Comprehensive utilization method of high-calcium-magnesium flotation tailings

Abstract

The invention discloses a comprehensive utilization method of high-calcium-magnesium flotation tailings, which realizes the following targets through tailings pretreatment and physical pre-enrichment, heavy product treatment, low-temperature selective activation roasting, dilute acid selective deep leaching, calcium-magnesium separation and product preparation: innovative centrifugal beneficiation and impurity removal-low-temperature roasting-pH regulation and control selective leaching technology; efficient separation of calcium and magnesium and deep removal of zinc and iron are realized, and a foundation is laid for preparation of high-purity products; high-activity magnesium oxide meeting the standards of light calcined magnesium oxide and light magnesium oxide and high-purity and high-whiteness calcium carbonate are prepared; the traditional high-temperature process is replaced by 700-750 DEG C low-temperature roasting, and the diluted acid selective leaching is combined, so that the energy consumption and the acid consumption are greatly reduced, and the process economy is improved; valuable metals such as zinc and iron are synchronously recovered, all-component green utilization of tailings is achieved, and solid waste emission is reduced. Therefore, an efficient, energy-saving, high-value and green technical route is constructed, and efficient separation and high-value conversion of elements such as calcium, magnesium, zinc and iron in the tailings are achieved.

Description

Technical Field

[0001] This invention belongs to the field of mineral processing and metallurgical chemical technology, specifically relating to a process for efficiently and comprehensively recovering calcium and magnesium resources from high-calcium and magnesium flotation tailings (zinc tailings, phosphorus tailings) and preparing high-value-added magnesium oxide and calcium carbonate products. Background Technology

[0002] In non-ferrous metal or phosphate rock flotation operations, high-calcium and magnesium tailings with dolomite (CaMg(CO3)2) as the main mineral phase are generated, often accompanied by small amounts of sulfur- and zinc-containing minerals that are not fully recovered. The stockpiling of such tailings not only occupies significant land resources but also results in resource waste due to the ineffective utilization of valuable elements. Currently, traditional processes for extracting calcium and magnesium from dolomite (such as high-temperature calcination and carbonization) generally suffer from drawbacks such as high energy consumption and lengthy process flows, and struggle to efficiently remove impurities such as zinc and iron, leading to low purity and limited added value in the final product, failing to achieve the goal of high-value utilization of tailings resources. Therefore, developing a comprehensive tailings utilization technology that can achieve efficient separation of calcium and magnesium, deep removal of harmful impurities, and simultaneous preparation of high-quality calcium and magnesium products has significant practical significance and application value for promoting resource recycling and reducing environmental pollution.

[0003] In summary, current mainstream processes have significant drawbacks: First, they struggle to achieve efficient and selective separation of calcium and magnesium, and cannot deeply remove impurities such as zinc and iron, resulting in cross-contamination of magnesium oxide and calcium carbonate products and low purity. Second, they either rely on complete calcination at temperatures above 800℃, leading to high energy consumption, or employ direct acid dissolution, which suffers from high acid consumption and severe equipment corrosion, both resulting in high production costs and insufficient economic feasibility. Third, they lack targeted recovery schemes for valuable impurities such as zinc and iron. Even with recovery processes, some technologies suffer from complex processes, leading to resource waste and potential secondary pollution. These combined shortcomings make it difficult for existing technologies to achieve high-value-added full utilization of such tailings in an economical and environmentally friendly manner. Summary of the Invention

[0004] The purpose of this invention is to address the technical pain points of existing technologies in processing high-calcium and magnesium, zinc-containing iron flotation tailings, such as low separation efficiency, high energy consumption, poor product purity, limited added value, and resource waste, and to provide a new, efficient, and green comprehensive utilization process, namely a comprehensive utilization method for high-calcium and magnesium flotation tailings. Through multi-process synergistic optimization, it achieves efficient separation and high-value conversion of elements such as calcium, magnesium, zinc, and iron in the tailings.

[0005] To achieve the above-mentioned objectives, the technical solution of the present invention is as follows:

[0006] The first step, tailings pretreatment and physical pre-enrichment, involves grinding the flotation tailings to -75μm. Based on the density difference between dolomite and zinc-bearing minerals and pyrite, centrifugal separation is used to separate the minerals, resulting in heavy products (pyrite and sphalerite) enriched in the underflow and light products (mainly dolomite) discharged in the overflow. This step can remove heavy impurities in advance, significantly reducing the load and acid consumption of subsequent chemical treatments, laying the foundation for the economic feasibility of the process.

[0007] The second step is the processing of heavy products: the heavy products obtained from centrifugal separation are separately processed for zinc-sulfur separation to achieve targeted recovery of zinc and sulfur resources.

[0008] The third step, low-temperature selective activation calcination: The light product obtained from centrifugation (enriched calcium and magnesium materials) is mixed with trace additives (such as CaCl2) and calcined at 600℃~800℃ in a weakly oxidizing atmosphere containing oxygen. The core innovation of this step lies in the synergistic effect of the additives and the controlled atmosphere, which selectively decomposes magnesium carbonate in dolomite into highly active magnesium oxide, while calcium carbonate remains stable, resulting in activated calcined material. Compared with the traditional high-temperature calcination process, this step reduces energy consumption by more than 30%, and the resulting porous composite structure can significantly improve the mass transfer efficiency and selectivity of subsequent reactions.

[0009] The fourth step, selective deep leaching with dilute acid: The activated roasted material is leached with dilute acid. By precisely controlling the leaching endpoint, the iron oxide and associated zinc oxide converted from pyrite are efficiently dissolved (leaching rate >95%), yielding a zinc-iron leaching solution. Meanwhile, the calcium component selectively remains in the slag as calcium carbonate or calcium oxide (calcium dissolution rate <5%), resulting in a calcium- and magnesium-rich leaching slag. This step is crucial for achieving deep impurity removal and ensuring the purity of the final product.

[0010] Step 5, Calcium and Magnesium Separation and Product Preparation: The calcium- and magnesium-rich leaching residue after deep impurity removal is separated using an ammonium salt cyclic leaching method. In the hot ammonium salt solution system, calcium preferentially forms a soluble complex and enters the liquid phase, while magnesium precipitates as magnesium hydroxide. After separation, CO2 is introduced into the calcium-containing liquid phase to prepare high-purity calcium carbonate. The magnesium hydroxide is washed, dried, and calcined to obtain highly active light-calcined magnesium oxide. This process has a high separation factor, excellent product purity, and the ammonium salt can be recycled.

[0011] The second step involves processing the heavy product obtained in the first step using a regrinding-flotation process, and recovering zinc concentrate and sulfur concentrate by adding selectivity modifiers and collectors.

[0012] The third step involves concentrating and dehydrating the calcium-magnesium enriched light product obtained in the first step to a moisture content of <15%. Anhydrous calcium chloride is weighed as an additive at 0.5% to 1.2% of the dry weight of the light product. The light product and anhydrous calcium chloride are then thoroughly and uniformly mixed in a mixer. The mixture is then fed into a controlled atmosphere calcination device, and a weak oxidizing mixed gas containing 1% to 5% oxygen by volume is introduced. The temperature is increased to the target temperature of 700℃ to 750℃ at a rate of ≤10℃ / min, and held at this temperature for 60 to 90 minutes. After the holding period, the temperature is cooled to <100℃ under a weak oxidizing atmosphere before being removed from the furnace to obtain the calcined product, i.e., the activated calcined material.

[0013] The fourth step specifically involves adding the cooled, low-temperature selectively activated calcined material obtained in the third step into an acid-resistant reactor equipped with stirring and heating functions; using a concentration of 1.5~2.5. A mol / L industrial sulfuric acid solution is used, with the liquid-to-solid mass ratio controlled at 4:1 to 6:1. The reaction temperature is maintained at 75℃~85℃ using waste heat from the flue gas. Precise pH control is achieved by using an online pH meter to monitor the pH of the reaction slurry in real time. Dilute sulfuric acid is slowly and continuously added via a metering pump, or by adding it manually in batches while closely monitoring the process. The final leaching endpoint pH is stably controlled within a narrow window of 4.2~4.8. When the pH reaches the set range and remains stable for 15~20 minutes without significant increase, the reaction endpoint is determined. The mixture is then kept warm and stirred for 30~45 minutes to ensure complete dissolution of zinc and iron oxides. Iron and zinc compounds are recovered through stepwise precipitation. After leaching, hot filtration is immediately performed using a plate and frame filter press or vacuum filtration device. The filter residue is then countercurrently washed 2~3 times with hot water at 60℃~80℃ to remove entrained soluble sulfates, yielding calcium-magnesium-rich leaching residue and a zinc-iron-containing leachate.

[0014] In the fifth step, calcium and magnesium are first separated by ammonium salt cyclic leaching: in the hot ammonium salt solution, calcium preferentially forms a soluble complex [Ca(NH3)n]. 2+The magnesium leaching residue enters the liquid phase, while it precipitates primarily as magnesium hydroxide (Mg(OH)2). The washed calcium-rich magnesium leaching residue is then added to a corrosion-resistant reactor, along with a 3.0–4.0 mol / L ammonium chloride (NH4Cl) solution as the leaching agent. The liquid-to-solid ratio is controlled at 5:1 to 7:1. Stirring and heating are initiated, maintaining the reaction temperature at 85°C–95°C for 90–120 minutes. After the reaction, the residue is filtered while hot using pressure or vacuum filtration to obtain a filtrate (A) rich in calcium chloride and ammonium chloride, and a crude magnesium hydroxide filter cake (A). For the preparation of calcium carbonate: the filtrate is transferred to a carbonation reactor. CO2 is introduced into the solution at room temperature or slightly heated (<40°C), controlling the aeration rate to ensure a stable reaction and avoid localized over-alkalinity or over-acidity. The reaction is monitored using a pH meter and conductivity meter. When the pH stabilizes at 7.0–7.5, the conductivity decreases significantly. When the carbonization stops changing, it is determined to be complete. The slurry is filtered, and the resulting filter cake is washed with deionized water until no chloride ions are present. The filter cake is dried at 105℃~120℃ to obtain a high-purity precipitated calcium carbonate product. The mother liquor is mainly an ammonium chloride solution. Then, the preparation of lightly calcined magnesium oxide is carried out: the previously obtained filter cake A, i.e., crude magnesium hydroxide, is slurried and washed with hot water at 60℃~80℃ to fully wash away the entrained ammonium chloride and calcium chloride. It is filtered again to obtain a relatively pure magnesium hydroxide filter cake. The washed magnesium hydroxide filter cake is thoroughly dried at 110℃~120℃. The dried magnesium hydroxide is sent to a calcining furnace and calcined at 800℃~880℃ for 1.5~2.5 hours. After cooling, the lightly calcined magnesium oxide product is obtained.

[0015] The beneficial effects of this invention are mainly reflected in the following aspects: (1) High separation precision and guaranteed product purity: Through the cascade design of selective activation and pH precise control leaching, the four elements of calcium, magnesium, zinc and iron are effectively separated and controlled, solving the problem of cross-contamination of products in traditional processes and providing a guarantee for the preparation of high-quality products.

[0016] (2) Green and low-carbon, with significant environmental benefits: The low-temperature selective roasting process significantly reduces energy consumption; the selective leaching of dilute acid and CO2 carbonization process reduce reagent consumption and carbon emissions; and the ammonium salt recycling system effectively reduces waste emissions, which is in line with the concept of green development.

[0017] (3) Achieve internal carbon recycling and practice waste treatment: The CO2 generated in the roasting process is directly recycled to the carbonization process to build a closed-loop carbon recycling system within the process. This not only improves the utilization rate of CO2 resources, but also achieves the synergistic benefits of waste treatment and further reduces the environmental burden.

[0018] (4) Energy recycling and cost advantage: By adding flue gas heat exchangers and / or high-temperature material cooling heat exchangers, the initial temperature of the leachate can be increased to above 60°C, significantly reducing the external heating energy consumption in the leaching process, realizing the recycling of internal energy, and further reducing production costs and carbon emissions.

[0019] (5) High-value utilization of all components and maximization of resource benefits: The main products are high-value magnesium oxide and calcium carbonate that meet industry standards. At the same time, by-products such as zinc and iron can be recycled in a targeted manner, truly realizing the resource utilization and high-value utilization of all components of tailings, and achieving the resource recycling goal of turning waste into treasure.

[0020] The process constructs a synergistic process chain of physical pre-enrichment, low-temperature selective activation, precise leaching separation, and high-value conversion to achieve the following objectives: (1) Innovate low-temperature roasting and pH-controlled leaching technology to achieve efficient separation of calcium and magnesium and deep removal of zinc and iron, with dolomite purity greater than 91% (MgO content greater than 20%) and total heavy metal impurities such as Zn and Fe less than 2%, laying the foundation for the preparation of high-purity products; (2) Prepare high-activity magnesium oxide that meets the standards of "Light-burned Magnesium Oxide" (YB / T 5206-2023) and "Light Magnesium Oxide" (HG / T 2573-2012), as well as high-purity and high-whiteness calcium carbonate, with CaCO3 purity and whiteness greater than 99.2% and 95%, respectively; (3) Replace the traditional high-temperature process with low-temperature roasting at 600℃~800℃, combined with dilute acid selective leaching, to significantly reduce energy consumption and acid consumption and improve process economy; (4) Simultaneously recover valuable metals such as zinc and iron, realize the green utilization of all components of tailings, and reduce solid waste emissions.

[0021] In summary, this invention aims to construct an efficient, energy-saving, high-value, and green technical route to break through the industry bottleneck of economically and efficiently utilizing complex tailings. Detailed Implementation

[0022] Step 1: Tailings pretreatment and physical pre-enrichment (1) Take high-calcium magnesium flotation tailings (mainly dolomite, with a small amount of pyrite and sphalerite) from the lead-zinc ore beneficiation plant; feed the tailings into a wet ball mill and add process water for grinding. Strictly control the grinding fineness so that the final pass rate of the material to -75μm is ≥95% to form a uniform slurry.

[0023] (2) The ground slurry is pumped into a high-efficiency centrifugal concentrator to separate dolomite (specific gravity approximately 2.8~2.9 g / cm³) from pyrite (specific gravity 5.0 g / cm³) and sphalerite (specific gravity 4.0 g / cm³). The drum speed is adjusted to 400~680 rpm, the feed concentration to 15%~20%, and the flushing water flow rate to 1.6~3.75 L / min. Heavy products are effectively enriched in the underflow, containing pyrite (i.e., sulfur resources) and sphalerite, while light products are discharged from the overflow, primarily consisting of dolomite. The zinc content in the light products should be reduced to <0.3%, and the iron content significantly reduced, achieving preliminary enrichment of calcium and magnesium materials and removal of impurities.

[0024] Step 2: Heavy product processing (zinc and sulfur resource recovery) (1) The heavy products obtained in the first step (rich in zinc and sulfur) are not included in the subsequent calcium and magnesium main process and are collected separately.

[0025] (2) Conventional regrinding-flotation processes can be used for processing. By adding selective modifiers (such as lime to suppress pyrite) and collectors (such as xanthate), zinc concentrate and sulfur concentrate can be recovered by flotation, thereby achieving targeted recovery and sale of zinc and sulfur resources. This part is a mature mineral processing procedure, and will not be elaborated in this plan.

[0026] Step 3: Low-temperature selective activation calcination (1) Take the calcium-magnesium enriched light product obtained in the first step, concentrate and dehydrate it to a moisture content of <15%. Weigh anhydrous calcium chloride as an additive at 0.5%~1.2% of the dry weight of the light product. Mix the light product and anhydrous calcium chloride thoroughly and evenly in a mixer.

[0027] (2) Feed the mixture into a controlled atmosphere calcination apparatus. Introduce a weakly oxidizing mixed gas containing 1%~5% oxygen by volume. Heat to the target temperature of 700℃~750℃ at a rate of ≤10℃ / min, and hold at this temperature for 60~90min. After the holding period, cool to <100℃ under a weakly oxidizing atmosphere and remove from the furnace to obtain the calcined product, i.e., the activated calcined material.

[0028] (3) The roasted product, when analyzed by XRD, should show clear periclase and calcite main peaks, while the characteristic peak of calcium oxide should be extremely weak or not visible. Thermogravimetric analysis should be used to calculate a magnesium carbonate decomposition rate ≥95% and a calcium carbonate decomposition rate ≤8%.

[0029] CaMg(CO3)2→ MgO·CaCO3+ CO2↑ Step 4: Selective deep leaching with dilute acid (1) The cooled low-temperature selectively activated calcined material obtained in the third step is put into an acid-resistant reactor equipped with stirring and heating functions. The main components of the activated calcined material are porous active magnesium oxide and stable calcium carbonate.

[0030] (2) Use an industrial sulfuric acid solution with a concentration of 1.5~2.5 mol / L. Sulfuric acid is inexpensive, and calcium sulfate has low solubility, which is beneficial for calcium retention. Control the liquid volume to solid mass ratio to be 4:1 to 6:1. This range ensures sufficient reaction space and mass transfer efficiency, and facilitates subsequent solid-liquid separation. Utilize the waste heat of the flue gas to maintain the reaction temperature at 75℃~85℃. Increasing the temperature can significantly accelerate the reaction rate and improve leaching efficiency. Maintain moderate stirring intensity to ensure that the solid particles are suspended and the acid solution is evenly distributed.

[0031] (3) Precise pH control. In the initial stage of leaching, active MgO and CaCO3 rapidly consume the acid, causing the system pH to rise rapidly. As acid is continuously added, MgO and CaCO3 gradually react, and once most of the MgO is neutralized, the pH begins to steadily decrease. Impurities (ZnO, Fe2O3) can dissolve efficiently in the pH range of 4-6, while a large amount of Ca... 2+ The leaching requires a lower pH. An online pH meter is used to monitor the pH of the reaction slurry in real time. Dilute sulfuric acid is added slowly and continuously using a metering pump, or manually in batches with close monitoring. The final leaching endpoint pH is strictly controlled within a narrow window of 4.2–4.8. The reaction endpoint is determined when the pH reaches the set range and remains stable for 15–20 minutes without significant increase. Continue stirring and maintaining the temperature for 30–45 minutes to ensure complete dissolution of zinc and iron oxides. Iron and zinc compounds are then recovered through stepwise precipitation.

[0032] (4) After leaching, immediately perform hot filtration using a plate and frame filter press or vacuum filtration device to prevent impurity ions from hydrolyzing and reprecipitating after the solution cools. Wash the filter residue countercurrently 2-3 times with hot water at 60℃~80℃ to remove entrained soluble sulfates. The wash water can be returned for the preparation of leaching acid or for other treatments. Calcium- and magnesium-rich leaching residue (mainly containing CaCO3, unreacted MgO, and possibly MgSO4) and zinc- and iron-containing leachate are obtained.

[0033] Step 5: Calcium and magnesium separation and product preparation Step 1: Ammonium salt cyclic leaching to separate calcium and magnesium (1) In hot ammonium salt solutions, calcium preferentially forms soluble complexes [Ca(NH3)n]. 2+ It enters the liquid phase, while magnesium precipitates primarily as magnesium hydroxide (Mg(OH)2). The main reactions are as follows: CaCO3+ 2NH4Cl → CaCl2+ 2NH3↑+ CO2↑ + H2O Ca² + + nNH3 [Ca(NH3)n]² + MgO + H2O → Mg(OH)2↓ The washed calcium- and magnesium-rich leaching residue was added to a corrosion-resistant reactor. A 3.0–4.0 mol / L ammonium chloride (NH₄Cl) solution was added as the leaching agent. The liquid-to-solid ratio was controlled at 5:1 to 7:1. Stirring and heating were started, and the reaction temperature was maintained at 85℃–95℃ for 90–120 minutes.

[0034] (2) After the reaction is complete, filter by pressure or vacuum while hot. Filtrate A, rich in calcium chloride and ammonium chloride, and crude magnesium hydroxide filter cake A are obtained.

[0035] Step 2: Preparation of calcium carbonate (carbonation process) (1) Transfer the filtrate to the carbonization reactor. At room temperature or slightly heated (<40℃), introduce CO2 into the solution. Preferably use CO2 from purified flue gas obtained from the low-temperature selective roasting process. Control the aeration rate to ensure a stable reaction and avoid localized over-alkali or over-acidity. Main reactions: CaCl2+ CO2+ 2NH3·H2O → CaCO3↓ + 2NH4Cl + H2O (2) The reaction was monitored using a pH meter and conductivity meter. Carbonation was considered complete when the pH value stabilized at 7.0-7.5 and the conductivity decreased significantly and no longer changed. The slurry was filtered, and the resulting filter cake was washed with deionized water until no chloride ions were present (tested with AgNO3 solution). The filter cake was dried at 105℃-120℃ to obtain high-purity precipitated calcium carbonate product. The mother liquor, mainly ammonium chloride solution, was returned to step one as a leaching agent for recycling, and the evaporation loss was replenished periodically.

[0036] Step 3: Preparation of lightly calcined magnesium oxide (1) The filter cake A (crude magnesium hydroxide) obtained in step one is slurried and washed with hot water at 60℃~80℃ to thoroughly wash away the entrained ammonium chloride and calcium chloride. Filter again to obtain a relatively pure magnesium hydroxide filter cake.

[0037] (2) The washed magnesium hydroxide filter cake is thoroughly dried at 110℃~120℃. The dried magnesium hydroxide is then placed in a calcining furnace and calcined at 800℃~880℃ for 1.5~2.5h. Calcination reaction: Mg(OH)2→ MgO + H2O↑. After cooling, the lightly calcined magnesium oxide product is obtained.

[0038] The embodiments described above are merely illustrative of the technical ideas and features of the present invention, and are intended to enable those skilled in the art to understand the content of the present invention and implement it accordingly. However, they do not limit the patent scope of the present invention. All equivalent changes or modifications made in accordance with the spirit disclosed in the present invention should still be covered within the patent scope of the present invention.

Claims

1. A method for comprehensive utilization of high-calcium-magnesium flotation tailings, characterized in that... The steps are as follows: Step 1: Tailings pretreatment and physical pre-enrichment The flotation tailings were ground to -75μm. Based on the specific gravity difference between dolomite and zinc-containing minerals and pyrite, centrifugal separation was used to separate the minerals, resulting in heavy products in the underflow and light products in the overflow. The heavy products were rich in pyrite and sphalerite, while the light products were mainly dolomite. Step 2: Heavy Product Processing The heavy products obtained from the first centrifugal separation are separately treated for zinc-sulfur separation to achieve targeted recovery of zinc and sulfur resources. Step 3: Low-temperature selective activation calcination The light product obtained from the first centrifugal separation is mixed with the trace additive CaCl2 and calcined at 600℃~800℃ in a weak oxidizing atmosphere containing oxygen. This allows the magnesium carbonate in the dolomite to be selectively decomposed into highly active magnesium oxide, while the calcium carbonate remains stable, resulting in activated calcined material. Step 4: Selective deep leaching with dilute acid The activated roasted material in the third step is leached with dilute acid. By precisely controlling the leaching endpoint, the iron oxide and associated zinc oxide converted from pyrite are efficiently dissolved to obtain a zinc-iron leaching solution. Meanwhile, the calcium component is selectively retained in the slag in the form of calcium carbonate or calcium oxide to obtain calcium-magnesium-rich leaching slag. Step 5: Calcium and magnesium separation and product preparation The calcium- and magnesium-rich leaching residue from the fourth step was separated using an ammonium salt cyclic leaching method. In the hot ammonium salt solution system, calcium preferentially forms a soluble complex and enters the liquid phase, while magnesium precipitates as magnesium hydroxide. After separation, CO2 is introduced into the calcium-containing liquid phase to prepare high-purity calcium carbonate. The magnesium hydroxide is then washed, dried, and calcined to obtain highly active lightly calcined magnesium oxide.

2. The method for comprehensive utilization of high-calcium-magnesium flotation tailings as described in claim 1, characterized in that: The second step involves processing the heavy products obtained in the first step using a regrinding-flotation process. By adding selectivity modifiers and collectors, zinc concentrate and sulfur concentrate are recovered by flotation.

3. The method for comprehensive utilization of high-calcium-magnesium flotation tailings as described in claim 1, characterized in that: The third step involves concentrating and dehydrating the calcium and magnesium enriched light product obtained in the first step until the water content is <15%. Anhydrous calcium chloride is weighed out as an additive at 0.5% to 1.2% of the dry weight of the light product. The light product and anhydrous calcium chloride are then thoroughly and evenly mixed in a mixer. The mixture is fed into a controlled atmosphere calcination device, and a weak oxidizing mixed gas containing 1% to 5% oxygen by volume is introduced. The temperature is raised to the target temperature of 700℃ to 750℃ at a rate of ≤10℃ / min, and held at this temperature for 60 to 90 minutes. After the holding period, the mixture is cooled to <100℃ under a weak oxidizing atmosphere and removed from the furnace to obtain the calcined product, i.e., the activated calcined material.

4. The method for comprehensive utilization of high-calcium-magnesium flotation tailings as described in claim 1, characterized in that: The fourth step involves adding the cooled, low-temperature selectively activated calcined material obtained in the third step into an acid-resistant reactor equipped with stirring and heating functions. Use an industrial sulfuric acid solution with a concentration of 1.5~2.5 mol / L, control the liquid volume to solid mass ratio to be 4:1 to 6:1, and use the waste heat of flue gas to maintain the reaction temperature at 75℃~85℃; Precise pH control: The pH value of the reaction slurry is monitored in real time using an online pH meter. Dilute sulfuric acid is added slowly and continuously using a metering pump, or manually added in batches with close monitoring. The final leaching endpoint pH value is stably controlled within a narrow window of 4.2 to 4.

8. When the pH value reaches the set range and stabilizes for 15 to 20 minutes without significant increase, it is determined to be the reaction endpoint. Continue to keep warm and stir for 30 to 45 minutes to ensure complete dissolution of zinc and iron oxides. Iron and zinc compounds are recovered through stepwise precipitation. After leaching, immediately use a plate and frame filter press or vacuum filtration device for hot filtration to prevent impurity ions from hydrolyzing and reprecipitating after the solution cools. Wash the filter residue countercurrently 2-3 times with hot water at 60℃~80℃ to remove entrained soluble sulfates, and obtain calcium- and magnesium-rich leaching residue and zinc- and iron-containing leachate.

5. The method for comprehensive utilization of high-calcium-magnesium flotation tailings as described in claim 4, characterized in that: In the fourth step, the wash water after leaching is completed is returned for the preparation of leaching acid or for further treatment.

6. The method for comprehensive utilization of high-calcium-magnesium flotation tailings as described in claim 1, characterized in that: In the fifth step, Step 1: Ammonium salt cyclic leaching to separate calcium and magnesium (1) In hot ammonium salt solutions, calcium preferentially forms soluble complexes [Ca(NH3)n]. 2+ It enters the liquid phase, while magnesium precipitates primarily as magnesium hydroxide Mg(OH)2. The main reactions are as follows: CaCO3+ 2NH4Cl → CaCl2+ 2NH3↑ + CO2↑ + H2O Ca² + + nNH3 [Ca(NH3)n]² + MgO + H2O → Mg(OH)2↓ The washed calcium- and magnesium-rich leaching residue was put into a corrosion-resistant reactor, and 3.0-4.0 mol / L ammonium chloride (NH4Cl) solution was added as the leaching agent. The liquid-to-solid ratio was controlled at 5:1 to 7:

1. Stirring and heating were started, and the reaction temperature was maintained at 85℃-95℃ for 90-120 min. (2) After the reaction is complete, filter the solution while it is hot by pressure filtration or vacuum filtration to obtain filtrate A rich in calcium chloride and ammonium chloride and crude magnesium hydroxide filter cake A. Step 2: Preparation of calcium carbonate (1) Transfer the filtrate to the carbonization reactor. At room temperature or slightly heated (<40℃), introduce CO2 into the solution, control the aeration rate to ensure a stable reaction, and avoid local over-alkali or over-acidity. The main reaction is as follows: CaCl2+ CO2+ 2NH3·H2O → CaCO3↓+ 2NH4Cl + H2O (2) The reaction was monitored by pH meter and conductivity meter. When the pH value stabilized at 7.0~7.5 and the conductivity decreased significantly and no longer changed, carbonation was determined to be complete. The slurry was filtered and the resulting filter cake was washed with deionized water until there were no chloride ions. The filter cake was dried at 105℃~120℃ to obtain high-purity precipitated calcium carbonate product. The mother liquor was mainly ammonium chloride solution. Step 3: Preparation of lightly calcined magnesium oxide (1) The filter cake A obtained in step one, i.e. crude magnesium hydroxide, is slurried and washed with hot water at 60℃~80℃ to fully wash away the entrained ammonium chloride and calcium chloride; then filtered again to obtain a relatively pure magnesium hydroxide filter cake. (2) Dry the washed magnesium hydroxide filter cake thoroughly at 110℃~120℃; send the dried magnesium hydroxide into a calcining furnace and calcine at 800℃~880℃ for 1.5~2.5h; calcination reaction: Mg(OH)2→ MgO + H2O↑, after cooling, the lightly calcined magnesium oxide product is obtained.

7. A method for comprehensive utilization of high-calcium-magnesium flotation tailings as described in claim 6, characterized in that: In step two, the preparation of calcium carbonate, the carbonation reactor uses purified flue gas CO2 from the low-temperature selective roasting process.

8. A method for comprehensive utilization of high-calcium-magnesium flotation tailings as described in claim 6, characterized in that: In step two, during the preparation of calcium carbonate, the mother liquor from the slurry after complete carbonation is returned to step one as a leaching agent for recycling, and the evaporation loss is replenished periodically.