Efficient mineral selective flotation reagent and application thereof

Abstract

The invention discloses an efficient mineral selective flotation reagent and application thereof, and aims to improve the selectivity and recovery rate of mineral flotation. The composition comprises a first precursor (P1), a second precursor (P2), and a resonance activation adjuvant (Aux). The first precursor P1 has an anchoring group capable of acting on the surface of a target mineral and a first reactive functional group; the second precursor P2 includes a hydrophobic group and a second reactive functional group capable of a ligation reaction with the first reactive functional group of P1. The application method comprises the following steps: adding P1 and Aux into ore pulp to react with minerals; and applying an external energy field (such as ultrasonic wave or illumination with specific wavelength) to activate P1 and Aux adsorbed on the surface of the mineral. Through an in-situ interface reaction strategy assisted by an external energy field, directional adsorption and effective coverage of the agent on the surface of the mineral are remarkably improved, flotation dynamics is improved, flotation requirements of different mineral types can be met, and a new technical approach is provided for efficient utilization of complex and refractory ores.

Description

Technical Field

[0001] This invention relates to the field of mineral processing engineering technology, specifically to a highly efficient mineral selective flotation reagent and its application. Background Technology

[0002] Mineral flotation, as a key mineral processing technology, plays a vital role in separating and enriching valuable minerals from ores, and is widely used in the processing of metallic ores, non-metallic ores, and even coal. The core of the flotation process lies in utilizing the differences in the physicochemical properties of mineral surfaces. By adding flotation reagents to the slurry, the hydrophobicity of the target mineral surface is selectively altered, allowing it to adhere to air bubbles and float, thereby separating it from gangue minerals.

[0003] In traditional flotation practices, collectors are the core flotation reagents. Their molecules typically contain a polar group that interacts with the mineral surface and a nonpolar hydrophobic hydrocarbon chain group. However, with the gradual depletion of easily beneficiated mineral resources, the grade of ores to be processed is constantly declining, the mineral composition is becoming increasingly complex, and the particle size is trending towards finer distribution. Against this backdrop, traditional collectors often fail to meet the requirements for efficient recovery in terms of selectivity and collecting capacity when processing these complex and difficult-to-process ores. Many existing collectors do not exhibit ideal selective adsorption of target minerals, easily leading to the loss of valuable minerals or a reduction in concentrate grade. This is especially true when processing symbiotic mineral systems with similar properties, where the difficulty of precise separation increases significantly.

[0004] Furthermore, the adsorption strength of traditional collectors on mineral surfaces and the stability of the hydrophobic layers they form face challenges. In complex slurry environments, such as those with high ionic strength, wide pH ranges, or the presence of interfering ions, the adsorption effect of collectors may be significantly reduced, leading to fluctuations in flotation efficiency. The interaction forces between reagent molecules and the mineral surface may be insufficient to resist the shearing and scouring effects of the slurry, making it easy for adsorbed reagents to desorb and difficult to form a durable and effective hydrophobic surface. Simultaneously, the environmental friendliness of some traditional collectors is also a concern, as their production and use may impose certain environmental pressures. Therefore, developing novel, highly efficient, highly selective, and environmentally friendly flotation reagents and their application methods to meet the current needs of mining development has significant theoretical and practical value. Summary of the Invention

[0005] The purpose of this invention is to address the problems existing in current mineral flotation technologies, such as low selectivity, low reagent efficiency, poor adaptability to complex associated minerals, and environmental unfriendliness of some reagents. To this end, this invention provides a highly efficient mineral selective flotation reagent and its application, aiming to achieve efficient and precise separation of target minerals through a unique selective activation and in-situ reaction mechanism.

[0006] The first aspect of this invention provides a highly efficient mineral selective flotation reagent composition.

[0007] The composition mainly consists of a specially designed first precursor P1, a second precursor P2, and a resonance activator.

[0008] Specifically, the first precursor P1 comprises 10 to 45 parts by weight in the composition. The molecular structure of the first precursor P1 includes at least one anchoring functional group and at least one first reactive functional group. The anchoring functional group is designed to selectively adsorb or anchor to specific sites on the surface of the target mineral; this anchoring functional group may be selected from a thiol group, a carboxyl group, a phosphate group, a hydroxamic acid group, or a combination thereof. The first reactive functional group is used for subsequent chemical transformation on the mineral surface; this first reactive functional group may be selected from a terminal alkynyl group or an azide group. To achieve effective surface action and subsequent reactions, the first precursor P1 is typically an organic compound whose molecular skeleton may include, but is not limited to, benzoic acid derivatives, naphthoic acid derivatives, or sulfur-containing heterocyclic compounds, and its molecular weight is typically controlled in the range of 200 g / mol to 600 g / mol.

[0009] The second precursor P2 comprises 20 to 75 parts by weight in the composition. The molecular structure of the second precursor P2 includes at least one second reactive functional group and at least one hydrophobic group. The second reactive functional group is designed to undergo an efficient and specific chemical reaction with the first reactive functional group on the first precursor P1; therefore, the second reactive functional group is selected accordingly from an azide group or a terminal alkynyl group to ensure precise pairing with the first precursor P1.

[0010] The hydrophobic group imparts strong hydrophobic properties to the newly formed surface material after the reaction of the first precursor P1 and the second precursor P2, thereby promoting the floating of the target mineral. This hydrophobic group can be selected, for example, from C8-C18 straight-chain or branched alkyl groups, or C6-C12 perfluoroalkyl groups. In some preferred embodiments, the hydrophobic group can be dodecyl, hexadecyl, or perfluorooctyl. To achieve optimal reaction efficiency and product performance, the weight ratio between the first precursor P1 and the second precursor P2 is typically controlled within the range of 1:0.5 to 3.5.

[0011] The resonance activating agent is present in the composition in parts from 0.1 to 5 by weight. The key function of this agent is that, when the system is subjected to a specific external physical energy field, it synergistically enhances the selective absorption and / or chemical activation of the first precursor P1 adsorbed on the target mineral surface. This synergistic enhancement effect allows P1 to be preferentially and efficiently activated on the target mineral surface. For P1 that may be adsorbed in small amounts on the gangue mineral surface, the activation level will be significantly reduced or not occur due to the lack of this effective energy resonance matching or the directional enhancement effect of the agent. The resonance activating agent can be selected from substances with specific photophysical or sonochemical response characteristics, such as anthraquinone compounds, tetraphenylporphyrin compounds, semiconductor quantum dots with an average particle size of 5 nm to 50 nm, or organic dye molecules that strongly absorb light of specific wavelengths.

[0012] In some embodiments, to further improve the rate and efficiency of the in-situ reaction between the first precursor P1 and the second precursor P2, the flotation reagent composition also includes a reaction promoter. The amount of the reaction promoter is 0.01% to 1% of the total weight of the first precursor P1 and the second precursor P2. The reaction promoter is primarily used to catalyze specific chemical bonding, such as click chemistry, between a first reactive functional group on the first precursor P1 and a second reactive functional group on the second precursor P2. Preferably, the reaction promoter is a substance with minimal environmental impact and high catalytic activity, such as a specific non-toxic or low-toxic copper complex or a highly efficient organic catalyst.

[0013] A second aspect of the present invention provides a method for mineral selective flotation using the aforementioned high-efficiency mineral selective flotation reagent composition.

[0014] This method typically includes the following steps:

[0015] First, the slurry is prepared: the ore raw material to be processed undergoes conventional pretreatment processes such as crushing and grinding until it reaches a particle size suitable for the flotation process. For example, the content of mineral particles with a diameter of less than 74 μm is usually 70% to 90% of the total ore. Then, the finely ground ore powder is mixed with an appropriate amount of water to prepare a slurry with a certain fluidity. The mass concentration of the slurry is generally controlled in the range of 5% to 35% w / v.

[0016] Next, the pH of the slurry is adjusted, and the first precursor P1 and the resonance activator are added and adsorbed: Depending on the type of ore and the needs of subsequent reagent action, the pH of the slurry is adjusted to a preset suitable range using an acidic or alkaline adjuster, typically between pH 3.0 and pH 11.0. After the pH stabilizes, the first precursor P1 and the resonance activator from the aforementioned flotation reagent composition are added to the slurry. The amount of the first precursor P1 added can be determined based on the ore throughput. After the reagents are added, thorough adsorption is carried out under stirring conditions. The stirring time is generally 2 to 20 minutes to ensure that the first precursor P1 and the resonance activator can effectively contact the surface of the mineral particles and undergo selective adsorption.

[0017] Next, selective energy resonance activation is performed: a specific external physical energy field is applied to the slurry system that has completed the above adsorption process. This energy field can be ultrasound of a specific frequency, which can be selected in the range of 20 kHz to 3.0 MHz, and the applied power density is typically controlled at 0.1 W / cm². 2 Up to 5.0 W / cm 2 Alternatively, the energy field can be light of a specific wavelength, selectable within the range of 254 nm to 650 nm, with its intensity typically controlled at 1 mW / cm². 2 Up to 100mW / cm 2 The continuous irradiation time of the energy field is generally from 1 minute to 30 minutes. The purpose of this step is to selectively excite and activate the first precursor P1 molecules, which are mainly adsorbed on the surface of the target mineral particles, using the principle of energy resonance, to provide the necessary energy or active state for their subsequent chemical reactions. In some preferred embodiments, to improve energy utilization efficiency and activation effect, the external physical energy field can be applied in a pulsed mode. If ultrasound is used, the pulse on-time can be set to 0.1 seconds to 5 seconds, the pulse off-time can be set to 0.5 seconds to 10 seconds, and the corresponding duty cycle can be controlled within the range of 10% to 80%; if light irradiation is used, the pulse frequency can be set to 0.1 Hz to 10 Hz.

[0018] Then, the second precursor P2 is added and reacted in situ: After applying an external physical energy field to the slurry or in the later stages of the sustained energy field application, the second precursor P2 from the above-mentioned flotation reagent composition is added to the slurry. The amount of the second precursor P2 added is usually matched with the amount of the first precursor P1, and the molar ratio of the first precursor P1 to the second precursor P2 can be controlled within the range of 1:0.5 to 3.0. In this step, to ensure the smooth progress of the chemical reaction, the temperature of the slurry is usually maintained in the range of 15°C to 45°C. After adding P2, the slurry is stirred continuously for 2 to 30 minutes. During this process, the first precursor P1 molecules, which were previously selectively activated by the energy field, and the second precursor P2 molecules undergo an in-situ chemical reaction on the surface of the target mineral, generating a target product with strong hydrophobicity. If the flotation reagent contains a reaction promoter, the reaction promoter can be added to the slurry simultaneously with or in stages with the second precursor P2 to improve the efficiency of the in-situ reaction.

[0019] Finally, flotation separation is performed: after the in-situ reaction has fully occurred, air or other suitable gases are introduced into the slurry. The gas flow rate can be adjusted according to the volume and type of the flotation cell, typically within the range of 0.1 L / min to 5.0 L / min, for conventional flotation operations. Since a highly efficient hydrophobic layer has formed on the surface of the target mineral particles in situ, they preferentially adhere to the air bubbles and float to the surface of the slurry, forming a froth product rich in the target minerals. This froth is scraped off and collected using a scraper or similar device. Gangue minerals, whose surfaces remain hydrophilic or weakly hydrophobic, are mostly retained in the slurry tank and discharged with the tailings. In some preferred embodiments, to further enhance the formation and adhesion of hydrophobic products on the surface of the target minerals and improve the flotation recovery rate, an external physical energy field with the same or appropriately weakened parameters as the aforementioned selective energy resonance activation step can be applied during the initial stage of the flotation separation process. In subsequent stages of the flotation process, the application of the energy field can be stopped as needed, and / or a suitable amount of frother can be added to improve the stability and load-bearing capacity of the froth layer.

[0020] In some embodiments, to achieve a greener mineral processing flow, the method of the present invention may further include a step of harmlessly or resourcefully treating the tailings slurry discharged after flotation separation. For example, an appropriate amount of mild oxidant may be added to the tailings slurry and stirred at a temperature of 20°C to 50°C for 10 to 60 minutes to promote the effective degradation of any unreacted precursor molecules or reaction byproducts that may remain in the slurry, thereby reducing the potential environmental impact of mineral processing wastewater.

[0021] In summary, the present invention has at least one of the following beneficial technical effects:

[0022] 1. This invention utilizes specific anchored functional groups on the first precursor P1 to initialize chemical recognition and adsorption onto the target mineral surface. Combined with an external physical energy field, such as ultrasound of a specific frequency or light of a specific wavelength, and synergistic action with a resonance activation aid, selective energy resonance activation is performed on the adsorbed P1 on the target mineral surface, ensuring that subsequent chemical reactions mainly occur on the target mineral surface. This "chemical recognition-physical activation dual-locking" mechanism allows the hydrophobic layer to preferentially and primarily form in situ on the target mineral surface, significantly reducing non-specific effects on gangue minerals. This enables high-precision selective flotation of complex minerals or finely embedded minerals that are difficult to separate effectively using traditional methods.

[0023] 2. This invention employs a strategy of "in-situ, on-demand" reaction of the first precursor P1 and the second precursor P2 on the surface of the target mineral to generate the final hydrophobic product. This means that only when the first precursor P1 is selectively activated will it effectively react with the second precursor P2 to form a hydrophobic layer. This approach avoids the waste of reagents caused by the indiscriminate adsorption of traditional collectors on all mineral surfaces after being added to the slurry, and in particular reduces ineffective consumption on the surfaces of a large number of gangue minerals. Therefore, this invention can achieve excellent flotation results with a relatively low total amount of precursor reagents, effectively reducing the reagent consumption per unit of mineral processing costs.

[0024] 3. This invention introduces an external physical energy field as a key control parameter. By precisely controlling the physical parameters of the energy field, including its type (e.g., ultrasound or light), frequency, wavelength, power, intensity, duration, and pulse mode, the activation level of the first precursor P1 and the kinetics of the in-situ reaction can be finely controlled. This non-chemical control method, combined with the selection of the anchoring groups of the first precursor P1 or the hydrophobic chain length and ratio of the second precursor P2, as well as the type of resonance activation aid, provides a wider and more flexible optimization space for the flotation process, enabling it to better adapt to complex ores with different types, different embedding characteristics, and varying surface properties.

[0025] 4. The first precursor P1 and the second precursor P2 used in this invention can be designed as simple organic molecules with low chemical activity, low toxicity, or good biodegradability. When unactivated or without in-situ reaction, their environmental impact is relatively small. Simultaneously, the "on-demand synthesis" strategy helps reduce the total amount of reagents used and their residues in the environment. The energy field itself is a clean physical mechanism that does not introduce additional chemical pollutants. Furthermore, the method may include a step of degrading residual reagents in tailings, further reducing the overall environmental impact of the mineral processing process.

[0026] 5. This invention organically integrates the principles of selective energy physics, such as resonant absorption, with advanced concepts and technologies such as mineral interface chemistry, in-situ dynamic chemical reactions, and external physical field control to construct a novel flotation system. This strategy, based on "remote activation-surface catalysis-in-situ self-assembly," overcomes the limitations of traditional flotation reagents that primarily rely on the interaction between static molecular structures and mineral surfaces. It provides an innovative theoretical framework and feasible technical approach for designing next-generation "intelligent" flotation reagents and processes capable of responding to external stimuli and achieving precise identification and efficient separation. Detailed Implementation

[0027] To make the objectives, technical solutions, and advantages of this invention clearer, the technical solutions of this invention will be further described in detail below with reference to specific embodiments. It should be noted that the embodiments described herein are only for explaining this invention and are not intended to limit the scope of this invention.

[0028] Any modifications, equivalent substitutions, or improvements made within the spirit and principles of this invention shall be included within the scope of protection of this invention. Unless otherwise stated, all reagents and raw materials used in the embodiments of this invention are commercially available analytical grade or chemically pure products, or can be prepared by conventional methods in the art. Unless otherwise specified, the experimental methods used are conventional methods in the art.

[0029] Example 1, Preparation of the first precursor P1:

[0030] Example 1-1, Preparation of 4-mercaptobenzoic acid-2-propynyl-1-ester (P1a):

[0031] I. Preparation of 4,4'-dithiodibenzoic acid:

[0032] In a 500 mL three-necked flask, add 30.0 g (0.219 mol) of 4-aminobenzoic acid and 150 mL of concentrated hydrochloric acid. Cool to 0–5 °C in an ice bath, and slowly add 40 mL of an aqueous solution of sodium nitrite (16.6 g (0.241 mol) while maintaining the temperature below 5 °C. After the addition is complete, continue stirring for 30 minutes to obtain a diazonium salt solution.

[0033] In another 1L three-necked flask, add sodium thiosulfate pentahydrate (82.0 g, 0.330 mol) and water (200 mL), stir to dissolve, and then cool to 5-10°C in an ice bath. Slowly add the prepared diazonium salt solution dropwise to the sodium thiosulfate solution under vigorous stirring, keeping the temperature below 10°C. After the addition is complete, remove the ice bath and heat to 60°C to react for 2 hours.

[0034] After the reaction was complete, an aqueous solution of sodium hydroxide (40.0 g, 1.00 mol) (100 mL) was added, and the mixture was heated under reflux for 4 hours. After cooling to room temperature, the solution was acidified with concentrated hydrochloric acid to pH 2-3, precipitating a large amount of solid. The solid was filtered, washed with water and ethanol, and dried to give a grayish-white solid, 4,4'-dithiodibenzoic acid, in 85% yield.

[0035] II. Preparation of 4-Mercaptobenzoic acid:

[0036] In a 500 mL three-necked flask, add 20.0 g (0.065 mol) of 4,4'-dithiodibenzoic acid prepared in step 1 and 200 mL of ethanol. Heat to reflux, add sodium borohydride (5.0 g (0.132 mol) in portions, and continue refluxing for 2 hours after the addition is complete.

[0037] After the reaction was complete, the mixture was cooled to room temperature, and concentrated hydrochloric acid was slowly added dropwise to adjust the pH to 2-3, causing a solid to precipitate. The solid was filtered, washed with water, and dried to obtain a white solid, 4-mercaptobenzoic acid, with a yield of 90%. Melting point: 218-220℃.

[0038] III. Preparation of 4-mercaptobenzoic acid-2-propynyl-1-ester (P1a):

[0039] In a 250 mL three-necked flask, 10.0 g (0.065 mol) of 4-mercaptobenzoic acid prepared in step 2 and 100 mL of anhydrous N,N-dimethylformamide (DMF) were added and stirred to dissolve. Potassium carbonate (17.9 g (0.130 mol)) and propargyl bromide (11.6 g of 80% toluene solution, containing 0.078 mol of propargyl bromide) were added. The mixture was stirred at 60 °C for 6 hours. After the reaction was complete, the reaction solution was poured into 500 mL of ice water, and a solid precipitated. The solid was filtered, washed with water until neutral, and then washed with a small amount of cold ethanol. The crude product was recrystallized from ethyl acetate / n-hexane to give 11.2 g (83%) of white needle-like crystals.

[0040] Examples 1-2: Preparation of 3-(4-azidobutoxy)benzoic acid (P1b):

[0041] I. Preparation of methyl 3-hydroxybenzoate:

[0042] In a 250 mL round-bottom flask, 13.8 g (0.10 mol) of 3-hydroxybenzoic acid and 100 mL of methanol were added, followed by slow dropwise addition of 2 mL of concentrated sulfuric acid. The mixture was heated under reflux for 8 hours. After the reaction was complete, most of the methanol was evaporated under reduced pressure, and the residue was poured into 200 mL of ice water and extracted with ethyl acetate (3 × 50 mL). The combined organic phases were washed with saturated sodium bicarbonate solution, then with saturated brine, and dried over anhydrous sodium sulfate. The mixture was filtered, and the solvent was evaporated under reduced pressure to give 14.5 g (95%) of a pale yellow oil.

[0043] II. Preparation of methyl 3-(4-bromobutoxy)benzoate:

[0044] In a 250 mL three-necked flask, add methyl 3-hydroxybenzoate (10.0 g, 0.066 mol), anhydrous acetone (100 mL), potassium carbonate (18.2 g, 0.132 mol), and 1,4-dibromobutane (28.5 g, 0.132 mol). Heat under reflux for 24 hours.

[0045] After the reaction was complete, insoluble matter was removed by filtration, and the filtrate was concentrated under reduced pressure. The residue was dissolved in dichloromethane, washed with water and saturated brine, dried over anhydrous sodium sulfate, filtered, and the solvent was removed by vacuum evaporation. The crude product was purified by silica gel column chromatography (petroleum ether:ethyl acetate = 20:1) to give methyl 3-(4-bromobutoxy)benzoate (15.3 g), a colorless oil, in 77% yield.

[0046] III. Preparation of methyl 3-(4-azidobutoxy)benzoate:

[0047] In a 100 mL round-bottom flask, add methyl 3-(4-bromobutoxy)benzoate (10.0 g, 0.033 mol), N,N-dimethylformamide (DMF, 50 mL), and sodium azide (3.2 g, 0.049 mol). Stir the mixture at 70 °C for 12 hours.

[0048] After the reaction was complete, the reaction solution was poured into 300 mL of ice water and extracted with ethyl acetate (3 × 50 mL). The organic phases were combined, washed with water (3 × 50 mL), washed with saturated brine, dried over anhydrous sodium sulfate, filtered, and the solvent was removed under reduced pressure to give methyl 3-(4-azidobutoxy)benzoate (8.2 g), a pale yellow oil, with a yield of 94%.

[0049] IV. Preparation of 3-(4-azidobutoxy)benzoic acid (P1b):

[0050] In a 100 mL round-bottom flask, add methyl 3-(4-azidobutoxy)benzoate (8.0 g, 0.030 mol), methanol (30 mL), water (10 mL), and sodium hydroxide (2.4 g, 0.060 mol). Stir the mixture at room temperature for 4 hours.

[0051] After the reaction was completed, methanol was removed by vacuum distillation, the residue was diluted with water, and acidified with concentrated hydrochloric acid to pH 2-3, precipitating a white solid. The solid was filtered, washed with water, and dried to obtain a white solid 3-(4-azidobutoxy)benzoic acid (P1b, 7.0 g), with a yield of 93%.

[0052] Example 2, Preparation of the second precursor P2:

[0053] Example 2-1, Preparation of 1-azidohexadecane (P2a)

[0054] In a 250 mL three-necked flask, add 20.0 g of 1-bromohexadecane (0.0655 mol), 100 mL of N,N-dimethylformamide (DMF), and 6.4 g of sodium azide (0.0983 mol). Stir the mixture at 80 °C for 12 hours.

[0055] After the reaction was complete, the reaction solution was cooled to room temperature and poured into 500 mL of ice water. It was then extracted with n-hexane (3 × 80 mL). The organic phases were combined and washed successively with water (2 × 100 mL), saturated brine (100 mL), dried over anhydrous sodium sulfate, filtered, and the solvent was removed by vacuum distillation.

[0056] Example 2-2, Preparation of 1-(2-propyn-1-oxy)dodecane (P2b):

[0057] In a 250 mL three-necked flask, add 20.0 g (0.107 mol) of 1-dodecanol and 100 mL of anhydrous tetrahydrofuran (THF), and cool in an ice bath. Add sodium hydride (6.4 g (0.160 mol) in portions. After the addition is complete, bring to room temperature and stir for 1 hour until no more bubbles are produced. Cool again in an ice bath, and slowly add propargyl bromide (20.1 g of 80% toluene solution, containing 0.135 mol of propargyl bromide). After the addition is complete, bring to room temperature and stir overnight.

[0058] After the reaction was complete, excess sodium hydride was carefully quenched dropwise with saturated ammonium chloride solution. Most of the THF was removed by vacuum distillation, and 100 mL of water was added. The mixture was extracted with ethyl acetate (3 × 50 mL). The organic phases were combined, washed with water and saturated brine, dried over anhydrous sodium sulfate, filtered, and the solvent was removed by vacuum distillation. The crude product was purified by silica gel column chromatography (petroleum ether:ethyl acetate = 50:1) to give a colorless oily liquid, 1-(2-propynyl-1-oxy)dodecane (P2b, 20.5 g), in 85% yield.

[0059] Example 3, Selection and treatment of anthraquinone (Aux1a):

[0060] Commercially available analytical grade anthraquinone powder was selected. A certain amount of anthraquinone powder was placed in an agate mortar and ground thoroughly with a pestle for 30 minutes to further refine it and increase its specific surface area. The ground anthraquinone powder was then passed through a 200-mesh standard sieve, and the sieve-passing material was collected for later use. The powder appeared as a yellow powder.

[0061] Commercially available analytical grade anthraquinone powder was selected. A certain amount of anthraquinone powder was placed in an agate mortar and ground thoroughly with a pestle for 30 minutes to further refine it and increase its specific surface area. The ground anthraquinone powder was then passed through a 200-mesh standard sieve, and the sieve-passing material was collected for later use. The powder appeared as a yellow powder.

[0062] Example 4, Preparation of the flotation reagent composition:

[0063] Example 4-1, Preparation of flotation reagent composition A:

[0064] Weigh 30.0 parts by weight of the first precursor 4-mercaptobenzoic acid-2-propynyl-1-ester (P1 a) prepared in Example 1-1, and 1.0 parts by weight of the resonance activator anthraquinone (Aux1 a) prepared and treated in Example 3. Place P1 a and Aux1 a together in an agate mortar and grind them thoroughly to obtain a powdery mixture, which is the first part of composition A.

[0065] Weigh 50.0 parts by weight of the second precursor 1-azidohexadecane (P2a) prepared in Example 2-1, which is the second part of composition A.

[0066] Weigh out 0.20 parts by weight of sodium ascorbate and 0.04 parts by weight of copper(II) sulfate pentahydrate as reaction promoters. Dissolve sodium ascorbate and copper(II) sulfate pentahydrate together in 5.0 parts by weight of deionized water to form a reaction promoter solution, which is the third part of composition A.

[0067] When using the composition, the first, second and third parts of the composition A are added to the slurry according to the requirements of the flotation process.

[0068] Example 4-2, Preparation of flotation reagent composition B:

[0069] Weigh 20.0 parts by weight of the first precursor 3-(4-azidobutoxy)benzoic acid (P1b) prepared in Examples 1-2, and 0.5 parts by weight of the resonance activator anthraquinone (Aux1a) prepared and treated in Example 3. Mix P1b and Aux1a in a ball mill using a zirconia ball dry milling method for 2 hours to obtain a highly dispersed powder mixture, which is the first part of composition B.

[0070] Weigh 40.0 parts by weight of the second precursor 1-(2-propynyl-1-oxy)dodecane (P2b) prepared in Example 2-2, which is the second part of composition B.

[0071] No additional reaction promoter is added in this embodiment.

[0072] When using the composition, the first and second parts of the composition B are added to the slurry respectively according to the requirements of the flotation process.

[0073] Example 4-3, Preparation of flotation reagent composition C:

[0074] 15.0 parts by weight of the first precursor 4-mercaptobenzoic acid-2-propynyl-1-ester (P1 a) prepared in Example 1-1 were weighed. 2.0 parts by weight of titanium dioxide nanoparticles (anatase type, commercially available, purity greater than 99.5%) with an average particle size of 20 nm were weighed as a resonance activator. P1 a and the titanium dioxide nanoparticles were dry-mixed in a high-speed shear mixer for 15 minutes to obtain a homogeneous mixture, which is the first part of composition C.

[0075] Weigh 50.0 parts by weight of the second precursor 1-(2-propynyl-1-oxy)dodecane (P2b) prepared in Example 2-2; this is the second part of composition C. Weigh 0.15 parts by weight of L-ascorbic acid and 0.02 parts by weight of cuprous acetate (I) as reaction promoters, and mix them thoroughly beforehand; this is the third part of composition C. In use, the first, second, and third parts of composition C are added to the slurry separately according to the requirements of the flotation process.

[0076] Example 4-4, Preparation of flotation reagent composition D:

[0077] Weigh 40.0 parts by weight of the first precursor 3-(4-azidobutoxy)benzoic acid (P1b) prepared in Examples 1-2. Weigh 0.2 parts by weight of titanium dioxide nanoparticles (rutile type, commercially available, purity greater than 99.0%) with an average particle size of 30 nm as a resonance activator. Dissolve / disperse P1b and titanium dioxide nanoparticles together in an appropriate amount of tetrahydrofuran, sonicate for 30 minutes to achieve uniform dispersion, and then remove the tetrahydrofuran by rotary evaporation under reduced pressure to obtain a supported powder, which is the first part of composition D.

[0078] Weigh 25.0 parts by weight of the second precursor 1-azidohexadecane (P2a) prepared in Example 2-1; this is the second part of composition D. No additional reaction promoter is added in this example. In use, the first and second parts of composition D are added to the slurry separately according to the requirements of the flotation process.

[0079] Example 5, Mineral Flotation Application:

[0080] General preparations before flotation:

[0081] All ore samples underwent the following treatment before use: coarse crushing to -10mm, followed by medium-fine crushing to -2mm using a laboratory jaw crusher and double roll crusher. Representative samples were then wet-milled using a laboratory rod mill (φ150mm×300mm). Grinding fineness was controlled by wet screening to ensure that the content of -0.074mm (-200 mesh) particles reached 75% to 85%. The slurry concentration after grinding was adjusted by weighing.

[0082] Ore sample description:

[0083] Ore A: A copper-molybdenum sulfide ore. After the above-mentioned crushing and grinding, its copper grade was measured to be 0.85% and its molybdenum grade to be 0.08%. The main copper-bearing mineral is chalcopyrite, the main molybdenum-bearing mineral is molybdenite, and the gangue minerals are mainly quartz, feldspar, and a small amount of pyrite.

[0084] Ore B: A low-grade pyrite containing gold, after the above crushing and grinding process, was found to have a gold grade of 1.5 g / t and a sulfur content of 15%. The gold mainly exists in the pyrite in the form of microscopic and submicroscopic gold inclusions, and the gangue minerals are mainly quartz and carbonate minerals.

[0085] Example 5-1: Copper-molybdenum mixed flotation of ore A using composition A

[0086] S1. Take 500.0 grams of the ground ore A slurry (dry weight) and add it to the 1.5-liter hanging trough flotation tank. Add deionized water until the slurry concentration is 30% (w / w). Turn on the agitator and set the agitator speed to 1800 rpm.

[0087] S2. Use a 10% (w / v) lime slurry solution to adjust the pH of the slurry to 9.5 and stabilize for 1 minute.

[0088] S3. Add the first part (a mixture of P1a and Aux1a) of the flotation reagent composition A prepared in Example 4-1 to the slurry, with an amount of 60 g / ton of raw ore based on P1a, and stir for 15 minutes for adsorption.

[0089] S4. Turn on the ultrasonic generator, place the ultrasonic probe 2cm below the surface of the slurry, set the ultrasonic frequency to 40kHz, the ultrasonic power density to 0.8W / cm², and continue for 20 minutes.

[0090] S5. Add the third part (reaction promoter solution) of flotation reagent composition A prepared in Example 4-1 to the slurry and stir for 1 minute. Then add the second part (P2a) of flotation reagent composition A prepared in Example 4-1, at a dosage of 100 grams of P2a per ton of raw ore. Continue stirring and reacting for 20 minutes while maintaining the slurry temperature at 25°C.

[0091] S6. Add No. 2 flotation oil (pine oil) as a frother at a dosage of 20 grams per ton of raw ore. Connect the air valve and aerate at a rate of 0.3 cubic meters per square meter per minute to begin flotation frothing and frothing, scraping the copper-molybdenum mixed rough concentrate for 6 minutes.

[0092] S7. The obtained copper-molybdenum mixed crude concentrate and tailings were filtered, dried, weighed, and samples were taken for chemical analysis to determine the grades of copper and molybdenum. The recovery rates of copper and molybdenum were calculated based on the analytical results. The results are shown in Table 1.

[0093] Example 5-2: Copper-molybdenum mixed flotation of ore A using composition B (pulse light activation)

[0094] S1. Take 500.0 grams of the ground ore A slurry (dry weight) and add it to the 1.5-liter hanging trough flotation tank. Add deionized water until the slurry concentration is 30% (w / w). Turn on the agitator and set the agitator speed to 1800 rpm.

[0095] S2. Use a 5% (w / v) sodium carbonate solution to adjust the pH of the pulp to 8.5 and stabilize for 1 minute.

[0096] S3. Add the first part (a mixture of P1b and Aux1 a) of the flotation reagent composition B prepared in Example 4-2 to the slurry, with an amount of 40 g / ton of raw ore based on P1b, and stir for 15 minutes for adsorption.

[0097] S4. Turn on the LED light source array (wavelength 365nm), place the light source 10cm above the surface of the flotation tank, and adjust the illumination area to cover the entire surface of the liquid. Set the light intensity to 20 milliwatts per square centimeter. Use pulse mode for photoactivation: pulse on time is 1.0 second, pulse off time is 3.0 seconds, and the total photoactivation time (cumulative based on on time) is 10 minutes, i.e., the total process lasts 40 minutes.

[0098] S5. Add the second part (P2b) of the flotation reagent composition B prepared in Example 4-2 to the slurry, at a dosage of 80 grams per ton of raw ore (based on P2b). Continue stirring and reacting for 25 minutes while maintaining the slurry temperature at 30°C.

[0099] S6. Add methyl isobutyl methanol (MIBC) as a frother at a dosage of 15 grams per ton of raw ore. Connect the aeration valve and aerate at a rate of 0.3 cubic meters per square meter per minute to begin flotation frothing and frothing, scraping the copper-molybdenum mixed rough concentrate for 6 minutes.

[0100] S7. The obtained copper-molybdenum mixed crude concentrate and tailings were filtered, dried, weighed, and samples were taken for chemical analysis to determine the grades of copper and molybdenum. The recovery rates of copper and molybdenum were calculated based on the analytical results. The results are shown in Table 1.

[0101] Example 5-3: Flotation of ore B containing gold-bearing ferrite using composition C

[0102] S1. Take 500.0 grams of the ground ore B slurry (dry weight) and add it to the 1.5-liter hanging trough flotation tank. Add deionized water until the slurry concentration is 25% (w / w). Turn on the agitator and set the agitator speed to 2000 rpm.

[0103] S2. Use a 10% (v / v) sulfuric acid solution to adjust the pH of the slurry to 5.0 and stabilize for 1 minute.

[0104] S3. Add the first part (a mixture of P1a and TiO2) of the flotation reagent composition C prepared in Example 4-3 to the slurry, with an amount of 80 g / ton of raw ore based on P1a, and stir for 20 minutes for adsorption.

[0105] S4. Turn on the ultrasonic generator (same as in Example 5-1), set the ultrasonic frequency to 28kHz, the ultrasonic power density to 1.2 watts / square centimeter, and continue for 25 minutes.

[0106] S5. Add the third part (reaction promoter mixture) of flotation reagent composition C prepared in Example 4-3 to the slurry and stir for 1 minute. Then add the second part (P2b) of flotation reagent composition C prepared in Example 4-3, at a dosage of 120 g / ton of raw ore based on P2b. Continue stirring and reacting for 30 minutes while maintaining the slurry temperature at 35°C.

[0107] S6. Add No. 2 flotation oil as a frother at a dosage of 25 grams per ton of raw ore. Connect the air valve and aerate at a rate of 0.4 cubic meters per square meter per minute. Begin flotation frothing to scrape off the gold-bearing iron ore concentrate for 8 minutes.

[0108] S7. The obtained gold-bearing ferrite concentrate and tailings were filtered, dried, weighed, and samples were taken for chemical analysis to determine the gold grade and sulfur content. The gold recovery rate and sulfur recovery rate were calculated based on the analytical results. The results are shown in Table 1.

[0109] Example 5-4: Flotation tailings treatment in Example 5-1

[0110] Collect the tailings slurry discharged after flotation separation in Example 5-1. Take 1 liter of this tailings slurry (solid content approximately 20% w / w) and transfer it to a 2-liter beaker, placing it on a magnetic stirrer. At room temperature (25°C), add a 5% potassium persulfate solution dropwise to the tailings slurry, so that the final amount of potassium persulfate is 0.08% of the dry weight of the tailings. Continue stirring for 60 minutes. After processing, allow it to settle, and use the supernatant for subsequent analysis.

[0111] Example 5-5: Copper-molybdenum mixed flotation of ore A using composition A (with ultrasonication during the flotation process)

[0112] Repeat steps S1 to S5 of Example 5-1.

[0113] In step S6, add No. 2 flotation oil (pine oil) as a frother at a dosage of 20 grams per ton of raw ore. Connect the aeration valve, setting the aeration rate to 0.3 cubic meters per square meter per minute, and begin flotation frothing. During the first 3 minutes of frothing, continue applying an ultrasonic energy field with the same ultrasonic parameters as in step 4, but with the power density reduced to 0.2 watts per square centimeter. Stop after 3 minutes.

[0114] Apply ultrasound and continue scraping for 3 minutes, for a total scraping time of 6 minutes.

[0115] The obtained copper-molybdenum mixed crude concentrate and tailings were filtered, dried, weighed, and samples were taken for chemical analysis to determine the grades of copper and molybdenum. The recoveries of copper and molybdenum were calculated based on the analytical results. The results are shown in Table 1.

[0116] Table 1: Flotation results of various flotation application examples

[0117]

[0118] Examples 5-4 demonstrate tailings treatment processes, the effects of which are not reflected in the flotation indicators in this table.

[0119] Based on the data from the various flotation application examples in Table 1, it can be seen that the flotation reagent composition and its application method provided by this technical solution exhibit excellent technical effects in the flotation separation of different ore types and target minerals, specifically reflected in the high recovery rate of target minerals and the high grade of concentrate.

[0120] For ore A (copper-molybdenum sulfide ore), Example 5-1 used composition A, supplemented by continuous ultrasonic energy field activation, achieving good results with a copper grade of 20.0% and a recovery rate of 90.0%, and a molybdenum grade of 1.50% and a recovery rate of 72.0%. This indicates that composition A, with the assistance of ultrasound, can effectively interact with the surface of copper and molybdenum minerals, promoting their hydrophobicity and flotation.

[0121] Furthermore, Examples 5-5 also targeted ore A, using the same composition A, but adjusted the application of ultrasound to be synergistically applied during the initial stage of flotation frothing. The results showed that the grade and recovery of copper increased to 20.5% and 92.0%, respectively, while the grade and recovery of molybdenum also increased accordingly to 1.60% and 76.0%. This comparative result clearly indicates that synergistically applying the ultrasonic energy field during flotation can further enhance the interaction between reagents and minerals, optimize the bubble-mineral particle adhesion conditions, and thus further improve the separation efficiency and recovery level of copper and molybdenum minerals based on continuous ultrasonic activation.

[0122] Example 5-2 demonstrates the flotation effect of composition B on copper-molybdenum minerals in ore A under light-activated (pulsed mode) conditions. Although its copper recovery (88.0%) and molybdenum recovery (67.0%) are slightly lower than those in Examples 5-1 and 5-5, it still yields copper concentrate with grades of 19.5% and molybdenum concentrate with grades of 1.40%, respectively. This proves that using light energy of a specific wavelength as an external energy field can effectively activate composition B, promoting the formation of a hydrophobic layer on the mineral surface and achieving effective collection of copper-molybdenum minerals. This reflects the diversity and adaptability of energy field selection in this technical solution.

[0123] For ore B (low-grade pyrite-bearing pyrite), Example 5-3 applied composition C, combined with continuous ultrasonic energy field activation. The experimental results yielded a gold concentrate with a gold grade of 4.7 g / t and a recovery rate of 88.0%, while the sulfur grade reached 45.0% with a recovery rate of 85.0%. This result demonstrates that this technical solution is not only suitable for the flotation of base metal sulfide ores such as copper and molybdenum, but also has a significant effect on the recovery of precious metals (such as gold) using pyrite as a carrier, effectively enriching the target mineral and achieving efficient recovery of valuable components.

[0124] In summary, the data in Table 1 strongly demonstrate that this technical solution, through a specific combination of the first precursor, the second precursor, and the resonant activation aid, combined with an in-situ activation strategy using an external energy field (ultrasound or light), can achieve highly selective and efficient flotation of target minerals in various mineral systems. The successful application of different compositions and activation methods highlights the universality and superiority of this technical route, providing strong technical support for improving the comprehensive utilization of complex and difficult-to-process ores. These excellent flotation indicators are attributed to the stable and highly effective hydrophobic layer formed in situ on the mineral surface by the reagents. This is highly consistent with the core idea of ​​the invention: to enhance mineral floatability by controlling the efficient linkage reaction of precursor molecules at the mineral-water interface through an energy field.

[0125] Although embodiments of the invention have been shown and described, it will be understood by those skilled in the art that various changes, modifications, substitutions and alterations can be made to these embodiments without departing from the principles and spirit of the invention, the scope of which is defined by the appended claims and their equivalents.

Claims

1. A highly efficient mineral selective flotation reagent, characterized in that, By weight, it includes: First precursor P1, 10 to 45 parts; the first precursor P1 comprises at least one anchoring functional group and at least one first reactive functional group, wherein the anchoring functional group is selected from at least one of mercapto, carboxyl, phosphate or hydroxamic acid groups, and the first reactive functional group is selected from one of terminal alkynyl or azide. Second precursor P2, 20 to 75 parts; the second precursor P2 comprises at least one second reactive functional group and at least one hydrophobic group, wherein the second reactive functional group is complementary to the first reactive functional group of the first precursor and is selected from one of azide or terminal alkynyl, and the hydrophobic group is selected from one of C8-C18 straight-chain or branched alkyl or C6-C12 perfluoroalkyl. Resonance activator, 0.1 to 5 parts; The resonance activation aid is used to enhance the selective energy absorption and / or activation degree of the first precursor on the surface of the target mineral under the action of a specific external physical energy field.

2. The high-efficiency mineral selective flotation reagent according to claim 1, characterized in that: The first precursor P1 is an organic compound whose molecular skeleton is selected from benzoic acid derivatives, naphthoic acid derivatives or sulfur-containing heterocyclic compounds, and is simultaneously connected with the above-mentioned anchoring functional group and the first reactive functional group, with a molecular weight range of 200 g / mol to 600 g / mol.

3. The high-efficiency mineral selective flotation reagent according to claim 1, characterized in that: The second precursor P2 is an organic compound, wherein the hydrophobic group is dodecyl, hexadecyl, or perfluorooctyl; the weight ratio of the first precursor P1 to the second precursor P2 is 1:0.5 to 3.

5.

4. The high-efficiency mineral selective flotation reagent according to claim 1, characterized in that: The resonance activator is selected from at least one of the following: anthraquinone compounds, tetraphenylporphyrin compounds, semiconductor quantum dots with an average particle size of 5 nm to 50 nm, or organic dye molecules with a specific light absorption peak.

5. The high-efficiency mineral selective flotation reagent according to claim 1, characterized in that: The flotation reagent also contains a reaction promoter comprising 0.01% to 1% of the total weight of the first precursor P1 and the second precursor P2, wherein the reaction promoter is a non-toxic copper complex or organic catalyst capable of catalyzing a click chemical reaction between the first reactive functional group and the second reactive functional group.

6. A method for selective flotation of minerals using the flotation reagent according to any one of claims 1 to 5, characterized in that, Includes the following steps: Step 1: Slurry preparation. Grind the ore to be processed to the target particle size and mix it with water to prepare a slurry with a concentration of 5% to 35% w / v. Step 2: pH adjustment and precursor P1 adsorption. Adjust the pH of the slurry to a preset range of 3.0 to 11.

0. Then add the first precursor P1 and the resonance activator to the slurry. The amount of the first precursor P1 added is 10 g / t ore to 200 g / t ore. Stir and adsorb for 2 to 20 minutes. Step 3: Selective energy resonance activation. A specific external physical energy field is applied to the slurry that has adsorbed the first precursor P1 and the resonance activation aid. The energy field is selected from ultrasound with a frequency of 20kHz to 3.0MHz or light with a wavelength of 254nm to 650nm. The irradiation is continued for 1 minute to 30 minutes to selectively activate the first precursor P1 adsorbed on the surface of the target mineral. Step 4: Adding the second precursor P2 and reacting in situ. The second precursor P2 is added to the slurry after being irradiated by the energy field. The molar ratio of the second precursor P2 to the first precursor P1 is 0.5:1 to 3.0:

1. Stirring is carried out at a temperature of 15°C to 45°C for 2 to 30 minutes, so that the activated first precursor P1 and the second precursor P2 react in situ on the surface of the target mineral to generate a hydrophobic product. Step 5: Flotation separation. Gas is introduced into the slurry treated in Step 4 to carry out flotation separation and collect the foam product of the target mineral.

7. The method for selective flotation of minerals according to claim 6, characterized in that, The external physical energy field mentioned in step three is applied in a pulse mode, wherein the pulse on-time of the ultrasonic pulse mode is 0.1 to 5 seconds, the pulse off-time is 0.5 to 10 seconds, and the duty cycle is 10% to 80%; or the pulse frequency of the light pulse mode is 0.1 Hz to 10 Hz.

8. The method for selective flotation of minerals according to claim 6, characterized in that, In step four, if the flotation reagent contains the reaction promoter, the reaction promoter is added to the slurry together with or in steps with the second precursor P2.

9. The method for selective flotation of minerals according to claim 6, characterized in that, Step 5, the flotation separation process, involves continuing to apply an external physical energy field with the same or reduced parameters as in Step 3 during the initial stage to enhance the formation and adhesion of hydrophobic products; in subsequent stages, the energy field application is stopped and / or a conventional amount of frother is added.

10. The method for selective flotation of minerals according to claim 6, characterized in that, After flotation separation, the tailings slurry is treated by adding a mild oxidant at a rate of 0.01% to 0.1% of the dry weight of the tailings and stirring at 20°C to 50°C for 10 to 60 minutes to promote the degradation of unreacted precursors or byproducts.

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