitation supernatant (V=0.17 g / L). The vanadium precipitation rate was 99.0%.

[0039] Lime slurry was added to the vanadium supernatant (480 mL) to adjust the pH to 9.0. After stirring for 30 min, solid-liquid separation was performed to obtain gypsum residue (30.6 g dry basis, Ca=35.8%, Mn=10.5%, V=0.27%) and 410 mL filtrate (Mn=0.8 g / L). The filtrate and primary residue were mixed at a liquid-to-solid ratio of 1.5:1 (mL:g), and sulfuric acid was added to adjust the pH to 0.5 for a second acid leaching. The reaction was stirred at room temperature for 5 min, and solid-liquid separation was performed to obtain a secondary leachate (TV=5.28 g / L, Mn=1.6 g / L, Fe=2.1 g / L, sulfate 27 g / L) and secondary residue (TV=1.26%, Mg=0.92%, Mn=4.88%). The total vanadium leaching rate was 85.9%, the total manganese leaching rate was 26.68%, and the total magnesium leaching rate was 26.07%.

[0040] Comparative Example 2: (Mixed with manganese sulfate and roasted)

[0041] Take 200 g of high-calcium vanadium slag (CaO=18.76%, TV=8.06%, Mn=6.4%, Mg=1.2%, CaO / V2O5 mass ratio=1.3), add 65 g of manganese sulfate monohydrate, mix well, oxidize and roast at 600 ℃ for 30 min, then raise the temperature to 870 ℃ and continue roasting for 60 min. After roasting, cool and grind to obtain 268 g of roasted material (TV=6.01%, Mn=12.67%).

[0042] 200 g of roasted material was added to 360 mL of water, and sulfuric acid was added to control the pH to 3.0-3.5 for the first acid leaching. The mixture was stirred for 90 min, and the solid and liquid were separated. The residue was washed to obtain a leachate (V=27.4 g / L, volume 338 mL), a primary residue (TV=0.75%, Mn=12.01%), and a washing filtrate (V=6.4 g / L, volume 200 mL). The primary leaching rate of vanadium was 88.14%, and the primary leaching rate of manganese was 9.95%. 330 mL of the leachate was taken, and 12 g of ammonium sulfate was added. The pH was adjusted to 1.8 with sulfuric acid, and vanadium was precipitated under boiling conditions to obtain ammonium polyvanadate and a vanadium precipitation supernatant (V=0.12 g / L). The vanadium precipitation rate was 99.6%.

[0043] The vanadium precipitation supernatant was mixed with the primary residue at a liquid-to-solid ratio of 1.5:1 (mL:g). Sulfuric acid was added to adjust the pH of the system to 0.5 for a second acid leaching. The mixture was stirred at room temperature for 5 min. Solid-liquid separation was performed to obtain a secondary leachate (TV=2.4 g / L, Mn=11.1 g / L, Mg=1.5 g / L, Fe=2.3 g / L, sulfate 64 g / L) and a secondary residue (TV=0.35%, Mg=0.71%, Mn=11.87%). The total vanadium leaching rate was 94.8%, and the total manganese leaching rate was 15.68%.

[0044] Comparative Example 3: (Sodium sulfate was added and the mixture was roasted)

[0045] Take 200 g of high-calcium vanadium slag (CaO=18.76%, TV=8.06%, Mn=6.4%, Mg=1.2%, CaO / V2O5 mass ratio=1.3), add 54.3 g of anhydrous sodium sulfate, mix evenly, oxidize and roast at 600 ℃ for 30 min, then raise the temperature to 900 ℃ and continue oxidizing and roasting for 60 min. After roasting, cool and grind to obtain 267 g of roasted material (TV=6.04%, Na=3.29%).

[0046] 200 g of roasted material was added to 360 mL of water, and sulfuric acid was added to control the pH to 3.0-3.5 for the first acid leaching. The mixture was stirred for 90 min, and the solid and liquid were separated. The residue was washed to obtain a leachate (V=20.4 g / L, volume 342 mL), a primary residue (TV=3.06%, Na=1.05%), and a washing filtrate (V=4.9 g / L, volume 200 mL). The primary leaching rate of vanadium was 51.87%, and the primary leaching rate of sodium was 69.68%. 330 mL of the leachate was taken, and 12 g of ammonium sulfate was added. The pH was adjusted to 1.8 with sulfuric acid, and vanadium was precipitated under boiling conditions to obtain ammonium polyvanadate and a vanadium precipitation supernatant (V=0.15 g / L). The vanadium precipitation rate was 99.3%.

[0047] The vanadium supernatant was mixed with the primary residue at a liquid-to-solid ratio of 1.5:1 (mL:g). Sulfuric acid was added to adjust the pH of the system to 0.5 for a second acid leaching. The mixture was stirred at room temperature for 5 min. Solid-liquid separation was performed to obtain a secondary leachate (Na=3.4 g / L, TV=5.57 g / L, Mn=5.7 g / L, Fe=2.2 g / L, sulfate 51 g / L) and a secondary residue (TV=2.24%, Na=0.44%, Mg=0.71%, Mn=4.25%). The total vanadium leaching rate was 80.4%, and the total sodium leaching rate was 84.13%.

[0048] Comparative Example 4: (Zinc sulfate was added and roasted)

[0049] Take 200 g of high-calcium vanadium slag (CaO=18.76%, TV=8.06%, Mn=6.4%, Mg=1.2%, CaO / V2O5 mass ratio=1.3), add 110 g of zinc sulfate heptahydrate, mix evenly, oxidize and roast at 500 ℃ for 60 min, then raise the temperature to 900 ℃ and continue roasting for 60 min. After roasting, cool and grind to obtain 277 g of roasted material (TV=5.82%, Zn=9.03%).

[0050] 200 g of roasted material was added to 400 mL of water, and sulfuric acid was added to control the pH to 2.8-3.0 for the first acid leaching. The mixture was stirred for 60 min, and the solid and liquid were separated. The residue was washed to obtain a leachate (V=24.27 g / L, volume 370 mL), a primary residue (TV=0.57%, Zn=9.05%), and a washing filtrate (V=7.7 g / L, volume 205 mL). The primary leaching rate of vanadium was 90.9%, and the primary leaching rate of zinc was 4.78%. 360 mL of the leachate was taken, 15 g of ammonium sulfate was added, and the pH was adjusted to 1.8 with sulfuric acid. Vanadium was precipitated under boiling conditions to obtain ammonium polyvanadate and a vanadium precipitation supernatant (V=0.17 g / L). The vanadium precipitation rate was 99.3%.

[0051] The vanadium supernatant was mixed with the primary residue at a liquid-to-solid ratio of 1.8:1 (mL:g). Sulfuric acid was added to adjust the pH of the system to 0.6 for a second acid leaching. The mixture was stirred at room temperature for 10 min. Solid-liquid separation was performed to obtain a secondary leaching solution (Zn=3.71 g / L, TV=0.83 g / L, Mn=8.24 g / L, Fe=2.12 g / L, sulfate 71 g / L) and secondary residue (TV=0.42%, Zn=9.08%, Mn=3.17%). The total vanadium leaching rate was 93.5%, and the total zinc leaching rate was 9.5%.

[0052] Comparative Example 5: (Mixed and roasted with magnesium oxide)

[0053] Take 200 g of high-calcium vanadium slag (CaO=18.76%, TV=8.06%, Mn=6.4%, Mg=1.2%, CaO / V2O5 mass ratio=1.3), add 16 g of magnesium oxide, mix evenly, oxidize and roast at 500 ℃ for 60 min, then raise the temperature to 900 ℃ and continue roasting for 60 min. After roasting, cool and grind to obtain 228 g of roasted material (TV=7.07%).

[0054] 200 g of roasted material was added to 600 mL of water, and sulfuric acid was added to control the pH to 3.0-3.5 for the first acid leaching. The mixture was stirred for 90 min, and the solid and liquid were separated. The residue was washed to obtain a leachate (V=16.58 g / L, volume 590 mL), a primary residue (TV=1.87%), and a washing filtrate (V=4.74 g / L, volume 210 mL). The vanadium primary leaching rate was 76.2%. 550 mL of the leachate was taken, 12 g of ammonium sulfate was added, and the pH was adjusted to 1.8 with sulfuric acid. Vanadium was precipitated under boiling conditions to obtain ammonium polyvanadate and a vanadium precipitation supernatant.

[0055] The vanadium precipitation supernatant was mixed with the primary residue at a liquid-solid ratio of 1.5:1 (mL:g). Sulfuric acid was added to adjust the pH of the system to 0.5 for a second acid leaching. The mixture was stirred at room temperature for 5 min. Solid-liquid separation was performed to obtain a secondary leachate and a secondary residue (TV=1.15%). The total vanadium leaching rate was 85.36%.

[0056] The ammonium polyvanadate obtained in the above examples and comparative examples was calcined at 550-580℃. The quality of the vanadium pentoxide product obtained is shown in Table 1 below, which meets the requirements of YB / T 5304-2017 standard. The vanadium extraction effect of each example and comparative example is shown in Table 2.

[0057] Table 1 Vanadium pentoxide product composition / %

[0058]

[0059] Table 2 Comparison of Vanadium Extraction Effects

[0060]

[0061] In the above embodiments, the use of magnesium sulfate / manganese sulfate / secondary leaching solution as additives significantly improved the vanadium leaching effect. Comparison with Examples 1-4:

[0062] In Comparative Example 1, the total vanadium leaching rate during direct roasting and acid leaching of the clinker was only 85.9%, lower than that of Examples 1-4. It should be noted that, without the addition of manganese, due to the large amount of manganese present in the vanadium slag, some manganese will enter the solution during the roasting-acid leaching process. To avoid manganese accumulation in the system, the conventional method is to treat the vanadium precipitation supernatant using lime, but this process generates a large amount of waste gypsum slag. Examples 1-4 of this patent first use the supernatant as a secondary leaching residue to obtain a secondary leaching solution containing vanadium, magnesium, and manganese, which is then returned to be mixed with high-calcium slag for roasting, achieving resource recycling of the supernatant.

[0063] In Comparative Example 2, the total vanadium leaching rate in the clinker was slightly lower when manganese sulfate was added during roasting. Furthermore, a large amount of the added manganese remained in the secondary residue after roasting and acid leaching. For example, the Mn content in the secondary residue of Comparative Example 2 was as high as 11.87%, and the total manganese leaching rate was only 15.68%. Some manganese was oxidized to a higher valence state during the high-temperature roasting process, making it difficult to leach into the solution and resulting in a waste of manganese resources. However, in Examples 1-4, when magnesium sulfate solid and / or a secondary leaching solution containing magnesium sulfate were added, magnesium re-entered the solution during the acid leaching process after high-temperature roasting and circulated as magnesium sulfate. For example, the magnesium content in the secondary residue was as low as 0.41%, and the total magnesium leaching rate reached as high as 77.36%.

[0064] In Comparative Example 3, after adding sodium sulfate and calcining, the sodium leaching rate in the clinker reached 84.13%, but the total vanadium leaching rate was only 80.4%, significantly lower than in Examples 1-4. This demonstrates that adding sodium sulfate is detrimental to the conversion and leaching of vanadium in high-calcium vanadium slag.

[0065] Comparative Example 4, after adding zinc sulfate and calcining, achieved a total vanadium leaching rate of 93.5%. Under the same vanadium slag raw material conditions, this significantly improved the vanadium leaching effect compared to Comparative Example 1 (direct calcination, 85.9%), and was 2.9% lower than when magnesium sulfate was added (Example 1, 96.4%). However, a large amount of the added zinc remained in the secondary residue, reaching 9.08%, resulting in a total zinc leaching rate of only 9.5%, significantly lower than that of magnesium sulfate. Similarly, adding other sulfates, such as aluminum sulfate, also resulted in a large amount remaining in the secondary residue, similar to zinc sulfate and manganese sulfate, leading to a very low leaching rate.

[0066] Comparative Example 5, after adding magnesium oxide and calcining, achieved a total vanadium leaching rate of 85.36%. Under the same vanadium slag raw material conditions, this was 0.54% lower than Comparative Example 1 (direct calcination, 85.9%) and 11.04% lower than when magnesium sulfate was added (Example 1, 96.4%). This demonstrates that adding magnesium oxide to high-calcium vanadium slag followed by calcination does not improve the vanadium leaching rate in the clinker; the effect is far less than that of magnesium sulfate.

[0067] In summary, the vanadium pentoxide products obtained by roasting sulfate and high-calcium vanadium slag using the method of this invention all meet the requirements of YB / T 5304-2017 standard. However, the total vanadium leaching rate obtained by roasting sodium sulfate and high-calcium vanadium slag is only 80.4%, which is significantly lower than that of other sulfates. The total leaching rate of manganese or zinc after roasting manganese sulfate and zinc sulfate is significantly lower than that of magnesium sulfate, at only 15.68% and 9.5% respectively. Moreover, a large amount of manganese or zinc remains in the secondary residue, which is difficult to recycle and results in waste. By mixing and roasting solid magnesium sulfate and / or magnesium sulfate-containing solutions with high-calcium vanadium slag, the vanadium leaching rate can be significantly increased (reaching 94.65-96.4%). At the same time, a large amount of magnesium enters the solution, and the secondary leachate contains a large amount of magnesium sulfate. The magnesium sulfate leaching rate is significantly higher than that of other sulfates, reaching up to 77.36% (the Mg content in the secondary residue is only 0.41%). Furthermore, this patent mixes solid magnesium sulfate and / or magnesium sulfate-containing secondary leachate with high-calcium vanadium slag, achieving the recycling of magnesium sulfate and the recovery of vanadium in the secondary leachate without introducing new impurities.

[0068] The conventional techniques and solutions not described in detail in the above embodiments are all well known in the art, and therefore will not be elaborated upon here. The above embodiments and / or experimental examples describe the preferred embodiments of the present invention in detail. However, the present invention is not limited to the specific details in the above embodiments. Within the scope of the technical concept of the present invention, various simple modifications can be made to the technical solutions of the present invention, and these simple modifications all fall within the protection scope of the present invention.

Claims

1. A method for cleanly extracting vanadium from high-calcium vanadium slag in a converter, characterized in that, Includes the following steps: 1) Mix the high-calcium vanadium slag with the additives evenly, oxidize and roast at 500-700 ℃ for 30-60 min, then raise the temperature to 850-900 ℃ and oxidize and roast for another 30-90 min. After cooling, grind finely to obtain the roasted material. 2) Add the roasted material to water, add sulfuric acid for the first acid leaching, stir the reaction, separate the solid and liquid, and obtain the first leachate and the first residue; 3) Add ammonium sulfate to the primary leachate for high-temperature vanadium precipitation to obtain ammonium polyvanadate and vanadium precipitation supernatant. Calcining the ammonium polyvanadate yields vanadium pentoxide. 4) Mix the vanadium precipitation supernatant with the primary residue, add sulfuric acid for a second acid leaching, stir the reaction, and separate the solid and liquid to obtain a secondary leachate and a secondary residue. The additive mentioned in step 1) is solid magnesium sulfate and / or a solution containing magnesium sulfate; the molar ratio of sulfate ions in the additive to calcium in the high-calcium vanadium slag is 0.2-0.8:1; the secondary leachate mentioned in step 4) is returned to step 1) for recycling as the solution containing magnesium sulfate.

2. The method as described in claim 1, characterized in that, The mass ratio of CaO / V2O5 in the high-calcium vanadium slag mentioned in step 1) is 0.96-2.

23.

3. The method as described in claim 1, characterized in that, The ratio of water to calcined material in step 2) is 1.8-2.5 mL:1 g.

4. The method as described in claim 1, characterized in that, The pH value of the first acid leaching described in step 2) is 2.5-3.

5.

5. The method as described in claim 1, characterized in that, The reaction time described in step 2) is 45-90 min.

6. The method as described in claim 1, characterized in that, In step 4), the ratio of the vanadium precipitation supernatant to the primary residue is 1-2 mL: 1 g.

7. The method as described in claim 1, characterized in that, The pH value of the second acid leaching described in step 4) is 0.5-0.

8.

8. The method as described in claim 1, characterized in that, The reaction time described in step 4) is 5-20 min.

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