Selective directional conversion and reduction method of iron alum slag in low-temperature molten salt

Abstract

The present invention belongs to the field of iron alum slag reduction and comprehensive utilization of solid waste in the smelting and recycling industries, and specifically provides a molten salt conversion method for an iron alum phase, wherein an iron alum phase having a chemical formula of MFe3(SO4)2(OH)6 is mixed with molten salt and then subjected to molten salt roasting at a temperature of 200-450°C to convert the iron alum phase into M3Fe(SO4)3; wherein M is Na, K, NH4 + At least one of the following: the molten salt is in a molten salt state at the roasting temperature, and the molar ratio of the molten salt to the iron vitriol phase is greater than 5. The present invention also provides a method for utilizing slag containing an iron-vanadium phase, such as a typical iron vitriol slag. The method of the present invention can achieve gentle and highly selective conversion of the iron-vanadium phase.

Description

Technical Field

[0001] The invention belongs to the technical field of comprehensive recycling of solid waste, and particularly relates to a method for reducing iron alum slag. Technical Background

[0002] The yellow potassium ferroalloy method is often used to remove iron in the hydrometallurgical smelting of metals such as zinc, copper, nickel, cobalt, and manganese, as well as in battery recycling. This produces a large amount of potassium ferroalloy slag, which mainly contains iron-vanadium phase. Its properties are stable, and it is difficult to reduce and recycle it.

[0003] The main existing methods for disposing of iron alum residue include pyrolysis, wet separation, flotation, and magnetic separation. For example, Chinese patent publication CN113930621A reports on placing iron alum residue in an oxalic acid solution for agitation, followed by solid-liquid separation to obtain a reaction solution and residual residue; the reaction solution is then irradiated with sunlight, followed by solid-liquid separation to obtain ferrous oxalate. Another example is Chinese patent publication CN114242977A, which discloses a method for preparing high-performance Fe3O4 / ZnFe2O4 / C composite negative electrode materials using sucrose-assisted hydrochloric acid leaching of iron alum residue.

[0004] Although the existing treatment methods can achieve the treatment of iron alum slag to a certain extent, there are still some technical problems that need to be overcome, such as: the main problem of SO2 emission pollution in pyrometallurgical treatment is the wet process, which has lower energy consumption and better selectivity than the pyrometallurgical process, but there is wastewater such as leachate that needs to be treated; the mineral processing process is simple and low in cost, but the flotation method mainly targets the silver in the iron alum slag, and there is still a large amount of tailings, and the iron recovered by magnetic separation is difficult to meet the smelting requirements; the cost of production materials is high; and although the iron alum slag is safely disposed of in the solidification treatment, the valuable elements are not reasonably recovered.

[0005] The comprehensive recycling of ferroalloy residue requires consideration not only of its properties but also of the environmental impact of the treatment process, minimizing SO2 and wastewater emissions, avoiding heavy metal contamination, and reducing the amount of waste residue stored. Current research focuses on recycling ferroalloy residue without causing secondary pollution, comprehensively considering resource, environmental, and economic considerations. Summary of the Invention

[0006] The first purpose of the present invention is to provide a molten salt conversion method for the iron alum phase, which aims to efficiently convert the iron alum phase, which is stable in nature and difficult to be water-soaked, into M3Fe(SO4)3, thereby reducing the difficulty of its recovery.

[0007] The second object of the present invention is to provide a method for treating ferroalite slag, aiming to utilize the conversion method to highly selectively convert the ferroalite phase in the ferroalite slag.

[0008] A molten salt conversion method for a ferroalite phase comprises mixing a ferroalite phase having a chemical formula of MFe3(SO4)2(OH)6 with a molten salt and then subjecting the mixture to a molten salt roasting treatment at a temperature of 200 to 450° C. to convert the ferroalite phase into M3Fe(SO4)3;

[0009] The M is at least one of Na, K, and NH4+;

[0010] The molten salt is in a molten salt state at the roasting temperature, and the molar ratio of the molten salt to the iron alum phase is above 5.

[0011] The present invention innovatively subjects the ferroalite phase to molten salt roasting treatment, and further coordinates the roasting temperature and the molten salt ratio with the joint control, so as to achieve synergy. Based on the solid-liquid reaction idea, the ferroalite phase can be efficiently and selectively converted into the M3Fe(SO4)3 phase. The converted phase has excellent water immersion behavior, so that the water immersion and efficient recycling of the ferroalite phase can be unexpectedly achieved.

[0012] In the present invention, the molten salt roasting method and the combined control of the amount of molten salt and the roasting temperature are the keys to synergistically improving the selective conversion of the iron alum phase into M3Fe(SO4)3.

[0013] In the present invention, the molten salt may be in a molten liquid state at the roasting temperature, and is preferably at least one salt selected from sodium bisulfate, potassium bisulfate, ammonium sulfate, and ammonium bisulfate, or a hydrate thereof, preferably sodium bisulfate and a hydrate thereof (salt with crystal water).

[0014] In the present invention, the amount of the molten salt is not less than the required limit of the present invention. Theoretically, further increasing the amount of molten salt will improve the molten salt roasting effect to a certain extent, but considering the preparation cost, the molar ratio of the molten salt to the iron alum phase is 5.5 to 10:1; further, it can be 6 to 10:1, and further, it can be 6 to 7:1.

[0015] In the present invention, the atmosphere in the calcination stage is an atmosphere containing at least one of oxygen, nitrogen, and an inert gas; considering the simplicity and cost of the process, the atmosphere in the calcination stage can be air.

[0016] Preferably, the calcination temperature is 250-400° C.; at the preferred temperature, highly selective molten salt conversion of the iron-vanadium phase can be achieved, and the release of sulfur oxides during the treatment process can be effectively avoided.

[0017] In the present invention, the calcination time is more than 1 hour, and can be further extended to 2 to 5 hours in consideration of the preparation efficiency.

[0018] As the same inventive concept, the present invention also provides a molten salt conversion-water leaching recovery method of iron alum phase, wherein the calcined material is subjected to water leaching to obtain Fe3+ The water soak is, for example, a ferric sulfate solution.

[0019] As the same inventive concept, the present invention also provides a solution for the selective conversion of the ferroalloy phase in a material containing ferroalloy phase components. For example, the present invention also provides a method for treating ferroalloy slag, which adopts a molten salt conversion method for the ferroalloy phase to perform molten salt roasting on the ferroalloy slag, and selectively convert the ferroalloy phase in the ferroalloy slag into M3Fe(SO4)3. That is, after the ferroalloy slag and the molten salt described in the present invention are mixed, the molten salt roasting is performed at a temperature of 200 to 450°C to convert the ferroalloy phase in the ferroalloy slag into M3Fe(SO4)3. The present invention has found that by innovatively combining the ferroalloy slag and the molten salt, under the described molten salt dosage and roasting temperature, the ferroalloy phase in the ferroalloy slag can be selectively converted into M3Fe(SO4)3 with high selectivity.

[0020] In the method for treating ferroalloy slag of the present invention, the molar ratio of the molten salt to the ferroalloy phase in the ferroalloy slag is above 5:1, and can further be 5.5-10:1; further can be 6-10:1, and further can be 6-7:1.

[0021] In the method for treating iron alum slag of the present invention, the atmosphere during the molten salt roasting stage is an atmosphere containing at least one of oxygen, nitrogen, and an inert gas. The roasting temperature is preferably 250 to 400° C., and the roasting time is preferably 2 to 5 hours.

[0022] The method for treating iron alum slag of the present invention is to subject the roasted material treated by molten salt roasting to water leaching to obtain water leaching slag and iron alum enriched with iron 3+ In the present invention, thanks to the selective conversion method of the ferrovanadium phase, the transformation extraction of the ferrovanadium phase can be achieved based on a simple water leaching treatment, and the efficient reduction of the iron alum slag and the efficient enrichment of the water-insoluble elements can be achieved.

[0023] In the present invention, there are no specific requirements for the water immersion process. For example, the liquid-to-solid ratio during the water immersion stage is 5-50 ml / g, and can be further increased to 15-25 ml / g for efficiency and cost considerations. The temperature during the water immersion stage is below 60°C, and can further be 25-60°C. The water immersion time can be at least 0.5 hours, and can further be 1-3 hours.

[0024] A typical iron alum slag treatment scheme of the present invention is that the iron alum slag also contains M element, and the M element includes at least one of zinc, manganese, nickel, cobalt, and copper; the iron alum slag is subjected to molten salt conversion treatment and water immersion treatment, and the M element therein is enriched in the water immersion liquid.

[0025] A typical iron alum slag treatment scheme of the present invention also contains N element, and the N element includes at least one of lead, silver, silicon and calcium; the iron alum slag is subjected to molten salt conversion treatment and water leaching treatment, and the N element is enriched in the water-leached slag.

[0026] The present invention provides a typical iron alum slag treatment scheme. When the iron alum slag also contains M element and N element, the iron alum slag is subjected to molten salt conversion treatment and water immersion treatment, wherein the M element is enriched in the water immersion liquid and the N element is enriched in the water immersion slag.

[0027] In the present invention, the M element and the N element in the iron alum slag are mainly present in the form of sulfate phase. The silicon is mainly present in the form of oxide.

[0028] Beneficial effects:

[0029] 1. The present invention innovatively subjects the ferroalite phase to molten salt roasting, further combining the roasting temperature with the combined control of the molten salt ratio. This allows for a synergistic, efficient, and highly selective conversion of the ferroalite phase into the M3Fe(SO4)3 phase based on a solid-liquid reaction. The converted phase exhibits excellent water-leaching behavior, thus unexpectedly enabling efficient water-leaching recovery of the ferroalite phase. Furthermore, the treatment process is mild, and no sulfur-oxidizing atmosphere is generated.

[0030] 2. The present invention innovatively performs the molten salt roasting treatment on the iron alum slag, and further cooperates with the joint control of the molten salt dosage and temperature to achieve synergy, and can highly selectively convert the iron alum phase in the iron alum slag into the M3Fe(SO4)3 phase, and then utilize it for water leaching. Not only that, it is also beneficial to the accompanying leaching of the M element that may be present therein, and the efficient enrichment of the N element.

[0031] For example, in the present invention, efficient leaching of elements such as iron, zinc and manganese can be achieved through simple water immersion, and the enrichment of lead and silver in the slag can be achieved. At the same time, the reduction rate of the ferroalloy slag can ultimately reach 73%. The actual reduction rate can be determined based on the proportion of the ferroalloy phase in the ferroalloy slag. The reaction process has low energy consumption, does not produce gases such as sulfur dioxide and carbon dioxide, and is easy to apply industrially.

[0032] 3. The process of the present invention has good effects on metal extraction and enrichment. In addition, the energy consumption is low, and compared with conventional high-temperature pyrolysis reduction treatment, the energy consumption and carbon emissions can be greatly reduced.

[0033] The present invention can achieve large-scale reduction of iron alum slag, thereby eliminating environmental risks and realizing comprehensive utilization thereof. BRIEF DESCRIPTION OF THE DRAWINGS

[0034] Figure 1The phase analysis results of the leached residue before and after molten salt roasting of the iron alum slag and after water washing in Example 1 to 4 are shown, wherein the spectrum at 25°C is the XRD pattern of the leached residue before molten salt roasting of the iron alum slag, and the leached residue refers to the XRD pattern of the water-leached residue of Example 4;

[0035] Figure 2 Phase analysis results of the iron alum slag molten salt before and after calcination at 900°C in Comparative Example 1;

[0036] Figure 3 This is the appearance of the calcined product of Comparative Example 2;

[0037] Figure 4 This is the phase analysis result of the red product calcined in Comparative Example 2;

[0038] Figure 5 Phase analysis of the product of molten salt roasting of iron alum slag in comparative example 3-6; DETAILED DESCRIPTION

[0039] The iron alum slag described in the present invention can be any iron alum slag in the industry, for example, it can be zinc smelting, battery materials, etc., which contains sodium iron alum (sodium iron alum, NaFe3(SO4)2(OH)6), and is also allowed to contain phase components such as lead sulfate, manganese sulfate, and zinc sulfate. There is no special requirement for the content of the iron vanadium phase in the iron alum slag. Considering the recovery economy of the process, the content of the iron vanadium phase in the slag is preferably above 50wt.%. In the following case, as a typical example, the content of the iron vanadium phase is 70-75wt.%. For example, in the following cases, the iron alum slag is all the iron alum slag with the component content described in Table 1.

[0040] Table 1 XRF analysis results of elemental components in iron alum slag / atm%

[0041] O S Fe Zn Mn Cu Na Ca 47.053 12.513 20.354 4.076 1.998 0.196 2.538 1.343 Al K Mg Si Pb Cd Cr other 1.037 0.511 0.174 3.851 3.726 0.066 0.025 0.440

[0042] In the present invention, considering the convenience and cost of the process, the atmosphere in the roasting stage is typically air.

[0043] The scale of the treatment process of the present invention is not particularly limited, and can be, for example, a small test, a pilot test, or an actual production scale. In the following cases, unless otherwise stated, the small test scale (500 g to 1 kg) is used as a typical example.

[0044] In the present invention, the molten salt is sufficient to be liquefiable at the roasting temperature. For example, it can be at least one salt selected from sodium bisulfate, potassium bisulfate, ammonium sulfate, and ammonium bisulfate, or a hydrate thereof. In the present invention, there is no particular requirement for whether the molten salt contains water of crystallization. Considering the simplicity of the processing process, it can directly use commercial-scale components. For example, when the molten salt is sodium bisulfate, it can use sodium bisulfate monohydrate, which is commonly used in the industry.

[0045] In the present invention, in the following cases, the molar amount of the ferrovanadium slag is calculated based on the sodium ferrovanadium (NaFe3(SO4)2(OH)6) contained therein. In other words, the molar ratio of the molten salt to the ferrovanadium slag refers to the molar ratio of the sodium ferrovanadium in the molten salt to the ferrovanadium slag.

[0046] Example 1:

[0047] Sodium bisulfate monohydrate and ferroaluminum slag were uniformly mixed in a molar ratio of 6:1, and the mixed material was placed in a muffle furnace for calcination at a calcination temperature of 250°C (marked as T1) and a calcination time of 3 hours. Analysis of the physical phase of the calcined product showed that under these conditions, the sodium ferroaluminum phase in the ferroaluminum slag disappeared (see Table 1 for XRD), and was selectively converted into sodium ferric sulfate.

[0048] The roasted material was subjected to water leaching treatment, wherein the temperature of the water leaching stage was room temperature, the liquid-to-solid ratio was 15-20 ml / g, and the leaching time was 1 hour; wherein, the leaching rates of the main elements iron, zinc and manganese were all over 96%, while lead was enriched in the water leaching residue; the leaching residue phases were mainly lead sulfate and silicon dioxide phases.

[0049] Example 2:

[0050] Compared with Example 1, the only difference is that the calcination temperature T1 is 300° C., and other operations and parameters are the same as Example 1.

[0051] By analyzing the phase of the calcined product ( Figure 1 ) analysis showed that under this condition, the sodium ferroalite phase in the ferroalite slag disappeared, and its conversion product was mainly sodium ferric sulfate. The leaching rates of iron, zinc and manganese elements were all over 96%, and the leached slag phases were mainly lead sulfate, calcium sulfate and silicon dioxide phases.

[0052] Example 3:

[0053] Compared with Example 1, the only difference is that the calcination temperature T1 is 350° C., and other operations and parameters are the same as Example 1.

[0054] By analyzing the phase of the calcined product ( Figure 1 ) analysis showed that under this condition, the sodium ferroalite phase in the ferroalite slag disappeared, and its conversion product was mainly sodium ferric sulfate. The leaching rates of iron, zinc and manganese elements were all over 96%, and the leached slag phases were mainly lead sulfate and silicon dioxide phases.

[0055] Example 4:

[0056] Compared with Example 1, the only difference is that the calcination temperature T1 is 400° C., and other operations and parameters are the same as Example 1.

[0057] By analyzing the phase of the calcined product ( Figure 1 ) analysis showed that under this condition, the sodium ferroalite phase in the ferroalite slag disappeared, and its conversion product was mainly sodium ferric sulfate. The leaching rates of iron, zinc and manganese elements were all over 96%, and the leached slag phases were mainly lead sulfate and silicon dioxide phases.

[0058] Table 2 Ion concentration of leachate after calcination of iron alum slag in molten salt in Example 4

[0059] Fe Zn Mn Cu Al Ca K Mg Si 22.86 5.69 4.20 207.1 1.13 536.5 462.0 355.2 14.5 As Cd Pb Cr Ti Co La Li Sr 61.2 52.8 9.5 12.3 56.7 5.1 4.3 21.2 1.6

[0060] Note: Fe, Zn, Mn, Al concentrations are in g / L, and other concentrations are in mg / L

[0061] Table 3 XRF analysis results of elements in leached slag after washing and leaching of iron alum slag in Example 4

[0062] O S Fe Zn Mn Cu Na Ca Ag 49.17 10.63 3.02 0.75 0.66 0.04 5.33 4.69 0.03 Al K Si Pb Cr Sr La Ti Ba 0.53 0.27 13.54 10.40 0.02 0.34 0.07 0.09 0.26

[0063] It can be seen from the data in Tables 2 and 3 that iron, zinc and manganese in the iron alum slag are efficiently leached and enriched in the leachate, and their content in the metal slag is relatively low. Lead and silver in the leached slag are enriched compared with the original slag. Therefore, the selective leaching of iron, zinc and manganese can be achieved at a lower temperature through the directional selective transformation of molten salt, and the enrichment of lead and silver in the slag can be achieved at the same time. It is calculated that the amount of iron alum slag can be reduced by about 73%.

[0064] In addition, the present invention further provides a case in which the ratio of molten salt to iron alum slag exceeds 6:1, for example, 10:1. In addition, molten salt types such as potassium bisulfate are also used, which can realize the molten salt roasting treatment described above, and can also realize the highly selective conversion of the iron alum phase in the iron alum slag into sodium ferric sulfate, and can realize the water leaching recovery of more than 96% of the iron, zinc and manganese ions therein, without affecting the enrichment of lead in the slag.

[0065] Comparative Example 1:

[0066] Compared with Example 3, the only difference is that the calcination temperature T1 is 900°C, and the other operations and parameters are the same as Example 1. By analyzing the phase of the calcined product, it can be seen that ( Figure 2 ), at which temperature a ferric oxide phase will be produced, which will make it difficult for the iron element in the slag to be leached out with water, and sulfur dioxide or sulfur trioxide gas cannot be avoided during the treatment process.

[0067] Comparative Example 2:

[0068] Compared with Example 3, the only difference is that the amount of sodium bisulfate monohydrate is changed so that the molar ratio of iron alum slag to sodium bisulfate monohydrate is 1:6, and the roasting temperature is 500°C. Other operations and parameters are the same as those in Example 1. The roasted product is shown in FIG. Figure 3 .

[0069] Depend on Figure 3 It can be seen that the calcined product in contact with the bottom of the crucible produces a red substance. By analyzing the phase of this part of the calcined product ( Figure 4 ) It can be seen that at this temperature, a ferric oxide phase will be produced, and the production of this phase will make it difficult to leach the iron element in the slag.

[0070] Comparative Example 3

[0071] Compared with Example 3, the only difference is that the amount of sodium bisulfate monohydrate is changed so that the molar ratio of sodium bisulfate monohydrate to iron alum residue is 3:1. Other operations are the same as in Example 3.

[0072] The XRD pattern of the calcined product is shown in Figure 5 , the ferroalloy phase in the ferroalloy slag has not been able to react completely.

[0073] Comparative Example 4

[0074] Compared with Example 3, the only difference is that the amount of sodium bisulfate monohydrate is changed so that the molar ratio of sodium bisulfate monohydrate to iron alum residue is 4:1. Other operations are the same as in Example 3.

[0075] The XRD pattern of the calcined product is shown in Figure 5 , the ferroalloy phase in the ferroalloy slag has not been able to react completely.

[0076] Comparative Example 5

[0077] Compared with Example 2, the only difference is that the amount of sodium bisulfate monohydrate is changed so that the molar ratio of sodium bisulfate monohydrate to iron alum residue is 3:1. Other operations are the same as in Example 2.

[0078] The XRD pattern of the calcined product is shown in Figure 5 , the ferroalloy phase in the ferroalloy slag has not been able to react completely.

[0079] Comparative Example 6

[0080] Compared with Example 3, the only difference is that the roasting temperature T1 is 150° C., and other operations are the same as Example 3.

[0081] The XRD pattern of the calcined product is shown in Figure 5 , the ferroalloy phase in the ferroalloy slag has not been able to react completely.

Claims

1. A molten salt conversion method of iron alum phase, characterized in that: The chemical formula is MFe3(SO4)2(OH)6 The iron alum phase is mixed with molten salt and then subjected to molten salt roasting at a temperature of 250-400°C to convert the iron alum phase into M3Fe(SO4)3; The M is at least one of Na, K, and NH4+; The molten salt is in a molten salt state at the roasting temperature, and the molar ratio of the molten salt to the iron alum phase is greater than 5; The molten salt is at least one salt selected from sodium bisulfate, potassium bisulfate, ammonium sulfate, and ammonium bisulfate, or a hydrate thereof.

2. The molten salt conversion method of the iron alum phase according to claim 1, characterized in that: The molar ratio of the molten salt to the iron alum phase is 5.5-10:

1.

3. The molten salt conversion method of the iron alum phase according to claim 1, characterized in that: The molar ratio of the molten salt to the iron alum phase is 6 to 10:

1.

4. The molten salt conversion method of the iron alum phase according to claim 1, characterized in that: The molar ratio of the molten salt to the iron alum phase is 6-7:

1.

5. The molten salt conversion method of the iron alum phase according to claim 1, characterized in that: The atmosphere in the calcination stage is an atmosphere containing at least one of oxygen, nitrogen, and an inert gas.

6. The molten salt conversion method of the iron alum phase according to any one of claims 1 to 5, characterized in that: The calcination time is more than 1 hour.

7. The molten salt conversion method of the iron alum phase according to claim 6, characterized in that: The roasting time is 2~5h.

8. A molten salt conversion-water leaching recovery method for iron alum phase, characterized in that: The iron alum phase is subjected to molten salt roasting treatment using the molten salt conversion method of the iron alum phase according to any one of claims 1 to 7, and then the roasted material is subjected to water immersion treatment to obtain Fe 3+ water extract.

9. A method for treating iron vitriol slag, characterized in that: The molten salt conversion method according to any one of claims 1 to 7 is used to perform molten salt roasting on the iron alum slag to selectively convert the iron alum phase in the iron alum slag into M3Fe(SO4)3.

10. The method for treating iron alum slag according to claim 9, wherein: The molten salt roasted material is subjected to water leaching to obtain water leaching slag and Fe-enriched 3+ water extract.

11. The method for treating iron vitriol slag according to claim 10, wherein: The liquid-to-solid ratio in the water immersion stage is 5~50ml / g.

12. The method for treating iron vitriol slag according to claim 10, wherein: The temperature during the immersion phase is below 60°C.

13. The method for treating iron alum slag according to claim 12, wherein: The temperature during the immersion stage is 25°C-60°C.

14. The method for treating iron alum slag according to claim 10, wherein: The immersion time is more than 0.5h.

15. The method for treating iron vitriol slag according to claim 14, wherein: The immersion time is 1~3h.

16. The method for treating iron vitriol slag according to any one of claims 10 to 15, characterized in that: The iron alum slag also contains M element, and the M element includes at least one of zinc, manganese, nickel, cobalt and copper; the iron alum slag is subjected to molten salt conversion treatment and water immersion treatment, and the M element is enriched in the water immersion liquid.

17. The method for treating iron vitriol slag according to any one of claims 10 to 15, characterized in that: The iron alum slag also contains N element, and the N element includes at least one of lead, silver, silicon and calcium; the iron alum slag is subjected to molten salt conversion treatment and water leaching treatment, and the N element is enriched in the water-leached slag.

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