A method for efficiently separating iron and manganese components in iron-manganese ore based on step-by-step leaching

Abstract

The present invention relates to a method for efficiently separating iron and manganese components from ferromanganese ore based on step-by-step leaching. The method comprises: mixing the ferromanganese ore with an intermediate product or a vanadium product from the vanadium industry to obtain a mixed material; calcining the mixed material at high temperature to obtain a roasted clinker; subjecting the roasted clinker to primary and secondary leaching to obtain a manganese-containing leachate and a vanadium-containing and iron-containing leach residue; precipitating or drying and crystallizing the obtained manganese-containing leachate, and filtering to obtain a manganese product and a manganese-extracted waste liquid; leaching the vanadium-containing and iron-containing leach residue, controlling the pH of the leaching system with an alkaline solution, and filtering after leaching to obtain a vanadium-containing leachate and an iron-containing leach residue. The method of the present invention achieves efficient separation of iron and manganese from ferromanganese ore, as well as efficient recycling of vanadium-containing intermediate products or vanadium products from the vanadium industry.

Description

Technical Field

[0001] The invention belongs to the technical fields of ferromanganese ore resource utilization, vanadium chemical technology and resource recycling, and particularly relates to a method for efficiently separating iron and manganese components in ferromanganese ore based on step-by-step leaching. Background Art

[0002] Manganese, used in metallurgy, batteries, and agriculture, is a scarce and essential metal that sustains the national economy. my country's manganese ore resources are characterized by high iron impurity levels and a scarcity of high-grade manganese ore. Of the explored deposits, 73% have iron content exceeding the standard. To promote the sustainable development of my country's manganese ore industry, increasing the utilization rate of manganese resources is imperative.

[0003] Existing iron and manganese separation technologies mainly include mechanical separation, pyrolysis and wet separation. Mechanical separation cannot achieve ideal technical and economic indicators due to the complex composition and fine embedded particle size of iron and manganese ore. Pyrolysis separation requires coke as a reducing agent and smelting at high temperature, which is characterized by high energy consumption and high pollution. It is a common phenomenon for iron and manganese to be leached together in wet separation, which makes the subsequent removal of iron from manganese more difficult.

[0004] A large number of studies have shown that compared with various existing separation technologies, the leaching method can obtain better separation indicators, has high resource utilization, simple operation, and low energy consumption. However, the disadvantages of co-leaching of iron and manganese in traditional chemical leaching methods are often not solved, and the roasting medium loss rate is high. Therefore, a roasting medium with high selectivity for iron or manganese is sought, and under the condition of low roasting medium loss rate, the mineral phase transformation of iron and manganese is promoted to achieve efficient separation of iron and manganese. Summary of the Invention

[0005] (1) Technical issues to be solved

[0006] In order to solve the above-mentioned problems of the prior art, the present invention proposes a method for efficiently separating iron and manganese components in iron-manganese ore based on step-by-step leaching. The method utilizes vanadium to directionally reconstruct and transform the insoluble manganese minerals in the iron-manganese ore into acid-soluble manganese vanadate, and regulates the physical phase of the iron minerals. The iron and manganese components are then leached out through an acid solution to achieve efficient separation and recycling of the roasting medium.

[0007] (2) Technical solution

[0008] In order to achieve the above objectives, the main technical solutions adopted by the present invention include:

[0009] A method for efficiently separating iron and manganese components in iron-manganese ore based on step-by-step leaching comprises the following steps:

[0010] S1. Roasting: Mixing ferromanganese ore with an intermediate product of the vanadium industry or a vanadium product to obtain a mixed material, wherein the mixing ratio of the ferromanganese ore and the intermediate product of the vanadium industry or the vanadium product is calculated as n(MnO2) / n(V2O5), wherein MnO2 is manganese in the ferromanganese ore (calculated as MnO2), and V2O5 is vanadium in the vanadium product or the intermediate product of the vanadium industry or the vanadium product (calculated as V2O5), and n(MnO2) / n(V2O5) is 0.5 to 3:1; roasting the mixed material at a high temperature to obtain a roasted clinker;

[0011] S2, primary leaching: Leaching the roasted clinker obtained in step S1, using an acid solution to adjust the pH value of the leaching system to 2-3; during the actual roasting process, the vanadium element directionally reconstructs and transforms the insoluble manganese minerals in the iron-manganese ore into acid-soluble manganese vanadate, and regulates the physical phase of the iron minerals, and then leaching them out through the acid solution to achieve efficient separation of the iron and manganese components and recycling of the roasting medium;

[0012] S3, secondary leaching: using sulfuric acid solution to continue adjusting the pH value to below 1.8, leaching and filtering to obtain manganese-containing leachate and vanadium-containing and iron-containing leach residue;

[0013] S4, recovering manganese: precipitating or drying and crystallizing the manganese-containing leachate obtained in step S3, and filtering to obtain a manganese product and a manganese extraction waste liquid;

[0014] S5, separation of vanadium and iron: leaching the vanadium-containing and iron-containing leaching residue obtained in step S3, controlling the pH value of the leaching system to 7.0-14.0 using an alkaline solution, filtering after leaching to obtain a vanadium-containing leaching solution and an iron-containing leaching residue.

[0015] The method as described above preferably further comprises:

[0016] S6. Recovering vanadium: Adding a vanadium precipitation additive to the vanadium-containing leachate obtained in step S5, and adjusting the pH to 2-3 to precipitate vanadium. After precipitation, filtering to obtain a vanadium industry intermediate product or vanadium product and vanadium extraction waste liquid; the recovered vanadium industry intermediate product or vanadium product can be directly exported as a product or recycled as a roasting additive;

[0017] S7. Utilization of the intermediate product or vanadium product in the vanadium industry: The intermediate product or vanadium product obtained in step S6 is output as a product, or recycled back into the roasting system as an additive to participate in the next round of roasting;

[0018] S8. Utilization of iron-containing tailings: The iron-containing leached slag obtained in step S5 is used as feed material for a blast furnace or as a raw material for preparing ferroalloys. The iron-containing leached slag has a high iron content, which can meet the standards of iron concentrate (TFe>60%) and can be directly used as a raw material for blast furnace ironmaking or as a raw material for preparing ferroalloys.

[0019] S9, waste liquid recycling: After impurities are removed from the manganese extraction waste liquid and vanadium extraction waste liquid obtained in steps S4 and S6, they are recycled back to the leaching system as the leaching medium in step S2.

[0020] In the method as described above, preferably, in step S1, the ferromanganese ore is a manganese ore with a Mn / Fe mass ratio of less than 3, and can be at least one of an oxide mineral, a hydrogenated mineral, a carbonate mineral, a sulfide mineral, a silicate mineral and a borate mineral; and its particle size is 200-300 mesh.

[0021] Preferably, the vanadium industry intermediates or products include at least one of vanadates, vanadium oxides, and vanadium slag. Vanadium salts include polyvanadates, ammonium metavanadate, pyrovanadates, and orthovanadates, specifically ammonium polyvanadate, ammonium metavanadate, sodium metavanadate, sodium pyrovanadate, sodium orthovanadate, calcium metavanadate, calcium pyrovanadate, calcium orthovanadate, manganese metavanadate, manganese pyrovanadate, and manganese orthovanadate. Vanadium oxides include at least one of vanadium trioxide, vanadium dioxide, and vanadium pentoxide. Vanadium slag is an intermediate product obtained from the vanadium-titanium magnetite smelting process, and has a vanadium content of 1-15% as V2O5.

[0022] Preferably, in step S1, the high temperature calcination temperature is 700-1000°C, the constant temperature time is 10-300 min, and the heating rate is 2-10°C / min.

[0023] Preferably, in step S2, leaching is performed at a liquid-to-solid ratio of deionized water to roasted clinker of 2 to 20:1 (mL / g), at a temperature of 80 to 100°C, and for 60 to 180 minutes. The acid solution is at least one of sulfuric acid, hydrochloric acid, or nitric acid. The leached residue is repeatedly rinsed with water having a pH lower than 1.8 and a temperature of 20 to 100°C. Furthermore, leaching can be performed in a water bath or a heating jacket.

[0024] Preferably, in step S4, the manganese precipitate is obtained by carbonization precipitation or direct precipitation. The carbonization precipitation method is to add at least one of carbon dioxide gas, sodium carbonate solution, and ammonium carbonate solution to the manganese-containing supernatant, wherein the carbon dioxide is pure carbon dioxide gas or industrial carbon dioxide, or carbon dioxide-containing flue gas released when vanadium slag is mixed with calcium salt and manganese salt and roasted; during the carbonization process, the pH value of the carbonization system is maintained at 6 to 8 by adding an alkaline solution; the alkaline solution is at least one of sodium hydroxide solution, ammonia water, sodium carbonate, and ammonium carbonate solution; the carbonization temperature is 20 to 100° C.; the carbonization time is 10 to 120 minutes; the manganese salt obtained by carbonization precipitation is manganese carbonate; the direct precipitation method is to add an alkaline substance to the manganese-containing supernatant to maintain the pH value of the precipitation system at 7 to 10, the alkaline substance is at least one of sodium hydroxide solution and ammonia water; the direct precipitation temperature is 20 to 100° C.; the precipitation time is 10 to 120 minutes; the manganese salt obtained by direct precipitation is trimanganese tetraoxide.

[0025] In step S4, the drying and crystallization method is to place the manganese-containing supernatant in at least one of a rotary evaporator, a vacuum drying oven, an oven, and a centrifuge, the drying temperature is 20 to 110° C., and the drying time is 12 to 48 hours; and the manganese salt obtained by drying and crystallizing is obtained.

[0026] In the above method, preferably, in step S5, leaching is performed at a solid-liquid ratio of deionized water to vanadium- and iron-containing leaching residue of 2 to 20:1 (mL / g), the leaching temperature is 30 to 100°C, and the leaching time is 10 to 100 minutes. The alkaline solution is at least one of sodium hydroxide solution, ammonia water, or hydrogen peroxide solution, and the iron-containing leaching residue is repeatedly rinsed with water having a pH of 7.0 to 14.0 and a temperature of room temperature to 100°C. Furthermore, the leaching can be performed in a water bath or a heating jacket.

[0027] Preferably, in step S6, the vanadium precipitation additive is at least one of ammonium sulfate, ammonium carbonate or ammonium bicarbonate; the vanadium precipitation additive is added in an amount such that the molar ratio of ammonium in the ammonium salt to vanadium in the leachate is 1.0 to 3.0; during the vanadium precipitation process, the pH value of the system is maintained at 2 to 3 by adding acid; the acid is at least one of sulfuric acid, hydrochloric acid or nitric acid; the vanadium precipitation temperature is 60 to 100° C., and the vanadium precipitation time is 20 to 180 minutes.

[0028] Preferably, in step S9, the impurities removed are ammonia nitrogen elements.

[0029] (3) Beneficial effects

[0030] The reaction mechanism of the present invention is:

[0031] The method provided by the present invention uses ferromanganese ore as raw material and vanadium industry intermediates or products as the roasting medium. By leveraging vanadium's excellent selective binding and directional reduction ability for manganese, the method reconstitutes the manganese minerals in the ferromanganese ore into acid-soluble vanadates without altering the iron mineral phase. Furthermore, through acid leaching and adjusting the leachate to different pH values, efficient iron and manganese separation and high recovery rates of the vanadium roasting medium are achieved, as well as the recycling of the vanadium roasting medium, effectively avoiding the drawbacks of existing iron and manganese separation methods in ferromanganese ore.

[0032] The advantages are as follows:

[0033] (1) The present invention limits the molar ratio of n(MnO2) / n(V2O5) to 0.5-3:1, reflecting the synergistic effect of vanadium and manganese. This is because from the low-temperature stage of roasting (600°C), manganese and vanadium combine to form a stable manganese pyrovanadate phase. The phase does not change when the roasting temperature is further increased. The manganese pyrovanadate phase is easily soluble in acid solution, which can achieve efficient extraction of manganese and vanadium, separation of manganese and iron, and separation efficiency of iron and manganese in ferromanganese ore. It is much higher than the existing ferromanganese ore separation method; it avoids the problem of iron and manganese being leached together during ferromanganese ore leaching, and reduces the subsequent process of removing iron from manganese;

[0034] (2) The inventive method achieves a high recovery rate of vanadium roasting medium by adjusting the leaching solution to different pH values ​​and combining the characteristics of vanadium, and the separated vanadium precipitate can be directly used to prepare a high-purity vanadium pentoxide product;

[0035] (3) The method of the present invention does not release toxic gases during the roasting process;

[0036] (4) The iron-containing tailings of the present invention enrich the iron in the ferromanganese ore into one phase, and its content reaches that of iron concentrate (TFe>60%), which can be directly used in blast furnace smelting of various ferroalloys;

[0037] (5) The addition of vanadium industry intermediates or vanadium products reduces the high roasting reaction temperature during the separation of iron and manganese in traditional ferromanganese ore, further saving costs and improving the recovery rate of the valuable components of iron and manganese in ferromanganese ore;

[0038] (6) The intermediate products or vanadium products used as additives in the vanadium industry can be recycled. BRIEF DESCRIPTION OF THE DRAWINGS

[0039] Figure 1 Schematic diagram of the process of the present invention. DETAILED DESCRIPTION

[0040] In order to better explain the present invention and facilitate understanding, the present invention is described in detail below through specific implementation methods.

[0041] The present invention innovatively uses vanadium industry intermediates or products as additives and roasts them together with ferromanganese ore. Vanadium and manganese combine at a low temperature (600°C) to form acid-soluble manganese pyrovanadate. Further increasing the roasting temperature results in a relatively stable phase, maintaining the manganese pyrovanadate, and enabling the separation and extraction of iron and manganese during subsequent leaching. Manganese and vanadium act synergistically in this process, neither of which can be dispensable. The manganese in the ferromanganese ore is calculated as MnO2, and the vanadium in the vanadium industry intermediates or products is calculated as V2O5, with a molar ratio of n(MnO2) / n(V2O5) of 0.5 to 3:1. If n(MnO2) / n(V2O5) is less than 0.5, the vanadium will not react sufficiently with the elements in the ferromanganese ore, resulting in the formation of Fe4Mn3(VO4)6. If n(MnO2) / n(V2O5) is greater than 3, there will be too little vanadium, which is insufficient to completely react with the manganese in the ferromanganese ore to form manganese pyrovanadate. The remaining manganese will combine with iron to form FeMnO3.

[0042] Example 1

[0043] A method for efficiently separating iron and manganese components in iron-manganese ore based on step-by-step leaching, such as Figure 1 As shown, the following steps are included:

[0044] (1) Calcination: The ferromanganese ore is ground and sieved to obtain a ferromanganese ore with a particle size of 200-250 mesh. Vanadium pentoxide is added to the mixture and mixed. The mixture is then calcined. The particle size of the vanadium pentoxide is required to be 200-250 mesh so that the vanadium in the vanadium pentoxide can further combine with the ferromanganese ore. The manganese content in the ferromanganese ore is 26.97% as MnO2, and the vanadium content in the vanadium pentoxide is 56.02% as V2O5. The ferromanganese ore and vanadium pentoxide are mixed in a molar ratio of n(MnO2) / n(V2O5) of 2.5. The mixed mixture is calcined at 850°C for 120 minutes at a heating rate of 2° / min to obtain a calcined clinker. The calcined clinker is crushed to a particle size of 250-300 mesh.

[0045] (2) Single leaching: The leaching process is carried out in a water bath. Deionized water is added to the leaching container at a liquid-solid ratio of deionized water to roasted clinker of 5 mL / g. When heated to 70°C, roasted clinker is added and stirring is started. The pH value of the leaching system is controlled to be around 2.2 by adding sulfuric acid with a volume fraction of 20%. Both leaching processes are carried out in the same device. In the single leaching, a sulfuric acid system with a pH value of 2.2 is selected as the leaching medium to achieve the dissolution of manganese vanadate in the roasted clinker. The first stage leaching reaction ends after a stable reaction of 60 minutes. No separation operation is performed in this step to ensure the effective separation of manganese and iron;

[0046] (3) Secondary leaching: Continue to add 20% volume fraction of sulfuric acid solution to the system to adjust the pH value of the leaching system to 1.8, so as to achieve hydrolysis and precipitation of vanadium ions in the vanadium-manganese leachate, as shown in reaction equation (1). After 60 minutes of stable reaction, the second stage leaching reaction ends, and then the solid-liquid separation is carried out by filtration to obtain vanadium-containing iron-containing leaching residue and manganese-containing leachate. The leaching residue is repeatedly rinsed with hot water with a pH value of 1.8 and a temperature of 80°C. The leaching rate is calculated according to formula (2). After leaching, by measuring the content of each element in the leachate, it is found that the leaching rate of manganese is 85.97%, while the leaching rate of iron is only 0.00492%, and the leaching rate of vanadium is 5.74%. That is, the leachate is a manganese-containing leachate, and vanadium and iron are enriched in the leachate residue, achieving efficient separation of iron and manganese.

[0047] 2(VO2) + + H2O = V2O5 ↓ + 2 H + (1)

[0048] Leaching rate (%) = (m L / m0)×100(2)

[0049] Where: m is Mn, Fe or V; m L and m0 is the total mass of m in the leachate and the total mass of m in the roasted sample.

[0050] (4) Recovering manganese: carbon dioxide was introduced into the manganese-containing supernatant obtained in step (3) for carbonization at a gas flow rate of 0.1 L / min; during the carbonization process, the pH value of the carbonization system was maintained at 6.8 by adding 5 g / L sodium hydroxide solution; the carbonization temperature was 40° C.; the carbonization time was 60 min; and then the solid and liquid were separated by filtration to obtain a manganese carbonate precipitate and a manganese extraction waste liquid.

[0051] Alternatively, the manganese-containing supernatant obtained in step (3) is electrolyzed to obtain an electrolytic manganese product with a manganese content of >99.7%.

[0052] (5) Separation of vanadium and iron: The process is carried out in a water bath. Deionized water is added to the leaching container at a solid-liquid ratio of 5 mL / g to the vanadium-containing and iron-containing leaching residue. When the leaching residue is heated to 80°C, the vanadium-containing and iron-containing leaching residue is added. The pH value of the leaching system is controlled to be 14.0 by adding a hydrogen peroxide solution with a volume fraction of 20%. After leaching for 60 minutes at a leaching temperature of 70°C, the leaching residue is filtered for solid-liquid separation to obtain a vanadium-containing leaching solution and an iron-containing leaching residue. The leaching residue is repeatedly rinsed with hot water at a pH of 14.0 and a temperature of 70°C to achieve separation of vanadium and iron.

[0053] (6) Vanadium precipitation: The vanadium precipitation process is carried out in a water bath, and ammonium sulfate is used as a vanadium precipitation additive. The amount of vanadium precipitation additive added is based on a molar ratio of ammonium in the ammonium salt to vanadium in the leachate of 1.5. The calculation basis for the amount of ammonium salt supplementation is shown in equation (3). The vanadium precipitation process is carried out by adding a sulfuric acid solution with a volume fraction of 10% to maintain the system pH value at 2. After the precipitation temperature is 95°C and the precipitation is carried out for 60 minutes, vanadium is precipitated, and the solid and liquid are filtered to separate to obtain ammonium polyvanadate precipitate and vanadium extraction waste liquid. After vanadium precipitation, the vanadium content in the vanadium extraction waste liquid is measured, and the vanadium precipitation rate is obtained to be 99.28%. Among them, the vanadium precipitation rate is calculated according to the following formula: Vanadium precipitation rate = (mass of vanadium in the leachate - mass of vanadium in the manganese-containing supernatant) / mass of vanadium in the leachate. The vanadium circulation rate is more than 90%.

[0054] NH 4+ + H2O = NH3·H2O + H + (3)

[0055] (7) Obtaining a vanadium product: The ammonium polyvanadate precipitate is dried at 120°C for 12 hours, and then heated to 550°C for calcination for 100 minutes. The resulting vanadium pentoxide has a purity of 99.11% as measured by chemical analysis. The vanadium pentoxide can be output as a vanadium pentoxide product or returned to the calcination system of step (1) for the next round of calcination.

[0056] (8) Utilization of iron-containing tailings: The iron-containing leaching slag is iron concentrate (TFe>60%), which is directly put into blast furnace smelting. The smelting atmosphere and temperature are controlled to obtain high-quality ferroalloy.

[0057] (9) Impurity Removal: After impurities are removed from the manganese and vanadium waste liquids, they are recycled back into the leaching system as the leaching medium in step S2. The impurities to be removed are primarily ammonia nitrogen. Ammonia nitrogen is removed by chemical precipitation and, after impurities are removed, recycled back into the leaching system as the leaching medium for repeated use.

[0058] Example 2

[0059] A method for efficiently separating iron and manganese components in iron-manganese ore based on step-by-step leaching, such as Figure 1 As shown, the following steps are included:

[0060] (1) Calcination: The ferromanganese ore is ground and sieved to obtain a ferromanganese ore with a particle size of 200-250 mesh. Ammonium polyvanadate is added to the mixture and mixed, and then the mixture is calcined. The particle size of the ammonium polyvanadate is required to be 200-250 mesh so that the vanadium in the ammonium polyvanadate can further combine with the ferromanganese ore. The manganese content in the ferromanganese ore is 26.97% as MnO2, and the vanadium content in the ammonium polyvanadate is 50.50% as V2O5. The ferromanganese ore and the ammonium polyvanadate are mixed at a molar ratio of n(MnO2) / n(V2O5) of 2.25. The mixed mixture is calcined at 750°C for 180 minutes at a heating rate of 5° / min to obtain a calcined clinker. The calcined clinker is crushed to a particle size of 250-300 mesh.

[0061] (2) Primary leaching: The leaching process is carried out in a water bath. Deionized water is added to the leaching container at a liquid-solid ratio of deionized water to roasted clinker of 10 mL / g. When heated to 80°C, roasted clinker is added. The pH value of the leaching system is controlled to be around 2.5 by adding hydrochloric acid with a volume fraction of 50%. Both leaching stages are carried out in the same device. In the primary leaching, a hydrochloric acid system with a pH value of 2.5 is selected as the leaching medium to achieve the dissolution of manganese vanadate in the roasted clinker. The first stage leaching reaction ends after a stable reaction of 60 minutes. No separation operation is performed in this step to ensure the effective separation of manganese and iron.

[0062] (3) Secondary leaching: Continue to add 50% hydrochloric acid solution to the system to adjust the pH value of the leaching system to 1.7 to achieve hydrolysis and precipitation of vanadium ions in the vanadium-manganese leachate. After 60 minutes of stable reaction, the second stage leaching reaction ends. Filter to separate the solid and liquid to obtain vanadium-containing and iron-containing leaching residue and manganese-containing leachate. Repeatedly rinse the leaching residue with hot water at a pH of 1.7 and a temperature of 80°C. After leaching, the iron, manganese and vanadium contents in the manganese-containing leachate were measured, and the manganese leaching rate was 85.96%, while the iron leaching rate was only 0.00685%, and the vanadium leaching rate was only 5.22%, achieving efficient separation of iron and manganese.

[0063] (4) Recovering manganese: adding ammonia water to the manganese-containing leachate obtained in step (3) to precipitate manganese ions, maintaining the pH value of the precipitation system at 8.0; the precipitation temperature is 60° C.; the precipitation time is 30 min; then filtering and separating the solid and liquid to obtain manganese tetraoxide precipitate and manganese extraction waste liquid.

[0064] (5) Separation of vanadium and iron: The process is carried out in a water bath. Deionized water is added to the leaching container at a solid-liquid ratio of deionized water to vanadium-containing and iron-containing leaching residue of 10 mL / g. When the leaching residue is heated to 60°C, the vanadium-containing and iron-containing leaching residue is added. The pH value of the leaching system is controlled to 12.0 by adding a sodium hydroxide solution with a volume fraction of 20%. After leaching for 120 minutes at a leaching temperature of 60°C, the leaching residue is filtered for solid-liquid separation to obtain a vanadium-containing leaching solution and an iron-containing leaching residue. The leaching residue is repeatedly rinsed with hot water at a pH of 12.0 and a temperature of 60°C to achieve separation of vanadium and iron.

[0065] (6) Vanadium precipitation: The vanadium precipitation process is carried out in a water bath, and ammonium sulfate is used as a vanadium precipitation additive. The amount of vanadium precipitation additive is adjusted to a molar ratio of 1.3 between the ammonium in the ammonium salt and the vanadium in the leachate. The pH value of the system is maintained at 2 by adding a 20% by volume sulfuric acid solution during the vanadium precipitation process. The precipitation temperature is 100°C, and after 40 minutes of precipitation, vanadium is precipitated. The solid and liquid are filtered to separate the polyvanadate precipitate and the vanadium extraction waste liquid. After vanadium precipitation, the vanadium content in the vanadium extraction waste liquid is measured, and the vanadium precipitation rate is 98.36%. The purity is measured by chemical analysis and is 98.76%. The product can be output as polyvanadate product or returned to the roasting system in step (1) to participate in the next round of roasting process. The vanadium recycling rate reaches over 90%.

[0066] (7) Obtaining vanadium product: The ammonium polyvanadate precipitate is dried at 120°C for 12 hours, and then heated to 550°C and calcined for 120 minutes. The obtained vanadium pentoxide has a purity of 99.23% as measured by chemical analysis. The vanadium pentoxide can be output as a vanadium pentoxide product or returned to the calcination system of step (1) to participate in the next round of calcination.

[0067] (8) Utilization of iron-containing tailings: The iron-containing leaching slag is iron concentrate (TFe>60%), which is directly put into blast furnace smelting. The smelting atmosphere and temperature are controlled to obtain high-quality ferroalloy.

[0068] (9) Impurity Removal: The impurities that need to be removed from the obtained manganese extraction waste liquid and vanadium extraction waste liquid are mainly ammonia nitrogen. The ammonia nitrogen is removed by conventional chemical precipitation method, and after impurity removal, it is recycled back to the leaching system as the leaching medium in step (2).

[0069] Example 3

[0070] A method for efficiently separating iron and manganese components in iron-manganese ore based on step-by-step leaching, such as Figure 1 As shown, the following steps are included:

[0071] (1) Calcination: The ferromanganese ore is ground and sieved to obtain a ferromanganese ore with a particle size of 200-250 mesh. Vanadium slag is added for mixing, and then the mixture is roasted. The particle size of the vanadium slag is required to be 200-250 mesh so that the vanadium in the vanadium slag can further combine with the ferromanganese ore. The manganese content in the ferromanganese ore is 26.97% as MnO2, and the vanadium content in the vanadium slag is 18.22% as V2O5. The ferromanganese ore and vanadium pentoxide are mixed in a molar ratio of n(MnO2) / n(V2O5) of 1.0. The mixed mixture is roasted at 1000°C for 120 minutes at a heating rate of 10° / min to obtain a roasted clinker. The roasted clinker is crushed to a particle size of 250-300 mesh.

[0072] (2) Primary leaching: The leaching process is carried out in a heating jacket. Deionized water is added to the leaching container at a liquid-solid ratio of deionized water to roasted clinker of 3 mL / g. Roasted clinker is added at 90°C and stirring is started. The pH value of the leaching system is controlled to about 2.8 by adding nitric acid with a volume fraction of 30%. For the primary leaching, a nitric acid system with a pH value of 2.8 is selected as the leaching medium to achieve the dissolution of manganese vanadate in the roasted clinker. The first stage of the leaching reaction ends after a stable reaction of 60 minutes. No separation operation is performed in this step to ensure the effective separation of manganese and iron;

[0073] (3) Secondary leaching: Continue to add 30% volume fraction of nitric acid solution to the system to adjust the pH value of the leaching system to 1.7 to achieve hydrolysis and precipitation of vanadium ions in the vanadium-manganese leaching solution. After 60 minutes of stable reaction, the second stage leaching reaction is completed. Filter to separate the solid and liquid to obtain vanadium-containing iron-containing leaching slag and manganese-containing leaching solution. Repeatedly rinse the leaching slag with hot water at a pH of 1.7 and a temperature of 90°C. After leaching, the content of iron, manganese and vanadium in the manganese-containing leaching solution was measured, and the manganese leaching rate was 170.68%, while the iron leaching rate was only 0.00412%, and the vanadium leaching rate was only 6.54%, achieving efficient separation of iron and manganese. In addition, 7.2% of the manganese in the original vanadium slag was transferred to the manganese-containing and vanadium-containing leaching solution after composite roasting and acid leaching. This part of manganese also combines with vanadium in the vanadium slag to form manganese pyrovanadate, and enters the solution through leaching, thereby achieving the recovery and utilization of part of the manganese in the vanadium slag.

[0074] (4) Recovering manganese: adding 20% ​​sodium hydroxide solution to the manganese-containing leachate obtained in step (3) to precipitate manganese ions, maintaining the pH value of the precipitation system at about 7.0; the precipitation temperature is 40° C.; the precipitation time is 120 minutes; and the solid and liquid are separated by filtration to obtain manganese carbonate precipitate and vanadium extraction waste liquid.

[0075] Manganese salt utilization: Manganese carbonate precipitate is oxidatively roasted at 400°C for 12 hours, during which oxygen is introduced at a flow rate of 1 L / min to produce a crude manganese dioxide product. This product is then placed in a 15% by volume sulfuric acid solution and a sodium chlorate solution to react for 120 minutes to obtain refined manganese dioxide with a manganese dioxide content of 91.98%. This product can be used as an industrial oxidant and catalyst.

[0076] (5) Separation of vanadium and iron: The process is carried out in a water bath. Deionized water is added to the leaching container at a solid-liquid ratio of 3 mL / g to the vanadium-containing and iron-containing leaching residue. When the leaching residue is heated to 90°C, the vanadium-containing and iron-containing leaching residue is added. The pH value of the leaching system is controlled to 13.0 by adding an ammonia solution with a volume fraction of 50%. After leaching for 60 minutes at a leaching temperature of 70°C, the leaching residue is filtered for solid-liquid separation to obtain a vanadium-containing leaching solution and an iron-containing leaching residue. The leaching residue is repeatedly rinsed with hot water at a pH of 13.0 and a temperature of 90°C to achieve separation of vanadium and iron.

[0077] (6) Vanadium precipitation: The vanadium precipitation process is carried out in a water bath, and ammonium sulfate is used as a vanadium precipitation additive. The amount of vanadium precipitation additive added is based on a molar ratio of ammonium in the ammonium salt to vanadium in the leachate of 2.0. The pH value of the system is maintained at 2.2 by adding a 50% by volume sulfuric acid solution during the vanadium precipitation process. After precipitation at a temperature of 90°C and precipitation for 90 minutes, vanadium is precipitated, and the solid and liquid are separated by filtration to obtain ammonium polyvanadate precipitate and vanadium extraction waste liquid. After vanadium precipitation, the vanadium content in the vanadium extraction waste liquid is measured, and the vanadium precipitation rate is obtained to be 97.35%. The purity is measured by chemical analysis to be 97.54%. The product can be output as ammonium polyvanadate product or returned to the roasting system of step (1) to participate in the next round of roasting process.

[0078] (7) Obtaining vanadium product: The ammonium polyvanadate precipitate is dried at 120°C for 12 hours, and then heated to 550°C and roasted for 150 minutes. The obtained vanadium pentoxide has a purity of 99.47% as measured by chemical analysis. The vanadium pentoxide can be output as a vanadium pentoxide product or returned to the roasting system of step (1) to participate in the next round of roasting process.

[0079] (8) Utilization of iron-containing tailings: The iron-containing leaching slag is iron concentrate (TFe>60%), which is directly put into blast furnace smelting. The smelting atmosphere and temperature are controlled to obtain high-quality ferroalloy.

[0080] (9) Impurity Removal: The impurities that need to be removed from the obtained manganese extraction waste liquid and vanadium extraction waste liquid are mainly ammonia nitrogen. The ammonia nitrogen is removed by chemical precipitation and recycled back to the leaching system as the leaching medium in step (2) after impurity removal.

[0081] Comparative Example 1

[0082] When a conventional sulfate roasting-water leaching method is used to separate iron and manganese from low-grade ferromanganese ore (TFe: 30.35%, Mn: 14.30%), not only SO2 but also a large amount of N2 needs to be introduced during the roasting process, resulting in a large amount of gas waste; during leaching, iron and manganese are inevitably leached together, which brings a lot of difficulties to the subsequent removal of iron from manganese.

[0083] Comparative Example 2

[0084] Endophytic fungi were used to bioleach ferromanganese ore (Fe2O3: 67.90%, MnO2: 8.79%). This method required a 20-day experiment to separate the iron and manganese from the ferromanganese ore, but the manganese leaching rate was only 79%, far lower than the manganese leaching rate of the present invention. However, different ferromanganese ores require different bacterial species, and the entire leaching process is slow, which greatly limits this method.

[0085] Comparative Example 3

[0086] A high-iron manganese oxide ore (TFe: 21.28%, Mn: 39.05%) is treated with 7 wt% Na2S2O3 as a reducing agent and subjected to reduction roasting and magnetic separation to separate the iron and manganese from the ore. This method requires not only high energy consumption at 1100°C but also requires a shaft furnace, resulting in a complex process and high production and maintenance costs. The manganese recovery rate is only 85.96%, far lower than that of the present invention.

[0087] The above description is merely a preferred embodiment of the present invention and is not intended to limit the present invention in any other manner. Any person skilled in the art may utilize the above-disclosed technical content to modify or modify the above-disclosed embodiments into equivalent embodiments. However, any simple modifications, equivalent variations, and modifications to the above-disclosed embodiments that do not depart from the technical content of the present invention and are based on the technical essence of the present invention shall remain within the scope of protection of the present invention.

Claims

1. A method for efficiently separating iron and manganese components in iron-manganese ore based on step-by-step leaching, characterized in that: It includes the following steps: S1. Roasting: Mix the ferromanganese ore with the intermediate product of the vanadium industry or the vanadium product to obtain a mixed material. The mixing ratio of the ferromanganese ore and the intermediate product of the vanadium industry or the vanadium product is n (MnO2) / n (V2O5) calculation, n (MnO2) / n (V2O5) is 0.5~3:1; the mixed material is roasted at high temperature to obtain roasted clinker; the ferromanganese ore is a manganese ore with a Mn / Fe mass ratio of <3; S2, primary leaching: leaching the roasted clinker obtained in step S1, and adjusting the pH value of the leaching system to 2-3 using an acid solution; S3, secondary leaching: using sulfuric acid solution to continue to adjust the pH value to below 1.8, leaching and filtering to obtain manganese-containing leachate and vanadium-containing and iron-containing leach residue; both leaching processes are carried out in the same device; S4, recovering manganese: precipitating or drying and crystallizing the manganese-containing leachate obtained in step S3, and filtering to obtain a manganese product and a manganese extraction waste liquid; S5, separating vanadium and iron: leaching the vanadium-containing and iron-containing leaching residue obtained in step S3, controlling the pH value of the leaching system to 7.0-14.0 using an alkaline solution, and filtering after leaching to obtain a vanadium-containing leachate and an iron-containing leaching residue; S6, recovering vanadium: adding a vanadium precipitation additive to the vanadium-containing leachate obtained in step S5, and adjusting the pH value to 2-3 to precipitate vanadium, and filtering after precipitation to obtain a vanadium industry intermediate product or vanadium product and vanadium extraction waste liquid; S7. Utilization of the intermediate product or vanadium product in the vanadium industry: The intermediate product or vanadium product obtained in step S6 is output as a product, or recycled back into the roasting system as an additive to participate in the next round of roasting; S8, Utilization of Iron-Containing Tailings: The iron-containing leached slag obtained in step S5 is used as feed material for a blast furnace, or as a raw material for preparing ferroalloys; S9, waste liquid recycling: After impurities are removed from the manganese extraction waste liquid and vanadium extraction waste liquid obtained in steps S4 and S6, they are recycled back to the leaching system as the leaching medium in step S2.

2. The method according to claim 1, wherein In step S1, the ferromanganese ore is at least one of an oxide mineral, a hydrogenated mineral, a carbonate mineral, a sulfide mineral, a silicate mineral and a borate mineral; and has a particle size of 200 to 300 meshes.

3. The method according to claim 1, wherein In step S1, the vanadium industry intermediate product or vanadium product includes at least one of vanadate, vanadium oxide, and vanadium slag.

4. The method according to claim 1, wherein In step S1 , the high temperature calcination temperature is 700-1000° C., the constant temperature time is 10-300 min, and the heating rate is 2-10° C. / min.

5. The method according to claim 1, wherein In step S2, leaching is performed at a liquid-to-solid ratio of deionized water to roasted clinker of 2 to 20:1 in mL / g. The leaching temperature is 80 to 100° C. and the leaching time is 60 to 180 min.

6. The method according to claim 1, wherein In step S4, the precipitation is carried out by carbonization precipitation or direct precipitation. The carbonization precipitation method is to add at least one of carbon dioxide gas, sodium carbonate solution, and ammonium carbonate solution to the manganese-containing supernatant, and to maintain the pH value of the carbonization system at 6 to 8 by adding an alkaline solution during the carbonization process. The carbonization temperature is 20 to 100° C., and the carbonization time is 10 to 120 minutes. The manganese salt obtained by carbonization precipitation is manganese carbonate. The direct precipitation method is to add an alkaline substance to the manganese-containing supernatant to maintain the pH value of the precipitation system at 7 to 10. The direct precipitation temperature is 20 to 100° C., and the precipitation time is 10 to 120 minutes. The manganese salt obtained by direct precipitation is trimanganese tetraoxide. The drying and crystallization is performed by placing the manganese-containing supernatant in at least one of a rotary evaporator, a vacuum drying oven, an oven, and a centrifuge, the drying temperature being 20-110° C. and the drying time being 12-48 hours; and drying and crystallizing to obtain the manganese salt.

7. The method according to claim 1, wherein In step S5, leaching is performed at a solid-liquid ratio of deionized water to vanadium-containing and iron-containing leaching residue of 2-20:1 in mL / g, the leaching temperature is 30-100° C., and the leaching time is 10-100 min.

8. The method according to claim 2, wherein In step S6, the vanadium precipitation additive is an ammonium salt, which is at least one of ammonium sulfate, ammonium carbonate, or ammonium bicarbonate; the vanadium precipitation additive is added in an amount such that the molar ratio of ammonium in the ammonium salt to vanadium in the leachate is 1.0 to 3.0; during the vanadium precipitation process, the pH value of the system is maintained at 2 to 3 by adding acid; the vanadium precipitation temperature is 60 to 100° C., and the vanadium precipitation time is 20 to 180 minutes.

9. The method according to claim 1, wherein In step S9, the impurities removed are ammonia and nitrogen elements.

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