Process method for comprehensively recovering copper from iron tailing slag

Abstract

The invention relates to the technical field of comprehensive recovery of mineral resources, in particular to a process method for comprehensively recovering copper from iron tailings, which comprises the following steps: crushing and grinding iron tailings, mixing to obtain ore pulp, and sequentially adding an iron inhibitor and a copper collector into the ore pulp for flotation to obtain copper concentrate; carrying out primary roasting on the copper concentrate in a nitrogen atmosphere, and carrying out ore grinding and magnetic separation on the roasted product to obtain high-grade copper concentrate; mixing the high-grade copper concentrate with sodium chloride or potassium chloride, then carrying out secondary roasting, and carrying out water leaching on the roasted product to obtain a copper-containing leachate; and the copper-containing leachate is subjected to electro-deposition, and deposited metal copper is obtained at a cathode. The invention provides a process method for comprehensively recovering copper from iron tailing slag, which takes phenyl phosphoric acid as an iron inhibitor, and adopts a flotation-roasting-magnetic separation-chloridizing roasting-water leaching-electrodeposition full-flow process to comprehensively recover copper from the iron tailing slag, so that efficient enrichment of copper resources in the iron tailing slag is realized.

Description

Technical Field

[0001] This invention relates to the field of comprehensive mineral resource recovery technology, and is a process for comprehensively recovering copper from iron tailings slag. Background Technology

[0002] Iron tailings are a large amount of solid waste generated during iron ore beneficiation. my country's accumulated iron tailings stockpile exceeds several billion tons, and continues to grow by hundreds of millions of tons annually. Currently, iron tailings are mainly disposed of through stockpiling, which not only occupies a large amount of land resources but also prevents the effective utilization of valuable metals such as iron, copper, and zinc contained within, resulting in resource waste and potential environmental risks.

[0003] The copper grade in iron tailings is generally low, and the copper grade in some tailings is close to that of low-grade copper ore. Moreover, copper minerals are mostly present in the form of fine-grained intergrowth or complex intergrowth with iron minerals. It is difficult to achieve efficient separation of copper minerals from iron minerals and gangue minerals by using a single physical beneficiation method.

[0004] Currently, there are various technical routes for recovering copper from secondary resources such as iron tailings. Based on processing principles, these can be divided into five main categories: physical beneficiation, roasting-magnetic separation, hydrometallurgy, pyrometallurgy, and combined processes. Physical beneficiation, with flotation as its core, traditional processes commonly use xanthate and black reagent collectors along with depressants such as lime, which suffers from insufficient collector selectivity and problems such as high lime consumption and scaling. Roasting-magnetic separation technology achieves separation by altering the magnetic properties of minerals through reduction roasting or magnetization roasting. Recent developments include microwave-assisted selective heating to reduce energy consumption and the addition of calcium-based additives to disrupt gangue structure and promote the dissociation of target minerals. Hydrometallurgy primarily uses acid leaching and ammonia leaching, followed by extraction and electrowinning to prepare cathode copper. Related cutting-edge technologies include microbial leaching, mechanically activated enhanced leaching, and composite acid leaching for treating complex materials. Pyrometallurgy, such as oxygen-enriched smelting and flash smelting, is mainly used for processing high-grade copper sulfide concentrates. Cutting-edge directions include short-process technologies and smelting modification technologies, which promote metal droplet settling by adjusting slag composition. Combined processes often employ combinations such as flotation-roasting-leaching and magnetic separation-flotation. Cutting-edge developments are trending towards systematic design, such as graded flotation to process materials of different particle sizes, coupling pyrometallurgy-hydrometallurgy-mineral processing to achieve cascade extraction of multiple valuable elements, and co-processing of solid waste to build a resource recycling network.

[0005] Wang Changtao (Wang Changtao. Fundamental and Applied Research on Flotation of Magnetore-Rich Copper Sulfide Ores Based on Slurry Aeration Regulation [D]. Central South University, 2023.) proposed a new low-carbon flotation method based on slurry aeration regulation to address the problem of good floatability and difficulty in suppressing pyrrhotite in pyrrhotite-rich copper sulfide ores. This method selectively oxidizes the surface of pyrrhotite under low-alkaline conditions (pH≈8~9) through slurry aeration, causing it to form a hydrophilic oxide layer for passivation. Combined with the highly selective copper collector ML8 (strong collecting ability for chalcopyrite, weak collecting ability for pyrrhotite and pyrite), efficient separation of chalcopyrite from pyrrhotite and pyrite is achieved. Industrial application results show that at Zhurihe Copper Industry and Dongguashan Copper Mine, copper recovery rates increased from 82.08% and 79.85% to 86.55% and 87.76%, respectively, while significantly reducing the amount of lime and sulfuric acid used, thus lowering reagent costs and environmental pollution. However, this method has high requirements for equipment and control, has limited effect on inhibiting pyrite, requires optimization of process parameters according to local conditions, and still requires the use of environmentally harmful agents such as lime and sulfuric acid.

[0006] Guo Weiyang (Guo Weiyang. Discussion on the extraction technology of precious metals from slag waste [J]. World Nonferrous Metals, 2025, (06): 214-216.) pointed out that hydrometallurgical technology is a wet metallurgical process using water as a medium. Its core lies in selectively dissolving precious metals in slag using solutions such as acids, alkalis, or cyanides, so that they enter the liquid phase in the form of complexes. Subsequently, the precious metals are separated from the solution by means of solvent extraction, ion exchange, or activated carbon adsorption, and finally, high-purity elemental metals are obtained by precipitation or electrolytic reduction. However, its main challenge is that the leaching agent is dangerous, and environmental risks need to be strictly controlled. At the same time, the reaction time is relatively long, and the process control requirements are relatively precise.

[0007] Wang Na (Wang Na. Study on the recovery of copper from acidic etching wastewater by electrolytic deposition [D]. Dalian Jiaotong University, 2022.) recovered copper from acidic etching wastewater of printed circuit boards by electrolytic deposition. Using a self-designed through-type coaxial electrode electrolysis device (stainless steel cathode, graphite anode), the effects of electrolysis voltage, electrolysis time, initial copper concentration, and initial pH on copper removal rate and product quality were systematically investigated. The results showed that under optimal process conditions (electrolysis voltage 7V, initial copper concentration 30 g / L, initial pH 4.0, electrolysis time 3h), the copper removal rate reached 99.45%. The recovered copper product was mainly elemental copper with a face-centered cubic structure, exhibiting a microstructure that transitioned from dense blocky to dendritic, with particle sizes ranging from 37.45 to 57.77 nm, and a copper content reaching a maximum of 88.91%. However, the electrolysis process consumes a lot of energy, and the products may contain impurities such as cuprous chloride or cuprous oxide. In addition, other metal ions in the actual wastewater may compete with each other, affecting the purity of copper and the recovery efficiency. Further optimization of the process is still needed to adapt to complex wastewater systems.

[0008] Overall, existing technologies for recovering copper from iron tailings slag generally suffer from problems such as low copper recovery rate, low concentrate grade, and incomplete process flow, making it difficult to meet the actual needs of efficient resource utilization of iron tailings slag. Summary of the Invention

[0009] This invention provides a process for the comprehensive recovery of copper from iron tailings slag, which overcomes the shortcomings of the prior art and can effectively solve the problems of low recovery rate, low concentrate grade, and incomplete process flow in the recovery of copper from iron tailings slag in the prior art.

[0010] One of the technical solutions of this invention is achieved through the following measures: a process for comprehensively recovering copper from iron tailings slag, comprising the following steps: Step 1: After crushing and grinding the iron ore tailings, the slurry is prepared. Iron inhibitor and copper collector are added to the slurry in sequence for flotation to obtain copper concentrate. Step 2: The copper concentrate is roasted once under a nitrogen atmosphere. The roasted product is then ground and magnetically separated to obtain high-grade copper concentrate. Step 3: High-grade copper concentrate is mixed with sodium chloride or potassium chloride and then roasted a second time. The roasted product is then leached with water to obtain a copper-containing leachate. Step four: Electrodeposition is performed on the copper-containing leaching solution to obtain deposited metallic copper at the cathode.

[0011] The following are further optimizations and / or improvements to one of the above-mentioned inventive technical solutions: In step one above, the copper collector is one or more of sodium butyl xanthate, butylated black powder, and ethyl thiocyanate.

[0012] In step one above, the ratio of the amount of copper collector added to the iron ore tailings is 50 g / t to 120 g / t.

[0013] In step one above, the iron inhibitor is phenylphosphine.

[0014] In step one above, the ratio of the amount of iron inhibitor added to the iron ore tailings in the slurry is 10 g / t to 60 g / t.

[0015] In step one above, the pulp pH is 5 to 7, the pulp concentration is 40 g / L to 80 g / L, the number of flotation times is 2 to 7, and the particle size of the iron tailings slag after grinding is 100 mesh to 200 mesh.

[0016] In step two above, the heating program for the first calcination is to heat to 400°C to 900°C at a rate of 5°C / min to 10°C / min, hold for 0.5h to 2.5h, and use nitrogen at a flow rate of 0.3L / min to 0.5L / min.

[0017] In step three above, the mass ratio of high-grade copper concentrate to sodium chloride or potassium chloride during the secondary roasting is 1:1 to 3:1, the secondary roasting temperature is 300℃ to 500℃, and the roasting time is 1h to 2h.

[0018] In step three above, the water immersion temperature is 60℃ to 100℃, the immersion time is 1h to 2h, and the liquid-to-solid ratio of the water immersion is 5:1 to 10:1.

[0019] In step four above, electrodeposition uses a three-electrode system: the working electrode is porous carbon KNHC-3, the auxiliary electrode is a carbon rod, the reference electrode is Ag / AgCl, the deposition potential is -0.5V to -1V, the deposition time is 500s to 1000s, the deposition temperature is 15℃ to 35℃, and no stirring is performed.

[0020] This invention provides a process for the comprehensive recovery of copper from iron tailings slag. It uses phenylphosphoric acid as an iron inhibitor and adopts a full-process process of flotation-roasting-magnetic separation-chlorination roasting-water leaching-electrodeposition to comprehensively recover copper from iron tailings slag, achieving efficient enrichment of copper resources in iron tailings slag. Attached Figure Description

[0021] Appendix Figure 1 This is the current-time curve (IT curve) of the electrodeposition process in Embodiment 11 of the present invention.

[0022] Appendix Figure 2 This is the mapping spectrum of the metallic copper sample obtained in Example 11 of the present invention. Detailed Implementation

[0023] This invention is not limited to the following embodiments, and specific implementation methods can be determined according to the technical solutions and actual conditions of this invention. Unless otherwise specified, all chemical reagents and chemicals mentioned in this invention are well-known and commonly used chemical reagents and chemicals in the prior art; unless otherwise specified, all percentages in this invention are mass percentages; unless otherwise specified, all solutions in this invention are aqueous solutions with water as the solvent, for example, hydrochloric acid solution is an aqueous solution of hydrochloric acid; room temperature in this invention generally refers to a temperature between 15°C and 25°C, generally defined as 25°C.

[0024] The present invention will be further described below with reference to embodiments: Example 1: The process for comprehensively recovering copper from iron tailings slag includes the following steps: Step 1: After crushing and grinding the iron ore tailings, the slurry is prepared. Iron inhibitor and copper collector are added to the slurry in sequence for flotation to obtain copper concentrate. Step 2: The copper concentrate is roasted once under a nitrogen atmosphere. The roasted product is then ground and magnetically separated to obtain high-grade copper concentrate. Step 3: High-grade copper concentrate is mixed with sodium chloride or potassium chloride and then roasted a second time. The roasted product is then leached with water to obtain a copper-containing leachate. Step four: Electrodeposition is performed on the copper-containing leaching solution to obtain deposited metallic copper at the cathode.

[0025] Example 2: As an optimization of the above example, in step one, the copper collector is one or more of sodium butyl xanthate, butylammonium black powder, and ethyl thiocyanate.

[0026] Example 3: As an optimization of the above example, in step one, the ratio of the amount of copper collector added to the iron ore tailings in the slurry to the mass of 50 g / t to 120 g / t.

[0027] Example 4: As an optimization of the above example, in step one, the iron inhibitor is phenylphosphine.

[0028] Example 5: As an optimization of the above example, in step one, the ratio of the amount of iron inhibitor added to the iron ore tailings in the slurry is 10 g / t to 60 g / t.

[0029] Example 6: As an optimization of the above example, in step one, the slurry pH is 5 to 7, the slurry concentration is 40 g / L to 80 g / L, the number of flotation times is 2 to 7, and the particle size of the iron tailings slag after grinding is 100 mesh to 200 mesh.

[0030] Example 7: As an optimization of the above example, in step two, the heating program for the first calcination is to heat to 400°C to 900°C at a rate of 5°C / min to 10°C / min, hold for 0.5h to 2.5h, and use nitrogen at a flow rate of 0.3L / min to 0.5L / min.

[0031] Example 8: As an optimization of the above example, in step three, the mass ratio of high-grade copper concentrate to sodium chloride or potassium chloride during secondary roasting is 1:1 to 3:1, the secondary roasting temperature is 300℃ to 500℃, and the roasting time is 1h to 2h.

[0032] Example 9: As an optimization of the above example, in step three, the water immersion temperature is 60°C to 100°C, the immersion time is 1h to 2h, and the liquid-solid ratio of the water immersion is 5:1 to 10:1.

[0033] Example 10: As an optimization of the above example, in step four, the electrodeposition adopts a three-electrode system, the working electrode is porous carbon KNHC-3, the auxiliary electrode is a carbon rod, the reference electrode is Ag / AgCl, the deposition potential is -0.5V to -1V, the deposition time is 500s to 1000s, the deposition temperature is 15℃ to 35℃, and no stirring is performed.

[0034] Example 11: A process for the comprehensive recovery of copper from iron tailings slag, comprising the following steps: Step 1: Flotation enrichment Iron tailings slag was crushed, screened, and slurry-adjusted to obtain a slurry with a concentration of 50 g / L. This slurry was then subjected to flotation in a flotation machine at a slurry temperature of 25 ± 1℃, a stirring speed of 2000 r / min, and an aeration rate of 0.2 m³ / min. 3 / (m 2 After stirring for 2 minutes, the pH was adjusted to 6. Then, 50 g / t of phenyl phosphate and 100 g / t of ethyl thiocyanate were added sequentially. The reaction time of each reagent was 2 minutes, the frothing time was 4 minutes, and the flotation was performed 3 times to obtain copper concentrate. The calculated yield was 8.35%, and the copper grade was 8469 mg / kg.

[0035] Step 2: Nitrogen roasting-magnetic separation purification 50g of copper concentrate was weighed and roasted in a tubular roasting furnace. Nitrogen gas (flow rate 0.5L / min) was introduced to replace the air in the furnace, and the temperature was increased to 600℃ at a rate of 10℃ / min, held for 60min, and then cooled to room temperature with the furnace. The roasted product was ground and then subjected to wet magnetic separation to obtain high-grade copper concentrate. Analysis showed that the copper content of the high-grade copper concentrate was 10368mg / kg, which was 22.4% higher than that of the copper concentrate obtained through flotation.

[0036] Step 3: Chlorination roasting - water immersion High-grade copper concentrate was mixed with sodium chloride or potassium chloride at a mass ratio of 1:1 and placed in a corundum boat. The mixture was heated to 400°C at a rate of 10°C / min under air atmosphere and held at that temperature for 90 min. The roasted product was leached with deionized water at 90°C for 60 min at a liquid-to-solid ratio of 7:1, and the solution was filtered to obtain a copper-containing leachate.

[0037] Step 4: Electrodeposition to recover copper The electrodeposition experiment used a three-electrode system: the working electrode was porous carbon KNHC-3, the auxiliary electrode was a carbon rod, the reference electrode was Ag / AgCl, the electrolyte was a copper-containing leaching solution, the deposition potential was -0.8V, the deposition time was 900s, the temperature was 25℃, and no stirring was performed.

[0038] During electrodeposition, the current-time curve (IT curve, see electrochemical workstation) is recorded. Figure 1 The results showed that the current decayed rapidly in the initial stage (0-300s), slowed down in the middle stage (300-600s), and stabilized in the final stage (600-900s). After deposition, the cathode plate was removed, washed with deionized water, and dried at room temperature. A uniform, dense, reddish-brown deposition layer covered the surface of the cathode plate, with a bright color and local areas exhibiting the typical rose-red luster of metallic copper. The deposition layer was tightly bonded to the porous carbon, and no peeling or lifting was observed after rinsing with deionized water and natural drying, indicating that metallic copper has good adhesion to the cathode.

[0039] The electrodeposition products were subjected to mapping analysis to obtain elemental surface distribution analysis maps (see...). Figure 2 This spectrum was obtained through surface scanning analysis of the cathode deposition layer using scanning electron microscopy combined with energy dispersive spectroscopy (SEM-EDS), visually demonstrating the distribution of major elements in the deposition product. The image shows that copper is uniformly distributed throughout the scanned area, without significant segregation or local enrichment, indicating a stable electrodeposition process and uniform reduction deposition of copper ions on the cathode surface. Simultaneously, the carbon signal mainly corresponds to the porous carbon KNHC-3 cathode substrate, forming a clear spatial stratification with the copper deposition layer, proving good adhesion of metallic copper to the cathode surface. Furthermore, no significant enrichment of impurities such as iron and chlorine was detected in the spectrum, indicating that copper and impurities were efficiently separated after chlorination roasting-water leaching, resulting in a high-purity electrodeposition product. The mapping results are consistent with the description of the deposition layer exhibiting a rose-red metallic luster and tight bonding, further verifying the effectiveness and stability of the "chlorination roasting-water leaching-electrodeposition" process constructed in this invention for preparing high-purity metallic copper.

[0040] Example 12: A process for the comprehensive recovery of copper from iron tailings slag, comprising the following steps: Step 1: Flotation enrichment Iron tailings slag was crushed, screened, and slurry-adjusted to obtain a slurry with a concentration of 80 g / L. This slurry was then subjected to flotation in a flotation machine at a slurry temperature of 25 ± 1℃, a stirring speed of 2000 r / min, and an aeration rate of 0.2 m³ / min. 3 / (m 2 After stirring for 2 minutes, the pH was adjusted to 7. Then, 60 g / t of phenyl phosphate and 120 g / t of ethyl thiocyanate were added sequentially. The reaction time of each reagent was 2 minutes, the frothing time was 4 minutes, and the flotation was performed twice to obtain copper concentrate. The calculated yield was 9.15%, and the copper grade was 8106 mg / kg.

[0041] Step 2: Nitrogen roasting-magnetic separation purification Copper concentrate was roasted in a tubular roaster. Nitrogen gas (flow rate 0.5 L / min) was introduced to replace the air in the furnace, and the temperature was increased to 900℃ at a rate of 5℃ / min, held for 120 min, and then cooled to room temperature with the furnace. The roasted product was ground and then subjected to wet magnetic separation to obtain high-grade copper concentrate. Analysis showed that the copper content of the high-grade copper concentrate was 11059 mg / kg, which was 36.4% higher than that of the copper concentrate obtained through flotation.

[0042] Step 3: Chlorination roasting - water immersion High-grade copper concentrate was mixed with sodium chloride or potassium chloride at a mass ratio of 3:1 and placed in a corundum boat. The mixture was heated to 500°C at a rate of 10°C / min under air atmosphere and held for 120 min. The roasted product was leached with deionized water at 100°C for 120 min at a liquid-to-solid ratio of 10:1. The leaching solution was then filtered to obtain a copper-containing leachate.

[0043] Step 4: Electrodeposition to recover copper The electrodeposition experiment used a three-electrode system: the working electrode was porous carbon KNHC-3, the auxiliary electrode was a carbon rod, the reference electrode was Ag / AgCl, the electrolyte was a copper-containing leaching solution, the deposition potential was -1V, the deposition time was 1000s, the temperature was 25℃, and no stirring was performed.

[0044] Example 13: A process for the comprehensive recovery of copper from iron tailings slag, comprising the following steps: Step 1: Flotation enrichment Iron tailings slag was crushed, screened, and slurry-adjusted to obtain a slurry with a concentration of 40 g / L. This slurry was then subjected to flotation in a flotation machine at a temperature of 25 ± 1℃, a stirring speed of 2000 r / min, and an aeration rate of 0.2 m³ / min. 3 / (m 2 After stirring for 2 minutes, the pH was adjusted to 5. Then, 40 g / t of phenyl phosphate and 80 g / t of ethyl thiocyanate were added sequentially. The reaction time of each reagent was 2 minutes, the frothing time was 4 minutes, and the flotation was performed 7 times to obtain copper concentrate. The calculated yield was 4.41%, and the copper grade was 8904 mg / kg.

[0045] Step 2: Nitrogen roasting-magnetic separation purification 50g of copper concentrate was weighed and roasted in a tubular roasting furnace. Nitrogen gas (flow rate 0.5L / min) was introduced to replace the air in the furnace, and the temperature was increased to 400℃ at a rate of 5℃ / min, held for 150min, and then cooled to room temperature with the furnace. The roasted product was ground and then subjected to wet magnetic separation to obtain high-grade copper concentrate. Analysis showed that the copper content of the high-grade copper concentrate was 9964mg / kg, an increase of 11.9% compared to the copper concentrate obtained through flotation.

[0046] Step 3: Chlorination roasting - water immersion High-grade copper concentrate was mixed with sodium chloride or potassium chloride at a mass ratio of 2:1 and placed in a corundum boat. The mixture was heated to 300°C at a rate of 10°C / min under air atmosphere and held for 60 min. The roasted product was leached with deionized water at 80°C for 90 min at a liquid-to-solid ratio of 5:1. The leaching solution was then filtered to obtain a copper-containing leachate.

[0047] Step 4: Electrodeposition to recover copper The electrodeposition experiment used a three-electrode system: the working electrode was porous carbon KNHC-3, the auxiliary electrode was a carbon rod, the reference electrode was Ag / AgCl, the electrolyte was a copper-containing leaching solution, the deposition potential was -0.5V, the deposition time was 500s, the temperature was 25℃, and no stirring was performed.

[0048] Example 14: The difference from Example 11 is that in step one, sodium butyl xanthate and butylammonium black powder (mass ratio 2:1) are used to replace ethylthiocyanate, and the total addition amount of sodium butyl xanthate and butylammonium black powder is 120 g / t. In this example, the flotation yield after step one is 5.63%, and the copper grade of the obtained copper concentrate is 6160 mg / kg.

[0049] Example 15: The difference from Example 11 is that in step one, sodium butyl xanthate and butylammonium black powder (mass ratio 2:1) replaced ethylthiocyanate. The total addition amount of sodium butyl xanthate and butylammonium black powder was 120 g / t, and the flotation was performed twice. In this example, the flotation yield after step one was 7.59%, and the copper grade of the obtained copper concentrate was 5573 mg / kg.

[0050] Example 16: The difference from Example 11 is that in step one, sodium butyl xanthate and butylammonium black powder (mass ratio 2:1) replaced ethylthiocyanate. The total addition amount of sodium butyl xanthate and butylammonium black powder was 120 g / t, and the pulp pH was 7. In this example, the flotation yield after step one was 5.94%, and the copper grade of the obtained copper concentrate was 6253 mg / kg.

[0051] Example 17: The difference from Example 11 is that in step one, sodium butyl xanthate and butylammonium black powder (mass ratio 2:1) replaced ethylthiocyanate. The total addition amount of sodium butyl xanthate and butylammonium black powder was 120 g / t, and the pulp concentration was 80 g / L. In this example, the flotation yield after step one was 26.07%, and the copper grade of the obtained copper concentrate was 5369 mg / kg.

[0052] Comparative Example 1: Unlike Example 11, no phenylphosphine inhibitor was added. In this example, the flotation yield after step one was 18.36%, and the copper concentrate obtained had a copper content of 3596 mg / kg.

[0053] Comparative Example 2: Unlike Example 11, lime was used as an iron inhibitor instead of phenylphosphoric acid. The results showed that the copper flotation grade was 4359 mg / kg, lower than 8469 mg / kg in Example 11, and the lime consumption was high, leading to scaling and a poorer operating environment.

[0054] Comparative Example 3: Unlike Example 11, in step one, carboxymethyl cellulose was used as an iron inhibitor instead of phenylphosphine. The results showed that the copper flotation grade was 3568 mg / kg, lower than the 8469 mg / kg in Example 11. Furthermore, carboxymethyl cellulose exhibited poor dispersibility in the pulp, easily leading to flocculation and decreased flotation foam stability, resulting in significant fluctuations in copper recovery. This demonstrates that the phenylphosphine system possesses both selective inhibition and foam control effects in copper-iron separation.

[0055] The raw material, iron tailings from a beneficiation plant in Xinjiang, used in Examples 11 to 17 and Comparative Examples 1 to 3, was analyzed by ICP-MS of magnetite tailings. The mass fraction of Cu was 0.05% to 0.14%. Copper mainly exists as a complex mineral of copper sulfide and iron sulfide, with a relatively fine particle size. XRD analysis of the magnetite tailings showed that the minerals in the tailings mainly exist in the form of quartz, pyrite, and chlorite.

[0056] Compared with the prior art, the process for comprehensive copper recovery from iron tailings slag of the present invention has the following beneficial effects: This invention is the first to use phenylphosphoric acid as a highly selective inhibitor for iron-bearing gangue. The phosphonic acid group (-PO(OH)2) in the phenylphosphoric acid molecule forms a stable coordination bond with iron ions on the surface of iron minerals, forming a hydrophilic adsorption layer on the gangue surface and preventing collector adsorption. However, its adsorption capacity on the surface of copper minerals is weak, and it does not affect the collecting effect of the collector. This achieves highly efficient separation of copper minerals by "floating and inhibiting iron minerals," increasing the flotation concentrate grade by more than 26% compared to the traditional lime system. The use of phenylphosphoric acid as an iron flotation inhibitor, synergistically with the collector, significantly improves the flotation selectivity and concentrate grade of copper.

[0057] In terms of process integration, this invention constructs a complete process of "flotation-roasting-magnetic separation-chlorination roasting-water leaching-electrodeposition", which deeply integrates physical mineral processing with pyrometallurgy / hydrometallurgy, and directly produces metallic copper deposits from iron tailings slag, realizing a one-step conversion from waste to metal products, which surpasses the limitation of existing technologies that produce more intermediate products such as copper concentrate.

[0058] Regarding the deep separation mechanism, this invention adopts a tiered separation strategy: flotation achieves selective separation of iron tailings slag; nitrogen roasting-magnetic separation utilizes thermal decomposition to convert copper sulfide minerals into Cu2S and iron minerals into magnetite, achieving synergy between chemical transformation and physical separation; chlorination roasting-water leaching converts copper into soluble chlorides that enter the solution, achieving further separation of copper from other components step by step.

[0059] In terms of the final product form, this invention produces a metallic copper deposition layer that can be directly used as a raw material for electrolytic copper production. The product has high purity and a rose-red appearance, realizing the transformation of low-value solid waste into high-value-added metal products. Through a strategy combining key reagent innovation with systematic process integration, this invention outperforms existing methods in terms of selectivity, recovery rate, product purity, and environmental friendliness, demonstrating significant innovation and promising industrial application prospects.

[0060] In summary, this invention employs a synergistic flotation process using iron inhibitor phenylphosphine and copper collector, significantly improving the selectivity of copper-iron separation and resulting in copper concentrate grade superior to traditional reagent systems. Through a cascade purification process involving flotation, nitrogen roasting-magnetic separation, chlorination roasting-water leaching, and electrodeposition, the grade and recovery rate of copper are effectively increased. The resulting metallic copper deposit is dense, uniform, and of high purity, making it suitable for direct use as a raw material for electrolytic copper production. The entire process is environmentally friendly, free of toxic heavy metal reagents, and features a complete, simple, and easily industrialized process, providing an efficient and feasible technical solution for the resource utilization of iron tailings slag.

[0061] The above technical features constitute various embodiments of the present invention, which have strong adaptability and implementation effect. Unnecessary technical features can be added or removed according to actual needs to meet the needs of different situations.

Claims

1. A process for the comprehensive recovery of copper from iron tailings slag, characterized in that... Includes the following steps: Step 1: After crushing and grinding the iron ore tailings, the slurry is prepared. Iron inhibitor and copper collector are added to the slurry in sequence for flotation to obtain copper concentrate. Step 2: The copper concentrate is roasted once under a nitrogen atmosphere. The roasted product is then ground and magnetically separated to obtain high-grade copper concentrate. Step 3: High-grade copper concentrate is mixed with sodium chloride or potassium chloride and then roasted a second time. The roasted product is then leached with water to obtain a copper-containing leachate. Step four: Electrodeposition is performed on the copper-containing leaching solution to obtain deposited metallic copper at the cathode.

2. The process for comprehensively recovering copper from iron tailings slag according to claim 1, characterized in that... In step one, the copper collector is one or more of sodium butyl xanthate, butylammonium black powder, and ethyl thiocyanate.

3. The process for comprehensively recovering copper from iron tailings slag according to claim 1 or 2, characterized in that... In step one, the ratio of the amount of copper collector added to the slurry to the mass of iron ore tailings is 50 g / t to 120 g / t.

4. The process for comprehensively recovering copper from iron tailings slag according to any one of claims 1 to 3, characterized in that... In step one, the iron inhibitor is phenylphosphine.

5. The process for comprehensively recovering copper from iron tailings slag according to any one of claims 1 to 4, characterized in that... In step one, the ratio of the amount of iron inhibitor added to the iron ore tailings in the slurry is 10 g / t to 60 g / t.

6. The process for comprehensively recovering copper from iron tailings slag according to any one of claims 1 to 5, characterized in that... In step one, the pulp pH is 5 to 7, the pulp concentration is 40 g / L to 80 g / L, and the number of flotation times is 2 to 7; or / and, the particle size of the iron tailings after grinding is 100 mesh to 200 mesh.

7. The process for comprehensively recovering copper from iron tailings slag according to any one of claims 1 to 6, characterized in that... In step two, the heating program for the first calcination is to raise the temperature to 400°C to 900°C at a rate of 5°C / min to 10°C / min, hold for 0.5h to 2.5h, and use nitrogen at a flow rate of 0.3L / min to 0.5L / min.

8. The process for comprehensively recovering copper from iron tailings slag according to any one of claims 1 to 7, characterized in that... In step three, the mass ratio of high-grade copper concentrate to sodium chloride or potassium chloride during the secondary roasting is 1:1 to 3:1, the secondary roasting temperature is 300℃ to 500℃, and the roasting time is 1h to 2h.

9. The process for comprehensively recovering copper from iron tailings slag according to any one of claims 1 to 8, characterized in that... In step three, the water immersion temperature is 60℃ to 100℃, the immersion time is 1h to 2h, and the liquid-to-solid ratio of the water immersion is 5:1 to 10:

1.

10. The process for comprehensively recovering copper from iron tailings slag according to any one of claims 1 to 9, characterized in that... In step four, electrodeposition uses a three-electrode system: the working electrode is porous carbon KNHC-3, the auxiliary electrode is a carbon rod, the reference electrode is Ag / AgCl, the deposition potential is -0.5V to -1V, the deposition time is 500s to 1000s, the deposition temperature is 15℃ to 35℃, and no stirring is performed.

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