Flotation Separation Methods for High-Sulfur Copper-Zinc Ores

Abstract

This invention discloses a flotation separation method for high-sulfur copper-zinc ore beneficiation, comprising the following steps: Step a. Classifying and grinding the high-sulfur copper-zinc ore, controlling the grinding fineness to ensure that the product particle size P80 is 120μm to 150μm; Step b. Performing a first-stage copper-zinc mixed flotation operation on the obtained grinding product: adding lime during the flotation process at a rate of 100g / t to 200g / t to adjust the pH value of the pulp to the range of 7.5 to 8.5, forming surface statics between chalcopyrite, sphalerite, and pyrite. Potential difference; an aerated flotation machine is used as the flotation equipment, and the selective collector O-ethyl-N-benzoyl thiocarbamate is added at a dosage of 25g / t to 30g / t for aerated flotation to obtain a copper-zinc-sulfur mixed concentrate; step c. the obtained copper-zinc-sulfur mixed concentrate is fed into a vertical spiral stirred mill for regrinding, and the regrinding fineness is controlled to P90≤20μm to achieve full mineral liberation and surface desorption; this invention can achieve precise separation of copper, zinc and sulfur, reduce costs and improve efficiency and increase the comprehensive utilization rate of resources.

Description

Technical Field

[0001] This invention relates to the field of mineral processing methods, specifically to a flotation separation method for high-sulfur copper-zinc ore. Background Technology

[0002] In high-sulfur copper-zinc ores (containing more than 30% sulfur and more than 50% pyrite), the intergrowth relationships of chalcopyrite, sphalerite, and pyrite are complex, and their floatability is similar. Achieving efficient separation of copper, zinc, and sulfur has always been a challenge in the mineral processing field. Traditional processes mainly employ two technical routes: "copper-preferred flotation under high-alkali conditions" or "copper-zinc mixed flotation followed by separation." These methods continue the mindset of "heavy pressure and heavy pulling": to suppress large amounts of pyrite and sphalerite activated by copper ions, a large amount of lime (3-8 kg / t) is added during grinding and flotation to raise the pulp pH to above 11.8, while simultaneously using inhibitors such as zinc sulfate and high doses of collectors to "strongly pull" copper minerals. This method of separation, which relies on the accumulation of chemical reagents, not only leads to high reagent costs, but also results in a lengthy process and a large equipment configuration (usually 3 to 5 times that of a low-alkalinity process). Furthermore, the high-alkalinity environment easily causes subsequent pipeline calcium buildup and difficulties in treating wastewater containing heavy metal ions (zinc ions), resulting in enormous environmental pressure.

[0003] To improve the drawbacks of the traditional "heavy pressure and heavy pulling" process, the industry has explored many approaches in recent years regarding low-toxicity reagents and process simplification. For example, CN120502431A discloses a copper sulfate-free flotation separation method for copper-zinc sulfide ores. This method uses ethyl thiocyanate as a collector and lime as a depressant for mixed copper-zinc flotation. Copper concentrate and zinc concentrate are then obtained through subsequent operations involving zinc sulfate suppression, ethyl thiocyanate collection, and lime assistance. Although this method avoids heavy metal pollution to some extent, it still uses the traditional lime suppression approach in the mixed copper-zinc flotation section. The ratio of ethyl thiocyanate to lime is difficult to optimize precisely, resulting in large fluctuations in copper-zinc separation performance and making it difficult to control beneficiation costs. For example, CN109158214A discloses a flotation separation process for copper-zinc sulfide ores. By using a combination of zinc sulfate and sodium sulfite for inhibition and a combination of collectors such as ethyl thiocyanate for collection, the separation of chalcopyrite and sphalerite is achieved. The process is simple and the reagent cost is low. However, this method still relies on the combined use of multiple inhibitors and collectors. The dosage and ratio of these agents need to be frequently adjusted according to the ore properties, making it difficult to achieve a stable and precise "compression-pull" balance. Especially when processing high-sulfur copper-zinc ores, the selective inhibition effect on pyrite is limited, the quality of the copper-zinc mixed concentrate is poor, and the subsequent separation load is still heavy.

[0004] Overall, existing improvement technologies mostly focus on replacing and optimizing reagent combinations, still failing to break free from the traditional "heavy pressure and heavy pulling" process framework, and lacking proactive control and precise utilization of the slurry electrochemical environment. In the flotation separation of high-sulfur copper-zinc ores, existing methods generally suffer from the following common problems: First, they fail to achieve selective flotation at low alkalinity by constructing surface electrostatic potential differences between chalcopyrite, sphalerite, and pyrite, resulting in the inability to completely eliminate high-dose depressants such as lime and zinc sulfate; second, the selection and parameter control of flotation equipment are relatively crude, lacking quantitative basis for key parameters such as aeration volume, turbulence intensity, and foam layer thickness, making it difficult to match the rapid flotation characteristics after mineral liberation, preventing the timely priority separation of high-grade copper minerals, resulting in a lengthy subsequent copper-zinc separation process and a single product structure; third, process optimization is mostly limited to technical indicators, failing to incorporate reagent costs and product pricing coefficients into comprehensive consideration, making it difficult to maximize overall economic benefits. Therefore, developing a low-alkalinity flotation separation method for high-sulfur copper-zinc ores based on electrochemical regulation, precise equipment matching, and economic efficiency has become a pressing technical challenge in this field. Summary of the Invention

[0005] To address the shortcomings of existing technologies, this invention provides a flotation separation method for high-sulfur copper-zinc ore, comprising the following steps: Step a. Classify and grind the high-sulfur copper-zinc ore to control the grinding fineness so that the product particle size P80 is 120μm to 150μm; Step b. Perform the first stage of copper-zinc mixed flotation on the grinding product obtained in step a: Add lime during the flotation process at a rate of 100 g / t to 200 g / t to adjust the pH of the pulp to the range of 7.5 to 8.5, forming a surface electrostatic potential difference between chalcopyrite, sphalerite, and pyrite; use a Smith / Otutec column-type aerated flotation machine as the flotation equipment, and add the selective collector O-ethyl-N-benzoyl thiocarbamate at a rate of 25 g / t to 30 g / t for aerated flotation to obtain a copper-zinc-sulfur mixed concentrate; Step c. Feed the copper-zinc-sulfur mixed concentrate obtained in step b into a vertical spiral stirred mill for regrinding. The regrinding fineness is controlled to P90≤20μm to achieve full mineral dissociation and surface desorption. Step d. Feed the regrinded product obtained in step c into a JAMESON flotation machine for rapid flotation: without adding any inhibitors, only add O-ethyl-N-benzoyl thiocarbamate as a collector at a dosage of 3 g / t to 5 g / t. Utilize the jet entrainment and microbubble precipitation mechanism of the JAMESON flotation machine to separate high-grade copper concentrate, which accounts for 60% to 75% of the total copper recovery of the raw ore. The high-grade copper concentrate contains Cu > 25% and Zn ≤ 1%. Step e. Feed the tailings from the rapid flotation in step d into the flotation column for the second stage of copper-zinc mixed flotation: Add lime at a rate of 100 g / t to 200 g / t and add O-ethyl-N-benzoyl thiocarbamate at a rate of 5 g / t to 10 g / t. By controlling the thickness of the froth layer and the aeration rate of the flotation column, copper and zinc are separated from sulfur. After multiple (at least two) cleaning processes, the flotation column concentrate is a copper-zinc mixed concentrate, and the flotation column tailings are a sulfur concentrate. The copper-zinc mixed concentrate contains 4% to 6% Cu and >45% Zn.

[0006] Furthermore, the graded grinding in step a is based on the mineral distribution characteristics of copper, zinc, and sulfur in the raw ore. The unit power consumption required to achieve the target particle size P80 is calculated using the Bond work index formula to optimize the grinding parameters. The calculation process for the unit power consumption W is as follows: ; Where W is the unit power consumption (kWh / t) required to achieve the target grinding fineness, Wi is the ore power index, P80,feed is the feed particle size, and P80,product is the target product particle size in step a.

[0007] Furthermore, the formation of the surface electrostatic potential difference between chalcopyrite, sphalerite and pyrite in step b is achieved by online monitoring of the pH value and redox potential (ORP) of the slurry, and feeding the monitoring data back to the lime addition system to dynamically adjust the amount of lime added so that the pH value and ORP value of the slurry meet the preset electrochemical window. The method for constructing the preset electrochemical window is as follows: based on the thermodynamic E-pH diagram, the surface electrostatic potentials of chalcopyrite, sphalerite, and pyrite under specific pH conditions are calculated using the Nernst equation. , and The specific calculation process is as follows: ; ; ; in, , , These are the standard electrode potentials for chalcopyrite, sphalerite, and pyrite, respectively. Based on the calculated surface electrostatic potentials of chalcopyrite, sphalerite, and pyrite under specific pH conditions, an electrochemical control model was established to ensure that the electrostatic potential of pyrite is lower than the pulp potential, while the electrostatic potentials of chalcopyrite and sphalerite are higher than the pulp potential. The determination criteria are as follows: ; ; And ensure that the measured potential E of the slurry meets the following requirements: .

[0008] Furthermore, in step b, the aeration rate of the aerated flotation machine is determined by acquiring bubble size distribution data on the foam surface through an online image analysis system, and then based on the bubble coalescence rate constant k. coal Dynamic adjustment is used to ensure the effective loading of the target mineral on the bubble surface and to avoid bubble coalescence leading to mineralization bubble instability; the bubble coalescence rate constant k coal The result is obtained by analyzing the sequence of foam images and calculating the rate of decrease in the number of bubbles per unit time.

[0009] Furthermore, the jet entrainment and microbubble precipitation mechanism of the JAMESON flotation machine in step d is achieved by controlling the jet Reynolds number; Based on the matching relationship between jet velocity and lower conduit inner diameter determined by computational fluid dynamics (CFD) simulation, a strong turbulent environment is formed in the mixing zone of the lower conduit by controlling the feed pressure, enabling the collector to achieve high-speed collision and adsorption on the mineral surface. The jet Reynolds number Re is defined as the ratio of the inertial force to the viscous force of the slurry flow in the lower conduit of the JAMESON flotation machine, and is a dimensionless number characterizing the intensity of turbulence. The specific calculation process is as follows: ; Where ρ is the slurry density, v is the jet velocity of the slurry through the lower conduit, D is the inner diameter of the lower conduit, and μ is the dynamic viscosity of the slurry. By controlling the feed pressure, the jet velocity v is made to satisfy the Reynolds number Re≥50000 in order to maintain sufficient turbulence intensity and achieve rapid flotation.

[0010] Furthermore, the high-grade copper concentrate, accounting for 60% to 75% of the total copper recovery from the raw ore, separated in step d, is achieved through the multi-stage series separation characteristics of the Jameson flotation machine. Specifically, a rapid flotation kinetic model based on mineral liberation and floatability is constructed within the Jameson flotation machine. The model determines the flotation rate constant k of copper minerals at different particle sizes through batch flotation rate experiments. Combined with the particle size distribution of the regrinding product, the theoretical recovery rate R within a short-time flotation period in the Jameson flotation machine is calculated. The calculation formula is as follows: ; Where mi is the mass fraction of the i-th particle size, ki is the flotation rate constant of the i-th particle size, and t is the flotation time. By adjusting the flotation time t, the calculated recovery rate R(t) is made to reach 60% to 75% of the total copper recovery rate, thereby locking in this part of the high floatability and high grade copper minerals.

[0011] Furthermore, in step e, controlling the thickness and aeration rate of the flotation column's foam layer is achieved by constructing an evaluation index for the stability of the foam layer, namely, the axial gas holding rate of the foam layer. In conjunction with the grade requirements of foam products, a relationship model between foam layer thickness H and concentrate grade β was established. By adjusting the flushing water volume and aeration volume, the cleaning effect of the foam layer was optimized, and the separation of copper, zinc and sulfur was enhanced. The calculation process of the axial gas holding rate of the foam layer is as follows: ; in, The gas holding rate at a distance h from the foam overflow weir. H represents the initial gas holding rate at the bottom of the foam. c The characteristic attenuation length is determined by controlling H. c Maintain the stability of the foam layer and the cleaning effect.

[0012] Furthermore, this method, through the combination of roughing-rapid flotation-column flotation under low alkalinity conditions, achieves a significant reduction in flotation reagent usage and optimization of product structure. The method further includes a step of optimizing and adjusting process parameters based on an economic benefit model: calculating the total reagent cost C. total and comprehensive economic benefits V total Furthermore, by optimizing the reagent dosage and flotation equipment parameters in steps b, d, and e, the total reagent cost C was reduced. total and comprehensive economic benefits V total The calculation process is as follows: ; ; Where Qi is the dosage of the i-th reagent, Pi is the unit price of the i-th reagent, Mj is the output of the j-th concentrate, Gj is the grade of the j-th concentrate, and Pricej is the price determined by the pricing coefficient corresponding to the grade.

[0013] The present invention has the following advantages over the prior art: 1. Breaking through the traditional framework of "heavy pressure and heavy pulling" to achieve selective separation at low alkalinity; This invention abandons the extensive method of traditional high-alkali processes that involve the heavy addition of lime (3-8 kg / t) and inhibitors such as zinc sulfate to "forcefully suppress" pyrite and sphalerite. By precisely controlling the pulp pH within a low-alkalinity range of 7.5-8.5, it fully utilizes the inherent surface electrostatic potential differences between chalcopyrite, sphalerite, and pyrite to construct an electrochemical window that favors the flotation of copper-zinc minerals while inhibiting pyrite. Combined with O-ethyl-N-benzoyl thiocarbamate collectors, which have excellent selectivity for copper-zinc minerals, low-reagent tailings removal of mixed copper-zinc-sulfur concentrates is achieved in an aerated flotation machine. Compared with traditional processes, the total reagent usage is only 1 / 5 to 1 / 8, and the use of heavy metal reagents such as zinc sulfate and copper sulfate is completely eliminated, thus eliminating the risk of heavy metal pollution at the source and achieving significant environmental benefits.

[0014] 2. Differentiated and rapid sorting, with streamlined and efficient equipment configuration; This invention utilizes a unique equipment combination of "aerated flotation machine roughing—vertical spiral stirred mill regrinding—JAMESON flotation machine rapid flotation—flotation column cleaning," fully leveraging the hydrodynamic advantages of various flotation equipment. With the core objective of amplifying the differences in mineral flotation rates, it achieves highly efficient separation of copper, zinc, and sulfur. Compared to similar mines, the amount of equipment used is reduced by approximately two-thirds, and the process flow is significantly simplified. In particular, the JAMESON flotation machine, by controlling the jet Reynolds number to generate strong turbulence and microbubble precipitation, combined with a rapid flotation kinetic model based on mineral liberation and floatability, preferentially separates 60%–75% of the total high-grade copper concentrate (Cu > 25%, Zn ≤ 1%) without adding any inhibitors. This enables the early recovery of high-value copper minerals and significantly reduces the workload of subsequent operations.

[0015] 3. Product structure optimization has significantly improved economic benefits; The product structure produced by this invention is more competitive in the market than that of traditional processes: on the one hand, it yields high-grade copper concentrate (Cu > 25%, Zn ≤ 1%), with a copper valuation coefficient far exceeding that of conventional copper concentrate (Cu 18%); on the other hand, it yields zinc-rich copper-zinc mixed concentrate (Cu 4%~6%, Zn > 45%), in which copper can still be valued, significantly enhancing the overall economic value. Simultaneously, this invention incorporates an economic benefit model into process parameter optimization, dynamically adjusting reagent dosages and equipment parameters for each operation based on total reagent costs and overall economic benefits. This ensures the process operates within the range of maximized economic benefits, achieving high-value utilization of mineral resources.

[0016] 4. Precise process control and model support enhance process stability and adaptability; This invention utilizes online monitoring of slurry pH and redox potential, combined with an electrochemical control model constructed using the Nernst equation, to dynamically adjust the amount of lime added, ensuring stable and controllable differences in electrostatic potential on the mineral surface. An image analysis system is used to acquire bubble size distribution, and the aeration rate is adjusted in real time based on the bubble coalescence rate constant to maintain the stability of mineralized bubbles. The flotation column establishes a correlation between froth layer thickness and concentrate grade using an axial gas holdup model, optimizing wash water and aeration rates to enhance the separation of copper, zinc, and sulfur. These quantitative control methods make the process more adaptable to fluctuations in different ore sources, and the separation accuracy and efficiency are significantly better than traditional empirical operations. Attached Figure Description

[0017] Figure 1 This is a flowchart of the present invention. Detailed Implementation

[0018] To further illustrate the technical means and effects of the present invention in achieving its intended purpose, the following detailed description of the specific implementation methods, structures, features, and effects of the present invention, in conjunction with the accompanying drawings and preferred embodiments, is provided below.

[0019] like Figure 1 As shown, the flotation separation method for high-sulfur copper-zinc ore includes the following steps: Step a. Classify and grind the high-sulfur copper-zinc ore to control the grinding fineness so that the product particle size P80 is 120μm to 150μm; Step b. Perform the first stage of copper-zinc mixed flotation on the grinding product obtained in step a: Add lime during the flotation process at a rate of 100 g / t to 200 g / t to adjust the pH of the pulp to the range of 7.5 to 8.5, forming a surface electrostatic potential difference between chalcopyrite, sphalerite, and pyrite; use an aerated flotation machine as the flotation equipment, and add the selective collector O-ethyl-N-benzoyl thiocarbamate at a rate of 25 g / t to 30 g / t to perform aerated flotation to obtain a copper-zinc-sulfur mixed concentrate; Step c. Feed the copper-zinc-sulfur mixed concentrate obtained in step b into a vertical spiral stirred mill for regrinding. The regrinding fineness is controlled to P90≤20μm to achieve full mineral dissociation and surface desorption. Step d. Feed the regrinded product obtained in step c into a JAMESON flotation machine for rapid flotation: without adding any inhibitors, only add O-ethyl-N-benzoyl thiocarbamate as a collector at a dosage of 3 g / t to 5 g / t. Utilize the jet entrainment and microbubble precipitation mechanism of the JAMESON flotation machine to separate high-grade copper concentrate, which accounts for 60% to 75% of the total copper recovery of the raw ore. The high-grade copper concentrate contains Cu > 25% and Zn ≤ 1%. Step e. Feed the tailings from the rapid flotation in step d into the flotation column for the second stage of copper-zinc mixed flotation: Add lime at a rate of 100 g / t to 200 g / t and add O-ethyl-N-benzoyl thiocarbamate at a rate of 5 g / t to 10 g / t. By controlling the thickness of the froth layer and the aeration rate of the flotation column, copper and zinc are separated from sulfur. After multiple cleaning processes, the flotation column concentrate is a copper-zinc mixed concentrate, and the flotation column tailings are a sulfur concentrate. The copper-zinc mixed concentrate contains 4% to 6% Cu and >45% Zn.

[0020] Step a, the graded grinding, is based on the mineral distribution characteristics of copper, zinc, and sulfur in the raw ore. The required unit power consumption to achieve the target particle size P80 is calculated using the Bond work index formula to optimize grinding parameters. The calculation process for the unit power consumption W is as follows: ; Where W is the unit power consumption (kWh / t) required to achieve the target grinding fineness, Wi P is the work index of ore. 80,feed For feed particle size, P 80,product The target product granularity for step a; Based on the mineral distribution characteristics of copper, zinc, and sulfur in the raw ore, the grinding parameters are optimized by calculating the unit power consumption required to achieve the target particle size using the Bond work index formula. This allows the grinding parameters to be set more closely to the actual mineral characteristics of the raw ore, avoiding blind grinding and accurately controlling the unit energy consumption and grinding fineness. This ensures that the particle size of the ground product meets the process requirements of subsequent flotation operations, laying a suitable raw material foundation for subsequent flotation separation. At the same time, the optimized grinding parameters can improve the efficiency of classification grinding, reduce energy consumption and resource waste caused by ineffective grinding, and balance the economy and process adaptability of grinding operations.

[0021] The formation of the surface electrostatic potential difference between chalcopyrite, sphalerite and pyrite in step b is achieved by online monitoring of the pH value and redox potential (ORP) of the slurry, and feeding the monitoring data back to the lime addition system to dynamically adjust the amount of lime added so that the pH value and ORP value of the slurry meet the preset electrochemical window. The method for constructing the preset electrochemical window is as follows: based on the thermodynamic E-pH diagram, the surface electrostatic potentials of chalcopyrite, sphalerite, and pyrite under specific pH conditions are calculated using the Nernst equation. , and The specific calculation process is as follows: ; ; ; in, , , These are the standard electrode potentials for chalcopyrite, sphalerite, and pyrite, respectively. Based on the calculated surface electrostatic potentials of chalcopyrite, sphalerite, and pyrite under specific pH conditions, an electrochemical control model was established to ensure that the electrostatic potential of pyrite is lower than the pulp potential, while the electrostatic potentials of chalcopyrite and sphalerite are higher than the pulp potential. The determination criteria are as follows: ; ; And ensure that the measured potential E of the slurry meets the following requirements: ; By monitoring the slurry pH and redox potential online and feeding the data back to the lime addition system, the amount of lime added can be dynamically adjusted. This allows for precise control of the slurry's electrochemical conditions, ensuring that the slurry pH and ORP values ​​stably match the preset electrochemical window. This guarantees the stable formation of the expected surface electrostatic potential difference between chalcopyrite, sphalerite, and pyrite, laying a crucial mineral surface electrochemical foundation for the effective separation of copper, zinc, and sulfur. Simultaneously, the electrostatic potential of each mineral surface is calculated using thermodynamic E-pH diagrams and the Nernst equation, and electrochemical control with clearly defined criteria is established. The model provides a scientific theoretical basis for the regulation of electrostatic potential on the mineral surface, replacing empirical manual regulation and significantly improving the accuracy, controllability, and stability of mineral surface regulation before flotation. In addition, by strictly controlling the electrostatic potential difference between copper-zinc minerals and pyrite, as well as the measured potential range of the pulp, this electrochemical control model can effectively ensure the floatability of chalcopyrite and sphalerite while inhibiting the floatability of pyrite. This achieves efficient preliminary separation of copper-zinc and sulfur, reduces the interference of sulfur minerals on subsequent copper-zinc mixed flotation, and improves the separation effect and concentrate quality of mixed flotation operations.

[0022] In step b, the aerated flotation machine obtains the bubble size distribution data on the foam surface through an online image analysis system, and then calculates the aeration rate based on the bubble coalescence rate constant k. coal Dynamic adjustment is used to ensure the effective loading of the target mineral on the bubble surface and to avoid bubble coalescence leading to mineralization bubble instability; the bubble coalescence rate constant k coal The result is obtained by analyzing a sequence of foam images and calculating the rate of decrease in the number of bubbles per unit time. By acquiring bubble size distribution data on the froth surface through an online image analysis system and dynamically adjusting the aeration volume of the aerated flotation machine in conjunction with the bubble coalescence rate constant, the aeration volume control is freed from empirical operation and achieves real-time and precise control. Simultaneously, it effectively suppresses bubble coalescence, prevents mineralized bubble instability, ensures effective loading of target minerals on the bubble surface, and improves the efficiency and stability of mineralization. Furthermore, a stable and suitable bubble state allows target minerals to combine with bubbles more efficiently, improving the separation effect of the first stage of copper-zinc mixed flotation, providing higher-quality copper-zinc-sulfur mixed concentrate for subsequent regrinding and flotation processes, and ensuring the smooth operation of subsequent stages of the overall flotation process.

[0023] The jet entrainment and microbubble precipitation mechanism of the JAMESON flotation machine in step d is achieved by controlling the jet Reynolds number; Based on the matching relationship between jet velocity and lower conduit inner diameter determined by computational fluid dynamics (CFD) simulation, a strong turbulent environment is formed in the mixing zone of the lower conduit by controlling the feed pressure, enabling the collector to achieve high-speed collision and adsorption on the mineral surface. The jet Reynolds number Re is defined as the ratio of the inertial force to the viscous force of the slurry flow in the lower conduit of the JAMESON flotation machine, and is a dimensionless number characterizing the intensity of turbulence. The specific calculation process is as follows: ; Where ρ is the slurry density, v is the jet velocity of the slurry through the lower conduit, D is the inner diameter of the lower conduit, and μ is the dynamic viscosity of the slurry. By controlling the feed pressure, the jet velocity v is made to satisfy the Reynolds number Re≥50000 in order to maintain sufficient turbulence intensity and achieve rapid flotation. By controlling the jet Reynolds number, the JAMESON flotation machine achieves efficient utilization of the jet entrainment and microbubble precipitation mechanisms. Furthermore, computational fluid dynamics simulations determine the matching relationship between jet velocity and the inner diameter of the lower conduit, providing scientific theoretical support for the regulation of core operating parameters and eliminating reliance on empirical operation, enabling precise control of the jet state. Simultaneously, by controlling the feed pressure to create a strong turbulent environment in the mixing zone of the lower conduit, the collision frequency and adsorption efficiency between the collector and the mineral surface are significantly increased. This allows the collector to be quickly and efficiently adsorbed onto the target mineral surface, meeting the process requirements of rapid flotation and significantly improving the separation efficiency of the JAMESON flotation machine in rapid flotation operations. In addition, precise control of turbulence intensity ensures more uniform and sufficient adsorption of the collector on the mineral surface, enhancing the floatability of the target mineral and laying a crucial foundation for the subsequent efficient separation of high-grade copper concentrate. This ensures the separation effect of the rapid flotation stage and ultimately improves the overall copper mineral recovery efficiency in the flotation process.

[0024] The high-grade copper concentrate, accounting for 60% to 75% of the total copper recovery from the raw ore, separated in step d, is achieved through the multi-stage series separation characteristics of the Jameson flotation machine. Specifically, a rapid flotation kinetic model based on mineral liberation and floatability is constructed within the Jameson flotation machine. The model determines the flotation rate constant k of copper minerals at different particle sizes through batch flotation rate experiments. Combined with the particle size distribution of the regrinding product, the theoretical recovery rate R within a short-time flotation period in the Jameson flotation machine is calculated. The calculation formula is as follows: ; Where mi is the mass fraction of the i-th particle size, ki is the flotation rate constant of the i-th particle size, and t is the flotation time. By adjusting the flotation time t, the calculated recovery rate R(t) is made to reach 60% to 75% of the total copper recovery rate, thereby locking in this part of the high floatability and high grade copper minerals. Leveraging the multi-stage series separation characteristics of the Jameson flotation machine, a rapid flotation kinetic model based on mineral liberation and floatability was constructed. Through batch flotation rate experiments, the flotation rate constants for copper minerals at different particle sizes were determined, and the theoretical recovery rate was calculated by combining this with the particle size distribution of the regrinding product. This allows the separation process of high-grade copper concentrate to move beyond empirical judgment, providing a scientific theoretical and experimental basis. It enables precise control of flotation time, accurately targeting highly floatable, high-grade copper minerals, significantly improving the accuracy and controllability of high-grade copper concentrate separation. Simultaneously, the model calculation results are used to control... The flotation process can efficiently separate the target copper concentrate within a short time, meeting the process requirements of rapid flotation, improving the separation efficiency of the JAMESON flotation machine's rapid flotation operation, and avoiding the waste of reagents and energy caused by ineffective flotation. In addition, this method can accurately achieve the separation of high-grade copper concentrate with the target copper recovery rate, lock in the core high-value copper minerals in advance, effectively reduce the processing load of subsequent flotation processes, optimize the overall flotation process operation rhythm, and make the subsequent separation of copper, zinc and sulfur more targeted, further ensuring the separation effect and product quality of the overall flotation separation process.

[0025] In step e, controlling the thickness and aeration rate of the flotation column's foam layer is achieved by constructing an evaluation index for the stability of the foam layer, namely, the axial gas holding rate of the foam layer. In conjunction with the grade requirements of foam products, a relationship model between foam layer thickness H and concentrate grade β was established. By adjusting the flushing water volume and aeration volume, the cleaning effect of the foam layer was optimized, and the separation of copper, zinc and sulfur was enhanced. The calculation process of the axial gas holding rate of the foam layer is as follows: ; in, The gas holding rate at a distance h from the foam overflow weir. H represents the initial gas holding rate at the bottom of the foam. c The characteristic attenuation length is determined by controlling H. c Maintain the stability of the foam layer and the cleaning effect; By constructing a dedicated evaluation index for foam layer stability—the axial gas holding rate—a scientific and quantitative basis for determining the state of the flotation column foam layer is established, eliminating reliance on empirical operational judgments and improving the precision and scientific nature of foam layer control. Simultaneously, a model relating foam layer thickness to concentrate grade is established based on foam product grade requirements. This allows for precise adjustment of flushing water and aeration rates according to the target concentrate grade, optimizing the cleaning effect of the foam layer and effectively removing sulfur minerals embedded in the foam, significantly enhancing the separation of copper and zinc from sulfur. Furthermore, achieving optimal matching between foam layer stability and cleaning effect through precise control improves the quality of the concentrate obtained from the second-stage copper-zinc mixed flotation, ensuring the separation accuracy of both copper-zinc mixed concentrate and sulfur concentrate. Reasonable control of aeration and flushing water also avoids ineffective resource consumption, balancing the separation effect and economy of the flotation column operation. This provides a crucial guarantee for the efficient separation and enrichment of copper, zinc, and sulfur in the overall flotation process, improving the comprehensive recovery and utilization of mineral resources.

[0026] This method, through a combination of roughing, rapid flotation, and column flotation under low alkalinity conditions, significantly reduces flotation reagent usage and optimizes product structure. The method further includes steps for optimizing process parameters based on an economic benefit model: calculating the total reagent cost C. total and comprehensive economic benefits V total Furthermore, by optimizing the reagent dosage and flotation equipment parameters in steps b, d, and e, the total reagent cost C was reduced. total and comprehensive economic benefits V total The calculation process is as follows: ; ; Where Qi is the dosage of the i-th reagent, Pi is the unit price of the i-th reagent, Mj is the output of the j-th concentrate, Gj is the grade of the j-th concentrate, and Pricej is the price determined by the pricing coefficient for the corresponding grade. By designing a combined process of roughing-rapid flotation-column flotation under low alkalinity conditions, the use of flotation reagents is significantly reduced at the process level, directly decreasing reagent costs. Simultaneously, the product structure is optimized, achieving graded enrichment and efficient separation of copper, zinc, and sulfur minerals, thus enhancing the utilization value of mineral resources. Furthermore, an economically-based process parameter optimization step is introduced, quantifying the total reagent cost and overall economic benefits. This ensures that process parameter adjustments are not solely driven by technical indicators but also consider technical feasibility and economic rationality, providing a clear economic standard for parameter optimization. In addition, by specifically optimizing reagent usage and equipment parameters at each flotation stage, cost inputs during production can be precisely controlled, maximizing overall economic benefits. This makes the flotation separation process for high-sulfur copper-zinc ore more economical and market-adaptable, increasing the overall profitability of mineral processing operations. Moreover, this optimization method allows for dynamic adjustment of process parameters based on actual production costs and product prices, enhancing the flexibility and adaptability of the entire flotation process and making it more practically valuable in real-world applications.

[0027] The above are merely preferred embodiments of the present invention and are not intended to limit the present invention in any way. Although the present invention has been disclosed above with reference to preferred embodiments, it is not intended to limit the present invention. Any person skilled in the art can make some modifications or alterations to the above-disclosed technical content to create equivalent embodiments without departing from the scope of the present invention. Any simple modifications, equivalent changes and alterations made to the above embodiments based on the technical essence of the present invention without departing from the scope of the present invention shall still fall within the scope of the present invention.

Claims

1. A flotation separation method for high-sulfur copper-zinc ore beneficiation, characterized in that, Includes the following steps: Step a. Classify and grind the high-sulfur copper-zinc ore to control the grinding fineness so that the product particle size P80 is 120μm to 150μm; Step b. The grinding product obtained in step a is subjected to the first stage of copper-zinc mixed flotation. Lime is added during the flotation process, and an aerated flotation machine is used as the flotation equipment to carry out aerated flotation to obtain copper-zinc-sulfur mixed concentrate and waste tailings. Step c. Feed the copper-zinc-sulfur mixed concentrate obtained in step b into a vertical spiral stirred mill for regrinding. The regrinding fineness is controlled to P90≤20μm to achieve full mineral dissociation and surface desorption. Step d. Feed the regrinding product obtained in step c into a flotation machine for flotation, adding only O-ethyl-N-benzoyl thiocarbamate as a collector at a dosage of 3 g / t to 5 g / t, to separate copper concentrate; Step e. Feed the tailings from step d into a flotation column for a second stage of copper-zinc mixed flotation. By controlling the thickness of the froth layer and the aeration rate of the flotation column, copper and zinc are separated from sulfur. After multiple fine selections, the flotation column concentrate is a copper-zinc mixed concentrate, and the flotation column tailings are a sulfur concentrate.

2. The flotation separation method for high-sulfur copper-zinc ore according to claim 1, characterized in that: Step a involves graded grinding. Based on the mineral distribution characteristics of copper, zinc, and sulfur in the raw ore, the unit power consumption required to achieve the target particle size P80 is calculated using the Bond work index formula to optimize grinding parameters.

3. The flotation separation method for high-sulfur copper-zinc ore according to claim 2, characterized in that: In step b, the amount of lime added is 100 g / t to 200 g / t to adjust the pH of the slurry to the range of 7.5 to 8.5, forming a surface electrostatic potential difference between chalcopyrite, sphalerite and pyrite; and the selective collector O-ethyl-N-benzoyl thiocarbamate is added at a dosage of 25 g / t to 30 g / t.

4. The flotation separation method for high-sulfur copper-zinc ore according to claim 3, characterized in that: In step b, the difference in surface electrostatic potential between chalcopyrite, sphalerite and pyrite is formed by online monitoring of the pH value and redox potential of the slurry and feeding the monitoring data back to the lime addition system to dynamically adjust the amount of lime added so that the pH value and redox potential value of the slurry meet the preset electrochemical window. The method for constructing the preset electrochemical window is as follows: based on the thermodynamic E-pH diagram, the surface electrostatic potentials of chalcopyrite, sphalerite, and pyrite under preset pH conditions are calculated using the Nernst equation; Based on the calculated surface electrostatic potentials of chalcopyrite, sphalerite, and pyrite under preset pH conditions, an electrochemical control model was established to ensure that the electrostatic potential of pyrite is lower than the pulp potential, while the electrostatic potentials of chalcopyrite and sphalerite are higher than the pulp potential.

5. The flotation separation method for high-sulfur copper-zinc ore according to claim 4, characterized in that: In step b, the aerated flotation machine obtains the bubble size distribution data on the foam surface through an online image analysis system, and then calculates the aeration rate based on the bubble coalescence rate constant k. coal Dynamic adjustment; bubble coalescence rate constant k coal The result is obtained by analyzing the sequence of foam images and calculating the rate of decrease in the number of bubbles per unit time.

6. The flotation separation method for high-sulfur copper-zinc ore according to claim 5, characterized in that: In step d, copper concentrate is separated using the jet entrainment and microbubble precipitation mechanism of the flotation machine. The jet entrainment and microbubble precipitation mechanism is achieved by controlling the jet Reynolds number. Based on the matching relationship between the jet velocity and the inner diameter of the lower conduit determined by computational fluid dynamics simulation, a turbulent environment is formed in the mixing zone of the lower conduit by controlling the feed pressure, so that the collector can achieve high-speed collision and adsorption on the mineral surface; the jet Reynolds number is the ratio of the inertial force to the viscous force of the slurry in the lower conduit of the flotation machine, and is a dimensionless number characterizing the intensity of turbulence.

7. The flotation separation method for high-sulfur copper-zinc ore according to claim 6, characterized in that: The copper concentrate separated in step d accounts for 60% to 75% of the total copper recovery of the raw ore. This is achieved through the multi-stage series separation characteristics of the flotation machine. A rapid flotation kinetic model based on mineral liberation degree and floatability is constructed in the flotation machine. The rapid flotation kinetic model determines the flotation rate constant k of copper minerals at different particle sizes through batch flotation rate tests. Combined with the particle size distribution of the regrinding product, the theoretical recovery rate in the short-time flotation of the flotation machine is calculated.

8. The flotation separation method for high-sulfur copper-zinc ore according to claim 7, characterized in that: In step e, lime and O-ethyl-N-benzoyl thiocarbamate are added during the flotation process, with dosages of 100 g / t to 200 g / t and 5 g / t to 10 g / t, respectively.

9. The flotation separation method for high-sulfur copper-zinc ore according to claim 8, characterized in that: In step e, by controlling the thickness of the froth layer and the aeration rate of the flotation column, an evaluation index for the stability of the froth layer, namely the axial gas holding rate of the froth layer, is constructed. Combined with the grade requirements of the froth product, a relationship model between the thickness of the froth layer and the grade of the concentrate is established. By adjusting the amount of flushing water and the amount of aeration, the separation of copper, zinc and sulfur is enhanced.

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