Arsenic-zinc polymetallic ore processing method using waste residue

Abstract

This invention relates to a method for treating arsenic-zinc polymetallic ores using waste residue, applicable to complex polymetallic sulfide ores with high pyrrhotite and arsenopyrite content, and dense intergrowth of sphalerite and chalcopyrite. This method utilizes the synergistic effects of modified calcium carbide slag slow-release pH adjustment, hot-dip galvanizing slag acid leaching solution replacing zinc sulfate, optimized ratio of sodium humate-citric acid-water glass for synergistic inhibition, EDTA-disodium complexing activator, sodium carbide-CMC-guar gum organic inhibitor, and sweet potato starch-guar gum selective flocculant. The total lime consumption throughout the process is reduced to 1300-2000 g / t, calcium carbide slag utilization is improved, zinc sulfate raw material costs are reduced, and the arsenic content in zinc concentrate is reduced to below 0.3%, with a zinc grade >48% and an arsenic concentrate grade >40%. This achieves the beneficial effects of low alkali consumption, high selectivity, high comprehensive recovery rate, and high-value utilization of industrial solid waste.

Description

Technical Field

[0001] This invention belongs to the field of mineral processing technology, and specifically relates to a method for processing arsenic-zinc polymetallic ores using waste residue. Background Technology

[0002] A primary sulfide polymetallic quartz vein-type arsenic-zinc deposit in Inner Mongolia contains chalcopyrite and sphalerite as its core valuable metal minerals, along with recoverable components such as silver and sulfur. The main harmful impurities are arsenopyrite, pyrrhotite, and pyrite. The original production process of the concentrator was a single-stage grinding-step flotation process: the raw ore was ground to -0.074mm (65%–68%), followed by two copper roughing processes, and then further grinding and cleaning to produce copper concentrate. The copper flotation tailings were then subjected to zinc roughing, multiple scavenging, and multiple cleaning processes to produce zinc concentrate. The zinc flotation tailings were then further ground and flotated to produce arsenic concentrate. This process revealed several core technical defects during actual production and experimental research, specifically: high consumption of high-alkali reagents, with the total lime consumption exceeding 4000g / t throughout the entire process. The high-alkali slurry system presents challenges for the subsequent activation of zinc minerals. Arsenic flotation requires the addition of large amounts of strong acid to neutralize the high-alkali slurry, resulting in high costs for reagent procurement, wastewater treatment, and equipment maintenance, leading to extremely poor economic efficiency. The zinc concentrate exhibits excessive arsenic content, poor zinc-sulfur-arsenic separation, and significant fluctuations and instability in zinc concentrate grade. Due to the intergrowth of fine-grained arsenopyrite (0.05mm) with sphalerite, the direct flotation concentrate grade is only 20%–32%. The reagent system suffers from poor selectivity and insufficient adaptability: the use of traditional single inhibitors and collectors lacks specificity for inhibiting pyrrhotite and arsenopyrite, resulting in incomplete separation of copper-arsenic, zinc-arsenic, and zinc-sulfur, leading to low overall recovery rates of valuable metals. Furthermore, raw material costs are high; lime and other raw materials requiring low impurities and high quality are necessary, and raw materials such as zinc sulfate are also expensive.

[0003] To address the aforementioned shortcomings, there is an urgent need to develop a treatment method with low alkali consumption and low cost. Summary of the Invention

[0004] To overcome the shortcomings of existing technologies, this application provides a method for processing arsenic-zinc polymetallic ores, addressing problems such as high lime consumption and price, high arsenic content in zinc concentrate, low grade of arsenic concentrate, and difficulty in separating sphalerite and pyrrhotite. This invention reduces costs by utilizing waste residue resources such as calcium carbide slag and hot-dip galvanizing slag in weakly alkaline copper flotation, and achieves highly efficient and selective separation of complex minerals under low-alkali conditions through the synergistic effect of a multi-reagent system.

[0005] The embodiments of this application are implemented as follows: In a first aspect, this application provides a method for processing arsenic-zinc polymetallic ore containing waste residue, comprising the following steps: S1. Staged grinding and classification of arsenic-zinc polymetallic ore; wherein the raw ore is subjected to two stages of crushing, followed by a first stage of grinding and a first stage of classification using a first hydrocyclone, with the underflow returned to the first stage of grinding and the overflow entering a weak-alkali copper flotation operation; the copper flotation tailings enter a second stage of grinding and a second stage of classification using a second hydrocyclone, with the underflow returned to a second stage ball mill for further grinding and the overflow entering a zinc-sulfur flotation operation; S2. The overflow slurry from the first stage classification is subjected to weak-alkali copper flotation to obtain low-arsenic copper concentrate and copper. Tailings; S3. The second stage graded overflow slurry is aerated and activated and semi-flotated. After slurry conditioning, semi-flotation roughing is performed to obtain semi-flotation froth and semi-flotation tailings; S4. Pyrrhotite is magnetically separated and recovered as sulfur concentrate product, yielding magnetic tailings; the semi-flotation froth is processed by S5 and S6 to obtain high-grade low-arsenic zinc concentrate; S5. The semi-flotation froth undergoes a first zinc-sulfur separation to obtain zinc rough concentrate and first sulfur concentrate; S6. The zinc rough concentrate undergoes a second zinc-sulfur separation to obtain high-grade low-arsenic zinc concentrate; S7. Fine-particle arsenic flocculation flotation is used to treat the magnetic tailings to obtain high-grade arsenic concentrate; S2. Weak Alkali Copper Floating: A pH adjuster, a pyrite composite inhibitor, a zinc mineral composite inhibitor, and a copper composite collector are added to the first-stage classification overflow slurry. Copper roughing, two copper scavenging processes, and three copper cleaning processes are performed to obtain low-arsenic copper concentrate and copper tailings. The pH adjuster is a combination of modified calcium carbide slag and lime. The modified calcium carbide slag is calcium carbide slag that has been wet-milled to a particle size of -0.074mm or higher (85%), prepared into a slurry with a mass concentration of 30%~40%, aged for 24~48 hours, and then mixed with sodium humate at a dry basis mass ratio of (5~8):1. The process involves premixing and stirring at 40-60℃ for 30-60 min to obtain a slow-release calcium carbide slag loaded with sodium humate; the pyrite-based inhibitor is a combination of sodium humate, citric acid, and water glass in a mass ratio of (2.5-3.5):(3.5-4.5):(1.5-2.5); the zinc mineral inhibitor is a combination of zinc sulfate, sodium sulfite, and sodium carbonate, wherein the zinc sulfate is obtained by leaching hot-dip galvanized slag with dilute sulfuric acid and oxidizing to remove iron; and the copper collector is a combination of ethyl thiocyanate, isobutyl xanthate, and anisole. The specific steps for leaching and oxidizing hot-dip galvanized slag with dilute sulfuric acid to remove iron are as follows: crush the hot-dip galvanized slag to -2mm, add dilute sulfuric acid with a concentration of 10% to 15% at a liquid-solid ratio of 3:1 to 5:1, stir and leach for 1 to 2 hours at 40 to 60℃, and filter to obtain the leachate; add hydrogen peroxide to the leachate to oxidize and remove iron to obtain a zinc sulfate solution.

[0006] Optionally, in step S2, the pH of the slurry is controlled to be 7.9-8.3. The pH adjuster is a mixture of modified carbide slag and lime, with a mass ratio of 5-8:1 and a total dosage of 500-800 g / t. The pyrite combination inhibitor is a composition of sodium humate, citric acid, and water glass, with a mass ratio of (2.5-3.5):(3.5-4.5):(1.5-2.5), and a total dosage of 250-400 g / t. The zinc ore... The zinc mineral combination inhibitor is a combination of zinc sulfate, sodium sulfite, and sodium carbonate in a mass ratio of (6~9):(1.5~3):(0.3~1); the total dosage of the zinc mineral combination inhibitor is 300~500 g / t; the copper combination collector is a combination of ethyl thiocyanate, isobutyl xanthate, and anisole in a mass ratio of (2.5~4):(2.5~4):(1~1.5); the total dosage of the copper combination collector is 50~90 g / t.

[0007] Optionally, S1. Staged grinding and classification of arsenic-zinc polymetallic ore: The raw ore is crushed to ≤10mm in two stages, and then ground in the first stage until the fineness of -0.074mm accounts for 70%~75%. The ore is then classified in the first stage by a first hydrocyclone. The underflow is returned to the first stage grinding, and the overflow enters the copper flotation operation. The copper flotation tailings enter the second stage grinding to the fineness of -0.074mm ≥80%. The ore is then classified in the second stage by a second hydrocyclone. The underflow is returned to the second stage ball mill for re-grinding, and the overflow enters the zinc-sulfur flotation operation.

[0008] Optionally, S3. Slurry aeration activation and semi-flotation: The overflow slurry from the second stage of classification is pumped into an aeration mixing tank and aerated with air; then, a first pyrite inhibitor, a complexing activator, a sphalerite activator, a sphalerite selective collector, and a frother are added. After slurry conditioning, semi-flotation roughing is performed to obtain semi-flotation froth and semi-flotation tailings; the first pyrite inhibitor is calcium hypochlorite, with a dosage of 200~400 g / t; the complexing activator is EDTA-disodium, with a dosage of 50~100 g / t; the sphalerite activator is copper sulfate, with a dosage of 100~200 g / t; the sphalerite selective collector is a complex of 2-mercaptobenzothiazole and isobutyl xanthate propylene ester, with a mass ratio of (1~2):(2~3); the dosage of the sphalerite selective collector is 20~40 g / t.

[0009] Optionally, S5. First zinc-sulfur separation: Add a second pyrite inhibitor, a sphalerite activator, and a sphalerite selective collector to the semi-mixed froth obtained in step S3 to perform the first zinc-sulfur separation roughing to obtain zinc rough concentrate and first sulfur concentrate; the second pyrite inhibitor is lime, with a dosage of 1000~1500 g / t; the sphalerite activator is copper sulfate, with a dosage of 100~150 g / t; the sphalerite selective collector is a complex of 2-mercaptobenzothiazole and isobutyl xanthate propylene ester, with a mass ratio of (1~2):(2~3), and a dosage of 10~20 g / t.

[0010] Optionally, S6. Second zinc-sulfur separation: The zinc concentrate obtained in step S5 is regrinded to a fineness of -0.044 mm accounting for 85%~90%, and then an organic combination inhibitor, auxiliary collector, and frother are added for a second zinc-sulfur separation. After two scavenging and three cleaning processes, a high-grade, low-arsenic zinc concentrate is obtained. The organic combination inhibitor is a mixture of sodium cyanurate carboxymethyl cellulose and guar gum in a mass ratio of (3~4):(3~4):(2~3). The auxiliary collector is a combination of thiocyanate and No. 25 black powder in a mass ratio of (1~1.5):(1~2). The total amount of the organic combination inhibitor is 350~600 g / t. The total amount of the auxiliary collector is 40~100 g / t. The pH adjuster is lime, with a dosage of 600~900 g / t. The pH of the pulp is controlled at 10.5~11.5.

[0011] Optionally, S7. Fine-particle arsenic flocculation flotation: The magnetic separation tailings obtained in step S4 are regrinded to a fineness of -0.038mm accounting for 80%~90%, concentrated to a concentration of 25%~30%, and then a slime dispersant, an arsenic mineral surface modifier, a selective flocculant, a second pH adjuster, and a tailings collector are added sequentially for arsenic roughing, two scavenging, and three cleaning processes to obtain a high-grade arsenic concentrate; the slime dispersant is a mixture of sodium pyrophosphate and water glass, with sodium pyrophosphate at a dosage of 30~50g / t and water glass at a dosage of 300~500g / t; the arsenic mineral surface modifier is wood The lignin sulfonate and gallic acid complex is used in a mixture of sodium lignin sulfonate and gallic acid, with sodium lignin sulfonate at a dosage of 40-60 g / t and gallic acid at a dosage of 20-30 g / t. The selective flocculant is a mixture of sweet potato starch and guar gum, with sweet potato starch at a dosage of 90-120 g / t and guar gum at a dosage of 30-40 g / t. The tailings collector is a combination of butyl xanthate, No. 25 black powder, and ethyl thiocyanate, with butyl xanthate at a dosage of 15-25 g / t, No. 25 black powder at a dosage of 15-30 g / t, and ethyl thiocyanate at a dosage of 5-10 g / t. The second pH adjuster is lime, which controls the pH of the slurry to 11-12.

[0012] Optionally, in step S1, the fineness of the first stage of graded overflow is -0.074mm ≥ 55% and the concentration is ≥ 35%; the fineness of the second stage of graded overflow is -0.074mm ≥ 75% and the concentration is ≥ 35%.

[0013] Optionally, S4. Magnetic separation and recovery of pyrrhotite: The semi-mixed tailings obtained in step S3 are subjected to weak magnetic separation to recover pyrrhotite as a sulfur concentrate product; the magnetic separation tailings are then used for subsequent arsenic recovery operations; the magnetic field strength is 0.1~0.2T.

[0014] Optionally, in step S3, the pH of the slurry is controlled to be 6.0~8.2, and aeration treatment is performed for 15~30 min with an aeration rate of 1.0~2.0 m³ / min; in step S7, the pH of the slurry is controlled to be 11.5~12.0.

[0015] The present invention provides a processing method for arsenic-zinc polymetallic ores, which has the following beneficial effects: (1) Modified slow-release calcium carbide slag achieves both alkali adjustment and inhibition, transforming calcium carbide slag from solid waste into a highly efficient flotation reagent with both slow-release pH adjustment and synergistic inhibition functions. The modified calcium carbide slag has fine particle size and high activity, and the alkalinity release is uniform and controllable, which can maintain the pH stability of copper flotation; at the same time, the loaded sodium humate is slowly released synchronously with the alkalinity, forming a continuously renewed hydrophilic chelate film on the surface of pyrite, which is conducive to achieving pH adjustment and pyrite inhibition, while avoiding the waste of reagents and abnormal increase in pulp viscosity caused by excessive addition of sodium humate.

[0016] (2) The amount of lime used is significantly reduced. The total amount of lime used in the entire process is reduced from more than 5,000 g / t in the traditional process to 1,300-2,000 g / t, a reduction of more than 60%. Through the slow-release buffering effect of the modified carbide slag-lime composite pH adjuster, the phenomenon of local over-alkaliness is further avoided.

[0017] (3) Replacing zinc sulfate with hot-dip galvanized slag acid leaching solution to reduce costs and increase efficiency: Hot-dip galvanized slag is leached with dilute sulfuric acid and oxidized with hydrogen peroxide to remove iron, and then replaced with industrial zinc sulfate as a zinc mineral composite inhibitor. Since trace amounts of Al in the hot-dip galvanized slag leaching solution have no negative impact on sphalerite inhibition, and can even synergistically enhance the reducing atmosphere with sodium sulfite, the cost of zinc sulfate raw materials is reduced. At the same time, it broadens the resource utilization pathways of zinc-containing solid waste and realizes the combination of mineral processing technology and solid waste treatment. Detailed Implementation

[0018] The technical solution of the present invention will be clearly and completely described below with reference to the embodiments. Obviously, the described embodiments are only some embodiments of the present invention, and not all embodiments. Based on the embodiments of the present invention, all other embodiments obtained by those skilled in the art without creative effort are within the scope of protection of the present invention.

[0019] Those skilled in the art will understand that the names of the chemical substances and preparation methods involved in the treatment methods are all known prior art, may be commercially available, may include relevant substances in patent applications, and the detection methods may refer to industry or national standards, which will not be described in detail here.

[0020] A polymetallic arsenic-zinc deposit in Inner Mongolia is a quartz vein-type primary sulfide ore. The ore's valuable metallic minerals are mainly chalcopyrite and sphalerite, with recoverable components such as silver and sulfur as byproducts. Harmful impurities are mainly arsenopyrite, pyrrhotite, and pyrite (e.g., the main metallic minerals are pyrrhotite (15%), arsenopyrite (8%), sphalerite (7%, containing 12%~15% Fe), chalcopyrite (3.5%), and pyrite (5%)). Multi-element chemical analysis of the ore shows: Cu 1.17%, Zn 4.66%, As 5.40%, S 7.82%, Fe 15.22%, Ag 94.00 g / t. Sulfur and iron contents are relatively high, and the mineral distribution is extremely complex. The iron sphalerite is severely altered and intergrowthed with chalcopyrite, arsenopyrite, and pyrrhotite. Fine-grained chalcopyrite is mostly embedded along the fissures and edges of the iron sphalerite, and some forms a solid solution separation structure. Arsenopyrite is mostly in the form of subhedral granules, intertwined with sulfide minerals and gangue minerals. It is difficult to separate individual minerals and belongs to a typical difficult-to-process high-arsenic polymetallic sulfide mineral.

[0021] To address the problems of high lime consumption, high arsenic content in zinc concentrate, low grade of arsenic concentrate, difficulty in separating sphalerite and pyrrhotite, and high cost in existing technologies, this invention provides a method for treating arsenic-containing zinc polymetallic ores using waste residue.

[0022] For example, a method for processing arsenic-zinc polymetallic ore is provided, comprising the following steps: S1. Segmented grinding and classification of arsenic-zinc polymetallic ore; S2. Weak alkali flotation of copper to obtain low-arsenic copper concentrate and copper tailings; S3. Slurry aeration activation and semi-mixed flotation to obtain semi-mixed flotation foam and semi-mixed flotation tailings; S4. Magnetic separation recovery of pyrrhotite to obtain magnetic separation tailings; Semi-mixed flotation foam is treated by S5 and S6 to obtain high-grade low-arsenic zinc concentrate, where S5 is the first zinc-sulfur separation; S6 is the second zinc-sulfur separation; S7. Fine-particle arsenic flocculation flotation treatment of magnetic separation tailings to obtain high-grade arsenic concentrate.

[0023] Among them, S1. Segmented grinding and classification of arsenic-zinc polymetallic ores: The raw ore is crushed to ≤10mm in two stages, and then ground in the first stage to a fineness of -0.074mm of 70%~75%. It is then classified in the first stage by a first hydrocyclone. The underflow is returned to the first stage of grinding, and the overflow (fineness of -0.074mm ≥55%, concentration ≥35%) enters the copper preferential flotation operation. The copper preferential flotation tailings enter the second stage of grinding to a fineness of -0.074mm ≥80%. It is then classified in the second stage by a second hydrocyclone. The underflow is returned to the second stage ball mill for regrinding, and the overflow (fineness of -0.074mm ≥75%, concentration ≥35%) enters the zinc-sulfur asynchronous flotation operation.

[0024] S2. Weak alkali preferentially floats copper: pH adjuster, pyrite assemblage inhibitor, zinc mineral assemblage inhibitor and copper assemblage collector were added to the first stage of graded overflow slurry to control the slurry pH to 7.9~8.3. Copper roughing, two copper scavenging and three copper cleaning were carried out to obtain low arsenic copper concentrate and copper tailings. in: The pH adjuster is a mixture of modified carbide slag and lime, with a mass ratio of 5~8:1 and a total dosage of 500~800g / t; the modified carbide slag is carbide slag that has been wet-milled to a particle size of -0.074mm accounting for more than 85%, prepared into a slurry with a mass concentration of 30%~40%, aged for 24~48h, and then premixed with sodium humate at a dry basis mass ratio of (5~8):1, and stirred and reacted at 40~60℃ for 30~60min to obtain a slow-release carbide slag loaded with sodium humate; The pyrite-based inhibitor is a composition of sodium humate, citric acid, and water glass (sodium silicate) in a mass ratio of (2.5~3.5):(3.5~4.5):(1.5~2.5), with a total dosage of 250~400 g / t, and the water glass modulus is 2.2~3.0. The zinc mineral combination inhibitor is a combination of zinc sulfate, sodium sulfite, and sodium carbonate in a mass ratio of (6~9):(1.5~3):(0.3~1), with a total dosage of 300~500 g / t. The zinc sulfate is obtained by leaching hot-dip galvanized slag with dilute sulfuric acid and then oxidizing to remove iron. The copper combination collector is a combination of ethyl thiocyanate, isobutyl xanthate, and anisole. The specific steps for leaching hot-dip galvanized slag with dilute sulfuric acid and oxidizing to remove iron are as follows: the hot-dip galvanized slag is crushed to -2 mm, and dilute sulfuric acid with a concentration of 10%~15% is added at a liquid-to-solid ratio of 3:1~5:1. The leaching is carried out by stirring at 40~60℃ for 1~2 hours, and the leachate is filtered to obtain the leachate. A zinc sulfate solution is obtained by oxidizing and removing iron with hydrogen peroxide in the leachate. The amount of hydrogen peroxide added is not specifically limited here, but depends on the Fe content in the leachate. 2+ Adjusting the amount of Fe in the leachate can improve the Fe content. 2+ 1.0 to 2.5 times the quality.

[0025] The copper-collecting agent is a mixture of ethyl thiocyanate, isobutyl xanthate and anisole in a mass ratio of (2.5~4):(2.5~4):(1~1.5), with a total dosage of 50~90 g / t.

[0026] Understandably: pH adjuster: In a weakly alkaline copper flotation system (pH 7.9~8.3), lime (CaO) alone poses a risk of localized over-alkalinity (micro-region pH > 9, leading to inhibition of sphalerite), while calcium carbide slag (Ca(OH)2) alone adjusts alkali slowly. Calcium carbide slag (proportionately used): mainly contains Ca(OH)2, with lower solubility than lime, exhibiting slow-release alkalinity characteristics. It can form an alkalinity buffer in the slurry, helping to maintain pH stability within 7.9~8.3±0.2 fluctuations and finely adjust the slurry pH. Simultaneously, calcium carbide slag helps activate chalcopyrite and inhibits the oxidation of pyrite surface to form hydrophobic polysulfides, reducing pyrite floatability. However, due to the large particle size and low activity of calcium carbide slag, and the presence of undigested CaO and trace amounts of CaC2, direct addition poses a safety hazard due to a sudden increase in localized alkalinity and the generation of acetylene gas. After wet ball milling to a particle size of -0.074 mm (over 85%), the specific surface area increased, and the reactivity significantly improved. Aging for 24-48 hours allowed residual CaO to fully hydrate and generate Ca(OH)₂, while trace amounts of CaC₂ were completely hydrolyzed, eliminating acetylene gas and resulting in a more uniform and controllable alkalinity release. Sodium humate is rich in phenolic hydroxyl and carboxyl groups, which bond with OH groups on the surface of calcium carbide slag via hydrogen bonds. - Its function, and through the carboxyl group and Ca 2+ Coordination bonds are formed, firmly loading the calcium carbide slag particles onto the surface and within their pores. This modified calcium carbide slag, loaded with sodium humate, facilitates the slow dissolution and release of OH groups from the calcium carbide slag particles after being added to the slurry. - The pH is maintained stable at 7.9~8.3±0.2. Simultaneously, sodium humate is slowly released along with the dissolution of the calcium carbide slag, forming a continuously renewing chelated hydrophilic film and a carboxyl electrostatic dispersion layer on the pyrite surface, achieving a better balance between alkali adjustment and inhibition. Compared to directly adding calcium carbide slag and sodium humate separately, this method improves the inhibition effect. Furthermore, because the modified calcium carbide slag already contains some sodium humate, the proportion of sodium humate in the separately added pyrite-based inhibitor can be adjusted accordingly, avoiding excessive addition of sodium humate which could lead to reagent waste and abnormally high slurry viscosity.

[0027] The main components of hot-dip galvanizing slag are metallic zinc, ZnO, and a small amount of Zn-Fe alloy. It has a high zinc grade and few impurities. After leaching with dilute sulfuric acid, the solution mainly contains Zn. 2+ Fe 2+ Because the flotation inhibitor does not require high purity zinc sulfate, residual Fe... 2+ It does not affect the sphalerite suppression effect, and even Fe 2+ It can synergistically enhance the reducing atmosphere with sodium sulfite. Therefore, it is even better to use hydrogen peroxide to remove Fe. 2+ Oxidized to Fe3+ Post-hydrolysis precipitation removes all or part of the iron, eliminating the need for complex purification processes such as zinc powder replacement, significantly simplifying the process and reducing costs. The leachate can be directly used as a source of zinc sulfate for flotation, increasing the utilization of hot-dip galvanizing slag. By using the above leaching process conditions—dilute sulfuric acid, suitable particle size, temperature, and reaction time—the content of impurities such as Al and Fe in the leached hot-dip galvanizing slag is reduced, thus avoiding affecting the utilization value of the leachate as zinc sulfate.

[0028] Lime (small proportion): provides rapid alkali adjustment capability, compensates for the slow initial reaction of carbide slag, and ensures that the pH of the roughing feed reaches above 7.9 quickly.

[0029] The two complement each other through the slow-release buffering and reduction inhibition of modified carbide slag and the rapid alkali adjustment of lime. This avoids local over-alkali inhibition of zinc minerals and inhibits pyrite oxidation through reducing components, thus reducing the amount of subsequent inhibitors required.

[0030] Pyrite inhibitor combination: A combination of sodium humate, citric acid, and water glass for inhibition. Sodium humate (moderate proportion, partially carried by modified carbide slag): As a natural macromolecular humic substance, it is rich in phenolic hydroxyl groups (-OH) and carboxyl groups (-COOH). The phenolic hydroxyl groups react with Fe on the pyrite surface. 3+ A stable five-membered ring chelate (similar to tannins but at a lower cost and with better results) is formed, creating an organic hydrophilic film on the surface of pyrite. At the same time, sodium humate generates a negative charge through carboxyl ionization, electrostatically dispersing fine gangue mud and eliminating its covering of arsenopyrite / chalcopyrite.

[0031] Citric acid (relatively high proportion): As a small-molecule tricarboxylic acid, it penetrates into the microcracks on the surface of pyrite, undertaking more surface modification and ion masking functions, and interacting with the internal Fe... 2+ / Fe 3+ The formation of carboxyl-iron coordination compounds alters the surface zeta potential, enhancing hydrophilicity; citric acid may react with free Cu. 2+ Complexation, blocking Cu 2+ Activation of pyrite (Cu) 2+ It is the main activator of pyrite, which helps to reduce the floating of pyrite after it is activated by copper ions.

[0032] Water glass (added in small amounts, modulus 2.2~3.0): Under pH 7.9~8.3, it partially hydrolyzes to generate orthosilicic acid (H4SiO4) and oligomeric silicate colloids (SiO2·nH2O), which are adsorbed onto the surface of pyrite through hydrogen bonding and van der Waals forces, forming an amorphous silicate gel layer that acts as a physical barrier to prevent collectors from forming. Simultaneously, water glass preferentially adsorbs onto the surface of silicate gangue (quartz, sericite), preventing its heterogeneous aggregation with arsenopyrite / chalcopyrite through homogeneous repulsion. The combination of water glass dispersed in gangue and sodium humate dispersed in slime helps reduce the non-selective consumption of collectors.

[0033] Sodium humate provides steric hindrance and dispersion in the outer layer, citric acid provides chemical modification and ion masking in the inner layer, and water glass provides an inorganic physical barrier. The three form a composite inhibition, which improves the inhibition efficiency and reduces the cost.

[0034] Zinc mineral combination inhibitor: A combination of zinc sulfate, sodium sulfite, and sodium carbonate is used for inhibition.

[0035] To inhibit iron sphalerite ((Zn,Fe)S) and zinc sphalerite (ZnS), under weakly alkaline conditions (pH 7.9~8.3): Zinc sulfate (the largest proportion) provided by hot-dip galvanizing slag leaching solution: provides Zn 2+ It partially hydrolyzes to form Zn(OH) at pH 7.9–8.3. + With trace amounts of Zn(OH)2, it readily adsorbs onto the surface of sphalerite to form a hydrophilic layer; simultaneously, Zn 2+ Fe in the crystal lattice 2+ Ion exchange occurs, occupying active sites and hindering subsequent Cu... 2+ activation.

[0036] Sodium sulfite (medium proportion): as a reducing agent (E°(SO3) 2- / S)=-0.93V), inhibiting the oxidation of sphalerite surface: In a weakly alkaline environment, sphalerite is easily oxidized by dissolved oxygen to generate hydrophobic elemental sulfur (ZnS+O2→ZnO+S). 0 Sodium sulfite consumes dissolved oxygen (2SO3). 2- +O2→2SO4 2- This reaction is blocked; at the same time, SO3 2- With residual Cu in the slurry 2+ Forming stable complexes and eliminating Cu 2+ Activation of sphalerite; citric acid complexation of Cu 2+ Cu complexed with sodium sulfite 2+ Synergistically, it completely eliminates the activation of pyrite and sphalerite by copper ions, reducing the floating of impurities.

[0037] Sodium carbonate (added in small amounts): introduces CO3. 2- , with Zn 2+ ZnCO3 precipitate is formed, with a solubility product lower than that of Zn(OH)2, etc. A Zn(OH)2 / ZnCO3 composite hydrophilic film is constructed on the surface of sphalerite, which is more stable and denser than that of zinc hydroxide alone; in addition, sodium carbonate complexes with Ca... 2+ (To prevent CaCO3 from depositing on the surface of sphalerite) and Cu 2+ It works synergistically with sodium sulfite to enhance the masking of copper ions.

[0038] Zinc sulfate provides a basic hydrophilic layer, sodium carbonate enhances the stability of the inhibition membrane by generating more insoluble zinc carbonate, and sodium sulfite provides a reducing environment to prevent membrane damage. The three form a precipitation-antioxidant-complex complex inhibition. When the pH is <8.5 (to avoid excessive dissolution of Zn(OH)2), the inhibition efficiency of sphalerite is improved compared with the traditional high-alkali system, and the amount of activator used in subsequent zinc flotation is saved.

[0039] Copper-based collector: a combination of ethyl thiocyanate, isobutyl xanthate, and anisole sulfide.

[0040] In a weakly alkaline environment (pH 7.9~8.3), chalcopyrite (CuFeS2) undergoes moderate oxidation to form a Cu2S / CuS active film, while pyrite undergoes oxidation to form a Fe(OH)3 hydrophilic film. The collector needs to achieve strong collection of chalcopyrite and selective collection of weakly reacting pyrite. Ethyl thiocyanate: Contains a sulfur-carbonyl group (C=S) and an ether oxygen bond (COC). In a weakly alkaline environment, Cu on the chalcopyrite surface... + It forms a stable five-membered chelate ring (Cu-SCNO) with the sulfur carbonyl group (S atom), exhibiting high binding energy and strong selectivity for chalcopyrite; however, it is less selective for pyrite (which has no Cu on its surface). + Only Fe 3+ C=S and Fe 3+ It has weak complexing ability and hardly adsorbs.

[0041] Isobutyl xanthate: The hydrophobic group is isobutyl, which is more hydrophobic than ethyl but less hydrophobic than n-butyl. It replaces sulfur on the chalcopyrite surface via ion exchange. 2- The formation of Cu(C4H9OCS2)2 provides basic collecting power; its selectivity lies in its affinity for chalcopyrite (containing Cu). + Strong adsorption, especially for pyrite (containing Fe). 3+ It reacts with xanthate to form unstable Fe(C4H9OCS2)3, which is easily hydrolyzed, resulting in weak adsorption and low cost, thus achieving cost reduction and efficiency improvement.

[0042] Benzyl sulfide: Contains a benzene ring (C6H5-) and a thioether bond (-S-CH3). Thioethers undergo accelerated hydrolysis under alkaline conditions, and are also catalyzed by metal ions in minerals. The thioether bond slowly hydrolyzes in weakly alkaline slurry, releasing trace amounts of HS-. - This process repairs oxidation defects on the chalcopyrite surface, restoring its hydrophobicity. The steric hindrance of the benzene ring (phenyl diameter 0.6 nm) prevents its adsorption on the tightly packed pyrite surface, further enhancing selectivity.

[0043] Ethylthiocyanate provides chemical chelation selectivity (to Cu) +(Specific recognition), isobutyl xanthate provides ion exchange coverage (reducing costs and treating coarse copper), and anisole provides hydrophobic enhancement (repairing oxide films and enhancing the stability of hydrophobic layers), forming a composite adsorption, which improves copper recovery compared to a single xanthate system and reduces sulfur content in copper concentrate.

[0044] S3. Slurry aeration activation and semi-flotation: The second stage overflow slurry is pumped into an aeration mixing tank, and air is introduced for aeration treatment. The aeration time is 15-30 minutes, and the aeration rate is 1.0-2.0 m³ / min. 3 / min; then add first pyrite inhibitor, complexing activator, sphalerite activator, sphalerite selective collector and frother, and after slurry conditioning, carry out semi-mixed flotation roughing to obtain semi-mixed flotation foam and semi-mixed flotation tailings; The first pyrite inhibitor is calcium hypochlorite, used at a dosage of 200-400 g / t; the complexing activator is EDTA-disodium, used at a dosage of 50-100 g / t; the sphalerite activator is copper sulfate, used at a dosage of 100-200 g / t; the sphalerite selective collector is a complex of 2-mercaptobenzothiazole (MBT) and isobutyl xanthate propylene, with a mass ratio of (1-2):(2-3), used at a dosage of 20-40 g / t; the slurry pH is controlled at 6.0-8.2, preferably 6.0-7.0. The foaming agent can be, for example, methyl isobutyl methanol, No. 2 oil, etc., used at a dosage of 0-25 g / t.

[0045] Understandably, aeration oxidizes pyrrhotite, which is beneficial for the subsequent inhibition of calcium hypochlorite. Aeration increases the Fe content on the surface of sphalerite. 2+ Moderate oxidation occurs, but the generated Fe(OH)3 hinders activation. EDTA-disodium (50~100 g / t) selectively chelates Fe on the surface of iron sphalerite. 3+ It may form [Fe(EDTA)] - The complex detaches from the surface, eliminating Fe 3+ It hinders activation while facilitating the production of Zn. 2 + Vacancies, followed by Cu in copper sulfate 2+ When combined with copper sulfate, a CuS activation film is formed, which improves the activation efficiency compared to copper sulfate alone.

[0046] Calcium hypochlorite: Calcium hypochlorite (Ca(ClO)2) reacts with water to form ClO - Beneficial for oxidizing Fe on the surface of pyrrhotite 2+ →Fe 3+ This generates Fe(OH)3 / FeOOH hydrophilic oxides; while iron sphalerite, having already undergone sulfidation and copperization and with its surface cleaned by EDTA-disodium, is not easily oxidized, thus facilitating asynchronous separation.

[0047] The selective collector for sphalerite is a complex of 2-mercaptobenzothiazole (MBT) and propylene isobutyl xanthate. 2-Mercaptobenzothiazole (MBT): Contains a benzothiazole ring, providing strong hydrophobicity; the thiol group (-SH) and the nitrogen atom on the thiazole ring can react with Cu. + A stable quaternary chelate ring is formed. After activation with copper sulfate, a CuS / Cu2S film forms on the surface of sphalerite, which is easily oxidized to a CuO / Cu(OH)2 hydrophilic film in alkaline slurry. MBT binds to Cu through its thiol groups. 2+ Forms a dense hydrophobic protective film (MBT-Cu complex layer) to prevent OH from forming. - The erosion of the CuS film by O2 maintains its active state; the hydrophobicity of the benzothiazole ring imparts an initial hydrophobic layer to the mineral surface, creating conditions for subsequent adsorption of xanthate collectors; and the oxidation of pyrrhotite (surface Fe) 3+ It has almost no adsorption and maintains high selectivity.

[0048] Isobutyl xanthate propylene ester: xanthate group (ROCSS) in the molecule - ) for Cu 2+ It exhibits strong complexing ability, and the unsaturated double bonds of the propenyl group enhance its chemisorption stability on the copper sulfide surface. This reagent is effective against Cu... 2+ The complexing ability is much higher than that for Fe. 3 + The complexation of Fe on the surface of pyrrhotite allows for efficient collection of iron sphalerite activated by copper sulfate, while oxidized pyrrhotite is collected due to the high efficiency of Fe on its surface. 3+ It cannot form a stable chelate with the collector and thus does not float.

[0049] S4. Magnetic separation and recovery of pyrrhotite: The semi-mixed tailings obtained in step S3 are subjected to weak magnetic separation (magnetic field strength 0.1~0.2T) to recover pyrrhotite as sulfur concentrate product; the magnetic separation tailings are then used for subsequent arsenic recovery operations.

[0050] S5. First zinc-sulfur separation (separation of pyrite and sphalerite): A second pyrite inhibitor, a sphalerite activator, and a sphalerite selective collector are added to the semi-mixed froth obtained in step S3 for the first zinc-sulfur separation roughing to obtain a zinc rough concentrate and a first sulfur concentrate (pyrite concentrate). The second pyrite inhibitor is lime, with a dosage of 1000-1500 g / t; the sphalerite activator is copper sulfate, with a dosage of 100-150 g / t; the sphalerite selective collector is a complex of 2-mercaptobenzothiazole (MBT) and isobutyl xanthate propylene ester, with a mass ratio of (1-2):(2-3), and a dosage of 10-20 g / t. This operation adopts an open-loop circulation, and the process water is independently circulated and not returned to the subsequent sulfur flotation system.

[0051] S6. Second zinc-sulfur separation (deep separation of sphalerite and residual pyrite): The zinc concentrate obtained in step S5 is regrinded to a fineness of -0.044 mm, accounting for 85%~90%. Then, an organic combination inhibitor, auxiliary collector, and frother are added for a second zinc-sulfur separation. After two scavenging and three cleaning processes, a high-grade, low-arsenic zinc concentrate (Zn grade >48%, As <0.5%) is obtained. Frothers can be, for example, methyl isobutyl methanol, No. 2 oil, etc., with a dosage of 0~25 g / t. The specific content is added as needed and is not specifically limited here.

[0052] The organic combination inhibitor is a mixture of sodium cyanurate, sodium carboxymethyl cellulose (CMC) and guar gum in a mass ratio of (3~4):(3~4):(2~3), with a total dosage of 350~600g / t; The collector is a combination of thiocyanate and No. 25 black powder (butyl sodium black powder) in a mass ratio of (1~1.5):(1~2), with a total dosage of 40~100g / t; Control the pH of the slurry to 10.5~11.5 and use 600~900g / t of lime.

[0053] Understandably, for the deep separation of sphalerite and pyrite, sodium cyanurate carbide (30%~40% by mass, based on 100% of the component) facilitates the formation of strong bonds between its surface groups and pyrite, while forming weak bonds with sphalerite, achieving selective inhibition due to coordination differences. Maintaining this component at a dosage of 120~200 g / t ensures targeted inhibition of pyrite.

[0054] Sodium carboxymethyl cellulose (CMC, 30%–40% by mass): A carboxymethyl etherified polysaccharide with cellulose as its backbone rather than starch. Its cellulose molecular chains are more linear and regular (compared to starch and similar substances). It exhibits higher viscosity stability and stronger film-forming properties in an alkaline environment of pH 10.5–11.5, and is less prone to starch gelatinization and degradation. Its carboxymethyl (-CH2COO) content... - It can react with Fe on the surface of pyrite. 3+ It forms ionic and coordinate bonds, and due to the rigidity of the cellulose chain, it provides a better steric hindrance effect than starch, thus more effectively preventing collectors from approaching the pyrite surface.

[0055] Guar gum (20%–30% by mass): a natural polymer of galactomannan with extremely high viscosity (a 1% solution can reach 5000–6000 mPa·s). Guar gum forms a dense hydrophilic gel film by forming a hydrogen-bonded hydration layer with the surface of pyrite through numerous hydroxyl groups (-OH) on its molecular chain. Its high viscosity enhances the viscosity of the slurry, increases the hydrodynamic repulsion between pyrite particles and air bubbles, and reduces mechanical entrainment. Guar gum is an extract from legumes, widely available, inexpensive, and has a fast dissolution rate, making it more suitable for industrial applications. In summary, sodium cyanurate carbide and pyrite Fe... 3+ Strong coordination is formed, which facilitates the fixation of inhibitor molecules on the mineral surface; the linear cellulose chains of CMC bind to Fe through carboxyl groups. 3+ The guar gum, with its rigid chain segments extending laterally, provides a mechanical barrier to prevent collector penetration. Its long-chain, flexible polysaccharides interweave within the CMC backbone, forming a three-dimensional network through hydrogen bonds. This network locks in water molecules, creating a high-viscosity hydration film that physically prevents bubble adhesion. Under the shearing force of the flotation machine, it becomes more stable, ensuring that pyrite is firmly suppressed when sphalerite (with a copper-activated hydrophobic surface) floats, thus improving separation efficiency and reducing reagent costs.

[0056] Auxiliary collector: Thioamino ester combined with No. 25 black drug.

[0057] For the harvesting of iron sphalerite (with a CuS film on the surface after activation by copper sulfate), thiourethane: its molecule contains a sulfur-carbonyl group (C=S) and an ether oxygen bond (COC) that reacts with the CuS film on the surface of iron sphalerite. + It forms a stable five-membered chelate ring (Cu-SCNO), which specifically recognizes Cu activation sites and attracts Fe from the surface of unactivated pyrite. 3+ Almost no adsorption.

[0058] The industrial standard reagent No. 25 (black powder) was selected to achieve cost reduction while maintaining selectivity. No. 25 is a dithiophosphate collector; the two butoxy groups (C4H9O-) in its molecule provide moderate steric hindrance, while the dithiophosphate groups ((RO)2PSS) - ) for Cu + The complexation stability constant is comparable to that of conventional agents such as xanthates, but for Fe... 3+ The affinity of [agent name] is significantly weaker than that of xanthate (due to the difference in electron cloud density between P=S and C=S, and greater steric hindrance). Therefore, [agent name] No. 25 exhibits strong adsorption on the surface of copper-activated sphalerite, but weaker adsorption on the surface of pyrite (even with slight oxidation), showing superior selectivity compared to conventional xanthates such as butyl xanthate, and at a lower cost. [Agent name] occupies the flat area of ​​the sphalerite surface via thiourethane (with a more compact molecular structure). 2+ Sites, forming a hydrophobic monolayer bonded to No. 25 black powder (with a relatively open molecular structure containing two hydrophobic groups), fill the surface micro-dimples and lattice defects with Cu.2+ The site cross-links with the adsorbed thiouric acid ester molecular chains through hydrophobic association, forming a denser and more stable hydrophobic multilayer film, thereby enhancing the bubble mineralization efficiency.

[0059] When the mass ratio of the two reagents is controlled at (1~1.5):(1~2), the recovery rate of iron sphalerite can be guaranteed while the collection of residual pyrite can be significantly reduced (because the two reagents affect the Fe). 3+ Both are weakly adsorbed, and pyrite has been deeply inhibited by sodium cyanurate carbide, etc.

[0060] The slurry pH is 10.5~11.5, and the lime dosage is 600~900 g / t. At this pH, both CMC and guar gum maintain stable viscosity and will not undergo molecular chain degradation due to excessive alkalinity (pH>12) (especially since guar gum is easily hydrolyzed in strong alkalis). This also facilitates the improvement of the reaction between sodium cyanurate carbonization and Fe. 3+ Coordination selectivity; Thioamino ester and No. 25 black medicine at this pH for Cu 2+ The chelating ability is maintained.

[0061] S7. Fine-particle arsenic flocculation flotation: The magnetic separation tailings obtained in step S4 are regrinded to a fineness of -0.038 mm accounting for 80%~90%, concentrated to a concentration of 25%~30%, and then mineral slime dispersant, arsenic mineral surface modifier, selective flocculant, pH adjuster, and tailings collector are added in sequence. Arsenic roughing, two scavenging and three cleaning processes are carried out to obtain high-grade arsenic concentrate (As grade >40%). The sludge dispersant is a mixture of sodium pyrophosphate and water glass, with sodium pyrophosphate at 30~50g / t and water glass at 300~500g / t; The arsenic mineral surface modifier is a complex of sodium lignosulfonate and gallic acid, with sodium lignosulfonate at 40-60 g / t and gallic acid at 20-30 g / t. The selective flocculant is a mixture of sweet potato starch and guar gum, with 90-120 g / t of sweet potato starch and 30-40 g / t of guar gum; The pH adjuster is 800~1200g / t of lime, controlling the pH to be 11~12; The tailings collector is a combination of butyl xanthate, No. 25 black powder, and ethyl thiocyanate, with butyl xanthate at 15~25 g / t, No. 25 black powder at 15~30 g / t, and ethyl thiocyanate at 5~10 g / t.

[0062] Understandably: regrinding and dissociation: 80% to 90% of the magnetic separation tailings are regrinded to -0.038mm, so that the arsenopyrite and gangue are fully dissociated.

[0063] Under highly alkaline conditions (pH 11-12), lime introduces a large amount of Ca. 2+It easily competes with dispersants for adsorption or forms precipitates. Sodium pyrophosphate (P₂O₇) ions... 4- ) for Ca 2+ The complexation stability constant is significantly higher than that of commonly used hexametaphosphate (compared to hexametaphosphate), enabling it to more effectively complex free Ca. 2+ Prevent Ca 2+ CaCO3 or Ca(OH)2 precipitates form on the surface of arsenopyrite, keeping the active sites clean and facilitating subsequent collector adsorption. Water glass (Na2SiO3) hydrolyzes to form silica colloids, which preferentially adsorb onto the silicate gangue surface to form a hydrophilic film. This film, combined with the electrostatic repulsion of sodium pyrophosphate, achieves chemical masking of CaCO3 on the surface of arsenopyrite. 2+ The physical shielding of the gangue by water glass provides dual dispersion.

[0064] Sodium lignosulfonate is a byproduct of papermaking. Its molecular structure contains phenolic hydroxyl groups, methoxy groups (-OCH3), and sulfonic acid groups (-SO3). - ), sulfonic acid group (-SO3) - The degree of ionization of arsenopyrite in a strongly alkaline environment of pH 11-12 is much higher than that of phenolic hydroxyl groups, which can reduce the Zeta potential on the surface of arsenopyrite and enhance electrostatic repulsion, effectively preventing heterogeneous aggregation of arsenopyrite and pyrite (which have similar isoelectric points) and improving separation selectivity; combined with gallic acid, the small molecule gallic acid penetrates into the microcracks of arsenopyrite and the internal Fe... 2+ / As 3+ Chelation: Sodium lignosulfonate provides electrostatic stability through sulfonic acid groups on the outer layer, forming a double-layer structure of inner anchoring and outer repulsion, which is more stable and has stronger resistance to stirring and shearing, preventing flocs from disintegrating during stirring; The macromolecular structure of sodium lignosulfonate can form a steric hindrance layer on the surface of pyrite, which has both dispersing and inhibiting effects, reducing the interference of pyrite on the flotation of arsenopyrite, while similar tannic acid and tannic acid have a weaker inhibitory effect on pyrite.

[0065] Sweet potato starch and guar gum form a complex flocculation: the amylose molecules of sweet potato starch bridge arsenopyrite particles through hydrogen bonds, forming primary flocs; guar gum, as a natural polymer of galactomannan, has extremely high viscosity (1% solution viscosity can reach 5000~6000 mPa·s). Its numerous hydroxyl groups can complex with iron ions on the surface of arsenopyrite. The rigid chains of guar gum interweave with the flexible chains of starch, forming a three-dimensional network with rigid-flexible interaction, increasing floc particle size and mechanical strength. The floc particle size reaches 0.27~0.43 mm, significantly increasing the specific surface area and providing more adsorption sites for butyl xanthate and No. 25 black powder. Guar gum is widely available, cheaper, dissolves faster, and has better industrial adaptability.

[0066] Butyl xanthate, as a traditional xanthate collector, has a strong chemisorption capacity on the surface of arsenopyrite (FeAsS) (through S...). - With surface Fe2+ / As 3+ (Forming ionic and covalent bonds), it can ensure the basic recovery rate of fine-grained arsenopyrite after flocculation; however, butyl xanthate also has a strong collecting ability for pyrite (FeS2), which easily causes the mixing of pyrite minerals. No. 25 black powder (the main component is sodium diaryl dithiophosphate), diaryl dithiophosphate ((ArO)2PSS) - In Ars(O-), the conjugation effect of the aryl group reduces the electron cloud density of the S atom, resulting in a significantly weaker collection ability for pyrite (which requires strong polar bonding) compared to butyl xanthate. However, it has a moderate affinity for As on the surface of Ars(O-), and when used in conjunction with butyl xanthate, it can preferentially occupy the active sites on the pyrite surface (competitive adsorption), preventing butyl xanthate from adsorbing on the pyrite surface, thereby diluting the collection of pyrite by butyl xanthate and facilitating selective separation.

[0067] The reinforcing effect of ethyl thiocyanate: As an auxiliary collector, ethyl thiocyanate (sodium diethyl dithiocarbamate) has a stronger chelating ability on the As sites on the surface of arsenopyrite than xanthate (forming a stable pentacyclic structure), which is beneficial for capturing fine arsenopyrite particles (<10μm), making up for the weak collecting power of No. 25 black powder. By controlling the mass ratio of the three, the arsenic recovery rate is effectively improved.

[0068] The present invention will be further described in detail below with reference to specific embodiments.

[0069] The experiment used an arsenic-zinc polymetallic ore from Inner Mongolia, characterized by the following main features: the raw ore contains 1.17% Cu, 4.66% Zn, 5.40% As, 15.22% Fe, and 7.82% S. It is characterized by high contents of pyrrhotite and arsenopyrite, and a dense intergrowth of sphalerite, chalcopyrite, and arsenic sulfide minerals. Understandably, the same batch of minerals was used in the examples and comparative examples.

[0070] Example 1 S1. Copper-Zinc Staged Grinding and Classification: The raw ore is crushed to -9mm in two stages. The first stage grinding reduces the fineness to -0.074mm (72%), and the ore is classified by the first hydrocyclone. The overflow fineness of -0.074mm (65%) and concentration of 38% enter the copper flotation. The copper tailings are ground in the second stage to -0.074mm (82%), and the ore is classified by the second hydrocyclone. The overflow fineness of -0.074mm (78%) and concentration of 36% enter the zinc-sulfur flotation.

[0071] S2. Weak Alkali Preferred Copper Floating: First, modified calcium carbide slag is prepared: the original calcium carbide slag is wet-milled with water until -0.074mm accounts for 88%, and a 35% concentration slurry is prepared and aged for 36 hours; sodium humate is added at a mass ratio of 6:1 (dry basis of calcium carbide slag to sodium humate), and the mixture is stirred and reacted at 50℃ for 45 minutes to obtain modified calcium carbide slag slurry (its mass is determined by solid content). Simultaneously, hot-dip galvanized slag is crushed to -2mm, and 12% dilute sulfuric acid is added at a liquid-to-solid ratio of 4:1. The mixture is stirred and leached at 50℃ for 1.5 hours. After filtration, the filtrate is oxidized with hydrogen peroxide to remove iron and then directly used as a source of zinc sulfate (based on the effective content, equivalent to adding 360g / t of pure zinc sulfate). Modified calcium carbide slag (620 g / t) and lime (100 g / t, mass ratio 6.2:1) were added to adjust the pH to 7.9; sodium humate (140 g / t), citric acid (160 g / t), and water glass (100 g / t, mass ratio 3.5:4:2.5) were added to suppress pyrite; zinc sulfate (360 g / t), sodium sulfite (90 g / t), and sodium carbonate (18 g / t, mass ratio 8:2:0.4) were added to suppress zinc minerals; and ethyl thiocyanate (38 g / t), isobutyl xanthate (38 g / t), and anisole (14 g / t, mass ratio 3:3:1.1) were added to collect copper minerals. Copper roughing, two scavenging, and three cleaning processes were performed to obtain copper concentrate.

[0072] S3. Slurry Aeration and Activation with Semi-Flotation: Copper tailings slurry was aerated for 25 min at an aeration rate of 1.5 m³ / min; 300 g / t of calcium hypochlorite was added to suppress pyrrhotite; 75 g / t of EDTA-disodium salt was added to complex and activate sphalerite; 150 g / t of copper sulfate was added to activate sphalerite; 12 g / t of 2-mercaptobenzothiazole and 18 g / t of isobutyl xanthate (mass ratio 1.5:2.3) were added for collection. The slurry pH was 6.5, and semi-flotation roughing was performed. The frother was methyl isobutyl methanol, 23 g / t. S4. Magnetic separation and recovery of pyrrhotite: The semi-mixed tailings are subjected to weak magnetic separation with a magnetic field strength of 0.15T to obtain pyrrhotite concentrate.

[0073] S5. First zinc-sulfur separation: Add 1200g / t of lime, 120g / t of copper sulfate, 8g / t of 2-mercaptobenzothiazole and 12g / t of isobutyl xanthate to the semi-mixed froth for zinc-sulfur separation roughing to obtain zinc crude concentrate and pyrite concentrate.

[0074] S6. Second zinc-sulfur separation: The zinc concentrate is regrinded to -0.044mm (88%), and 160g / t of sodium cyanurate carbide, 160g / t of CMC, and 114g / t of guar gum (mass ratio 3.5:3.5:2.5) are added. 30g / t of thiouric acid ester and 45g / t of No. 25 black powder (mass ratio 1.2:1.8) are added, and 750g / t of lime is added to adjust the pH to 11.0. After two scavenging and three cleaning processes, the zinc concentrate is obtained.

[0075] S7. Fine-grained arsenic flocculation flotation: Magnetic separation tailings are regrinded to -0.038mm (85%), concentrated to 28%, and then 40g / t of sodium pyrophosphate and 400g / t of water glass are added to disperse the slime; 50g / t of sodium lignosulfonate and 25g / t of gallic acid are added for modification; 100g / t of sweet potato starch and 30g / t of guar gum are added for selective flocculation; 1000g / t of lime is added to adjust the pH to 11.5; 20g / t of butyl xanthate, 22g / t of No. 25 black powder, and 7g / t of ethyl thiocyanate are added for collection. After two scavenging and three cleaning processes, arsenic concentrate is obtained.

[0076] Experimental results: Copper concentrate grade 30.31%, recovery rate 84.39%; zinc concentrate grade 48.53%, recovery rate 91.21%, arsenic content 0.29%; arsenic concentrate grade 41.42%, recovery rate 89.89%; total lime consumption throughout the process 1540 g / t.

[0077] Example 2 Unlike Example 1, in step S2, the mass ratio of modified calcium carbide slag to lime was adjusted to 5:1 (600 g / t of modified calcium carbide slag and 120 g / t of lime), and the pH of the slurry was controlled at 7.9-8.0. Before the copper flotation tailings entered the zinc flotation, the aeration time was extended to 30 min, and the amount of calcium hypochlorite was increased to 350 g / t.

[0078] Experimental results: Copper concentrate grade 29.91%, recovery rate 84.03%; zinc concentrate grade 48.23%, recovery rate 90.71%, arsenic content 0.31%; arsenic concentrate grade 40.92%, recovery rate 85.99%. Due to the reduced proportion of calcium carbide slag, the pH of copper flotation was relatively unstable, and the copper recovery rate decreased slightly, but remained within an acceptable range.

[0079] Example 3 Unlike Example 1, in step S6, the ratio of the organic combination inhibitors was adjusted to 138 g / t of sodium cyanurate carbide, 138 g / t of CMC, and 100 g / t of guar gum (mass ratio 3.3:3.3:2.4). The amount of guar gum was increased to improve the viscosity of the slurry and enhance the inhibition of fine-grained pyrite.

[0080] Experimental results: The zinc concentrate grade increased to 49.43%, the arsenic content decreased to 0.24%, and the recovery rate was 90.55%. Increasing the amount of guar gum significantly improved the zinc concentrate grade, but the increased pulp viscosity led to a higher load on the flotation machine, necessitating adjustments to the aeration rate.

[0081] Example 4 Unlike Example 1, in step S7, the selective flocculant was adjusted to 120 g / t of sweet potato starch and 35 g / t of guar gum, increasing the flocculant dosage and allowing the flocs to grow more fully.

[0082] Experimental results: The arsenic concentrate grade increased to 42.24%, with a recovery rate of 88.41%. Adding sweet potato starch and guar gum resulted in a more compact floc structure, improving the collection efficiency for fine arsenopyrite particles (<10μm), and the arsenic concentrate grade exceeded 42%.

[0083] Example 5 Unlike Example 1, step S3 uses two-stage aeration: first aerate for 15 minutes and add half of the calcium hypochlorite, then aerate for another 15 minutes and add the remaining reagent, for a total aeration time of 30 minutes; in step S6, the amount of lime is adjusted to 600 g / t, and the pH is controlled at 10.5.

[0084] Experimental results: Zinc concentrate grade 48.44%, recovery rate 91.62%, arsenic content 0.29%. Staged aeration resulted in more complete oxidation of pyrrhotite and better inhibition effect; low-alkali zinc separation reduced the risk of zinc mineral inhibition and slightly improved the recovery rate.

[0085] Example 6 Full-process closed-loop test: A closed-loop cycle test was conducted under the conditions of Example 1, with the intermediate ore returned in the conventional manner to examine process stability and reagent cumulative effect.

[0086] Experimental results: After 5 cycles, equilibrium was reached. The copper concentrate grade was 29.75%, with a recovery rate of 83.83%; the zinc concentrate grade was 48.54%, with a recovery rate of 90.92% and an arsenic content of 0.28%; and the arsenic concentrate grade was 40.72%, with a recovery rate of 86.23%. All product indicators were stable, and there were no negative effects from reagent accumulation.

[0087] Comparative Example 1 (Traditional High-Alkali Process) The traditional single-stage grinding-high-alkali process flow in the background technology is adopted: the first-stage grinding fineness is 0.074mm 65%, copper flotation lime 2000g / t (pH>12), zinc sulfate 1500g / t, and ethyl xanthate collection; zinc flotation lime 3000g / t, copper sulfate activation, and xanthate collection; after zinc tailings are re-grinded, the pH is adjusted to 6 with sulfuric acid to float arsenic.

[0088] Test results: Copper concentrate grade 28.45%, recovery rate 81.23%; zinc concentrate grade 43.85%, recovery rate 84.12%, arsenic content 0.48% (exceeding standard); arsenic concentrate grade 28.5%, recovery rate 65.3%; total lime consumption throughout the process was 5200g / t, subsequent acid adjustment costs were extremely high, and zinc concentrate consistently failed to meet standards.

[0089] Comparative Example 2 Same as in Example 1, but in step S3, EDTA-disodium is not added, and copper sulfate is used directly for activation at 150g / t.

[0090] Experimental results: The activation of sphalerite was insufficient, the zinc grade in the semi-mixed froth was only 19.2%, the pyrrhotite inclusions were serious, and the separation of zinc and sulfur was difficult; the final zinc concentrate grade was 46.33%, and the recovery rate dropped to 82.44%.

[0091] Comparative Example 3 Same as in Example 1, but in step S7, sweet potato starch and guar gum are not added, and conventional butyl xanthate flotation at 80g / t is used directly.

[0092] Experimental results: A large amount of fine-grained toxic arsine was lost, the arsenic concentrate grade was 39.61%, the recovery rate was 84.42%, and the arsenic content in the tailings was as high as 0.89%, resulting in a reduced arsenic resource recovery rate and significant environmental risks associated with the tailings.

[0093] Comparative Example 4 Comparative Example 4 is the same as Example 1, but in step S2, modified carbide slag is not used. Instead, 720 g / t of carbide slag and 180 g / t of lime (mass ratio 4:1) are used, along with 275 g / t of sodium humate, 183 g / t of citric acid, and 92 g / t of water glass (mass ratio 4.5:3:1.5). Experimental results: Copper concentrate grade 29.68%, recovery rate 83.45%; zinc concentrate grade 47.85%, recovery rate 90.12%, arsenic content 0.31%; arsenic concentrate grade 40.55%, recovery rate 88.92%. Compared with Example 1, the copper recovery rate decreased due to the coarse particle size and low activity of the directly added calcium carbide slag.

[0094] Comparative Example 5 Comparative Example 5 is basically the same as Example 1, except that the pyrite combination inhibitor is only sodium humate 200g / t + citric acid 0g / t + water glass 0g / t (i.e., only sodium humate).

[0095] Experimental results: The copper concentrate grade was 28.45%, and the recovery rate was 82.10%; the copper concentrate contained 8.5% sulfur and 0.42% arsenic. Compared with Example 1, the lack of citric acid significantly reduced the Fe content. 2+ / Fe 3+ The complexation and penetration of the silica gel barrier of the water-deficient glass, the incomplete hydrophilic film on the pyrite surface, and the decrease in arsenic removal rate.

[0096] Comparative Example 6 Comparative Example 6 is basically the same as Example 1, except that the specific steps of leaching hot-dip galvanized slag with dilute sulfuric acid and oxidizing to remove iron are as follows: the hot-dip galvanized slag is crushed to -2.4 mm, 25% dilute sulfuric acid is added, and the mixture is stirred and leached at 65°C for 3 hours, and then filtered to obtain the leachate.

[0097] Experimental results: Copper concentrate grade 30.24%, recovery rate 84.19%; zinc concentrate grade 48.41%, recovery rate 91.01%. Compared with Example 1, the hot-dip galvanizing slag leaching process was suboptimal, introducing more impurities and affecting the improvement of zinc grade.

[0098] Comparative Example 7 Comparative Example 7 is basically the same as Example 1, except that in step S2: 720g / t of lime is used to adjust the pH to 8.0; only 250g / t of sodium humate is added to inhibit pyrite; only 400g / t of industrial zinc sulfate is added to inhibit zinc minerals; and only 70g / t of isobutyl xanthate is added to collect copper minerals.

[0099] Experimental results: The copper concentrate grade was 26.82%, and the recovery rate was 78.15%; the zinc content in the copper tailings was as high as 1.85% (zinc was lost in the copper tailings). Compared with Example 1, the single reagent had poor selectivity, incomplete inhibition of zinc minerals, severe pyrite flotation, and deterioration of copper concentrate quality.

[0100] The preferred embodiments of the present invention have been described in detail above, and are only used to illustrate the technical solutions of the present invention, and are not intended to limit it. Although the present invention has been described in detail with reference to the foregoing embodiments, those skilled in the art should understand that modifications can still be made to the technical solutions described in the foregoing embodiments, or equivalent substitutions can be made to some or all of the technical features therein; and these modifications or substitutions do not cause the essence of the corresponding technical solutions to deviate from the scope of the technical solutions of the embodiments of the present invention.

Claims

1. A method for treating arsenic-zinc polymetallic ore containing waste residue, comprising the following steps: S1. Staged grinding and classification of arsenic-zinc polymetallic ore; wherein, the raw ore is crushed in two stages, followed by stage grinding and stage classification in a first hydrocyclone. The underflow is returned to the first stage grinding, and the overflow enters the weak-alkali copper flotation operation; the copper flotation tailings enter the second stage grinding and stage classification in a second hydrocyclone. The underflow is returned to the second stage ball mill for regrinding, and the overflow enters the zinc-sulfur flotation operation; S2. The weak-alkali copper flotation of the overflow slurry from the first stage classification yields low-arsenic copper. Concentrate and copper tailings; S3. The second stage graded overflow slurry is aerated and activated and semi-flotated, and after slurry conditioning, semi-flotation roughing is performed to obtain semi-flotation froth and semi-flotation tailings; S4. Pyrrhotite is magnetically separated and recovered as sulfur concentrate product to obtain magnetic tailings; the semi-flotation froth is sequentially processed by S5 and S6 to obtain high-grade low-arsenic zinc concentrate; wherein S5. The semi-flotation froth undergoes a first zinc-sulfur separation to obtain zinc rough concentrate and first sulfur concentrate; S6. The zinc rough concentrate undergoes a second zinc-sulfur separation to obtain high-grade low-arsenic zinc concentrate; S7. Fine-particle arsenic flocculation flotation treatment of magnetic tailings to obtain high-grade arsenic concentrate; characterized in that... S2. Weak Alkali Copper Floating: A pH adjuster, a pyrite composite inhibitor, a zinc mineral composite inhibitor, and a copper composite collector are added to the first-stage classification overflow slurry. Copper roughing, two copper scavenging processes, and three copper cleaning processes are performed to obtain low-arsenic copper concentrate and copper tailings. The pH adjuster is a combination of modified calcium carbide slag and lime. The modified calcium carbide slag is calcium carbide slag that has been wet-milled to a particle size of -0.074mm or higher (85%), prepared into a slurry with a mass concentration of 30%~40%, aged for 24~48 hours, and then mixed with sodium humate at a dry basis mass ratio of (5~8):

1. The process involves premixing and stirring at 40-60℃ for 30-60 min to obtain a slow-release calcium carbide slag loaded with sodium humate; the pyrite-based inhibitor is a combination of sodium humate, citric acid, and water glass in a mass ratio of (2.5-3.5):(3.5-4.5):(1.5-2.5); the zinc mineral inhibitor is a combination of zinc sulfate, sodium sulfite, and sodium carbonate, wherein the zinc sulfate is obtained by leaching hot-dip galvanized slag with dilute sulfuric acid and oxidizing to remove iron; and the copper collector is a combination of ethyl thiocyanate, isobutyl xanthate, and anisole. The specific steps for leaching and oxidizing hot-dip galvanized slag with dilute sulfuric acid to remove iron are as follows: crush the hot-dip galvanized slag to -2mm, add dilute sulfuric acid with a concentration of 10% to 15% at a liquid-solid ratio of 3:1 to 5:1, stir and leach for 1 to 2 hours at 40 to 60℃, and filter to obtain the leachate; add hydrogen peroxide to the leachate to oxidize and remove iron to obtain a zinc sulfate solution.

2. The method according to claim 1, characterized in that, In step S2, the pH of the slurry is controlled to be 7.9-8.

3. The pH adjuster is a mixture of modified carbide slag and lime, with a mass ratio of 5-8:1 and a total dosage of 500-800 g / t. The pyrite combination inhibitor is a composition of sodium humate, citric acid, and water glass, with a mass ratio of (2.5-3.5):(3.5-4.5):(1.5-2.5), and a total dosage of 250-400 g / t. The zinc mineral group... The inhibitor is a combination of zinc sulfate, sodium sulfite, and sodium carbonate in a mass ratio of (6~9):(1.5~3):(0.3~1); the total amount of the zinc mineral inhibitor is 300~500 g / t; the copper collector is a combination of ethyl thiocyanate, isobutyl xanthate, and anisole in a mass ratio of (2.5~4):(2.5~4):(1~1.5); the total amount of the copper collector is 50~90 g / t.

3. The method according to claim 1, characterized in that, S1. Segmented grinding and classification of arsenic-zinc polymetallic ore: The raw ore is crushed to ≤10mm in two stages, and then ground in the first stage to a fineness of -0.074mm accounting for 70%~75%. The ore is then classified in the first stage by a first hydrocyclone. The underflow is returned to the first stage grinding, and the overflow is used for copper flotation. The copper flotation tailings are ground in the second stage to a fineness of -0.074mm ≥80%. The ore is then classified in the second stage by a second hydrocyclone. The underflow is returned to the second stage ball mill for further grinding, and the overflow is used for zinc-sulfur flotation.

4. The method according to claim 1, characterized in that, S3. Slurry Aeration and Activation and Semi-Flotation: The overflow slurry from the second stage of classification is pumped into an aeration mixing tank and aerated with air. Then, a first pyrite inhibitor, a complexing activator, a sphalerite activator, a sphalerite selective collector, and a frother are added. After slurry conditioning, semi-flotation roughing is performed to obtain semi-flotation froth and semi-flotation tailings. The first pyrite inhibitor is calcium hypochlorite, with a dosage of 200-400 g / t. The complexing activator is EDTA-disodium, with a dosage of 50-100 g / t. The sphalerite activator is copper sulfate, with a dosage of 100-200 g / t. The sphalerite selective collector is a complex of 2-mercaptobenzothiazole and isobutyl xanthate propylene ester, with a mass ratio of (1-2):(2-3). The dosage of the sphalerite selective collector is 20-40 g / t.

5. The method according to claim 1, characterized in that, S5. First zinc-sulfur separation: Add a second pyrite inhibitor, a sphalerite activator, and a sphalerite selective collector to the semi-mixed froth obtained in step S3 to perform the first zinc-sulfur separation roughing to obtain zinc rough concentrate and first sulfur concentrate; the second pyrite inhibitor is lime, with a dosage of 1000~1500 g / t; the sphalerite activator is copper sulfate, with a dosage of 100~150 g / t; the sphalerite selective collector is a complex of 2-mercaptobenzothiazole and isobutyl xanthate propylene ester, with a mass ratio of (1~2):(2~3), and a dosage of 10~20 g / t.

6. The method according to claim 1, characterized in that, S6. Second Zinc-Sulfur Separation: The zinc concentrate obtained in step S5 is regrinded to a fineness of -0.044mm accounting for 85%~90%, and then an organic combination inhibitor, auxiliary collector, and frother are added for a second zinc-sulfur separation. After two scavenging and three cleaning processes, a high-grade, low-arsenic zinc concentrate is obtained. The organic combination inhibitor is a mixture of sodium cyanurate carboxymethyl cellulose sodium and guar gum in a mass ratio of (3~4):(3~4):(2~3). The auxiliary collector is a combination of thiocyanate and No. 25 black powder in a mass ratio of (1~1.5):(1~2). The total amount of the organic combination inhibitor is 350~600g / t. The total amount of the auxiliary collector is 40~100g / t. The pH adjuster is lime, with a dosage of 600~900g / t. The pH of the pulp is controlled at 10.5~11.

5.

7. The method according to claim 1, characterized in that, S7. Fine-particle arsenic flocculation flotation: The magnetic separation tailings obtained in step S4 are regrinded to a fineness of -0.038mm (80%~90%), concentrated to a concentration of 25%~30%, and then a slime dispersant, an arsenic mineral surface modifier, a selective flocculant, a second pH adjuster, and a tailings collector are added sequentially. Arsenic roughing, two scavenging, and three cleaning processes are performed to obtain high-grade arsenic concentrate. The slime dispersant is a mixture of sodium pyrophosphate and water glass, with sodium pyrophosphate at a dosage of 30~50 g / t and water glass at a dosage of 300~500 g / t. The arsenic mineral surface modifier is lignin. The complex of sodium sulfonate and gallic acid, with sodium lignin sulfonate at a dosage of 40-60 g / t and gallic acid at a dosage of 20-30 g / t; the selective flocculant is a mixture of sweet potato starch and guar gum, with sweet potato starch at a dosage of 90-120 g / t and guar gum at a dosage of 30-40 g / t; the tailings collector is a combination of butyl xanthate, No. 25 black powder, and ethyl thiocyanate, with butyl xanthate at a dosage of 15-25 g / t, No. 25 black powder at a dosage of 15-30 g / t, and ethyl thiocyanate at a dosage of 5-10 g / t; the second pH adjuster is lime, controlling the slurry pH to 11-12.

8. The method according to claim 1, characterized in that, In step S1, the first stage of overflow has a fineness of -0.074mm ≥ 55% and a concentration of ≥ 35%; the second stage of overflow has a fineness of -0.074mm ≥ 75% and a concentration of ≥ 35%.

9. The method according to claim 1, characterized in that, S4. Magnetic separation and recovery of pyrrhotite: The semi-mixed tailings obtained in step S3 are subjected to weak magnetic separation to recover pyrrhotite as a sulfur concentrate product; the magnetic separation tailings are then used for subsequent arsenic recovery operations; the magnetic field strength is 0.1~0.2T.

10. The method according to claim 1, characterized in that, Step S3 controls the slurry pH to 6.0~8.2, performs aeration treatment, with an aeration time of 15~30 min and an aeration rate of 1.0~2.0 m³ / min; Step S7 controls the slurry pH to 11.5~12.0.