Method for recovering iron resources from jarosite

Abstract

The invention discloses a method for recovering iron resources from jarosite, which comprises the following steps: S1, reacting ammonium oxalate with jarosite to realize jarosite phase dissolution: mixing jarosite with ammonium oxalate, and heating and reacting in a water reaction system to generate ammonia gas to obtain leached residues and leached liquid; s2, the leachate reacts with ammonia to precipitate and separate iron, and ammonium oxalate is recycled: S2.1, the leachate reacts with the ammonia: ammonia gas is introduced into the leachate or ammonia water is added into the leachate for full reaction, and reaction product slurry containing ferric hydroxide precipitate and ammonium oxalate is obtained; and S2.2, separating ferric hydroxide and ammonium oxalate from the reaction product slurry: filtering and separating to obtain a ferric hydroxide precipitate, and cooling to separate out crystals from ammonium oxalate in the filtrate so as to recover ammonium oxalate. According to the method, ammonium oxalate used for replacing iron in the iron vitriol can be recycled, the raw material cost is low, iron element separation and utilization are achieved, solid waste is remarkably reduced, and enrichment of other valuable elements in the iron vitriol is facilitated.

Description

Technical Field

[0001] The present invention relates to the technical field of solid waste treatment and iron resource recycling, and in particular to a method for recovering iron resources from iron alum. Background Art

[0002] Ferroalloy slag is an acidic slag formed in the process of iron removal in hydrometallurgy of nonferrous metals. Ferroalloy is the main component of ferroalloy slag, which mainly comes from the iron removal unit operation of zinc hydrometallurgy. In the process of iron removal by zinc roasting sand, valuable components such as zinc ferrite and lead sulfate are often entrained (Wang Zhengmin. Industrial practice of comprehensive recovery and harmless treatment of ferroalloy slag in hydrometallurgy of zinc [J]. World Nonferrous Metals, 2021(24):4-6.). The main component of ferroalloy is MFe3(SO4)2(OH)6, where M=K + 、Na + NH4 + 、H + Existing methods for treating ferroalloy include building tailings ponds for stacking, high-temperature volatilization roasting to recover lead and zinc, and wet acid leaching to recover valuable elements. Building tailings ponds for stacking brings environmental risks and is difficult to develop sustainably. Treating ferroalloy stored in tailings ponds is an inevitable trend in the industry in the future. High-temperature volatilization roasting treatment is to mix ferroalloy and coal in a certain proportion, and then fume them at high temperature to volatilize the lead and zinc components into the smoke dust as oxides for collection and recovery, but it faces the problems of high energy consumption and high desulfurization load (Zhao Yahui, Yang Bin, Lu Xingwu, Dai Xi. Oxygen-enriched reduction volatilization smelting of ferroalloy slag and leaching toxicity of smelting final slag [J]. Nonferrous Metals (Smelting Part), 2022, (9): 150-155). Wet acid leaching can dissolve zinc, iron and other components, which is conducive to separation in the liquid phase, but it requires a large amount of acid leaching, resulting in high operating reagent costs (Wang Xuewu, Research on the preparation of soft magnetic manganese-zinc ferrite using yellow potassium ferroalloy slag [M]. Lanzhou University of Technology, 2010). Chinese patent ZL 202110434643.7 proposes a wet treatment process for iron alum residue in wet zinc smelting. The process uses sulfur dioxide reduction to leach the iron alum residue, and solid-liquid separation is performed to obtain the leach residue and a leachate containing ferrous sulfate. The leachate is further hydrothermally oxidized to produce hematite for iron removal. Although this method can recover iron and achieve the goals of waste reduction and resource utilization, the production of hematite requires high-pressure acid-resistant equipment, which faces relatively expensive pressure-resistant equipment and energy consumption. Chinese patent ZL 202111192868.2 proposes a process for leaching iron alum with oxalic acid, and further decomposes the oxalic acid-complexed iron leachate under light to prepare a ferrous oxalate product. This process consumes the relatively expensive oxalic acid reagent and does not solve the problem of oxalic acid recovery. Ferrous oxalate has limited uses and unsatisfactory economic benefits.

[0003] In summary, the existing methods for treating iron alum slag are divided into two categories: pyrometallurgy and wet leaching. They can technically achieve the comprehensive utilization of iron alum, but they face problems such as high energy consumption, large equipment investment, high reagent cost, and low product value. Summary of the Invention

[0004] In view of the above technical problems existing in the prior art, the purpose of the present invention is to provide a new, low-cost method for recovering iron resources from iron vitriol.

[0005] In order to achieve its purpose, the present invention adopts the following technical solutions:

[0006] A method for recovering iron resources from iron vitriol, comprising the following steps:

[0007] S1. Ammonium oxalate reacts with ferrocyanide to achieve ferrocyanide phase dissolution

[0008] Mixing iron alum and ammonium oxalate, heating and reacting in a water reaction system to generate ammonia gas, and obtaining a reaction material, leaching and filtering the reaction material to obtain a leaching residue and a leachate, wherein the leachate contains trioxalate-complexed ammonium ferrite generated by the reaction; the iron alum is MFe3(SO4)2(OH)6;

[0009] S2. Leachate reacts with ammonia to separate iron through precipitation and recover ammonium oxalate

[0010] S2.1. Reaction of the leachate with ammonia: Adding ammonia gas or aqueous ammonia to the leachate obtained in step S1 to react sufficiently to obtain a reaction product slurry containing ferric hydroxide precipitate and ammonium oxalate;

[0011] S2.2. Separating ferric hydroxide and ammonium oxalate from the reaction product slurry: Separate the ferric hydroxide precipitate by filtration, and cool the filtrate to allow the ammonium oxalate in the filtrate to precipitate crystals to recover the ammonium oxalate, thereby obtaining the ferric hydroxide product and the ammonium oxalate.

[0012] Preferably, the method for recovering iron resources from iron vitriol further comprises step S3:

[0013] S3. Further recovery of oxalate

[0014] After step S2, the remaining filtrate still contains some unprecipitated ammonium oxalate;

[0015] A precipitant is added to the filtrate remaining after step S2, and the precipitant reacts with ammonium oxalate in the filtrate to obtain an oxalate precipitate, which is filtered and separated to obtain oxalate; the precipitant is ferrous sulfate, calcium sulfate, or calcium hydroxide, and the obtained oxalate precipitate is ferrous oxalate or calcium oxalate.

[0016] Preferably, when the precipitant is ferrous sulfate, the filtrate remaining after step S2 is heated until ammonia no longer overflows, and then ferrous sulfate is added to the filtrate.

[0017] Preferably, the water reaction system in step S1 is provided by passing water vapor into a reactor, or by adding raw materials into water and heating to the reaction temperature; the reaction temperature in step S1 is 90-160° C., and the reaction time is 10 min-60 min.

[0018] More preferably, the reaction temperature of step S1 is 95-130° C. or 95-110° C., and the reaction time is 10 min-30 min.

[0019] Preferably, the method for recovering iron resources from ferrocyanide comprises subjecting the leaching residue to secondary leaching to further improve the iron replacement rate in ferrocyanide: the leaching residue obtained in step S1 is added to a water reaction system for heating reaction; since the leaching residue contains unreacted ammonium oxalate and ferrocyanide, it reacts after being heated again in the water reaction system to generate ammonia gas, and leaching residue and leachate are obtained again; all the obtained leachates are combined and mixed and used in step S2.

[0020] Preferably, the ferrosite in step S1 is selected from yellow jarosite, yellow ammonium jarosite and yellow sodium jarosite, and the mass ratio of ammonium oxalate to ferrosite is 2.7 to 5:1.

[0021] Further preferably, the ammonium oxalate is (NH4)2C2O4·H2O, and the mass ratio of (NH4)2C2O4·H2O to ferroalite is 2.8 to 4:1.

[0022] Preferably, in step S2.1, ammonia gas is introduced into the leachate obtained in step S1 or ammonia water is added until the pH of the solution is 8.5-11.

[0023] Preferably, in step S2.2, ferric hydroxide and ammonium oxalate are separated by recrystallization: the reaction product slurry is filtered and separated, and the obtained solid precipitate is a mixture of ferric hydroxide and ammonium oxalate crystals, and the filtrate contains ammonium oxalate. The ammonium oxalate in the filtrate is allowed to precipitate crystals by cooling, and the ammonium oxalate crystals and the filtrate are collected by filtration; the solid precipitate is added to the filtrate from which the ammonium oxalate crystals are removed, heated, and the ammonium oxalate in the solid precipitate is dissolved in the filtrate, filtered and separated, and the obtained filtrate is cooled to precipitate ammonium oxalate crystals, and the ammonium oxalate crystals are collected by filtration. The remaining filtrate is used again to dissolve the ammonium oxalate in the solid precipitate next time, and the crystallization is repeated many times until all the ammonium oxalate in the solid precipitate is separated out, and the final precipitate is the ferric hydroxide filter cake from which the ammonium oxalate crystals are removed.

[0024] The beneficial effects of the present invention are:

[0025] 1) Ferric hydroxide products can be obtained, creating economic value.

[0026] 2) The separation and utilization of iron elements and the significant reduction of solid waste are achieved, which is beneficial to the enrichment of other valuable elements in iron alum.

[0027] 3) The ammonium oxalate used to replace the iron in the iron alum can be recycled, and the raw material cost is low.

[0028] 4) No expensive pressure-resistant or acid-resistant equipment is required, and a conventional reaction device can complete the process of the present invention. The requirements for the reaction device are not high, the equipment cost is low, and batch-scale production is possible.

[0029] 5) The reaction is carried out at a lower temperature, which results in lower energy consumption compared to desulfurization treatment at a high temperature. BRIEF DESCRIPTION OF THE DRAWINGS

[0030] Figure 1 It is a process flow chart of Example 1 of the present invention.

[0031] Figure 2 It is a process flow chart of Examples 2 and 3 of the present invention.

[0032] Figure 3 It is the XRD pattern of the iron alum raw material and product of the present invention.

[0033] Figure 4 This is a graph showing the thermogravimetric test results of the ferric hydroxide product of Example 1 of the present invention. DETAILED DESCRIPTION

[0034] The present invention will be further described below with reference to the embodiments, but the present invention is not limited thereto.

[0035] The ferroalcite raw materials used in the embodiment of the present invention include sodium ferroalcite NaFe3(SO4)2(OH)6, yellow potassium ferroalcite KFe3(SO4)2(OH)6 and yellow ammonium ferroalcite NH4Fe3(SO4)2(OH)6, which are artificially synthesized by reacting analytically pure bicarbonate with ferric sulfate reagent: a 50% ferric sulfate solution is heated to boiling, bicarbonate powder is slowly added and stirred, and the pH is adjusted to about 1.5. During the process, evaporated water is continuously added to maintain the liquid level. The reaction is carried out for 2 hours to obtain a slurry, the slurry is filtered, and the soluble impurity salts are washed to obtain an insoluble khaki ferroalcite filter cake, which is then dried to obtain the raw material product. The reaction is as follows:

[0036] 12MHCO3+3Fe2(SO4)3=2MFe3(SO4)2(OH)6↓+12CO2↑+5M2SO4

[0037] Where M = Na + , K + , NH4 + . Figure 3The XRD patterns of the synthesized ferrosites (sodium ferrosite, jarosite, and ammonium ferrosite) are shown in the figure, and the results show that the ferrosite phase is well-crystallized. The synthesized sodium ferrosite, jarosite, and ammonium ferrosite are used as raw materials for Examples 1-3.

[0038] During their research, the inventors discovered that placing ammonium oxalate solution in a household pressure cooker could pressure-leach iron alum and generate ammonia gas. Adding ammonia to the leachate precipitated ferric hydroxide. Quantitative testing was conducted using equipment rented from a company. The equipment primarily consisted of a steam generator, a reactor, an adjustable constant-pressure valve, and auxiliary equipment such as piping. High-temperature steam generated by the steam generator was passed through a reactor pre-filled with ammonium oxalate and iron alum for direct heating, or water was added to the reactor containing ammonium oxalate and iron alum for indirect heating. Stirring was maintained, and the constant-pressure valve connected to the reactor's exhaust port was adjusted to maintain the reactor temperature at a constant temperature of 90-160°C, eliminating ammonia.

[0039] Example 1 Recovery of iron resources from jarosite

[0040] Operation process as follows Figure 1 As shown, follow the steps below:

[0041] S1. Ammonium oxalate reacts with ferrocyanide to achieve ferrocyanide phase dissolution

[0042] S1.1. Primary leaching

[0043] Ammonium oxalate monohydrate (NH4)2C2O4·H2O) and jarosite KFe3(SO4)2(OH)6 were mixed in a mass ratio of 4:1, using high-temperature steam as the solvent. The jarosite and ammonium oxalate were mixed in a reactor. High-temperature steam was continuously introduced into the reactor through the bottom inlet until the solution reached a temperature of 160°C to evaporate ammonia. Steam was introduced continuously for 10 minutes. Ammonia generated by the reaction was discharged from the exhaust port at the top of the reactor, resulting in a post-reaction product containing a leachate residue and a leachate. The leachate residue primarily consisted of unreacted (NH4)2C2O4 and KFe3(SO4)2(OH)6 (due to the reversible reaction of ammonium oxalate leaching iron oxides, some of the raw materials were not completely reacted). The leachate primarily contained (NH4)3Fe(C2O4)3, (NH4)2SO4, and K2SO4.

[0044] S1.2, secondary leaching

[0045] The post-reaction material from step S1.1 was filtered and solid-liquid separated to obtain a leach residue and a leachate, respectively. The leach residue was added to water and heated to 95°C to react and generate ammonia gas. A secondary leaching was performed to obtain a secondary leachate and a secondary leach residue. The secondary leach residue was primarily composed of a small amount of unreacted KFe3(SO4)2(OH)6. The primary components of the secondary leachate were the same as those of the leachate from step S1.1. The leachates from the two leaching processes were combined for use in the next experiment.

[0046] The chemical reaction occurring in step S1 is:

[0047] KFe3(SO4)2(OH)6+9(NH4)2C2O4=3(NH4)3Fe(C2O4)3+1.5(NH4)2SO4+0.5K2SO4+6H2O+6NH3↑

[0048] Step S1.2 is repeated leaching to increase the replacement rate of iron in ferroalloy and improve the yield.

[0049] S2. Leachate reacts with ammonia to separate iron through precipitation and recover ammonium oxalate

[0050] S2.1. Reaction of leachate with ammonia: Ammonia gas is introduced into the mixed leachate obtained in step S1 until the pH of the solution reaches 8.5 (excess ammonia gas is introduced), and the mixture is fully reacted to obtain a reaction product slurry;

[0051] S2.2, recrystallization filtering separation ferric hydroxide and ammonium oxalate:By reaction product slurry filtering separation, the solid precipitation obtained is the mixture of ferric hydroxide and ammonium oxalate crystal (due to the limited solubility of ammonium oxalate, therefore part ammonium oxalate separates out crystal and ferric hydroxide precipitation is mixed together), filtrate contains ammonium oxalate, by cooling, the ammonium oxalate in filtrate is allowed to separate out crystal, ammonium oxalate crystal and filtrate are collected by filtering.Solid precipitation is added in the filtrate having removed ammonium oxalate crystal, heating, the ammonium oxalate in solid precipitation is dissolved in filtrate, filtering separation, separate out ammonium oxalate crystal after the filtrate obtained cools, ammonium oxalate crystal is collected by filtering, remaining filtrate is used again for the ammonium oxalate in next time dissolved solid precipitation, so repeatedly crystallizes until the ammonium oxalate in solid precipitation is all separated, the precipitation finally obtained is the ferric hydroxide filter cake having removed ammonium oxalate crystal, and final filtrate is used for step S3.

[0052] The chemical reaction occurring in step S2 is:

[0053] (NH4)3Fe(C2O4)3+3NH3·H2O=Fe(OH)3↓+3(NH4)2C2O4↓

[0054] S3. Further recovery of oxalate

[0055] Since ammonium oxalate is soluble in water, it is not possible to completely precipitate crystals by cooling. Therefore, the filtrate obtained in step S2 still contains some unprecipitated ammonium oxalate. Therefore, this step is used to recover the oxalate therein:

[0056] The filtrate was heated until no more ammonia was released (due to the excessive ammonia introduced in step S2.1, unreacted ammonia remained). A 20 wt.% solution of ferrous sulfate heptahydrate was then added until no more yellow precipitate was produced. A total of 150 g of ferrous sulfate heptahydrate was added, and the mixture was filtered to obtain a ferrous oxalate dihydrate filter cake and a sulfate solution. In this step, residual ammonia in the filtrate must be removed before adding ferrous sulfate to precipitate ammonium oxalate to prevent the formation of ferrous hydroxide precipitate, which would inhibit the formation of ferrous oxalate and hinder the recovery of oxalate.

[0057] The chemical reaction occurring in step S3 is:

[0058] FeSO4+(NH4)2C2O4=FeC2O4↓+(NH4)2SO4

[0059] Example 2 Recovery of iron resources from yellow ammonium ferroaluminate

[0060] Operation process as follows Figure 2 As shown, follow the steps below:

[0061] S1. Ammonium oxalate reacts with ferrocyanide to achieve ferrocyanide phase dissolution

[0062] S1.1. Primary leaching

[0063] Ammonium oxalate monohydrate and yellow ammonium ferrosite are placed in a reactor at a mass ratio of 2.8:1 and mixed. Water is added to the reactor as the reaction solvent system, and the reactor is heated to 110°C for 30 minutes to evaporate ammonia to generate ammonia gas. The reaction material comprises a leachate residue and a leachate. The leachate residue mainly contains unreacted ammonium oxalate and yellow ammonium ferrosite, and the leachate mainly contains (NH4)3Fe(C2O4)3 and (NH4)2SO4.

[0064] S1.2, secondary leaching

[0065] The post-reaction material from step S1.1 was filtered and solid-liquid separated to obtain a leachate and a leachate, respectively. The leachate was added to water and heated to 95°C to react and generate ammonia gas. A secondary leaching was performed to obtain a secondary leachate and a secondary leachate. The secondary leachate was primarily composed of a small amount of unreacted yellow ammonium ferrosite. The primary components of the secondary leachate were the same as those of the leachate from step S1.1. The leachates from the two leaching processes were combined for use in the next experiment.

[0066] The chemical reaction occurring in step S1 is:

[0067] NH4Fe3(SO4)2(OH)6+9(NH4)2C2O4=3(NH4)3Fe(C2O4)3+2(NH4)2SO4+6H2O+6NH3↑

[0068] S2. Leachate reacts with ammonia to separate iron through precipitation and recover ammonium oxalate

[0069] S2.1. Reaction of leachate with ammonia: Ammonia gas is introduced into the mixed leachate obtained in step S1 until the pH of the solution reaches 9.0 (excess ammonia gas is introduced), and the mixture is fully reacted to obtain a reaction product slurry;

[0070] S2.2. Recrystallization to separate ferric hydroxide and ammonium oxalate: same as in Example 1.

[0071] The reaction principle in step S2 is the same as that in Example 1.

[0072] S3. Further recovery of oxalate

[0073] 150 g of calcium sulfate dihydrate was added to the filtrate of S2, and the filter cake containing calcium oxalate was obtained by filtration. Ammonia was evaporated from the filtrate to obtain ammonium sulfate solution.

[0074] The chemical reaction occurring in step S3 is:

[0075] CaSO4+(NH4)2C2O4=CaC2O4↓+(NH4)2SO4

[0076] Example 3 Recovery of iron resources from sodium ferroaluminate

[0077] Operation process as follows Figure 2 As shown, follow the steps below:

[0078] S1. Ammonium oxalate reacts with ferrocyanide to achieve ferrocyanide phase dissolution

[0079] S1.1. Primary leaching

[0080] Ammonium oxalate monohydrate and natrolite are mixed in a reactor at a mass ratio of 3.3:1. Water is added to the reactor as the reaction solvent system, and the reactor is electrically heated to 95°C for 60 minutes to evaporate ammonia to generate ammonia gas. The reaction material comprises a leachate residue and a leachate. The leachate residue mainly contains unreacted ammonium oxalate and natrolite, and the leachate mainly contains (NH4)3Fe(C2O4)3, (NH4)2SO4, and Na2SO4.

[0081] S1.2, secondary leaching

[0082] The post-reaction material from step S1.1 was filtered and subjected to solid-liquid separation to obtain a leach residue and a leachate, respectively. The leach residue was added to water and heated to 95°C to react and generate ammonia gas. A secondary leaching was performed to obtain a secondary leachate and a secondary leach residue. The secondary leach residue mainly consisted of a small amount of iron hydroxide impurities produced during the synthesis of ferroalite and a small amount of unreacted natrona. The secondary leachate had the same main components as the leachate from step S1.1. The leachates from the two leaching processes were combined for use in the next experiment.

[0083] The chemical reaction occurring in step S1 is:

[0084] NaFe3(SO4)2(OH)6+9(NH4)2C2O4=3(NH4)3Fe(C2O4)3+1.5(NH4)2SO4+0.5Na2SO4+6H2O

[0085] +6NH3↑

[0086] S2. Leachate reacts with ammonia to separate iron through precipitation and recover ammonium oxalate

[0087] S2.1. Reaction of leachate with ammonia to separate iron: Aqueous ammonia is added to the mixed leachate obtained in step S1 until the pH of the solution reaches 11 (adding excess aqueous ammonia), and the mixture is fully reacted to obtain a reaction product slurry;

[0088] S2.2. Recrystallization to separate ferric hydroxide and ammonium oxalate: same as in Example 1.

[0089] S3. Further recovery of oxalate

[0090] Calcium hydroxide solid was added to the filtrate of S2, and a total of 40 g of calcium hydroxide was added to the filtrate. The mixture was fully reacted and filtered to obtain a calcium oxalate filter cake. Ammonia was evaporated from the filtrate to obtain a sulfate solution.

[0091] The chemical reaction occurring in step S3 is:

[0092] Ca(OH)2+(NH4)2C2O4=CaC2O4↓+2NH3·H2O

[0093] Product identification

[0094] The iron hydroxide recovered from step S2 of Example 1-3 and the oxalate precipitate recovered from step S3 were outsourced for XRD analysis of their phase and elemental composition. The results are as follows: Figure 3 As shown in , the product obtained in step S3 is oxalate precipitate. The thermogravimetric analysis results of the iron hydroxide in Example 1 are as follows Figure 4As shown, the temperature was increased at 10°C / min in an air atmosphere. When the temperature reached 100°C, water loss began and the water loss rate increased. When the temperature rose to 230°C, the water loss rate reached a peak. Thereafter, the water loss rate slowed down. When the temperature reached 400°C, there was no more weight loss. The residual mass after weight loss was 74.95%, which was close to the theoretical value of 74.71%. It can be inferred that the product of step S2 is Fe(OH)3.

[0095] The raw material dosage, iron recovery rate and ammonium oxalate recovery rate of each embodiment are shown in Table 1. It is worth noting that the purity of iron hydroxide in step S2 is based on the XRF results in Table 2 and Figure 4 The calculated value after deducting the adsorbed water from the thermogravimetric analysis is:

[0096] Table 1

[0097]

[0098]

[0099] Table 2 is the XRF elemental analysis results of the products of Examples 1-3

[0100] Fe K Na S Example 1 96.405％ 2.388％ 1.136％ 0.071％ Example 2 98.88％ 0.0537％ 0.546％ 0.520％ Example 3 97.453％ 0.666％ 1.575％ 0.306％

[0101] Iron recovery rate refers to the mass percentage of iron in ferric hydroxide to iron in iron alum raw material.

[0102] The ammonium oxalate recovery rate refers to the mass ratio of the ammonium oxalate obtained in step S2 to the raw material ammonium oxalate in step S1.

[0103] The experiments of Examples 1-3 show that the process of the present invention can be used to efficiently utilize iron alum. The ferrous oxalate dihydrate and calcium oxalate monohydrate products obtained in Examples 1 and 3 are both well-crystalline. The ferric hydroxide product of Example 1 is in an amorphous state due to rapid precipitation in the solution. In Example 2, oxalate is recovered by adding excess calcium sulfate, and the product is a mixture of calcium sulfate dihydrate and calcium oxalate monohydrate. There is almost no residual oxalate in the salt solution obtained in step S3 of Examples 2 and 3. In the process of the present invention, more than 95% of oxalate is recovered by steps S2 and S3, and oxalate can be recycled, which greatly reduces the cost of iron recovery.

Claims

1. A method for recovering iron resources from iron vitriol, characterized in that: The steps include: S1. Ammonium oxalate reacts with ferrocyanide to achieve ferrocyanide phase dissolution Mixing iron alum and ammonium oxalate, heating and reacting in a water reaction system to generate ammonia gas, and obtaining a reaction material, leaching and filtering the reaction material to obtain a leaching residue and a leachate, wherein the leachate contains trioxalate-complexed ammonium ferrite generated by the reaction; the iron alum is MFe3(SO4)2(OH)6; S2. Leachate reacts with ammonia to separate iron through precipitation and recover ammonium oxalate S2.

1. Reaction of the leachate with ammonia: Adding ammonia gas or aqueous ammonia to the leachate obtained in step S1 to react sufficiently to obtain a reaction product slurry containing ferric hydroxide precipitate and ammonium oxalate; S2.

2. Separating ferric hydroxide and ammonium oxalate from the reaction product slurry: Separate the ferric hydroxide precipitate by filtration, and cool the filtrate to allow the ammonium oxalate in the filtrate to precipitate crystals to recover the ammonium oxalate, thereby obtaining the ferric hydroxide product and the ammonium oxalate.

2. The method according to claim 1, characterized in that Also includes step S3: S3. Further recovery of oxalate After step S2, the remaining filtrate still contains some unprecipitated ammonium oxalate; A precipitant is added to the filtrate remaining after step S2, and the precipitant reacts with ammonium oxalate in the filtrate to obtain an oxalate precipitate, which is filtered and separated to obtain oxalate; the precipitant is ferrous sulfate, calcium sulfate, or calcium hydroxide, and the obtained oxalate precipitate is ferrous oxalate or calcium oxalate.

3. The method according to claim 2, wherein: When the precipitant is ferrous sulfate, the filtrate remaining after step S2 is heated until ammonia no longer overflows, and then ferrous sulfate is added to the filtrate.

4. The method according to claim 1, wherein: The water reaction system in step S1 is provided by passing water vapor into the reactor, or by adding raw materials into water and heating to the reaction temperature; the reaction temperature in step S1 is 90-160° C., and the reaction time is 10 min-60 min.

5. The method according to claim 4, characterized in that: The reaction temperature of step S1 is 95-130° C. or 95-110° C., and the reaction time is 10 min-30 min.

6. The method according to claim 1 or 4, characterized in that: The leaching residue is subjected to secondary leaching to further improve the iron replacement rate in the iron alum: the leaching residue obtained in step S1 is added to a water reaction system for heating reaction. Since the leaching residue contains unreacted ammonium oxalate and iron alum, it reacts after being heated again in the water reaction system to generate ammonia gas, and the leaching residue and leachate are obtained again. All the obtained leachates are combined and mixed and used in step S2.

7. The method according to claim 1, wherein: The ferrosite in step S1 is selected from yellow jarosite, yellow ammonium jarosite and yellow sodium jarosite, and the mass ratio of ammonium oxalate to ferrosite is 2.7 to 5:

1.

8. The method according to claim 7, wherein: The ammonium oxalate is (NH4)2C2O4·H2O, and the mass ratio of (NH4)2C2O4·H2O to ferroalite is 2.8-4:

1.

9. The method according to claim 1, wherein: In step S2.1, ammonia gas or ammonia water is introduced into the leachate obtained in step S1 until the pH of the solution is 8.5-11.

10. The method according to claim 1, wherein: In step S2.2, ferric hydroxide and ammonium oxalate are separated by recrystallization: the reaction product slurry is filtered and separated, and the obtained solid precipitate is a mixture of ferric hydroxide and ammonium oxalate crystals. The filtrate contains ammonium oxalate, and the ammonium oxalate in the filtrate is cooled to precipitate crystals, and the ammonium oxalate crystals and the filtrate are collected by filtration; the solid precipitate is added to the filtrate from which the ammonium oxalate crystals are removed, heated, and the ammonium oxalate in the solid precipitate is dissolved in the filtrate, filtered and separated, and the obtained filtrate is cooled to precipitate ammonium oxalate crystals, and the ammonium oxalate crystals are collected by filtration. The remaining filtrate is used again to dissolve the ammonium oxalate in the solid precipitate next time, and the crystallization is repeated many times until all the ammonium oxalate in the solid precipitate is separated out, and the final precipitate is the ferric hydroxide filter cake from which the ammonium oxalate crystals are removed.

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