A pre-treatment-potential-pH three-stage synergistic copper-lead-zinc pyrite flotation method

Abstract

This invention relates to the field of mineral flotation separation technology, and discloses a three-stage synergistic flotation method for copper-lead-zinc-pyrite ore, comprising: first, subjecting the copper-lead-zinc-pyrite ore to heavy medium cyclone pre-classification to obtain a pyrite-enriched slurry; adjusting the pH and redox potential of the pyrite-enriched slurry, adding a potential-responsive collector for chalcopyrite flotation to obtain copper scavenging tailings slurry and chalcopyrite concentrate slurry; raising the pH and lowering the redox potential of the copper scavenging tailings slurry for galena flotation to obtain lead scavenging tailings slurry and galena concentrate slurry; raising the pH and redox potential of the lead scavenging tailings slurry for sphalerite flotation to obtain zinc scavenging tailings slurry and sphalerite concentrate slurry; and lowering the pH of the zinc scavenging tailings slurry for sulfur concentrate flotation to obtain sulfur flotation tailings slurry and sulfur concentrate slurry. This method addresses the problems of cross-contamination, high reagent consumption, and difficulty in maintaining precise separation of copper, lead, and zinc in existing flotation processes.

Description

Technical Field

[0001] This invention relates to the field of mineral flotation and separation technology, specifically to a three-stage synergistic flotation method for copper-lead-zinc-pyrite ore using pretreatment-potential-pH. Background Technology

[0002] Copper-lead-zinc-pyrite is an important non-ferrous metal mineral resource in my country. The copper, lead, and zinc metals in the ore can be widely used in many core industrial fields such as metallurgical processing, machinery manufacturing, fine chemicals, new energy batteries, and electronics. In the beneficiation and processing of copper-lead-zinc-pyrite ores, flotation is the core process for achieving polymetallic separation, improving the quality and efficiency of ore resources, and efficiently recovering valuable metals. The process system, control methods, and reagent compatibility of flotation separation technology directly determine the separation effect and concentrate quality of copper-lead-zinc polymetallic ores.

[0003] Currently, the flotation of copper-lead-zinc pyrite mainly faces the following problems: First, traditional collectors (such as xanthates and black reagents) are not selective enough for galena, sphalerite, and chalcopyrite, and are prone to cross-contamination in complex polymetallic symbiotic ores, resulting in prominent quality problems such as lead in copper concentrate and zinc in lead concentrate; Second, conventional flotation processes rely on a large amount of inhibitors (cyanide, zinc sulfate, sulfite, etc.) to suppress non-target minerals, resulting in high reagent consumption and environmental unfriendliness; Third, the pulp pH and redox potential (Eh) fluctuate drastically during continuous flotation, leading to unstable adsorption selectivity and making it difficult to maintain accurate separation of copper, lead, and zinc under large-scale industrial conditions.

[0004] Solving the above problems presents the following difficulties: First, the matching window of the potential-responsive collector is narrow, meaning that the resting potential difference between chalcopyrite, galena, and sphalerite is only tens of millivolts (mV). To achieve highly selective collection of only one mineral while simultaneously suppressing the other two in a wide range of pulp fluctuations, extremely high requirements are placed on the molecular design and potential response sensitivity of the reagent. Second, there is interference from the dissolution and re-adsorption of ions on the mineral surface. That is, in the continuous flotation process, the already dissociated Cu... 2+ Pb 2+ Zn 2+ Ions may be adsorbed on the surface of non-target minerals in subsequent flotation stages, triggering cross-activation and destroying the segmented separation effect; thirdly, the online pH-Eh coupled control is not precise enough. Industrial-grade pH and ORP sensors are easily contaminated and drifted when operating in slurry for a long time, making it difficult to achieve long-term stable control. Once the adsorption coverage deviation exceeds ±3%, the selectivity deteriorates rapidly. This requires the control system to have both high precision, fast response and strong anti-interference ability.

[0005] Therefore, there is an urgent need for a highly selective preferential flotation method for copper-lead-zinc-pyrite ore to solve the problems of easy cross-contamination, high reagent consumption, and difficulty in maintaining accurate separation of copper, lead, and zinc under large-scale industrial conditions in existing copper-lead-zinc-pyrite flotation. Summary of the Invention

[0006] The present invention aims to provide a three-stage synergistic pretreatment-potential-pH flotation method for copper-lead-zinc-pyrite ore, in order to solve the problems of easy cross-contamination, high reagent consumption, and difficulty in maintaining accurate separation of copper, lead, and zinc under large-scale industrial conditions in existing copper-lead-zinc-pyrite flotation methods.

[0007] To achieve the above objectives, the present invention adopts the following technical solution: a three-stage synergistic flotation method for copper-lead-zinc-pyrite ore, comprising the following steps: S1. Heavy medium cyclone pre-classification treatment is carried out on copper-lead-zinc-pyrite to obtain pyrite mineral-enriched slurry; S2. Adjust the pH and redox potential of the pyrite-rich slurry, add a potential-responsive collector, and perform chalcopyrite flotation to obtain copper scavenging tailings slurry and chalcopyrite concentrate slurry. S3. Increase the pH value of the copper scavenging tailings slurry and decrease the redox potential to carry out galena flotation to obtain lead scavenging tailings slurry and galena concentrate slurry. S4. Increase the pH value and redox potential of the lead scavenging tailings slurry to carry out sphalerite flotation to obtain zinc scavenging tailings slurry and sphalerite concentrate slurry. S5. Lower the pH value of the zinc scavenging tailings slurry and carry out sulfur concentrate flotation to obtain sulfur flotation tailings slurry and sulfur concentrate slurry.

[0008] The beneficial effects of this scheme are as follows: 1. This scheme addresses the separation challenges of complex copper-lead-zinc pyrite ores by constructing a three-stage synergistic flotation system of pretreatment-potential-pH. This system uses heavy medium cyclone pre-classification as the first-stage physical enrichment method, removing a large amount of gangue minerals in advance, increasing the total grade of the target pyrite minerals to over 80%, reducing the competitive interference of gangue particles in the pulp on the collector adsorption sites, and fundamentally reducing the complexity of subsequent reagent identification and selective adsorption. The second stage adopts a preferential flotation process, sequentially producing chalcopyrite concentrate, galena concentrate, sphalerite concentrate, and pyrite concentrate, achieving stepwise recovery of multi-metal minerals and avoiding the mutual interference caused by the simultaneous flotation of multiple pyrite minerals, which is beneficial for obtaining qualified concentrate products for each valuable mineral. The third stage incorporates a potential-responsive collector and pH-Eh linkage control. A potential-responsive collector refers to a collector that reacts to a specific target mineral within a specific pH and redox potential window. Collectors with significant selective adsorption capacity and significantly reduced adsorption capacity for other minerals, including but not limited to 2-(benzothiazolylthio)acetic acid, 2-mercaptobenzothiazolium and polyethylene glycol combinations, are selected based on the reagents used in each flotation stage. Adding a potential-responsive collector allows for preferential selective adsorption of chalcopyrite within a set pH-Eh window, inhibiting the simultaneous flotation of galena and sphalerite. Each flotation stage is matched with a corresponding pH and Eh operating range. By utilizing pH gradient changes combined with potential control, the differences in floatability between different pyrite minerals are amplified, weakening the mutual activation effects between minerals and improving the selectivity of multimetal separation. Precise control of adsorption coverage is achieved under fluctuating pulp conditions, thus solving problems such as easy cross-contamination, high reagent consumption, and difficulty in maintaining precise separation of copper, lead, and zinc under large-scale industrial conditions in copper-lead-zinc pyrite flotation.

[0009] 2. The entire process design follows the principle of discarding and recovering as early as possible. It reduces the amount of gangue entering the flotation circuit by pre-enriching with heavy media at the front end, achieves preferential selective collection of copper minerals through potential-responsive reagents, and amplifies the difference in floatability of lead and zinc minerals by switching pH gradients. The entire process relies on the synergistic effect of pretreatment and segmented pH-Eh condition matching to improve the comprehensive recovery rate of valuable components such as copper, lead, zinc, and sulfur, reduce the process difficulty of polymetallic pyrite separation, improve the separation index of complex associated pyrite, and ultimately achieve efficient and precise separation of copper, lead, zinc, and pyrite.

[0010] Preferably, step S1 includes the following steps: crushing the raw copper-lead-zinc-pyrite ore to a particle size of less than or equal to 0.5 mm, performing wet screening and grading to obtain a screened slurry; A ferrosilicon heavy medium suspension is added to the screened slurry, and the slurry is subjected to heavy medium cyclone treatment to obtain a heavy product. The heavy product is ground until the proportion of copper-lead-zinc-pyrite powder with a particle size of -0.074 mm in the heavy product is 65-75%. The pH value of the heavy product is adjusted to 9.3-9.7, and then water is added to obtain a pyrite-rich slurry.

[0011] The beneficial effects of this scheme are as follows: by crushing, wet screening and heavy media cyclone pre-classification of copper-lead-zinc-pyrite ore, gangue and light product tailings can be removed in advance, and pyrite minerals entering the grinding and flotation circuit can be effectively enriched, increasing the total grade of pyrite minerals to more than 80%, which greatly reduces the difficulty of subsequent flotation reagent identification and separation; at the same time, the grinding fineness can be precisely controlled to have a particle size of -0.074mm accounting for 65~75%, which can fully dissociate the complex copper-lead-zinc-sulfur intergrowth and provide a material basis for subsequent efficient separation and enrichment.

[0012] Preferably, the copper-lead-zinc-pyrite ore comprises the following components: copper (Cu), lead (Pb), zinc (Zn), iron (Fe), sulfur (S), calcium oxide (CaO), aluminum oxide (Al2O3), silicon dioxide (SiO2), silver (Ag*), gold (Au*), and magnesium oxide (MgO), with the remainder being other trace impurities and loss on ignition; where * indicates the content of the corresponding element in g / t.

[0013] Preferably, the heavy medium suspension has a density of 6.8 g / cm³. 3 It is prepared from ferrosilicon powder, and the feed pressure of the heavy medium cyclone treatment is 0.08~0.12MPa.

[0014] Preferably, the ferrosilicon powder is an iron-silicon alloy powder, comprising iron, silicon, manganese, carbon, sulfur and phosphorus, wherein the mass fraction of iron is 80-83% and the mass fraction of silicon is 14-16%; the iron-silicon alloy powder is prepared by grinding and sieving after smelting in an electric furnace.

[0015] Preferably, the concentration of copper-lead-zinc pyrite in the pyrite-enriched slurry after adding water is 30-33%.

[0016] Preferably, the copper-lead-zinc-pyrite ore in the screening slurry has a particle size of 0.1~0.5mm.

[0017] Preferably, after the screened slurry is mixed with the heavy medium suspension, it is fed into an FHMC350 heavy medium hydrocyclone to obtain underflow and overflow. The underflow is the heavy product. After the heavy product is treated by a desliming screen, copper-lead-zinc-pyrite concentrate is obtained. After the copper-lead-zinc-pyrite concentrate is ground, water is added to obtain the pyrite mineral-enriched slurry.

[0018] Preferably, the overflow is a light product, which is discharged after being treated by a desliming sieve.

[0019] Preferably, after the medium removal sieve treatment, a medium suspension is obtained, and the medium suspension is then subjected to magnetic separation to achieve the recycling of ferrosilicon.

[0020] Preferably, the grinding process is ball milling, and the ball milling time is 10-12 minutes.

[0021] Preferably, step S2 includes the following steps: The pH of the pyrite-rich slurry was adjusted to be between 9.3 and 9.7, and the redox potential was adjusted to be between -110 and -90 mV. A potential-responsive collector, a foaming agent, and a lead-zinc inhibitor are added to the pyrite-rich slurry, and chalcopyrite roughing is carried out to obtain copper roughing concentrate slurry and copper roughing tailings slurry. The copper roughing concentrate slurry is subjected to chalcopyrite beneficiation to obtain chalcopyrite concentrate slurry; The copper roughing tailings slurry is subjected to chalcopyrite scavenging to obtain copper scavenged tailings slurry.

[0022] Preferably, the potential-responsive collector is selected from 2-(benzothiazolylthio)acetic acid.

[0023] The beneficial effects of this scheme are as follows: In the chalcopyrite flotation stage, a special 2-(benzothiazolylthio)acetic acid potential-responsive collector is used, which, together with the range of pH and Eh and the synergistic effect of frother and lead-zinc inhibitor, can achieve selective preferential collection of chalcopyrite while effectively inhibiting the simultaneous flotation of galena and sphalerite.

[0024] Preferably, the roughing of the chalcopyrite is performed 1 to 2 times, and the skimming time is 8 to 12 minutes.

[0025] Preferably, in the roughing of the chalcopyrite, the amount of 2-(benzothiazolylthio)acetic acid added is 15~35 g / t.

[0026] Preferably, the lead-zinc inhibitor is a combination of zinc sulfate (ZnSO4) and sodium sulfite (Na2SO3), and the mass ratio of ZnSO4 to Na2SO3 is (2~2.5):1.

[0027] Preferably, in the roughing of the chalcopyrite, the amount of ZnSO4 added is 200~500g / t, and the amount of Na2SO3 added is 100~200g / t.

[0028] Preferably, the chalcopyrite is selected using a blank, untreated selection method, and the chalcopyrite is selected 2 to 3 times, with a skimming and soaking time of 4 to 6 minutes.

[0029] Preferably, 2-(benzothiazolylthio)acetic acid is added as a potential-responsive collector to the copper roughing tailings slurry, and a frother is added to perform chalcopyrite scavenging. The chalcopyrite scavenging is performed 1 to 2 times, and the frothering time is 4 to 6 minutes.

[0030] Preferably, when the chalcopyrite is scavenged twice, the amount of 2-(benzothiazolylthio)acetic acid added in the first chalcopyrite scavenging is 4~10 g / t; and the amount of 2-(benzothiazolylthio)acetic acid added in the second chalcopyrite scavenging is 2~5 g / t.

[0031] Preferably, the foaming agent is methyl isobutyl methanol (MIBC), and the amount of MIBC added in the roughing of chalcopyrite is 15~25g / t.

[0032] Preferably, the frother used in the scavenging of chalcopyrite is MIBC, and the amount of MIBC added in the scavenging of chalcopyrite is 5~10g / t.

[0033] Preferably, step S3 includes the following steps: adjusting the pH value of the copper scavenging tailings slurry to between 10.3 and 10.7, and the redox potential to between -60 and -40 mV; then adding galena roughing collector, frother, lead activator and zinc inhibitor to the copper scavenging tailings slurry to perform galena roughing to obtain lead roughing concentrate slurry and lead roughing tailings slurry. The lead roughing concentrate slurry is subjected to galena beneficiation to obtain galena concentrate slurry; The lead roughing tailings slurry is subjected to galena scavenging to obtain lead scavenged tailings slurry.

[0034] Preferably, a combination of 2-mercaptobenzothiazole (MBT) and polyethylene glycol (PEG) is added to the copper scavenging tailings slurry as a roughing collector for galena, wherein the mass ratio of MBT to PEG in the roughing collector for galena is (1:0.5) to 1.

[0035] The beneficial effects of this scheme are as follows: In the galena flotation stage, the combined collector of 2-mercaptobenzothiazole and polyethylene glycol, along with the frother, lead activator, and zinc inhibitor, and the range of pH and Eh, can precisely enhance the flotation adsorption effect of galena, significantly widen the adsorption difference between galena and residual sphalerite and chalcopyrite, with an adsorption difference of over 55%, effectively achieve efficient separation and enrichment of galena, reduce the content of impurities such as zinc in lead concentrate, and improve the recovery rate of lead minerals.

[0036] Preferably, the roughing of galena is performed 1 to 2 times, and the frothing time is 8 to 12 minutes.

[0037] Preferably, in the roughing of galena, the amount of MBT added is 25~50g / t, and the amount of PEG added is 10~25g / t.

[0038] Preferably, the lead activator is sodium sulfide (Na2S), and the amount of Na2S added in the roughing of galena is 30~40g / t.

[0039] Preferably, the zinc inhibitor is selected from a combination of ZnSO4 and Na2SO3, and the mass ratio of ZnSO4 to Na2SO3 is 2 to (2.5:1).

[0040] Preferably, in the roughing of galena, the amount of ZnSO4 added is 200~500g / t, and the amount of Na2SO3 added is 100~200g / t.

[0041] Preferably, the galena beneficiation adopts a blank beneficiation method without adding reagents, and the galena beneficiation is carried out 2 to 3 times, with a skimming and bubbling time of 4 to 6 minutes.

[0042] Preferably, an MBT and PEG composition is added to the lead roughing tailings slurry as a galena scavenging collector, and a frother is added for galena scavenging. The galena scavenging is performed 1 to 2 times, and the frothering time is 4 to 6 minutes.

[0043] Preferably, in the galena scavenging collector, the mass ratio of MBT to PEG is (1:0.5) to 1.

[0044] Preferably, in the galena scavenging process, the amount of MBT added is 4~8 g / t, and the amount of PEG added is 2~4 g / t.

[0045] Preferably, the frother used in the roughing of galena is MIBC, and the amount of MIBC added in the roughing of galena is 20~30g / t.

[0046] Preferably, the frother used in the galena scavenging process is MIBC, and the amount of MIBC added in the galena scavenging process is 5~10g / t.

[0047] Preferably, step S4 includes the following steps: Add sphalerite activator to the lead scavenging tailings slurry to obtain activated lead scavenging tailings slurry; The pH value of the activated lead scavenging tailings slurry was adjusted to be between 11.3 and 11.7, and the redox potential was between -10 and 10mV. Then, sphalerite roughing collector and frother were added to carry out sphalerite roughing to obtain zinc roughing concentrate slurry and zinc roughing tailings slurry. The zinc roughing concentrate slurry is subjected to sphalerite beneficiation to obtain sphalerite concentrate slurry; The zinc roughing tailings slurry is subjected to sphalerite scavenging to obtain zinc scavenged tailings slurry.

[0048] Preferably, the zincblende activator is copper sulfate (CuSO4), and the amount of CuSO4 added is 200~350g / t.

[0049] Preferably, the sphalerite roughing is performed 1 to 2 times, and the skimming time is 8 to 12 minutes.

[0050] Preferably, in the roughing of sphalerite, the amount of xanthate added is 30~60g / t.

[0051] Preferably, in the roughing of sphalerite, the frother is MIBC, and the amount of MIBC added in the roughing of sphalerite is 15~25g / t.

[0052] Preferably, the sphalerite is selected using a blank, untreated selection method, and the galena is selected 2 to 3 times, with a skimming and bubbling time of 4 to 6 minutes.

[0053] Preferably, butyl xanthate and frother are added to the zinc roughing tailings slurry for sphalerite scavenging, and the sphalerite scavenging is performed 1 to 2 times, with a scavenging time of 4 to 6 minutes.

[0054] Preferably, when the sphalerite is scavenged twice, the amount of butyl xanthate added in the first sphalerite scavenging is 12~20g / t; and the amount of butyl xanthate added in the second sphalerite scavenging is 8~15g / t.

[0055] Preferably, the frother used in the sphalerite scavenging is MIBC, and the amount of MIBC added in the sphalerite scavenging is 5~10g / t.

[0056] Preferably, step S5 includes the following steps: Sulfur concentrate activator is added to zinc scavenging tailings slurry to obtain activated zinc scavenging tailings slurry; The pH value of the activated zinc scavenging tailings slurry was adjusted to between 6.3 and 6.7. Then, a sulfur concentrate roughing collector and a frother were added to carry out sulfur concentrate roughing to obtain sulfur flotation tailings slurry and sulfur concentrate slurry.

[0057] Preferably, the sulfur concentrate activator is CuSO4, and the amount of CuSO4 added is 100~200g / t.

[0058] Preferably, the sulfur concentrate is roughed 1 to 2 times and the frothing time is 8 to 12 minutes.

[0059] Preferably, in the roughing of the sulfur concentrate, the amount of xanthate added is 80~120g / t.

[0060] Preferably, the frother in the roughing of the sulfur concentrate is MIBC, and the amount of MIBC added in the roughing of the sulfur concentrate is 20~30g / t.

[0061] Preferably, in the chalcopyrite flotation process, a mixed aqueous solution of Na2CO3 and NaHCO3 is used to adjust the pH value of the pyrite-enriched slurry, and an oxidizing agent or reducing agent is used to adjust the Eh value of the pyrite-enriched slurry.

[0062] The beneficial effect of this scheme is that the combination of Na2CO3 and NaHCO3 forms CO3. 2- / HCO3 - A conjugate acid-base pair buffer system can neutralize newly added acids / bases in the pulp in both directions, locking the pH within an effective buffer zone around its pKa2 ≈ 10.3; when additional H+ is introduced into the pulp system... + or OH - At that time, H is consumed through proton transfer between conjugated components. + or OH - This suppresses drastic pH changes in the pulp, and with online pH feedback dosing control, it ensures minimal pH fluctuations in the pulp during flotation.

[0063] Preferably, in the galena flotation process, a mixed aqueous solution of Na2CO3 and NaHCO3 is used to adjust the pH value of the copper scavenging tailings slurry, and an oxidizing agent or reducing agent is used to adjust the Eh value of the copper scavenging tailings slurry.

[0064] Preferably, in the sphalerite flotation, lime slurry is used to adjust the pH value of the lead scavenging tailings slurry, and an oxidizing agent or reducing agent is used to adjust the Eh value of the lead scavenging tailings slurry.

[0065] Preferably, in the sphalerite flotation, sulfuric acid (H2SO4) is used to adjust the pH value of the zinc scavenging tailings slurry, and an oxidizing agent or reducing agent is used to adjust the Eh value of the zinc scavenging tailings slurry.

[0066] The beneficial effects of this solution are as follows: it matches a dedicated pH control system for each of the three flotation processes of copper, lead, and zinc. The copper and lead flotation uses a Na2CO3 / NaHCO3 buffer system to accurately stabilize the pulp pH, while the zinc flotation uses lime slurry to control the highly alkaline environment, forming a gradient pH separation system. This system can accurately adapt to the optimal flotation media conditions for different target minerals, effectively distinguish the flotation characteristics of each mineral, improve the multi-metal separation selectivity from the perspective of the media environment, and reduce the mutual interference between different minerals.

[0067] Preferably, in the flotation of chalcopyrite, the molar ratio of Na2CO3 to NaHCO3 in the mixed aqueous solution of Na2CO3 and NaHCO3 is 1:7, and the total concentration of Na2CO3 and NaHCO3 is 0.05~0.1mol / L.

[0068] Preferably, in the flotation of galena, the molar ratio of Na2CO3 to NaHCO3 in the mixed aqueous solution of Na2CO3 and NaHCO3 is 1:2.1, and the total concentration of Na2CO3 and NaHCO3 is 0.05~0.1mol / L.

[0069] Preferably, the method for adjusting Eh includes the following steps: when the value is lower than Eh (excessive reducing power), an oxidizing agent is added; when the value is higher than Eh (excessive oxidizing power), a reducing agent is added.

[0070] The beneficial effects of this scheme are as follows: it sets up exclusive and precise Eh potential ranges for the flotation of chalcopyrite, galena, and sphalerite, and dynamically fine-tunes the pulp potential by using oxidants and reducing agents to accurately match the potential response flotation characteristics of each mineral. This effectively avoids the problem of disordered flotation of non-target minerals caused by potential drift, further enhances the separation specificity of each mineral, significantly reduces the impurity mixing rate between concentrates, and ensures the quality of each metal concentrate.

[0071] Preferably, the oxidant is H2O2 and the reducing agent is Na2S.

[0072] Preferably, it also includes a pH-Eh online linkage feedback control flotation system, which synchronously feeds forward to compensate the Eh control output when adjusting the pH, thereby suppressing pH-Eh oscillations throughout the flotation process.

[0073] The beneficial effects of this scheme are as follows: by combining the Na2CO3 / NaHCO3 buffer system with online pH-Eh linkage feedback control, the pH-Eh control oscillation is suppressed; under the condition of small fluctuations in slurry conditions, the adsorption coverage deviation of the collector on the chalcopyrite surface can be controlled within ±3%, maintaining stable separation selectivity.

[0074] Preferably, the pH-Eh online linkage feedback control flotation system adopts a decoupling algorithm, which synchronously feeds forward to compensate the Eh control output when adjusting the pH based on the influence coefficient of pH on Eh.

[0075] Preferably, the influence coefficient of pH on Eh is -58 to -60 mV / pH.

[0076] Preferably, the pH-Eh online linkage feedback control flotation system includes an Eh sensor and a pH sensor, and uses a PLC (programmable logic controller) as the core control unit for real-time data acquisition.

[0077] Preferably, the data sampling frequency of the Eh sensor and pH sensor is once every 1 to 5 seconds; if pH or Eh adjustment of the corresponding slurry is required after the data is collected, the adjustment time of the corresponding slurry is 3 to 5 minutes.

[0078] Preferably, the method also includes a tailings treatment method, comprising the following steps: adding a flocculant to the sulfur flotation tailings slurry, and concentrating it for 8-10 hours to obtain thickened overflow and underflow tailings; the underflow tailings are dewatered to obtain tailings filter cake, which is then stored in a tailings dam; the thickened overflow is reused.

[0079] Preferably, the flocculant is polyacrylamide (PAM), and the amount of PAM added is 30~50g / t.

[0080] Preferably, the thickened overflow is treated and returned to the chalcopyrite flotation step.

[0081] Preferably, the moisture content of the tailings filter cake is less than or equal to 12%. Attached Figure Description

[0082] Figure 1 A schematic diagram illustrating the mechanism of action between the collector 2-(benzothiazolylthio)acetic acid and chalcopyrite; Figure 2 The measured open-circuit potential-time variation curves of electrostatic potentials for chalcopyrite, galena, and sphalerite are shown. Figure 3 A schematic diagram illustrating the linkage feedback control principle between the Na2CO3 / NaHCO3 buffer system and pH-Eh. Figure 4 The process flow diagram for the three-stage synergistic high-selectivity preferential flotation of copper, lead, zinc, pyrite and iron ore. Detailed Implementation

[0083] The following detailed description illustrates the specific implementation method: To make the objectives, technical solutions, and effects of this invention clearer and more explicit, the invention is further described in detail below through specific embodiments. It should be understood that the specific embodiments described herein are only for explaining the invention and are not intended to limit the invention. Unless otherwise defined, all technical and scientific terms used herein have the same meaning as commonly understood by one of ordinary skill in the art to which this invention pertains. The terminology used in this specification is for the purpose of describing particular embodiments only and is not intended to limit the invention.

[0084] First, the technical principles involved in this invention will be explained one by one: Regarding the heavy medium cyclone pre-classification treatment performed in the first step.

[0085] Studies have found that gangue minerals (quartz, calcite, etc.) in complex copper-lead-zinc pyrite deposits have a density of 2.6~2.8 g / cm³. 3 ) and pyrite (chalcopyrite has a density of 4.1~4.3 g / cm³) 3Galena has a density of 7.4~7.6 g / cm³. 3 Sphalerite has a density of 3.9~4.1 g / cm³. 3 There is a significant density difference between these two minerals. Heavy medium hydrocyclones can utilize this density difference to physically separate and remove 60-70% of the gangue minerals, achieving front-end physical enrichment. Specifically, centrifugal force is used to accelerate mineral sedimentation to a density of 2.8-3.0 g / cm³. 3 The heavy medium suspension is used as the separation medium, and the concentration referred to here is 2.8~3.0 g / cm³. 3 The working density of the separation medium (i.e., a heavy medium suspension prepared from ferrosilicon powder and water) determines the density of the flotation circuit. Pyrite minerals with a density greater than that of the medium move towards the wall and are discharged from the bottom flow, while gangue minerals with a density less than that of the medium are discharged with the overflow. This resolves the core contradictions of large-scale gangue dilution of the flotation circuit, increased reagent consumption, and increased identification difficulty. It ensures that subsequent flotation processes deal with high-grade, small-quantity, and relatively simple-composition materials, creating the preconditions for the selective adsorption of potential-responsive reagents. The heavy medium cyclone pre-classification directly determines the processing scale and stability of the subsequent flotation circuit.

[0086] The mechanism by which 2-(benzothiazolylthio)acetic acid, a gradient potential-responsive collector, selectively adsorbs chalcopyrite (CuFeS2) while inhibiting the activation of lead / zinc (Pb / Zn) minerals under the conditions of Eh=-100±10mV (i.e. -110~-90mV) and pH=9.5±0.2 (i.e. 9.3~9.7).

[0087] This scheme designs copper as the first separation stage in the preferential flotation process. The key lies in the embedded gradient potential-responsive collector 2-(benzothiazolylthio)acetic acid. The mechanism of interaction between this collector and CuFeS2 is as follows: Figure 1 As shown: In 2-(benzothiazolylthio)acetic acid, the benzothiazolyl ring provides an active center for coordination with copper atoms (Cu) on the CuFeS2 surface. The nitrogen (N) and sulfur (S) atoms on the benzothiazolyl ring can form a stable five-membered ring chelate. The carboxymethyl thioether group (-S-CH2-COOH) endows the collector with responsive properties under specific pH and potential conditions. The carboxyl group (-COOH) partially dissociates at pH=9.5 to form -COO. -This enhances the hydrophilic-hydrophobic balance of the collector; the thio-bridged bond (-S-) acts as a bridging group connecting the benzothiazole ring and the free carboxyl group (-COOH), and can undergo redox response at a specific potential, thereby regulating the adsorption / desorption behavior of the collector on the mineral surface. Specifically, under the conditions of pH=9.5±0.2 and Eh=-100±10mV, the collector exhibits preferential adsorption characteristics for CuFeS2, while simultaneously inhibiting galena (PbS) and zinc sphalerite (ZnS). The resting potential of CuFeS2 in this potential window is approximately -105 to -95mV (relative to SHE), which is highly matched with the response potential window of the collector system. At this potential, the thiobonds in the collector molecule undergo moderate reduction and activation, generating an active form with strong coordination ability to copper atoms. The spatial arrangement of copper atoms on the surface of CuFeS2 forms a geometric match with the chelate configuration of the collector molecule. The distance between the N and S atoms of the benzothiazole ring is close to the distance between copper atoms on the dissociated crystal plane (112) of chalcopyrite, which is conducive to the formation of stable bidentate chelates.

[0088] Meanwhile, under conditions of -100±10 mV, electrochemical adsorption of the collector was difficult to occur on the surfaces of galena (resting potential approximately +135 to +155 mV) and sphalerite (resting potential approximately -200 to -180 mV), as both deviated from the optimal response window of the collector, resulting in weak adsorption. Electrostatic potential measurements were performed on a mixed slurry of galena, chalcopyrite, and sphalerite in a three-electrode system. Specifically, the concentration of the collector 2-(benzothiazolylthio)acetic acid was set to 1 × 10⁻⁶ at pH 9.5. -4 The redox potential of the mixed slurry was adjusted to -110 mV by adding Na2S / H2O2. The working electrode was a mineral electrode (powder-pressed electrode of chalcopyrite, galena, and sphalerite), the reference electrode was a saturated calomel electrode, and the auxiliary electrode was a graphite electrode. The electrolyte was a 0.1 mol / L Na2SO4 solution, meaning the test solution contained electrolyte, collector, and potential and pH adjusters. Before testing, N2 was introduced into the mixed slurry to remove dissolved oxygen, and it was equilibrated in a set pH buffer. The measuring instrument was an electrochemical workstation, recording the open-circuit potential over time until the potential stabilized, which was the resting potential. The measured electrostatic potentials of chalcopyrite, galena, and sphalerite are shown below. Figure 2 As shown, Figure 2 The stable values ​​shown in the test results (approximately +150mV for galena, approximately -100mV for chalcopyrite, and approximately -180mV for sphalerite) correspond perfectly to the above description.

[0089] The principle of lead selection using a pH gradient switching + inhibitor synergistic system.

[0090] After copper preferential flotation, the difference between lead and zinc, as well as pyrite, has narrowed. At this point, a new selective method is needed to ensure efficient lead separation. This scheme designs a pH gradient switching system and a 2-mercaptobenzothiazole (MBT) + polyethylene glycol (PEG) combined inhibitor system to achieve galena flotation. The pH for copper flotation is 9.5, while lead-zinc separation requires a higher pH to amplify the difference in floatability between galena and sphalerite. When MBT is used alone, the adsorption difference between galena and sphalerite is low, which cannot meet the requirements for efficient lead-zinc separation. Adding PEG will significantly improve the difference.

[0091] Specifically, when the pH is increased to 10.5, the hydrophilicity of the sphalerite surface increases, forming Zn(OH)2, further reducing its natural floatability; galena maintains good floatability in this pH range, achieving first-layer selectivity. The -SH group (pKa≈6.9) of MBT partially dissociates into -S at pH 10.5. - , with Pb 2+ Strong Pb-S bonds are formed, while for Zn 2+ The chelation stability is relatively low, achieving a second layer of selectivity. Ultimately, PEG molecules form a hydrogen-bonded hydration layer with the surface of sphalerite through ether oxygen bonds, physically hindering the adsorption of MBT on the sphalerite surface, while not interfering with the adsorption of MBT on galena, significantly enhancing the difference in reagent adsorption on the two mineral surfaces.

[0092] The principle of online linkage feedback control of Na2CO3 / NaHCO3 buffer system + pH-Eh.

[0093] The buffering principle of the Na2CO3 / NaHCO3 buffer system is that Na2CO3 and NaHCO3 establish chemical equilibrium in water as shown in chemical formulas (Ⅰ) and (Ⅱ):

[0094] The effective buffer zone of the Na2CO3 / NaHCO3 buffer system is pH = 9.0~10.0 (corresponding to CO3). 2- / HCO3 - (Buffer pair), but by adjusting the molar ratio of Na2CO3 to NaHCO3, its buffer range can be extended to pH=8.0~10.5, covering the pH requirements of copper flotation and lead flotation sections.

[0095] The mechanism of action of the Na2CO3 / NaHCO3 buffer system with online pH-Eh linkage feedback control is as follows: Figure 3As shown, without precise pH-Eh control, the first three core elements will rapidly deteriorate under fluctuating slurry conditions. In actual production, fluctuations in raw ore grade, changes in feed rate, residual reagents in the return water, and mineral dissolution continuously disturb slurry conditions. The copper flotation stage requires pH=9.5±0.2 and Eh=-100±10mV; the potential-responsive collector remains effective only within a 0±10mV window. Once the deviation exceeds this window, the collector easily loses its selectivity. Similarly, the pH=10.5±0.2 in the lead flotation stage and the pH=11.5±0.2 in the zinc flotation stage also need to be maintained stably. The Na₂CO₃ / NaHCO₃ buffer pair has a strong buffering capacity within the pH=8.0~10.5 range, resisting pH drift caused by disturbances such as ore dissolution and reagent addition. An online ORP sensor periodically samples Eh data, and the PLC compares the measured value with the set value in real time: when Eh is too low (excessive reducing power), a trace amount of H2O2 oxidant is automatically added; when Eh is too high (excessive oxidizing power), a trace amount of Na2S reducing agent is automatically added. There is a coupling relationship between pH and Eh (for every 1 unit increase in pH, Eh decreases by approximately 59 mV). The system achieves stable adsorption coverage by decoupling control, prioritizing pH accuracy and using this as a basis for feedforward compensation of Eh control output.

[0096] Next, through specific examples and comparative examples, the specific method and effects of the three-stage synergistic high-selectivity flotation of copper, lead, zinc and pyrite by pretreatment-potential-pH will be described in detail.

[0097] Example 1

[0098] like Figure 4 As shown, a three-stage synergistic pretreatment-potential-pH flotation method for copper-lead-zinc-pyrite ore is described. All stages of the flotation operation (roughing, cleaning, and scavenging) are performed within the flotation cell. During the flotation process, the ore pulp is continuously agitated and aerated to generate bubbles. These bubbles carry mineral particles to the surface, forming a foam layer. The product is collected by skimming the bubbles. The method specifically includes the following steps: In this embodiment, the composition and content of the copper-lead-zinc-pyrite ore include: 0.57% copper (Cu), 1.01% lead (Pb), 1.81% zinc (Zn), 7.65% iron (Fe), 2.43% sulfur (S), 1.12% calcium oxide (CaO), 4.98% aluminum oxide (Al2O3), 68.74% silicon dioxide (SiO2), 95.15 g / t silver (Ag*), 0.88 g / t gold (Au*), and 11.2% magnesium oxide (MgO), with the remainder being other trace impurities and loss on ignition.

[0099] S1, Heavy Medium Swirl Pre-Classification: 1 kg of copper-lead-zinc-pyrite ore is taken and crushed to a particle size of -0.5 mm (i.e., particle size less than or equal to 0.5 mm). Then, wet screening and classification are performed. The slurry with a particle size of +0.1 mm obtained from screening is fed into the feed mixing tank of a heavy media hydrocyclone. A heavy media suspension is added to the feed mixing tank to mix with the slurry before feeding it into an FHMC350 heavy media hydrocyclone. The slurry with a particle size of -0.1 mm that has been screened is combined with the tailings. In this embodiment, the heavy media suspension has a density of 6.8 g / cm³. 3 Ferrosilicon powder (Fe-Si) was used as a weighting agent and mixed with water to prepare the suspension. The total mass percentage of ferrosilicon in the heavy medium suspension was 80-82%; the feed pressure was 0.08 MPa.

[0100] The underflow from the heavy medium hydrocyclone is the heavy product (i.e., the enriched product of pyrite minerals). After the medium is removed by the desizing screen, it enters the grinding-flotation loop. The overflow from the heavy medium hydrocyclone is the light product. After the medium is removed by the desizing screen, it is combined with the tailings. The heavy medium suspension under the desizing screen is used by the magnetic separator to recover ferrosilicon. The recovered ferrosilicon powder is recycled.

[0101] S2. Preparation of slurry: The pyrite enrichment product was ground until the proportion of copper-lead-zinc-pyrite ore at -0.074 mm reached 65%, resulting in a pyrite enrichment slurry. The grinding method was ball milling for 10 minutes. Water was added to adjust the concentration of copper-lead-zinc-pyrite ore powder in the pyrite enrichment slurry to 30%. A mixed aqueous solution of Na2CO3 and NaHCO3 with a total concentration of 0.1 mol / L (hereinafter referred to as Na2CO3 / NaHCO3 aqueous solution) was added to adjust the pH of the pyrite enrichment slurry to 9.5. The molar ratio of Na2CO3 to NaHCO3 was 1:7, and the Na2CO3 and NaHCO3 were mixed together to prepare the Na2CO3 / NaHCO3 aqueous solution.

[0102] S3, pH-Eh online linkage feedback control flotation system: A pH-Eh online linkage feedback control flotation system is installed in the feed pipes of the roughing and cleaning cells in each flotation section. The Eh sensor is a platinum electrode-Ag / AgCl reference electrode composite ORP sensor, and the pH sensor is an industrial-grade glass electrode pH sensor with automatic temperature compensation. A PLC (Programmable Logic Controller) is used as the core control unit for real-time data acquisition. The pH-Eh online linkage control employs a decoupling algorithm. Based on the influence coefficient of pH on Eh (empirical value approximately -59mV / pH, meaning that for every unit increase in pH, the system equilibrium Eh shifts negatively by approximately 59mV), the system synchronously feeds forward to compensate the Eh control output during pH adjustment. Specifically, when Eh is too low (i.e., excessive reducing power), a trace amount of H2O2 oxidant is automatically added; when Eh is too high (i.e., excessive oxidizing power), a trace amount of Na2S reducing agent is automatically added. The data sampling frequency of the pH and ORP sensors is set to once every 3 seconds, and the pulp conditioning time is 3 minutes, finely adjusting the potential of the pyrite-rich pulp to -90mV.

[0103] S4, chalcopyrite flotation: S41, chalcopyrite roughing: The above-mentioned pyrite-enriched slurry was transferred to the roughing cell of the copper flotation section. 20 g / t collector 2-(benzothiazolylthio)acetic acid (i.e., 20 g collector per ton of pyrite-enriched slurry), 15 g / t frother MIBC (methyl isobutyl methanol) (i.e., 15 g frother per ton of pyrite-enriched slurry), and lead-zinc inhibitor were added to the pyrite-enriched slurry. The lead-zinc inhibitor was a combination of zinc sulfate (ZnSO4) and sodium sulfite (Na2SO3), with ZnSO4 added at a rate of 200 g / t (i.e., 200 g ZnSO4 per ton of pyrite-enriched slurry) and Na2SO3 added at a rate of 100 g / t (i.e., 100 g Na2SO3 per ton of pyrite-enriched slurry). One roughing of chalcopyrite was performed, with a frothing time of 8 minutes. Copper roughing concentrate slurry and copper roughing tailings slurry were obtained.

[0104] S42, Chalcopyrite Refinement: The copper rougher concentrate pulp was transferred to the copper flotation section's cleaning cell, where a blank, reagent-free cleaning method was employed. This method means that no additional flotation reagents such as collectors, frothers, or depressants were added, and basic flotation operating conditions such as agitation and aeration were maintained as normal. Subsequent operations followed the same procedure. Two chalcopyrite cleaning processes were performed: a first chalcopyrite cleaning of the copper rougher concentrate pulp yielded a first-stage copper cleaning pulp and a first-stage copper cleaning middlings pulp; a second chalcopyrite cleaning of the first-stage copper cleaning pulp yielded a second-stage copper cleaning pulp and a second-stage copper cleaning middlings pulp. The skimming time for both cleaning processes was 4 minutes.

[0105] The middlings slurry from the first copper beneficiation process is returned to the pyrite-rich slurry, and the middlings slurry from the second copper beneficiation process is returned to the copper roughing concentrate slurry. The second copper beneficiation slurry is dried to obtain chalcopyrite concentrate.

[0106] S43, scavenging of chalcopyrite: The copper roughing tailings slurry was subjected to two scavenging processes: First, 6 g / t of collector 2-(benzothiazolylthio)acetic acid (6 g collector per ton of copper roughing tailings slurry) and 6 g / t of frother MIBC (6 g frother per ton of copper roughing tailings slurry) were added to the copper roughing tailings slurry for the first chalcopyrite scavenging, yielding the first copper scavenged concentrate slurry and the first copper scavenged tailings slurry. Then, 3 g / t of collector 2-(benzothiazolylthio)acetic acid (3 g collector per ton of first copper scavenged tailings slurry) and 6 g / t of frother MIBC (6 g frother per ton of first copper scavenged tailings slurry) were added to the first copper scavenged tailings slurry for the second chalcopyrite scavenging, yielding the second copper scavenged concentrate slurry and the second copper scavenged tailings slurry. The frothing time for both scavenging processes was 4 minutes.

[0107] The first copper scavenging concentrate pulp is returned to the pyrite-rich pulp, and the second copper scavenging concentrate pulp is returned to the copper roughing tailings pulp; the second copper scavenging tailings pulp enters the next step of galena flotation.

[0108] S5, Galena flotation: S51, Galena roughing: Before the second copper scavenging tailings slurry enters the galena roughing process, pH gradient switching and slurry condition adjustment are required. Specifically, a Na2CO3 / NaHCO3 aqueous solution with a total Na2CO3 / NaHCO3 concentration of 0.1 mol / L is added to the second copper scavenging tailings slurry to adjust the pH value to 10.5, and the molar ratio of Na2CO3 to NaHCO3 is 1:2.1. The aeration rate and Na2S addition are adjusted through the pH-Eh online linkage control system to adjust the Eh of the second copper scavenging tailings slurry to -40mV. Next, a collector, a combination of 2-mercaptobenzothiazole (MBT) and polyethylene glycol (PEG), was added to the second copper scavenging tailings slurry after Eh adjustment. The amount of MBT added was 25 g / t (i.e., 25 g MBT per ton of second copper scavenging tailings slurry), and the amount of PEG added was 25 g / t (i.e., 25 g PEG per ton of second copper scavenging tailings slurry). 30 g / t of lead activator Na2S was added (i.e., 30 g lead activator per ton of second copper scavenging tailings slurry). A zinc inhibitor, a combination of ZnSO4 and Na2SO3, was added. The amount of ZnSO4 added was 200 g / t (i.e., 200 g ZnSO4 per ton of second copper scavenging tailings slurry), and the amount of Na2SO3 added was 100 g / t (i.e., 100 g Na2SO3 per ton of second copper scavenging tailings slurry). Na2SO3); add 20g / t of frother MIBC (i.e., add 20g of frother to each ton of the second copper scavenging tailings slurry); carry out the first roughing of galena, with a frothing time of 8min; and obtain lead roughing concentrate slurry and lead roughing tailings slurry.

[0109] S52, Galena concentrate: The lead rougher concentrate slurry was transferred to the lead flotation section's cleaning cell, and a blank, reagent-free cleaning method was used to perform two galena cleaning processes. Specifically, the lead rougher concentrate slurry underwent a first galena cleaning process to obtain a first-cleaned lead slurry and a first-cleaned middlings slurry; then, the first-cleaned lead slurry underwent a second galena cleaning process to obtain a second-cleaned lead slurry and a second-cleaned middlings slurry; the skimming time for both cleaning processes was 4 minutes.

[0110] The middlings slurry from the first lead beneficiation process is returned to the tailings slurry from the second copper scavenging process, and the middlings slurry from the second lead beneficiation process is returned to the lead roughing concentrate slurry. The second lead beneficiation slurry is dried to obtain galena concentrate.

[0111] S53, Galena scavenging: The lead roughing tailings slurry undergoes two scavenging processes: First, a combination of MBT and PEG is added as a collector to the lead roughing tailings slurry, with MBT added at a rate of 4 g / t (i.e., 4 g MBT per ton of lead roughing tailings slurry) and PEG added at a rate of 2 g / t (i.e., 2 g PEG per ton of lead roughing tailings slurry). Then, 5 g / t of MIBC (i.e., 5 g frother per ton of lead roughing tailings slurry) is added for the first galena scavenging, yielding the first lead scavenged concentrate slurry and the first lead scavenged tailings slurry. Next, 4 g / t MBT, 2 g / t PEG, and 5 g / t MIBC (i.e., 4 g MBT, 2 g PEG, and 5 g MIBC per ton of the first lead scavenged tailings slurry) are added to the first lead scavenged tailings slurry. PEG and 5g of foaming agent were used to carry out a second galena scavenging to obtain a second lead scavenging concentrate slurry and a second lead scavenging tailings slurry; the foaming time for both scavenging was 4 minutes.

[0112] The lead scavenging concentrate slurry from the first scavenging is returned to the tailings slurry from the second copper scavenging, and the lead scavenging concentrate slurry from the second scavenging is returned to the lead roughing tailings slurry; the tailings slurry from the second scavenging proceeds to the next step of sphalerite flotation.

[0113] S6, Sphalerite flotation: S61, Sphalerite roughing: Before the second lead scavenging tailings slurry enters the sphalerite roughing process, it needs to undergo activation treatment. Specifically, sphalerite activator, copper sulfate (CuSO4), is added to the second lead scavenging tailings slurry at a rate of 200 g / t (i.e., 200 g of activator is added per ton of second lead scavenging tailings slurry). Lime slurry is added to adjust the slurry pH to 11.5. At the same time, the aeration rate and the amount of lime slurry added are adjusted through the pH-Eh linkage control system to adjust the Eh of the second lead scavenging tailings slurry to +10 mV.

[0114] Next, 30 g / t of butyl xanthate (sodium butyl xanthate) was added as a collector to the activated and conditioned second lead scavenging tailings slurry, and 15 g / t of MIBC was added as a frother; a first roughing of sphalerite was carried out, with a frothing time of 8 min; zinc roughing concentrate slurry and zinc roughing tailings slurry were obtained.

[0115] S62, Sphalerite Refined: The zinc rougher concentrate slurry was transferred to the zinc flotation section's cleaning cell, and two sphalerite cleaning processes were performed using a blank, reagent-free cleaning method. Specifically, the zinc rougher concentrate slurry underwent a first sphalerite cleaning to obtain a first-cleaned zinc slurry and a first-cleaned middlings slurry; then, the first-cleaned zinc slurry underwent a second sphalerite cleaning to obtain a second-cleaned zinc slurry and a second-cleaned middlings slurry; the skimming time for both cleaning processes was 4 minutes.

[0116] The middlings slurry from the first zinc beneficiation process is returned to the tailings slurry from the second lead scavenging process, and the middlings slurry from the second zinc beneficiation process is returned to the zinc roughing concentrate slurry. The second zinc beneficiation slurry is dried to obtain sphalerite concentrate.

[0117] S63, Sphalerite scavenging: The zinc roughing tailings slurry was subjected to two scavenging processes: 14 g / t xanthate was added to the zinc roughing tailings slurry as a collector (i.e., 14 g of collector was added per ton of zinc roughing tailings slurry); 5 g / t frother MIBC was added for the first sphalerite scavenging, resulting in the first zinc scavenged concentrate slurry and the first zinc scavenged tailings slurry; then, 10 g / t xanthate and 5 g / t MIBC were added to the first zinc scavenged tailings slurry for the second sphalerite scavenging, resulting in the second zinc scavenged concentrate slurry and the second zinc scavenged tailings slurry; the frothing time for both scavenging processes was 4 minutes.

[0118] The first zinc scavenging concentrate slurry is returned to the second lead scavenging tailings slurry, and the second zinc scavenging concentrate slurry is returned to the zinc roughing tailings slurry; the second zinc scavenging tailings slurry enters the next sulfur concentrate flotation step.

[0119] S7, Sulfur Concentrate Flotation: First, the tailings slurry from the second zinc scavenging process is activated. The activator for the sulfur concentrate is copper sulfate (CuSO4), and the amount of CuSO4 added is 100g / t (i.e., 100g of activator is added to each ton of tailings slurry from the second zinc scavenging process). Sulfuric acid (H2SO4) is added to adjust the pH of the slurry to 6.5.

[0120] Next, roughing of the sulfur concentrate is carried out. 80 g / t butyl xanthate is added as a collector and 20 g / t frother MIBC is added to the activated and conditioned tailings slurry from the second zinc flotation. A first roughing of the sulfur concentrate is then performed, with a frothing time of 8 minutes. This yields a sulfur roughing concentrate slurry and a sulfur flotation tailings slurry. The sulfur roughing concentrate slurry is then dried to obtain the sulfur concentrate.

[0121] S8. Tailings treatment: 30 g / t of polyacrylamide (PAM) was added as a flocculant to the sulfur flotation tailings slurry. After thickening in a high-efficiency thickener for 8 hours, the thickened overflow was mixed with the pyrite-enriched slurry and then fed into the chalcopyrite roughing operation. The residual reagent concentration in the reflux water was reduced to a level that would not affect the flotation selectivity after harmless treatment. The harmlessly treated reflux water was returned to the chalcopyrite flotation step to reduce the water consumption of the entire process. The underflow tailings were sent to a filter press for dewatering. The resulting tailings filter cake (with a moisture content of 12%) was dry-stacked and then sent to the tailings dam. The harmless treatment was carried out in accordance with GB 25466-2010, and will not be elaborated here.

[0122] Example 2

[0123] A pretreatment-potential-pH three-stage synergistic flotation method for copper-lead-zinc-pyrite ore is disclosed in this embodiment. The process flow for flotation of copper-lead-zinc-pyrite ore in this embodiment is the same as that in Example 1, and specifically includes the following steps: In this embodiment, the composition and content of the copper-lead-zinc-pyrite ore include: 0.56% copper (Cu), 1.06% lead (Pb), 1.84% zinc (Zn), 8.42% iron (Fe), 2.49% sulfur (S), 0.58% calcium oxide (CaO), 4.15% aluminum oxide (Al2O3), 69.98% silicon dioxide (SiO2), 90.33 g / t silver (Ag*), 0.85 g / t gold (Au*), and 10.43% magnesium oxide (MgO), with the remainder being other trace impurities and loss on ignition.

[0124] S1, Heavy Medium Swirl Pre-Classification: Take 1 kg of copper-lead-zinc-pyrite ore and crush it to a particle size of -0.5 mm. Then, perform wet screening and classification. The screening and classification method is the same as described in Example 1, except that the feed pressure is 0.10 MPa. The underflow of the FHMC350 heavy medium hydrocyclone, which is the heavy product (i.e., the pyrite mineral enrichment product), enters the grinding-flotation loop after the medium is removed by the demedium screen.

[0125] S2. Preparation of slurry: The pyrite-enriched product was ground until the proportion of copper-lead-zinc pyrite at -0.074 mm reached 70%, resulting in a pyrite-enriched slurry. The grinding method was ball milling, with a milling time of 11 min. Water was added to adjust the concentration of copper-lead-zinc pyrite powder in the pyrite-enriched slurry to 33%. A Na₂CO₃ / NaHCO₃ aqueous solution with a total concentration of 0.1 mol / L was added to adjust the pH of the pyrite-enriched slurry to 9.5; the molar ratio of Na₂CO₃ to NaHCO₃ was 1:7.

[0126] S3, pH-Eh online linkage feedback control flotation system: In each flotation stage, the roughing and cleaning cells are equipped with the same pH-Eh online linkage feedback control flotation system as described in Example 1. The pH-Eh online linkage control also employs a decoupled algorithm, synchronously feeding forward to compensate the Eh control output during pH adjustment based on the pH-Eh influence coefficient of -59mV / pH. The difference from Example 1 is that the data sampling frequency of the pH and ORP sensors is set to once every 4 seconds, the pulp conditioning time is 4 minutes, and the potential of the pyrite-rich pulp is fine-tuned to -100mV.

[0127] S4, chalcopyrite flotation: S41, chalcopyrite roughing: The pyrite-enriched slurry was transferred to the roughing cell of the copper flotation section. 30 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 30 g collector per ton of pyrite-enriched slurry) was added to the pyrite-enriched slurry; 20 g / t of frother MIBC (i.e., 20 g frother per ton of pyrite-enriched slurry) was added; and a lead-zinc depressant, a combination of ZnSO4 and Na2SO3, was added, with ZnSO4 added at 300 g / t (i.e., 300 g ZnSO4 per ton of pyrite-enriched slurry) and Na2SO3 added at 150 g / t (i.e., 150 g Na2SO3 per ton of pyrite-enriched slurry). One roughing of chalcopyrite was performed, with a frothing time of 10 minutes. Copper roughing concentrate slurry and copper roughing tailings slurry were obtained.

[0128] S42, Chalcopyrite Refinement: The same blank, untreated chalcopyrite selection method as in Example 1 was used for two chalcopyrite selections, the difference being that the skimming time for both selections was 5 minutes.

[0129] The middlings slurry from the first copper beneficiation process is returned to the pyrite-rich slurry, and the middlings slurry from the second copper beneficiation process is returned to the copper roughing concentrate slurry. The second copper beneficiation slurry is dried to obtain chalcopyrite concentrate.

[0130] S43, scavenging of chalcopyrite: The copper roughing tailings slurry was subjected to two scavenging processes: First, 8 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 8 g of collector per ton of copper roughing tailings slurry) and 8 g / t of frother MIBC (i.e., 8 g of frother per ton of copper roughing tailings slurry) were added to the copper roughing tailings slurry for the first chalcopyrite scavenging, yielding the first copper scavenged concentrate slurry and the first copper scavenged tailings slurry. Then, 4 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 4 g of collector per ton of the first copper scavenged tailings slurry) and 8 g / t of frother MIBC (i.e., 8 g of frother per ton of the first copper scavenged tailings slurry) and 8 g / t of frother MIBC (i.e., 8 g of frother per ton of the first copper scavenged tailings slurry) were added to the first copper scavenged tailings slurry for the second chalcopyrite scavenging, yielding the second copper scavenged concentrate slurry and the second copper scavenged tailings slurry. The frothing time for both scavenging processes was 5 minutes.

[0131] The first copper scavenging concentrate pulp is returned to the pyrite-rich pulp, and the second copper scavenging concentrate pulp is returned to the copper roughing tailings pulp; the second copper scavenging tailings pulp enters the next step of galena flotation.

[0132] S5, Galena flotation: S51, Galena roughing: Before the second copper scavenging tailings slurry enters the galena roughing process, Na2CO3 / NaHCO3 aqueous solution is added to the second copper scavenging tailings slurry to switch the pH gradient and adjust the slurry conditions, adjusting the pH value to 10.5. The difference from Example 1 is that the Eh of the second copper scavenging tailings slurry is adjusted to -50mV through the pH-Eh online linkage control system. Next, a collector, a combination of MBT and PEG, was added to the second copper scavenging tailings slurry after Eh adjustment. The MBT addition was 34 g / t (i.e., 34 g MBT per ton of the second copper scavenging tailings slurry), and the PEG addition was 17 g / t (i.e., 17 g PEG per ton of the second copper scavenging tailings slurry). 35 g / t of lead activator Na2S was added (i.e., 35 g lead activator per ton of the second copper scavenging tailings slurry). A zinc inhibitor, a combination of ZnSO4 and Na2SO3, was added. The ZnSO4 addition was 300 g / t (i.e., 300 g ZnSO4 per ton of the second copper scavenging tailings slurry), and the Na2SO3 addition was 150 g / t (i.e., 150 g Na2SO3 per ton of the second copper scavenging tailings slurry). Na2SO3); add 25g / t of frother MIBC (i.e., add 25g of frother to each ton of the second copper scavenging tailings slurry); carry out the first roughing of galena, with a frothing time of 10min; and obtain lead roughing concentrate slurry and lead roughing tailings slurry.

[0133] S52, Galena concentrate: The lead roughing concentrate slurry was transferred to the lead flotation section for cleaning. The same blank, reagent-free cleaning method as in Example 1 was used for two-stage cleaning of lead ore. The difference was that the skimming time for both cleaning stages in this example was 5 minutes.

[0134] The middlings slurry from the first lead beneficiation process is returned to the tailings slurry from the second copper scavenging process, and the middlings slurry from the second lead beneficiation process is returned to the lead roughing concentrate slurry. The second lead beneficiation slurry is dried to obtain galena concentrate.

[0135] S53, Galena scavenging: The lead roughing tailings slurry undergoes two scavenging processes: First, a combination of MBT and PEG is added as a collector to the lead roughing tailings slurry, with MBT added at a rate of 6 g / t (6 g MBT per ton of lead roughing tailings slurry) and PEG added at a rate of 3 g / t (3 g PEG per ton of lead roughing tailings slurry). Then, 8 g / t of MIBC (8 g frother per ton of lead roughing tailings slurry) is added for the first galena scavenging, yielding the first lead scavenged concentrate slurry and the first lead scavenged tailings slurry. Next, 6 g / t MBT, 3 g / t PEG, and 8 g / t MIBC (6 g MBT, 3 g PEG, and 8 g PEG per ton of first lead scavenged tailings slurry) are added to the first lead scavenged tailings slurry. PEG and 8g of foaming agent were used to carry out a second galena scavenging to obtain a second lead scavenging concentrate slurry and a second lead scavenging tailings slurry; the foaming time for both scavenging was 5 minutes.

[0136] The lead scavenging concentrate slurry from the first scavenging is returned to the tailings slurry from the second copper scavenging, and the lead scavenging concentrate slurry from the second scavenging is returned to the lead roughing tailings slurry; the tailings slurry from the second scavenging proceeds to the next step of sphalerite flotation.

[0137] S6, Sphalerite flotation: S61, Sphalerite roughing: Before the second lead scavenging tailings slurry enters the sphalerite roughing process, it needs to be activated. Specifically, 300g / t of activator CuSO4 is added to the second lead scavenging tailings slurry (i.e., 300g of activator is added per ton of the second lead scavenging tailings slurry); lime slurry is added to adjust the slurry pH to 11.5, and the aeration rate and lime slurry addition rate are adjusted through the pH-Eh linkage control system to adjust the Eh of the second lead scavenging tailings slurry to +5mV.

[0138] Next, 45 g / t butyl xanthate and 20 g / t MIBC were added to the activated and conditioned second lead scavenging tailings slurry; a first roughing of sphalerite was carried out, with a frothing time of 10 min; zinc roughing concentrate slurry and zinc roughing tailings slurry were obtained.

[0139] S62, Sphalerite Refined: The zinc roughing concentrate slurry was transferred to the zinc flotation section for two rounds of sphalerite cleaning, using the same blank, reagent-free cleaning method as in Example 1. The difference was that in this example, the skimming time for both cleaning rounds was 5 minutes.

[0140] The middlings slurry from the first zinc beneficiation process is returned to the tailings slurry from the second lead scavenging process, and the middlings slurry from the second zinc beneficiation process is returned to the zinc roughing concentrate slurry. The second zinc beneficiation slurry is dried to obtain sphalerite concentrate.

[0141] S63, Sphalerite scavenging: The zinc roughing tailings slurry was subjected to two scavenging processes: 16 g / t xanthate was added to the zinc roughing tailings slurry as a collector (i.e., 16 g of collector was added per ton of zinc roughing tailings slurry); 8 g / t frother MIBC was added for the first sphalerite scavenging, resulting in the first zinc scavenged concentrate slurry and the first zinc scavenged tailings slurry; then 12 g / t xanthate and 8 g / t MIBC were added to the first zinc scavenged tailings slurry for the second sphalerite scavenging, resulting in the second zinc scavenged concentrate slurry and the second zinc scavenged tailings slurry; the frothing time for both scavenging processes was 5 minutes.

[0142] The first zinc scavenging concentrate slurry is returned to the second lead scavenging tailings slurry, and the second zinc scavenging concentrate slurry is returned to the zinc roughing tailings slurry; the second zinc scavenging tailings slurry enters the next sulfur concentrate flotation step.

[0143] S7, Sulfur Concentrate Flotation: First, the tailings slurry from the second zinc scavenging process is activated by adding 150g / t of activator CuSO4 (i.e., 150g of activator is added to each ton of tailings slurry from the second zinc scavenging process); H2SO4 is added to adjust the pH of the slurry to 6.5.

[0144] Next, roughing of the sulfur concentrate is carried out. 100 g / t butyl xanthate is added as a collector and 25 g / t frother MIBC is added to the activated and conditioned tailings slurry from the second zinc flotation. A first roughing of the sulfur concentrate is then performed, with a frothing time of 10 minutes. This yields a sulfur roughing concentrate slurry and a sulfur flotation tailings slurry. The sulfur roughing concentrate slurry is then dried to obtain the sulfur concentrate.

[0145] S8. Tailings treatment: 40 g / t of flocculant PAM is added to the sulfur flotation tailings slurry, which is then concentrated in a high-efficiency thickener for 9 hours. The thickened overflow is mixed with the pyrite-rich slurry and then enters the chalcopyrite roughing operation. The residual reagent concentration in the reflux water is reduced to a level that does not affect the flotation selectivity after harmless treatment. The harmlessly treated reflux water is returned to the chalcopyrite flotation step to reduce the water consumption of the entire process. The underflow tailings are sent to a filter press for dewatering, and the resulting tailings filter cake (with a moisture content of 11%) is dry-stacked and then sent to the tailings dam.

[0146] Example 3

[0147] A pretreatment-potential-pH three-stage synergistic flotation method for copper-lead-zinc-pyrite ore is disclosed. The process flow for flotation of copper-lead-zinc-pyrite ore in this embodiment is the same as in Example 1, specifically including the following steps: In this embodiment, the composition and content of the copper-lead-zinc-pyrite ore include: 0.56% copper (Cu), 1.08% lead (Pb), 1.86% zinc (Zn), 8.32% iron (Fe), 2.47% sulfur (S), 0.64% calcium oxide (CaO), 4.23% aluminum oxide (Al2O3), 70.55% silicon dioxide (SiO2), 88.24 g / t silver (Ag*), 0.81 g / t gold (Au*), and 9.6% magnesium oxide (MgO), with the remainder being other trace impurities and loss on ignition.

[0148] S1, Heavy Medium Swirl Pre-Classification: Take 1 kg of copper-lead-zinc pyrite ore and crush it to a particle size of -0.5 mm. Then, perform wet screening and classification. The screening and classification method is the same as described in Example 1, except that the feed pressure is 0.12 MPa. The underflow of the FHMC350 heavy medium hydrocyclone, which is the heavy product (i.e., the pyrite enrichment product), enters the grinding-flotation loop after the medium is removed by the demedium screen.

[0149] S2. Preparation of slurry: The pyrite-enriched product was ground until the proportion of copper-lead-zinc pyrite at -0.074 mm reached 75%, yielding a pyrite-enriched slurry. The grinding method was ball milling, with a milling time of 11 min. Water was added to adjust the concentration of copper-lead-zinc pyrite powder in the pyrite-enriched slurry to 33%. A Na₂CO₃ / NaHCO₃ aqueous solution with a total concentration of 0.1 mol / L was added to adjust the pH of the pyrite-enriched slurry to 9.5; the molar ratio of Na₂CO₃ to NaHCO₃ was 1:7.

[0150] S3, pH-Eh online linkage feedback control flotation system: In each flotation section, the roughing and cleaning cells were equipped with the same pH-Eh online linkage feedback control flotation system as described in Example 1. The pH-Eh online linkage control also employed a decoupled algorithm, synchronously feeding forward to compensate the Eh control output during pH adjustment based on the pH-Eh influence coefficient of -59mV / pH. The difference from Example 1 is that when Eh is too low (i.e., excessively reducing), a trace amount of H2O2 oxidant is automatically added; when Eh is too high (i.e., excessively oxidizing), a trace amount of Na2S reducing agent is automatically added. The data sampling frequency of the pH and ORP sensors was set to once every 5 seconds, and the pulp conditioning time was 5 minutes, finely adjusting the potential of the pyrite-rich pulp to -110mV.

[0151] S4, chalcopyrite flotation: S41, chalcopyrite roughing: The pyrite-enriched slurry was transferred to the roughing cell of the copper flotation section. 35 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 35 g collector per ton of pyrite-enriched slurry) was added; 25 g / t of frother MIBC (i.e., 25 g frother per ton of pyrite-enriched slurry) was added; and a lead-zinc depressant, a combination of zinc sulfate (ZnSO4) and sodium sulfite (Na2SO3), was added. The amount of ZnSO4 added was 500 g / t (i.e., 500 g ZnSO4 per ton of pyrite-enriched slurry), and the amount of Na2SO3 added was 200 g / t (i.e., 200 g Na2SO3 per ton of pyrite-enriched slurry). One roughing of chalcopyrite was performed, with a frothing time of 12 minutes. Copper roughing concentrate slurry and copper roughing tailings slurry were obtained.

[0152] S42, Chalcopyrite Refinement: The same blank, untreated chalcopyrite selection method as in Example 1 was used for two chalcopyrite selections, the difference being that the skimming time for both selections was 6 minutes.

[0153] The middlings slurry from the first copper beneficiation process is returned to the pyrite-rich slurry, and the middlings slurry from the second copper beneficiation process is returned to the copper roughing concentrate slurry. The second copper beneficiation slurry is dried to obtain chalcopyrite concentrate.

[0154] S43, scavenging of chalcopyrite: The copper roughing tailings slurry undergoes two scavenging processes: First, 10 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 10 g of collector per ton of copper roughing tailings slurry) and 10 g / t of frother MIBC (i.e., 10 g of frother per ton of copper roughing tailings slurry) are added to the slurry for the first chalcopyrite scavenging, yielding the first copper scavenged concentrate slurry and the first copper scavenged tailings slurry; then... In the first copper scavenging tailings slurry, 5 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 5 g of collector per ton of the first copper scavenging tailings slurry) and 10 g / t of frother MIBC (i.e., 10 g of frother per ton of the first copper scavenging tailings slurry) were added to conduct the second chalcopyrite scavenging, resulting in the second copper scavenging concentrate slurry and the second copper scavenging tailings slurry; the frothing time for both scavenging processes was 6 minutes.

[0155] The first copper scavenging concentrate pulp is returned to the pyrite-rich pulp, and the second copper scavenging concentrate pulp is returned to the copper roughing tailings pulp; the second copper scavenging tailings pulp enters the next step of galena flotation.

[0156] S5, Galena flotation: S51, Galena roughing: Before the second copper scavenging tailings slurry enters the galena roughing process, Na2CO3 / NaHCO3 aqueous solution is added to the second copper scavenging tailings slurry to switch the pH gradient and adjust the slurry conditions, adjusting the pH value to 10.5. The difference from Example 1 is that the Eh of the second copper scavenging tailings slurry is adjusted to -60mV through the pH-Eh online linkage control system. Next, a collector, a combination of MBT and PEG, was added to the second copper scavenging tailings slurry after Eh adjustment. The MBT addition was 50 g / t (i.e., 50 g MBT per ton of second copper scavenging tailings slurry), and the PEG addition was 25 g / t (i.e., 25 g PEG per ton of second copper scavenging tailings slurry). 40 g / t of lead activator Na2S was added (i.e., 40 g lead activator per ton of second copper scavenging tailings slurry). A zinc inhibitor, a combination of ZnSO4 and Na2SO3, was added. The ZnSO4 addition was 500 g / t (i.e., 500 g ZnSO4 per ton of second copper scavenging tailings slurry), and the Na2SO3 addition was 200 g / t (i.e., 200 g Na2SO3 per ton of second copper scavenging tailings slurry). Na2SO3); add 30g / t of frother MIBC (i.e., add 30g of frother to each ton of the second copper scavenging tailings slurry); carry out the first roughing of galena, with a frothing time of 12min; and obtain lead roughing concentrate slurry and lead roughing tailings slurry.

[0157] S52, Galena concentrate: The lead roughing concentrate slurry was transferred to the lead flotation section for cleaning. The same blank, reagent-free cleaning method as in Example 1 was used for two-stage cleaning of lead ore. The difference was that the skimming time for both cleaning stages in this example was 6 minutes.

[0158] The middlings slurry from the first lead beneficiation process is returned to the tailings slurry from the second copper scavenging process, and the middlings slurry from the second lead beneficiation process is returned to the lead roughing concentrate slurry. The second lead beneficiation slurry is dried to obtain galena concentrate.

[0159] S53, Galena scavenging: The lead roughing tailings slurry undergoes two scavenging processes: First, a combination of MBT and PEG is added as a collector to the lead roughing tailings slurry, with MBT added at a rate of 8 g / t (i.e., 8 g MBT per ton of lead roughing tailings slurry) and PEG added at a rate of 4 g / t (i.e., 4 g PEG per ton of lead roughing tailings slurry). Then, 10 g / t of MIBC (i.e., 10 g frother per ton of lead roughing tailings slurry) is added for the first galena scavenging, yielding the first lead scavenged concentrate slurry and the first lead scavenged tailings slurry. Next, 8 g / t MBT, 4 g / t PEG, and 10 g / t MIBC (i.e., 8 g MBT, 4 g PEG, and 10 g MIBC per ton of the first lead scavenged tailings slurry) are added to the first lead scavenged tailings slurry. PEG and 10g of foaming agent were used to carry out a second galena scavenging to obtain a second lead scavenging concentrate slurry and a second lead scavenging tailings slurry; the foaming time for both scavenging was 6 minutes.

[0160] The lead scavenging concentrate slurry from the first scavenging is returned to the tailings slurry from the second copper scavenging, and the lead scavenging concentrate slurry from the second scavenging is returned to the lead roughing tailings slurry; the tailings slurry from the second scavenging proceeds to the next step of sphalerite flotation.

[0161] S6, Sphalerite flotation: S61, Sphalerite roughing: Before the second lead scavenging tailings slurry enters the sphalerite roughing process, it needs to be activated. Specifically, 350g / t of activator CuSO4 is added to the second lead scavenging tailings slurry (i.e., 350g of activator is added per ton of the second lead scavenging tailings slurry); lime slurry is added to adjust the slurry pH to 11.5, and the aeration rate and lime slurry addition rate are adjusted through the pH-Eh linkage control system to adjust the Eh of the second lead scavenging tailings slurry to -10mV.

[0162] Next, 60 g / t butyl xanthate and 25 g / t MIBC were added to the activated and conditioned second lead scavenging tailings slurry; a first roughing of sphalerite was carried out, with a frothing time of 12 min; zinc roughing concentrate slurry and zinc roughing tailings slurry were obtained.

[0163] S62, Sphalerite Refined: The zinc roughing concentrate slurry was transferred to the zinc flotation section for two rounds of sphalerite cleaning, using the same blank, reagent-free cleaning method as in Example 1. The difference was that in this example, the skimming time for both cleaning rounds was 6 minutes.

[0164] The middlings slurry from the first zinc beneficiation process is returned to the tailings slurry from the second lead scavenging process, and the middlings slurry from the second zinc beneficiation process is returned to the zinc roughing concentrate slurry. The second zinc beneficiation slurry is dried to obtain sphalerite concentrate.

[0165] S63, Sphalerite scavenging: The zinc roughing tailings slurry was subjected to two scavenging processes: 20 g / t xanthate was added to the zinc roughing tailings slurry as a collector (i.e., 20 g of collector was added per ton of zinc roughing tailings slurry); 10 g / t frother MIBC was added for the first sphalerite scavenging, resulting in the first zinc scavenged concentrate slurry and the first zinc scavenged tailings slurry; then, 15 g / t xanthate and 10 g / t MIBC were added to the first zinc scavenged tailings slurry for the second sphalerite scavenging, resulting in the second zinc scavenged concentrate slurry and the second zinc scavenged tailings slurry; the frothing time for both scavenging processes was 6 minutes.

[0166] The first zinc scavenging concentrate slurry is returned to the second lead scavenging tailings slurry, and the second zinc scavenging concentrate slurry is returned to the zinc roughing tailings slurry; the second zinc scavenging tailings slurry enters the next sulfur concentrate flotation step.

[0167] S7, Sulfur Concentrate Flotation: Specifically, the following steps are included: First, the tailings slurry from the second zinc scavenging process is activated by adding 200g / t of activator CuSO4 (i.e., 200g of activator is added to every ton of tailings slurry from the second zinc scavenging process); H2SO4 is added to adjust the pH of the slurry to 6.5.

[0168] Next, roughing of the sulfur concentrate is carried out. 120 g / t butyl xanthate is added as a collector and 30 g / t frother MIBC is added to the activated and conditioned tailings slurry from the second zinc flotation. A first roughing of the sulfur concentrate is then performed, with a frothing time of 12 minutes. This yields a sulfur roughing concentrate slurry and a sulfur flotation tailings slurry. The sulfur roughing concentrate slurry is then dried to obtain the sulfur concentrate.

[0169] S8. Tailings treatment: 50 g / t of flocculant PAM is added to the sulfur flotation tailings slurry, which is then concentrated in a high-efficiency thickener for 10 hours. The thickened overflow is mixed with the pyrite-rich slurry and then enters the chalcopyrite roughing operation. The residual reagent concentration in the reflux water is reduced to a level that does not affect the flotation selectivity after harmless treatment. The harmlessly treated reflux water is returned to the chalcopyrite flotation step to reduce the water consumption of the entire process. The underflow tailings are sent to a filter press for dewatering, and the resulting tailings filter cake (with a moisture content of 10%) is dry-stacked and then sent to the tailings dam.

[0170] Comparative Example 1 A flotation method for copper-lead-zinc-pyrite ore with pre-classification using a weightless medium is disclosed. The difference between this comparative example and Example 3 is the absence of a weightless medium for pre-classification; the remaining process flow is the same as in Example 3, specifically including the following steps: In this comparative example, the composition and content of the copper-lead-zinc pyrite ore include: 0.58% copper (Cu), 1.08% lead (Pb), 1.80% zinc (Zn), 8.14% iron (Fe), 2.51% sulfur (S), 0.88% calcium oxide (CaO), 4.12% aluminum oxide (Al2O3), 71.45% silicon dioxide (SiO2), 94.25 g / t silver (Ag*), 0.96 g / t gold (Au*), and 8.96% magnesium oxide (MgO), with the remainder being other trace impurities and loss on ignition.

[0171] S1. Preparation of slurry: Take 1 kg of copper-lead-zinc-pyrite ore and crush it to a particle size of -0.5 mm. Then, perform wet ball milling for 12 minutes until the proportion of copper-lead-zinc-pyrite ore with a particle size of -0.074 mm reaches 75%, thus obtaining a pyrite mineral slurry. Add water to adjust the concentration of copper-lead-zinc-pyrite ore powder in the pyrite mineral slurry to 33%. Add a Na2CO3 / NaHCO3 aqueous solution with a total concentration of 0.1 mol / L to adjust the pH of the pyrite mineral enrichment slurry to 9.5; the molar ratio of Na2CO3 to NaHCO3 is 1:7.

[0172] S2, pH-Eh online linkage feedback control flotation system: In each flotation section, the roughing and cleaning cells were equipped with the same pH-Eh online linkage feedback control flotation system as described in Example 3. The pH-Eh online linkage control also employed a decoupled algorithm. Based on the pH-Eh influence coefficient of -59mV / pH, the Eh control output was synchronously fed forward to compensate during pH adjustment. Specifically, when Eh was too low (i.e., excessively reducing), a trace amount of H2O2 oxidant was automatically added; when Eh was too high (i.e., excessively oxidizing), a trace amount of Na2S reducing agent was automatically added. The data sampling frequency of the pH and ORP sensors was set to once every 5 seconds, and the pulp conditioning time was 5 minutes, finely adjusting the pyrite pulp potential to -110mV.

[0173] S3, chalcopyrite flotation: S31, chalcopyrite roughing: The pyrite slurry was transferred to the roughing cell of the copper flotation section. 35 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 35 g of collector per ton of pyrite slurry) was added to the pyrite slurry; 25 g / t of frother MIBC (i.e., 25 g of frother per ton of pyrite slurry) was added; and a lead-zinc depressant, a combination of zinc sulfate (ZnSO4) and sodium sulfite (Na2SO3), was added, with ZnSO4 added at a rate of 500 g / t and Na2SO3 added at a rate of 200 g / t, i.e., 500 g of ZnSO4 and 200 g of Na2SO3 per ton of pyrite slurry, was added for one roughing of chalcopyrite. The frothing time for roughing was 12 minutes. Copper roughing concentrate slurry and copper roughing tailings slurry were obtained.

[0174] S32, Chalcopyrite Refinement: Two chalcopyrite beneficiation processes were performed using the same blank, reagent-free beneficiation method as in Example 3, resulting in chalcopyrite concentrate.

[0175] S33, scavenging of chalcopyrite: Following the method described in Example 3, the copper roughing tailings slurry was subjected to two scavenging processes: 10 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 10 g of collector per ton of copper roughing tailings slurry) and 10 g / t of frother MIBC (i.e., 10 g of frother per ton of copper roughing tailings slurry) were added to the copper roughing tailings slurry for the first chalcopyrite scavenging, yielding the first copper scavenged concentrate slurry and the first copper scavenged tailings slurry. The first copper scavenging tailings slurry was then subjected to a second chalcopyrite scavenging process. This involved adding 5 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 5 g of collector per ton of the first copper scavenging tailings slurry) and 10 g / t of frother MIBC (i.e., 10 g of frother per ton of the first copper scavenging tailings slurry). The frothing time for both scavenging processes was 6 minutes. The first copper scavenging concentrate slurry was returned to the pyrite-enriched slurry, and the second copper scavenging concentrate slurry was returned to the copper roughing tailings slurry. The second copper scavenging tailings slurry then proceeded to the next step: galena flotation.

[0176] S4, Galena flotation: S41, Galena roughing: Before the second copper scavenging tailings slurry enters the galena roughing process, Na2CO3 / NaHCO3 aqueous solution is added to the second copper scavenging tailings slurry to switch the pH gradient and adjust the slurry conditions, adjusting the pH value to 10.5; and the Eh of the second copper scavenging tailings slurry is adjusted to -60mV through the pH-Eh online linkage control system. Next, a collector, a combination of MBT and PEG, was added to the second copper scavenging tailings slurry after Eh adjustment. The MBT addition was 50 g / t (i.e., 50 g MBT per ton of second copper scavenging tailings slurry), and the PEG addition was 25 g / t (i.e., 25 g PEG per ton of second copper scavenging tailings slurry). 40 g / t of lead activator Na2S was added (i.e., 40 g lead activator per ton of second copper scavenging tailings slurry). A zinc inhibitor, a combination of ZnSO4 and Na2SO3, was added. The ZnSO4 addition was 500 g / t (i.e., 500 g ZnSO4 per ton of second copper scavenging tailings slurry), and the Na2SO3 addition was 200 g / t (i.e., 200 g Na2SO3 per ton of second copper scavenging tailings slurry). Na2SO3); add 30g / t of frother MIBC (i.e., add 30g of frother to each ton of the second copper scavenging tailings slurry); carry out the first roughing of galena, with a frothing time of 12min; and obtain lead roughing concentrate slurry and lead roughing tailings slurry.

[0177] S42, Galena concentrate: The lead rougher concentrate slurry was transferred to the lead flotation section for cleaning. The same blank, reagent-free cleaning method as in Example 3 was used to perform two galena cleaning processes to obtain galena concentrate.

[0178] S43, Galena scavenging: The lead roughing tailings slurry undergoes two scavenging processes: First, a combination of MBT and PEG is added as a collector to the lead roughing tailings slurry, with MBT added at a rate of 8 g / t (i.e., 8 g MBT per ton of lead roughing tailings slurry) and PEG added at a rate of 4 g / t (i.e., 4 g PEG per ton of lead roughing tailings slurry). Then, 10 g / t of MIBC (i.e., 10 g frother per ton of lead roughing tailings slurry) is added for the first galena scavenging, yielding the first lead scavenged concentrate slurry and the first lead scavenged tailings slurry. Next, 8 g / t MBT, 4 g / t PEG, and 10 g / t MIBC (i.e., 8 g MBT, 4 g PEG, and 10 g MIBC per ton of the first lead scavenged tailings slurry) are added to the first lead scavenged tailings slurry. PEG and 10g of frother were added to perform a second galena scavenging process, yielding a second lead scavenging concentrate slurry and a second lead scavenging tailings slurry. The frothing time for both scavenging processes was 6 minutes. The first lead scavenging concentrate slurry was returned to the second copper scavenging tailings slurry, and the second lead scavenging concentrate slurry was returned to the lead rougher tailings slurry. The second lead scavenging tailings slurry was then used in the next step, sphalerite flotation.

[0179] S5, Sphalerite flotation: S51, Sphalerite roughing: Before the second lead scavenging tailings slurry enters the sphalerite roughing process, it needs to undergo activation treatment, which is the same as in Example 3. Lime slurry is added to adjust the pH of the second lead scavenging tailings slurry to 11.5, and the aeration rate and lime slurry addition are adjusted using a pH-Eh linkage control system to adjust the Eh of the second lead scavenging tailings slurry to -10mV. Next, 60g / t butyl xanthate and 25g / t MIBC are added to the activated and adjusted second lead scavenging tailings slurry; a sphalerite roughing process is then performed, with a skimming time of 12 minutes; resulting in zinc roughing concentrate slurry and zinc roughing tailings slurry.

[0180] S52, Sphalerite Refined: The zinc roughing concentrate slurry was transferred to the zinc flotation section for cleaning, and two rounds of sphalerite cleaning were performed using the same blank, reagent-free cleaning method as in Example 3 to obtain sphalerite concentrate.

[0181] S53, Sphalerite scavenging: The zinc roughing tailings slurry undergoes two scavenging processes: 20 g / t butyl xanthate is added as a collector (i.e., 20 g of collector per ton of zinc roughing tailings slurry); 10 g / t frother MIBC is added for the first sphalerite scavenging, yielding the first zinc scavenged concentrate slurry and the first zinc scavenged tailings slurry; then, 15 g / t butyl xanthate and 10 g / t MIBC are added to the first zinc scavenged tailings slurry for the second sphalerite scavenging, yielding the second zinc scavenged concentrate slurry and the second zinc scavenged tailings slurry; the frothing time for both scavenging processes is 6 minutes. The first zinc scavenged concentrate slurry is returned to the second lead scavenged tailings slurry, and the second zinc scavenged concentrate slurry is returned to the zinc roughing tailings slurry; the second zinc scavenged tailings slurry proceeds to the next step, sulfur concentrate flotation.

[0182] S6, sulfur concentrate flotation: First, the tailings slurry from the second zinc scavenging process is activated by adding 200 g / t CuSO4 activator (i.e., 200 g of activator per ton of tailings slurry from the second zinc scavenging process); H2SO4 is added to adjust the pH of the slurry to 6.5. Next, sulfur concentrate roughing is performed by adding 120 g / t butyl xanthate as a collector and 30 g / t frother MIBC to the activated and adjusted tailings slurry for a first-stage sulfur concentrate roughing process with a frothing time of 12 minutes; this yields sulfur roughing concentrate slurry and sulfur flotation tailings slurry. The sulfur roughing concentrate slurry is then dried to obtain sulfur concentrate.

[0183] S7. Tailings Treatment: Add 50g / t of flocculant PAM to the sulfur flotation tailings slurry, and concentrate it for 10 hours using a high-efficiency thickener. The thickened overflow is then returned to the chalcopyrite flotation step after being treated to reduce the water consumption of the entire process. The underflow tailings are sent to a filter press for dewatering, and the resulting tailings filter cakes are then stored in a tailings dam.

[0184] Comparative Example 2 A flotation method for copper-lead-zinc-pyrite without online pH-Eh linkage is disclosed. The difference between this comparative example and Example 3 is that it lacks an online pH-Eh linkage feedback control system and instead employs only a traditional potential control method. The method specifically includes the following steps: In this comparative example, the composition and content of the copper-lead-zinc pyrite ore include: 0.58% copper (Cu), 1.03% lead (Pb), 1.88% zinc (Zn), 8.75% iron (Fe), 2.41% sulfur (S), 1.06% calcium oxide (CaO), 3.94% aluminum oxide (Al2O3), 68.86% silicon dioxide (SiO2), 90.65 g / t silver (Ag*), 0.79 g / t gold (Au*), and 11.02% magnesium oxide (MgO), with the remainder being other trace impurities and loss on ignition.

[0185] S1, Heavy Medium Swirl Pre-Classification: Take 1 kg of copper-lead-zinc pyrite ore and crush it to a particle size of -0.5 mm; then perform wet screening and classification, the screening and classification method is the same as described in Example 3. The heavy medium hydrocyclone underflow, which is the heavy product (i.e., the pyrite mineral enrichment product), enters the grinding-flotation loop after the medium is removed by the demedium screen.

[0186] S2. Preparation of slurry: The pyrite-enriched product was ground until the proportion of copper-lead-zinc-pyrite ore at -0.074 mm reached 75%, resulting in a pyrite-enriched slurry. The grinding method was ball milling, with a grinding time of 12 minutes. Water was added to adjust the concentration of copper-lead-zinc-pyrite ore powder in the pyrite-enriched slurry to 33%. Then, lime slurry was added to adjust the slurry pH to 9.5, and an appropriate amount of Na₂S was added to finely adjust the slurry potential to -110 mV. The conditioning time was 5 minutes.

[0187] S3, chalcopyrite flotation: S31, chalcopyrite roughing: The pyrite-enriched slurry was transferred to the roughing cell of the copper flotation section. 35 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 35 g collector per ton of pyrite-enriched slurry) was added; 25 g / t of frother MIBC (i.e., 25 g frother per ton of pyrite-enriched slurry) was added; and a lead-zinc depressant, a combination of zinc sulfate (ZnSO4) and sodium sulfite (Na2SO3), was added. The amount of ZnSO4 added was 500 g / t (i.e., 500 g ZnSO4 per ton of pyrite-enriched slurry), and the amount of Na2SO3 added was 200 g / t (i.e., 200 g Na2SO3 per ton of pyrite-enriched slurry). One roughing of chalcopyrite was performed, with a frothing time of 12 minutes. Copper roughing concentrate slurry and copper roughing tailings slurry were obtained.

[0188] S32, Chalcopyrite Refinement: Two chalcopyrite beneficiation processes were performed using the same blank, reagent-free beneficiation method as in Example 3 to obtain chalcopyrite concentrate.

[0189] S33, scavenging of chalcopyrite: The copper roughing tailings slurry undergoes two scavenging processes: First, 10 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 10 g of collector per ton of copper roughing tailings slurry) and 10 g / t of frother MIBC (i.e., 10 g of frother per ton of copper roughing tailings slurry) are added to the slurry for the first chalcopyrite scavenging, yielding the first copper scavenged concentrate slurry and the first copper scavenged tailings slurry; then... In the first copper scavenging tailings slurry, 5 g / t of collector 2-(benzothiazolylthio)acetic acid (i.e., 5 g of collector per ton of the first copper scavenging tailings slurry) and 10 g / t of frother MIBC (i.e., 10 g of frother per ton of the first copper scavenging tailings slurry) were added for the second chalcopyrite scavenging, yielding a second copper scavenging concentrate slurry and a second copper scavenging tailings slurry; the frothing time for both scavenging processes was 6 minutes. The first copper scavenging concentrate slurry was returned to the pyrite-enriched slurry, and the second copper scavenging concentrate slurry was returned to the copper roughing tailings slurry; the second copper scavenging tailings slurry proceeded to the next step of galena flotation.

[0190] S4, Galena flotation: S41, Galena roughing: Before the second copper scavenging tailings slurry enters the galena roughing process, pH gradient switching and slurry conditions are adjusted. Lime slurry is added to adjust the pH of the slurry to 10.5±0.2; Na2S is added to adjust the Eh of the second copper scavenging tailings slurry to -60mV. Next, a collector, a combination of MBT and PEG, was added to the second copper scavenging tailings slurry after Eh adjustment. The MBT addition was 50 g / t (i.e., 50 g MBT per ton of second copper scavenging tailings slurry), and the PEG addition was 25 g / t (i.e., 25 g PEG per ton of second copper scavenging tailings slurry). 40 g / t of lead activator Na2S was added (i.e., 40 g lead activator per ton of second copper scavenging tailings slurry). A zinc inhibitor, a combination of ZnSO4 and Na2SO3, was added. The ZnSO4 addition was 500 g / t (i.e., 500 g ZnSO4 per ton of second copper scavenging tailings slurry), and the Na2SO3 addition was 200 g / t (i.e., 200 g Na2SO3 per ton of second copper scavenging tailings slurry). Na2SO3); add 30g / t of frother MIBC (i.e., add 30g of frother to each ton of the second copper scavenging tailings slurry); carry out the first roughing of galena, with a frothing time of 12min; and obtain lead roughing concentrate slurry and lead roughing tailings slurry.

[0191] S42, Galena concentrate: The lead rougher concentrate slurry was transferred to the lead flotation section for cleaning. The same blank, reagent-free cleaning method as in Example 3 was used to perform two galena cleaning processes to obtain galena concentrate.

[0192] S43, Galena scavenging: The lead roughing tailings slurry undergoes two scavenging processes: First, a combination of MBT and PEG is added as a collector to the lead roughing tailings slurry, with MBT added at a rate of 8 g / t (i.e., 8 g MBT per ton of lead roughing tailings slurry) and PEG added at a rate of 4 g / t (i.e., 4 g PEG per ton of lead roughing tailings slurry). Then, 10 g / t of MIBC (i.e., 10 g frother per ton of lead roughing tailings slurry) is added for the first galena scavenging, yielding the first lead scavenged concentrate slurry and the first lead scavenged tailings slurry. Next, 8 g / t MBT, 4 g / t PEG, and 10 g / t MIBC are added to the first lead scavenged tailings slurry, i.e., 8 g MBT, 4 g PEG, and 10 g MIBC (i.e., 8 g MBT, 4 g PEG, and 10 g PEG per ton of first lead scavenged tailings slurry) are added to the first lead scavenged tailings slurry. PEG and 10g of frother were added to perform a second galena scavenging process, yielding a second lead scavenging concentrate slurry and a second lead scavenging tailings slurry. The frothing time for both scavenging processes was 6 minutes. The first lead scavenging concentrate slurry was returned to the second copper scavenging tailings slurry, and the second lead scavenging concentrate slurry was returned to the lead rougher tailings slurry. The second lead scavenging tailings slurry was then used in the next step, sphalerite flotation.

[0193] S5, Sphalerite flotation: S51, Sphalerite roughing: Before the second lead scavenging tailings slurry enters the sphalerite roughing process, 350g / t of activator CuSO4 is added to the second lead scavenging tailings slurry (i.e., 350g of activator is added per ton of second lead scavenging tailings slurry); lime slurry is added to adjust the slurry pH to 11.5±0.2, and Na2S is added to adjust the Eh of the second lead scavenging tailings slurry to -10mV.

[0194] Next, 60 g / t butyl xanthate and 25 g / t MIBC were added to the activated and conditioned second lead scavenging tailings slurry; a first roughing of sphalerite was carried out, with a frothing time of 12 min; zinc roughing concentrate slurry and zinc roughing tailings slurry were obtained.

[0195] S52, Sphalerite Refined: The zinc roughing concentrate slurry was transferred to the zinc flotation section for cleaning, and two rounds of sphalerite cleaning were performed using the same blank, reagent-free cleaning method as in Example 3 to obtain sphalerite concentrate.

[0196] S53, Sphalerite scavenging: The zinc roughing tailings slurry undergoes two scavenging processes: 20 g / t butyl xanthate is added as a collector (i.e., 20 g of collector per ton of zinc roughing tailings slurry); 10 g / t frother MIBC is added for the first sphalerite scavenging, yielding the first zinc scavenged concentrate slurry and the first zinc scavenged tailings slurry; then, 15 g / t butyl xanthate and 10 g / t MIBC are added to the first zinc scavenged tailings slurry for the second sphalerite scavenging, yielding the second zinc scavenged concentrate slurry and the second zinc scavenged tailings slurry; the frothing time for both scavenging processes is 6 minutes. The first zinc scavenged concentrate slurry is returned to the second lead scavenged tailings slurry, and the second zinc scavenged concentrate slurry is returned to the zinc roughing tailings slurry; the second zinc scavenged tailings slurry proceeds to the next step, sulfur concentrate flotation.

[0197] S6, sulfur concentrate flotation: First, the tailings slurry from the second zinc scavenging process is activated by adding 200g / t of activator CuSO4 (i.e., 200g of activator is added to each ton of tailings slurry from the second zinc scavenging process); H2SO4 is added to adjust the pH of the slurry to 6.5±0.2.

[0198] Next, roughing of the sulfur concentrate is carried out. 120 g / t butyl xanthate is added as a collector and 30 g / t frother MIBC is added to the activated and conditioned tailings slurry from the second zinc flotation. A first roughing of the sulfur concentrate is then performed, with a frothing time of 12 minutes. This yields a sulfur roughing concentrate slurry and a sulfur flotation tailings slurry. The sulfur roughing concentrate slurry is then dried to obtain the sulfur concentrate.

[0199] S7. Tailings Treatment: Add 50g / t of flocculant PAM to the sulfur flotation tailings slurry, and concentrate it for 10 hours using a high-efficiency thickener. The thickened overflow is then returned to the chalcopyrite flotation step after being treated to reduce the water consumption of the entire process. The underflow tailings are sent to a filter press for dewatering, and the resulting tailings filter cakes are then stored in a tailings dam.

[0200] Comparative Example 3 A flotation method for copper-lead-zinc-pyrite ore based on traditional collectors is presented in this comparative example. The traditional collectors are used in each flotation stage, and no potential control or heavy medium pre-classification is employed. The specific steps are as follows: In this embodiment, the composition and content of the copper-lead-zinc-pyrite ore include: 0.56% copper (Cu), 1.06% lead (Pb), 1.85% zinc (Zn), 9.12% iron (Fe), 2.41% sulfur (S), 0.74% calcium oxide (CaO), 4.21% aluminum oxide (Al2O3), 72.13% silicon dioxide (SiO2), 93.32 g / t silver (Ag*), 0.83 g / t gold (Au*), and 7.45% magnesium oxide (MgO), with the remainder being other trace impurities and loss on ignition.

[0201] S1. Preparation of slurry: Take 1 kg of copper-lead-zinc-pyrite ore and crush it to a particle size of -0.5 mm. Then, perform wet ball milling for 12 minutes until the proportion of copper-lead-zinc-pyrite ore with a particle size of -0.074 mm reaches 75%, thus obtaining a pyrite mineral slurry. Add water to adjust the concentration of copper-lead-zinc-pyrite ore powder in the pyrite mineral slurry to 33%. Then, add lime milk to adjust the pH value of the slurry to 9.5, and adjust the slurry time to 5 minutes.

[0202] S2, chalcopyrite flotation: S21, chalcopyrite roughing: The pyrite slurry was transferred to the roughing cell of the copper flotation section. 35 g / t xanthate was added to the pyrite slurry as a collector (i.e., 35 g collector per ton of pyrite slurry); 25 g / t 2 # Oil (pine oil) was used as a foaming agent (i.e., 25g of foaming agent was added per ton of pyrite slurry); lead-zinc inhibitors were added, which were a combination of zinc sulfate (ZnSO4) and sodium sulfite (Na2SO3), wherein the amount of ZnSO4 added was 500g / t and the amount of Na2SO3 added was 200g / t, that is, 500g of ZnSO4 and 200g of Na2SO3 were added per ton of pyrite slurry for one roughing of chalcopyrite, and the foaming time of the roughing was 12min; copper roughing concentrate slurry and copper roughing tailings slurry were obtained.

[0203] S22, Chalcopyrite Refinement: Two chalcopyrite beneficiation processes were performed using the same blank, reagent-free beneficiation method as in Example 3 to obtain chalcopyrite concentrate.

[0204] S23, scavenging of chalcopyrite: The copper roughing tailings slurry undergoes two scavenging processes: specifically, 10 g / t xanthate collector (i.e., 10 g collector per ton of copper roughing tailings slurry) and 10 g / t 2... # Oil was used as a frother (i.e., 10g of frother was added to each ton of copper roughing tailings slurry) for the first chalcopyrite scavenging, yielding the first copper scavenging concentrate slurry and the first copper scavenging tailings slurry; then, 5g / t xanthate and 10g / t 2... were added to the first copper scavenging tailings slurry. # Oil (i.e., 5g of butyl xanthate and 10g of 2 g of butyl xanthate are added to each ton of the first copper scavenging tailings slurry) # The first copper scavenging concentrate was returned to the pyrite-enriched slurry, and the second copper scavenging tailings slurry was returned to the copper roughing tailings slurry. The skimming time for both scavenging processes was 6 minutes. The first copper scavenging concentrate slurry was returned to the pyrite-enriched slurry, and the second copper scavenging concentrate slurry was returned to the copper roughing tailings slurry. The second copper scavenging tailings slurry was then used in the next step of galena flotation.

[0205] S3, Galena flotation: S31, Galena roughing: Before the second copper scavenging tailings slurry enters the galena roughing process, pH gradient switching and slurry conditions are adjusted by adding lime slurry to adjust the slurry pH to 10.5. Next, a collector, a combination of ethyl thiocyanate and butyl ammonium black powder, is added to the adjusted second copper scavenging tailings slurry. The ethyl thiocyanate is added at a rate of 50 g / t (i.e., 50 g of ethyl thiocyanate per ton of second copper scavenging tailings slurry), and the butyl ammonium black powder is added at a rate of 25 g / t (i.e., 25 g of butyl ammonium black powder per ton of second copper scavenging tailings slurry). 40 g / t of lead activator Na2S is added (i.e., 40 g of lead activator per ton of second copper scavenging tailings slurry). A zinc inhibitor, a combination of ZnSO4 and Na2SO3, is added. The ZnSO4 is added at a rate of 500 g / t (i.e., 500 g of ZnSO4 per ton of second copper scavenging tailings slurry), and the Na2SO3 is added at a rate of 200 g / t (i.e., 200 g of Na2SO3 per ton of second copper scavenging tailings slurry). Finally, 30 g / t of frother is added. # Add 30g of frother to each ton of the second copper scavenging tailings slurry; perform a first roughing of lead ore, with a frothing time of 12 minutes; to obtain lead roughing concentrate slurry and lead roughing tailings slurry.

[0206] S32, Galena concentrate: The lead rougher concentrate slurry was transferred to the lead flotation section for cleaning. The same blank, reagent-free cleaning method as in Example 3 was used to perform two galena cleaning processes to obtain galena concentrate.

[0207] S33, Galena scavenging: The lead roughing tailings slurry was subjected to two scavenging processes: a combination of ethyl thiocyanate and butyl ammonium black powder was added as a collector to the lead roughing tailings slurry, with ethyl thiocyanate added at a rate of 8 g / t and butyl ammonium black powder added at a rate of 4 g / t; 10 g / t of 2... # Oil, specifically, is produced by adding 8g of ethyl thiocyanate, 4g of butanone black powder, and 10g of 2 to each ton of lead roughing tailings slurry. # The oil was subjected to the first galena scavenging, yielding the first lead scavenging concentrate slurry and the first lead scavenging tailings slurry. Then, 8 g / t ethyl thiocyanate, 4 g / t butanone black powder, and 10 g / t 2... # Oil, specifically, is produced by adding 8g of ethyl thiocyanate, 4g of butyl ammonium black powder, and 10g of 2 to each ton of the first lead scavenging tailings slurry. # The oil undergoes a second galena scavenging process to obtain a second lead scavenging concentrate slurry and a second lead scavenging tailings slurry; the skimming time for both scavenging processes is 6 minutes. The first lead scavenging concentrate slurry is returned to the second copper scavenging tailings slurry, and the second lead scavenging concentrate slurry is returned to the lead rougher tailings slurry; the second lead scavenging tailings slurry proceeds to the next step, sphalerite flotation.

[0208] S4, Sphalerite flotation: S41, Sphalerite roughing: Before the second lead scavenging tailings slurry enters the sphalerite roughing process, 350g / t of activator CuSO4 is added to the second lead scavenging tailings slurry (i.e., 350g of activator is added per ton of the second lead scavenging tailings slurry); lime slurry is added to adjust the slurry pH to 11.5.

[0209] Next, 60 g / t butyl xanthate and 25 g / t 2 g / t xanthate were added to the activated and conditioned second lead scavenging tailings slurry. # Oil; perform one roughing of sphalerite, with a frothing time of 12 minutes; to obtain zinc roughing concentrate slurry and zinc roughing tailings slurry.

[0210] S42, Sphalerite Refined: The zinc roughing concentrate slurry was transferred to the zinc flotation section for cleaning, and two rounds of sphalerite cleaning were performed using the same blank, reagent-free cleaning method as in Example 3 to obtain sphalerite concentrate.

[0211] S43, Sphalerite scavenging: The zinc roughing tailings slurry is subjected to two scavenging processes, specifically: 20 g / t xanthate is added to the zinc roughing tailings slurry as a collector (i.e., 20 g of collector is added per ton of zinc roughing tailings slurry); 10 g / t frother is added. # The oil was subjected to the first sphalerite scavenging to obtain the first zinc scavenging concentrate slurry and the first zinc scavenging tailings slurry; then, 15 g / t butyl xanthate and 10 g / t 2 g / t scavenging agent were added to the first zinc scavenging tailings slurry. # The oil is subjected to a second sphalerite scavenging process to obtain a second zinc scavenging concentrate slurry and a second zinc scavenging tailings slurry; the skimming time for both scavenging processes is 6 minutes. The first zinc scavenging concentrate slurry is returned to the second lead scavenging tailings slurry, and the second zinc scavenging concentrate slurry is returned to the zinc rougher tailings slurry; the second zinc scavenging tailings slurry proceeds to the next step of sulfur concentrate flotation.

[0212] S5, sulfur concentrate flotation: First, the tailings slurry from the second zinc scavenging process is activated by adding 200g / t of activator CuSO4 (i.e., 200g of activator is added to each ton of tailings slurry from the second zinc scavenging process); H2SO4 is added to adjust the pH of the slurry to 6.5.

[0213] Next, roughing of the sulfur concentrate is carried out. 120 g / t butyl xanthate is added as a collector to the activated and conditioned second zinc scavenging tailings slurry, along with 30 g / t frother MIBC (i.e., 120 g butyl xanthate and 30 g frother are added per ton of second zinc scavenging tailings slurry). This roughing process is repeated for 12 minutes, resulting in a sulfur roughing concentrate slurry and a sulfur flotation tailings slurry. The sulfur roughing concentrate slurry is then dried to obtain the sulfur concentrate.

[0214] S6. Tailings treatment: Add 50g / t of flocculant PAM to the sulfur flotation tailings slurry, and concentrate it for 10 hours using a high-efficiency thickener. The thickened overflow is then returned to the chalcopyrite flotation step after being treated to reduce the water consumption of the entire process. The underflow tailings are sent to a filter press for dewatering, and the resulting tailings filter cakes are then stored in a tailings dam.

[0215] Next, the testing and analysis were conducted, and the specific details are as follows: The grades of Cu, Pb, Zn, and S in the copper-lead-zinc-pyrite ore powders used in Examples 1, 2, and 3, as well as Comparative Examples 1, 2, and 3, and the chalcopyrite concentrate, galena concentrate, sphalerite concentrate, sulfur concentrate, and tailings filter cake (tailings filter cake dried to powder) obtained by flotation were determined. The yields of chalcopyrite concentrate, galena concentrate, sphalerite concentrate, and sulfur concentrate, as well as the recoveries of Cu, Pb, Zn, and S were calculated respectively.

[0216] The grades of Cu, Pb, Zn, and S were determined by atomic absorption spectrophotometry (AAS), specifically: The grade of Cu was determined by atomic absorption spectrophotometry (AAS) in GB / T 14353.1-2010 for copper; the grade of Pb was determined by atomic absorption spectrophotometry (AAS) in GB / T 8151.5-2012 for lead; the grade of Zn was determined by atomic absorption spectrophotometry (AAS) in GB / T 8152.4-2006 for zinc; and the grade of S was determined by GB / T 2466 (Determination of sulfur content in pyrite and sulfur concentrate).

[0217] The formula for calculating the yield of chalcopyrite concentrate is: Chalcopyrite concentrate yield = (mass of chalcopyrite concentrate / mass of copper-lead-zinc-pyrite ore) × 100%; The formula for calculating the yield of galena concentrate is: Galena concentrate yield = (mass of galena concentrate / mass of copper-lead-zinc-pyrite ore) × 100%; The formula for calculating the yield of sphalerite concentrate is: Sphalerite concentrate yield = (mass of sphalerite concentrate / mass of copper-lead-zinc-pyrite ore) × 100%; The formula for calculating the yield of sulfur concentrate is: Sulfur concentrate yield = (mass of sulfur concentrate / mass of copper-lead-zinc-pyrite ore) × 100%; The formula for calculating Cu recovery rate is: Cu recovery rate = [(yield of chalcopyrite concentrate × grade of Cu in chalcopyrite concentrate) / grade of Cu in copper-lead-zinc-pyrite ore] × 100%; The formula for calculating the Pb recovery rate is: Pb recovery rate = [(yield of galena concentrate × grade of Pb in galena concentrate) / grade of Pb in copper-lead-zinc-pyrite ore] × 100%; The formula for calculating the Zn recovery rate is: Zn recovery rate = [(yield of sphalerite concentrate × grade of Zn in sphalerite concentrate) / grade of Zn in copper-lead-zinc-pyrite ore] × 100%.

[0218] The formula for calculating the recovery rate of sulfur is: Recovery rate of sulfur = [(Yield of sulfur concentrate × Grade of sulfur in sulfur concentrate) / Grade of sulfur in copper-lead-zinc-pyrite ore] × 100%.

[0219] The results are shown in Tables 1 through 6 below: Table 1. Statistical table of Cu, Pb, Zn, and S grades and recoveries in raw ore and various concentrates in Example 1.

[0220] Table 2. Statistical table of Cu, Pb, Zn, and S grades and recoveries in raw ore and various concentrates in Example 2.

[0221] Table 3. Statistical table of Cu, Pb, Zn, and S grades and recoveries in raw ore and various concentrates in Example 3.

[0222] Table 4. Statistical table of Cu, Pb, Zn, and S grades and recoveries in the raw ore and concentrates of Comparative Example 1.

[0223] Table 5. Statistics on Cu, Pb, Zn, and S grades and recoveries in raw ore and concentrates of Comparative Example 2.

[0224] Table 6. Statistics on Cu, Pb, Zn, and S grades and recoveries in raw ore and concentrates of Comparative Example 3.

[0225] Based on Tables 1-6, the results of Cu, Pb, Zn, and S grades and recoveries in Examples 1-3 and Comparative Examples 1-3 show that the pretreatment-potential-pH three-stage synergistic method is significantly superior to the traditional process. The recoveries of Cu, Pb, Zn, and S in the three examples are all close to or exceed 90%, achieving efficient separation and enrichment of the four minerals (copper, lead, zinc, and sulfur) with low product contamination. Example 3 is the optimal implementation scheme, with a Cu grade of 26.04% and a recovery rate of 90.02% in chalcopyrite concentrate, a Pb grade of 68.29% and a recovery rate of 91.72% in galena concentrate, a Zn grade of 55.77% and a recovery rate of 93.6% in sphalerite concentrate, and a S grade of 40.28% and a recovery rate of 90.45% in sulfur concentrate. All core indicators are at the highest levels.

[0226] Comparative Example 1 eliminated the heavy media cyclone pre-classification step. This change led to a significant decline in all indicators. The above data fully demonstrates that heavy media cyclone pre-classification can remove some of the light product tailings in advance, enriching pyrite minerals in the material entering the grinding-flotation loop. This is an important prerequisite for ensuring high-selectivity flotation; the absence of this step will lead to a simultaneous decrease in grade and recovery rate.

[0227] Comparative Example 2 did not employ an online pH-Eh feedback control system, relying solely on traditional potential regulation methods (using lime slurry to adjust pH, and H2O2 and Na2S to regulate slurry Eh). All indicators in this control group were lower than those in the example. The core issue was a significant increase in the content of dissimilar metals in each concentrate: chalcopyrite concentrate contained 4.15% Pb and 2.03% Zn, while galena concentrate contained as much as 8.01% Zn, resulting in poor product mixing and significantly insufficient selectivity. This demonstrates that the online pH-Eh feedback control system can accurately regulate slurry potential in real time, preventing the flotation of non-target minerals due to potential drift, while traditional potential regulation methods struggle to maintain a stable potential window, leading to a significant decrease in selectivity.

[0228] Comparative Example 3 employed a traditional flotation method without heavy media pre-classification, potential control, or using a traditional combination of butyl xanthate and ethyl thiocyanate / butyl ammonium black reagent. All indicators were unsatisfactory, with the most severe product contamination problem. The Cu recovery rate in its chalcopyrite concentrate was only 62.54%, meaning a significant amount of copper was lost to other products. The Pb recovery rate in galena concentrate was 74.15%, and the Zn recovery rate in sphalerite concentrate was 83.52%, both also showing substantial unrecovered metals. This clearly demonstrates that traditional collectors (butyl xanthate, ethyl thiocyanate, etc.) have poor selectivity for copper, lead, zinc, and pyrite. Without potential control and pre-classification, the flotation separation effect between minerals is extremely poor, resulting in severe product contamination.

[0229] The above descriptions are merely embodiments of the present invention, and common knowledge such as specific technical solutions and / or characteristics are not described in detail here. It should be noted that those skilled in the art can make various modifications and improvements without departing from the technical solutions of the present invention, and these should also be considered within the scope of protection of the present invention. These modifications and improvements will not affect the effectiveness of the implementation of the present invention or the practicality of the patent. The scope of protection claimed in this application should be determined by the content of its claims, and the specific embodiments described in the specification can be used to interpret the content of the claims.

Claims

1. A pretreatment-potential-pH three-stage synergistic flotation method for copper-lead-zinc-pyrite ore, characterized in that, Includes the following steps: S1. Heavy medium cyclone pre-classification treatment is carried out on copper-lead-zinc-pyrite to obtain pyrite mineral-enriched slurry; S2. Adjust the pH and redox potential of the pyrite-rich slurry, add a potential-responsive collector, and perform chalcopyrite flotation to obtain copper scavenging tailings slurry and chalcopyrite concentrate slurry. S3. Increase the pH value of the copper scavenging tailings slurry and decrease the redox potential to carry out galena flotation to obtain lead scavenging tailings slurry and galena concentrate slurry. S4. Increase the pH value and redox potential of the lead scavenging tailings slurry to carry out sphalerite flotation to obtain zinc scavenging tailings slurry and sphalerite concentrate slurry. S5. Lower the pH value of the zinc scavenging tailings slurry and carry out sulfur concentrate flotation to obtain sulfur flotation tailings slurry and sulfur concentrate slurry.

2. The flotation method for copper-lead-zinc-pyrite according to claim 1, characterized in that, S1 includes the following steps: The raw copper-lead-zinc-pyrite ore is crushed to a particle size of less than or equal to 0.5 mm, and then wet-screened and classified to obtain a screened slurry. A ferrosilicon heavy medium suspension is added to the screened slurry, and the slurry is subjected to heavy medium cyclone treatment to obtain a heavy product. The heavy product is ground until the proportion of copper-lead-zinc-pyrite powder with a particle size of -0.074 mm in the heavy product is 65-75%. The pH value of the heavy product is adjusted to 9.3-9.7, and then water is added to obtain a pyrite-rich slurry.

3. The flotation method for copper-lead-zinc-pyrite according to claim 2, characterized in that, S2 includes the following steps: The pH of the pyrite-rich slurry was adjusted to be between 9.3 and 9.7, and the redox potential was adjusted to be between -110 and -90 mV. A potential-responsive collector, a foaming agent, and a lead-zinc inhibitor are added to the pyrite-rich slurry, and chalcopyrite roughing is carried out to obtain copper roughing concentrate slurry and copper roughing tailings slurry. The copper roughing concentrate slurry is subjected to chalcopyrite beneficiation to obtain chalcopyrite concentrate slurry; The copper roughing tailings slurry is subjected to chalcopyrite scavenging to obtain copper scavenged tailings slurry.

4. The flotation method for copper-lead-zinc-pyrite according to claim 3, characterized in that, S3 includes the following steps: The pH value of the copper scavenging tailings slurry is adjusted to be between 10.3 and 10.7, and the redox potential is between -60 and -40 mV. Then, galena roughing collector, frother, lead activator and zinc inhibitor are added to the copper scavenging tailings slurry to carry out galena roughing to obtain lead roughing concentrate slurry and lead roughing tailings slurry. The lead roughing concentrate slurry is subjected to galena beneficiation to obtain galena concentrate slurry; The lead roughing tailings slurry is subjected to galena scavenging to obtain lead scavenged tailings slurry.

5. The flotation method for copper-lead-zinc-pyrite according to claim 4, characterized in that, S4 includes the following steps: Add sphalerite activator to the lead scavenging tailings slurry to obtain activated lead scavenging tailings slurry; The pH value of the activated lead scavenging tailings slurry was adjusted to be between 11.3 and 11.7, and the redox potential was between -10 and 10mV. Then, sphalerite roughing collector and frother were added to carry out sphalerite roughing to obtain zinc roughing concentrate slurry and zinc roughing tailings slurry. The zinc roughing concentrate slurry is subjected to sphalerite beneficiation to obtain sphalerite concentrate slurry; The zinc roughing tailings slurry is subjected to sphalerite scavenging to obtain zinc scavenged tailings slurry.

6. The flotation method for copper-lead-zinc-pyrite according to claim 5, characterized in that, S5 includes the following steps: Sulfur concentrate activator is added to zinc scavenging tailings slurry to obtain activated zinc scavenging tailings slurry; The pH value of the activated zinc scavenging tailings slurry was adjusted to between 6.3 and 6.

7. Then, a sulfur concentrate roughing collector and a frother were added to carry out sulfur concentrate roughing to obtain sulfur flotation tailings slurry and sulfur concentrate slurry.

7. The flotation method for copper-lead-zinc-pyrite according to claim 6, characterized in that, The potential-responsive collector is selected from 2-(benzothiazolylthio)acetic acid; The foaming agent is selected from methyl isobutyl methanol; The lead-zinc inhibitor is selected from a combination of zinc sulfate and sodium sulfite, and the mass ratio of zinc sulfate to sodium sulfite is (2~2.5):1; The lead activator is sodium sulfide; The zinc inhibitor is selected from a combination of zinc sulfate and sodium sulfite, and the mass ratio of zinc sulfate to sodium sulfite is 2 to (2.5:1).

8. The flotation method for copper-lead-zinc-pyrite according to claim 1, characterized in that, In the chalcopyrite flotation process, a mixed aqueous solution of Na2CO3 and NaHCO3 is used to adjust the pH value of the pyrite-enriched slurry, and an oxidizing agent or reducing agent is used to adjust the Eh value of the pyrite-enriched slurry. In the galena flotation process, a mixed aqueous solution of Na2CO3 and NaHCO3 is used to adjust the pH value of the copper scavenging tailings slurry, and an oxidizing agent or reducing agent is used to adjust the Eh value of the copper scavenging tailings slurry. In the sphalerite flotation process, lime slurry is used to adjust the pH value of the lead scavenging tailings slurry, and an oxidizing agent or reducing agent is used to adjust the Eh value of the lead scavenging tailings slurry. In the sulfur concentrate flotation process, sulfuric acid is used to adjust the pH value of the zinc scavenging tailings slurry, and an oxidizing agent or reducing agent is used to adjust the Eh value of the zinc scavenging tailings slurry.

9. The flotation method for copper-lead-zinc-pyrite according to claim 8, characterized in that, It also includes a pH-Eh online linkage feedback control flotation system, which synchronously feeds forward to compensate the Eh control output when adjusting the pH, thereby suppressing pH-Eh oscillations throughout the flotation process.

10. The flotation method for copper-lead-zinc-pyrite according to claim 1, characterized in that, It also includes methods for tailings treatment, comprising the following steps: Add flocculant to sulfur flotation tailings slurry, and after concentration treatment for 8-10 hours, obtain thick overflow and underflow tailings. The bottom tailings are dewatered to obtain tailings filter cake, which is then stored in a tailings dam. The concentrated overflow is reused.

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