Compound, preparation method and application thereof, and flotation method

Abstract

The invention discloses a compound, a preparation method and application thereof and a flotation method. The chemical structure of the compound is shown as a formula (I), and R1 and R2 are respectively and independently selected from hydrogen, C1-C6 alkyl or C2-C6 monoolefine. The compound can serve as a collecting agent to be applied to flotation of rare earth minerals, and the flotation effect is good.

Description

Technical Field

[0001] The present invention relates to a compound, a preparation method and application thereof, and a flotation method, and in particular to a compound or its salt, a preparation method thereof, and application of the salt of the compound as a rare earth mineral flotation collector in iron tailings, and a flotation method for rare earth minerals. Background Art

[0002] In the flotation and beneficiation process of rare earth minerals, collectors are the core agents that determine the effectiveness of mineral enrichment and separation. Commonly used rare earth flotation collectors on the market include hydroxamic acid collectors (such as benzohydroxamic acid (BHA), salicylic hydroxamic acid (SHA), naphthohydroxamic acid (H205), and isohydroxamic acid (H316)) and compound collectors (such as L102, P8, and GYB compounds).

[0003] Hydroxamic acid collectors contain amide and oxime functional groups, which readily chelate with metal cations, resulting in good selectivity. However, the flotation performance of existing hydroxamic acid collectors still needs to be improved. Compound collectors significantly enhance the capture capacity of rare earth minerals through the synergistic effect of multiple components. However, compound collectors are mostly composed of organic compounds, which can improve flotation recovery and selectivity. However, they have the following problems: 1. The performance of the reagent is easily affected by changes in its own composition, ratio, and external process conditions, resulting in large fluctuations and insufficient stability during application. Minor adjustments to process parameters often have a significant impact on the flotation process, making it difficult to stably control the grade and recovery rate of the final concentrate at the ideal level. 2. Due to the constraints of the ratio and process parameters, the flotation effect is poor when processing complex ores, increasing the difficulty and uncertainty of operation. 3. The cumbersome synthesis process limits its widespread application and stable performance in actual industry.

[0004] Therefore, developing a single rare earth mineral flotation collector that is simple to synthesize, has stable medicinal properties and is easy to use, so as to improve the flotation efficiency and the overall mineral processing effect, is of great significance for optimizing the existing rare earth flotation reagents and improving the economic benefits of mineral processing production. Summary of the Invention

[0005] One object of the present invention is to provide a compound or a salt thereof. Another object of the present invention is to provide a method for preparing the compound or a salt thereof. Yet another object of the present invention is to provide a use of the compound or a salt thereof. The use of this collector in rare earth flotation can simultaneously improve the grade and recovery rate of rare earth flotation. Yet another object of the present invention is to provide a method for flotation of rare earth minerals.

[0006] The present invention adopts the following technical solutions to achieve the above technical objectives.

[0007] In one aspect, the present invention provides a compound or a salt thereof, wherein the chemical structure of the compound is shown in formula (I):

[0008]

[0009] In the formula, R1 and R2 are independently selected from hydrogen, a C1-C6 alkyl group, or a C2-C6 monoolefin. The inventors of this application believe that the synergistic effect of the allyl and methoxy groups in such collectors imparts dual hydrophobicity to the agent, improving the stability and floatability of the foam during flotation. The allyl group enhances the affinity of the collector with the mineral surface by increasing its hydrophobicity, thereby improving the selective adsorption capacity of the collector. The methoxy group enhances its surface activity by controlling surface tension, improving the formation and stability of foam during flotation, thereby increasing the grade and recovery rate during flotation.

[0010] R1 can be independently selected from hydrogen, C1-C6 alkyl or C2-C6 monoolefin; preferably selected from hydrogen, C1-C3 alkyl or C2-C3 monoolefin; more preferably hydrogen.

[0011] R2 can be independently selected from hydrogen, C1-C6 alkyl or C2-C6 monoolefin; preferably selected from hydrogen, C1-C3 alkyl or C2-C3 monoolefin; more preferably hydrogen.

[0012] In the present invention, examples of C1-C6 alkyl groups include, but are not limited to, methyl, ethyl, n-propyl, isopropyl, n-butyl, isobutyl, sec-butyl, tert-butyl, n-pentyl, tert-pentyl, neopentyl, n-hexyl, etc. Examples of C2-C6 monoolefins include, but are not limited to, ethenyl, propenyl, 1-butenyl, (E)-2-butenyl, isobutenyl, 1-pentenyl, (E)-2-pentenyl, 2-methyl-1-butenyl, 1-hexenyl, (E)-2-hexenyl, (E)-3-hexenyl, 2,3-dimethyl-1-butenyl, etc.

[0013] In the present invention, the structure of the compound is preferably the structure shown in the following formula:

[0014]

[0015] According to the compound or salt thereof of the present invention, preferably, the salt is a potassium salt, a sodium salt or an ammonium salt.

[0016] In the present invention, the salt of the compound preferably has a structure shown in the following formula:

[0017]

[0018] Wherein, M is a +1 valence alkaline cation, preferably, M is K + 、Na+ or NH4 + The above salt is obtained by saponifying the compound represented by formula (I).

[0019] On the other hand, the present invention also provides a method for preparing the above-mentioned compound or a salt thereof, comprising the following steps:

[0020] (1) subjecting the compound represented by formula (II) to a reflux reaction with hydroxylamine under alkaline conditions to obtain a hydroxamate, which is then acidified to obtain an acidified product;

[0021]

[0022] In formula (II), R1 and R2 are independently selected from hydrogen, C1-C6 alkyl or C2-C6 monoolefin, and R is independently selected from C1-C6 alkyl;

[0023] (2) extracting the acidified product with an organic solvent, combining the extracted organic phases, and distilling to obtain a distillation concentrate; adding water to the distillation concentrate to precipitate a solid, filtering the obtained solid, and drying to obtain the compound.

[0024] According to the preparation method of the present invention, preferably, R1 and R2 are independently selected from hydrogen, and R is independently selected from methyl or ethyl.

[0025] In formula (II), R1 can be independently selected from hydrogen, C1-C6 alkyl or C2-C6 monoolefin, preferably selected from hydrogen, C1-C3 alkyl or C2-C3 monoolefin, more preferably hydrogen.

[0026] R2 can be independently selected from hydrogen, C1-C6 alkyl or C2-C6 monoolefin, preferably selected from hydrogen, C1-C3 alkyl or C2-C3 monoolefin, more preferably hydrogen.

[0027] R can be independently selected from C1 to C6 alkyl groups, preferably selected from C1 to C3 alkyl groups, and more preferably methyl or ethyl groups.

[0028] In the present invention, examples of C1-C6 alkyl groups include, but are not limited to, methyl, ethyl, n-propyl, isopropyl, n-butyl, isobutyl, sec-butyl, tert-butyl, n-pentyl, tert-pentyl, neopentyl, n-hexyl, etc. Examples of C2-C6 monoolefins include, but are not limited to, ethenyl, propenyl, 1-butenyl, (E)-2-butenyl, isobutenyl, 1-pentenyl, (E)-2-pentenyl, 2-methyl-1-butenyl, 1-hexenyl, (E)-2-hexenyl, (E)-3-hexenyl, 2,3-dimethyl-1-butenyl, etc.

[0029] According to the preparation method of the present invention, preferably, the mass ratio of the compound represented by formula (II) to hydroxylamine is 1.5 to 3:1; and the solvent used in the reaction is C1 to C4 alkyl alcohol or water.

[0030] The solvent used in the reaction can be selected from a C1-C4 alkyl alcohol or water. In certain embodiments, the solvent used in the reaction is a C1-C2 alkyl alcohol. In other embodiments, the solvent used in the reaction is water. The present invention has discovered that using a C1-C4 alkyl alcohol as a solvent can significantly increase the yield of the compound represented by formula (I) compared to water.

[0031] In the present invention, examples of C1-C4 alkyl alcohols include, but are not limited to, methanol, ethanol, isopropanol, n-butanol, sec-butanol, tert-butanol, and the like.

[0032] According to the preparation method of the present invention, preferably, the mass ratio of the compound represented by formula (II) to hydroxylamine is 2 to 2.3:1; and the solvent used in the reaction is C1 to C4 alkyl alcohol.

[0033] The preparation method according to the present invention preferably comprises the following specific steps:

[0034] (1) Add the compound represented by formula (II) to a reaction vessel, then add hydroxylamine hydrochloride and anhydrous methanol, and fully dissolve the reaction materials under stirring; after complete dissolution, adjust the pH value to 11-13 with a sodium hydroxide aqueous solution; then raise the temperature to 35-50°C and reflux the reaction at this temperature for 3-8 hours to obtain a hydroxamate; after the reaction is completed, reduce the temperature to 20-30°C, and adjust the pH value to 4-7 with hydrochloric acid to obtain an acidified product;

[0035]

[0036] In formula (II), R1 and R2 are independently selected from hydrogen, C1-C6 alkyl or C2-C6 monoolefin, and R is independently selected from C1-C6 alkyl;

[0037] (2) extracting the acidified product multiple times with ethyl acetate, combining the extracted organic phases, and distilling to obtain a distillation concentrate; adding 0-10°C water to the distillation concentrate to precipitate a solid, filtering the obtained solid, and drying it at 35-55°C for 6-18 hours to obtain the compound.

[0038] In the present invention, the ratio of the volume of the solvent used in the reaction to the mass of the compound represented by formula (II) is 5 to 50 mL / g, preferably 10 to 40 mL / g, and more preferably 15 to 35 mL / g.

[0039] In the present invention, an alkaline reagent is used to adjust the pH value of the reaction system to 8 to 13.5, preferably, the pH value is 11 to 13, and more preferably, the pH value is 12.5.

[0040] In the present invention, the alkaline reagent is an alkaline reagent, specifically sodium carbonate, potassium carbonate, sodium bicarbonate, potassium bicarbonate, sodium hydroxide, potassium hydroxide or ammonia. Preferably, the alkaline reagent is sodium hydroxide, potassium hydroxide or ammonia. The hydroxide anions contained in such alkaline reagents can activate reactants and stabilize reaction intermediates.

[0041] According to a specific embodiment of the present invention, the alkaline reagent is a sodium hydroxide solution, and the concentration of the sodium hydroxide solution can be 0.5 to 14.3 mol / L, preferably 1 to 10 mol / L, and more preferably 2.5 to 5 mol / L.

[0042] In the present invention, preferably, the compound represented by formula (II) is reacted with hydroxylamine under alkaline conditions for 2 to 12 hours, preferably 3 to 8 hours, more preferably 6 to 7 hours, at a reaction temperature of 20 to 50°C, preferably 35 to 50°C.

[0043] In the present invention, the pH value of the system after acidification can be 4 to 7, preferably, the pH value of the system is 5 to 6.8, and more preferably, the pH value of the system is 5.5 to 6.5. The acidifying agent can be selected from at least one of hydrochloric acid, sulfuric acid, nitric acid, boric acid, oxalic acid, acetic acid, and citric acid. Preferably, the acidifying agent is hydrochloric acid, sulfuric acid, or nitric acid. More preferably, the acidifying agent is 10 to 15 wt% hydrochloric acid.

[0044] In the present invention, the organic solvent can be selected from at least one of ethyl acetate, dichloromethane and chloroform, preferably ethyl acetate. The extraction times are two or more, for example, three times. This ensures that the product is completely transferred to the organic phase, avoiding product loss.

[0045] In the present invention, it is preferred to use a vacuum distillation apparatus to distill the combined organic phase. During distillation, the water bath temperature can be 40-60°C, preferably 45-50°C. When the product is about to precipitate, the vacuum distillation is stopped and water at 0-10°C (preferably cold water at 10°C) is immediately added to the distillation concentrate to precipitate the solid. Then, suction filtration is performed and rinsed with ultrapure water. The filter cake is placed in an oven for drying, and the oven temperature can be controlled to be 40-60°C, preferably 45-50°C.

[0046] In another aspect, the present invention further provides the use of the above-described compound or its salt as a collector in the flotation of rare earth minerals. Preferably, the compound or its salt is used for the flotation of rare earth minerals in iron ore tailings, and more preferably, for the flotation of rare earth minerals in the Bayan Obo iron ore tailings obtained from Baotou Steel. The specific application steps can be referred to the rare earth mineral flotation method described below.

[0047] In another aspect, the present invention further provides a flotation method for rare earth minerals, comprising the following steps:

[0048] (a) forming a slurry from the mineral to be processed;

[0049] (b) adding water glass, the salt as described above as a rare earth mineral flotation collector, a pH regulator and a frother to the ore pulp in sequence, and performing roughing to obtain a roughing concentrate and a roughing tailing;

[0050] (c) forming a rougher concentrate slurry from the rougher concentrate, and performing two or more rounds of ...

[0051] In step (a), the mineral to be selected may be iron ore tailings. The mineral to be selected is mixed with water to form a slurry having a concentration of 30 to 60 wt%, preferably 40 to 60 wt%, and more preferably 50 to 60 wt%. Before adding water glass to the slurry, the slurry is heated to 40 to 70° C., preferably 50 to 70° C., and more preferably 60 to 70° C.

[0052] In step (b), the amount of water glass used can be 1 to 4 kg / t, preferably 2.5 to 3.5 kg / t, based on the total mass of the mineral to be selected. The amount of the salt of the compound of the present invention used can be 0.001 to 1.5 kg / t, preferably 0.08 to 0.8 kg / t, based on the total mass of the mineral to be selected. The amount of the foaming agent used is 0.1 to 0.2 kg / t, preferably 0.10 to 0.15 kg / t, based on the total mass of the mineral to be selected.

[0053] In step (b), the pH value of the slurry after adding the salt of the compound of the present invention may be 7 to 10, preferably 8.5 to 10, and more preferably 9.1 or 9.2.

[0054] In the present invention, the pH adjuster can be sodium carbonate, potassium hydroxide, calcium oxide, sodium hydroxide, sulfuric acid, or hydrochloric acid, preferably sodium carbonate, potassium hydroxide, or hydrochloric acid, and more preferably sodium carbonate.

[0055] In the present invention, the foaming agent can be pine oil (2# oil), methyl isobutyl carbinol (MIBC), triethoxybutane (TEB), polyethylene glycol ether (such as Dowfroth TMseries), diethyl phthalate (DEP), fusel oil, 2-ethylhexanol, camphor oil, alkoxylated polypropylene glycol (APG), silicone polyether (such as SAG TM series) and Aerofroth TM Preferably, the foaming agent is 2# oil.

[0056] In the present invention, the slurry is heated under stirring for 2 to 4 minutes, then water glass is added and stirred for 2 to 4 minutes, then the salt of the compound of the present invention is added and stirred for 2 to 4 minutes, the pH value of the slurry is adjusted with a pH regulator, the reaction is carried out for 2 to 4 minutes, and then a foaming agent is added and stirred for 2 to 4 minutes.

[0057] In step (c), according to one embodiment of the present invention, the rougher concentrate and water are heated under stirring conditions to form a rougher concentrate slurry. The slurry temperature is 40-70°C, preferably 50-70°C, and more preferably 60-70°C. The number of rougher concentrate passes is preferably two to five, more preferably three to five, for example, three times.

[0058] According to the flotation method of the present invention, preferably, step (c) includes the following process:

[0059] (1) forming a rougher concentrate slurry from the rougher concentrate, sequentially adding water glass, the salt described above, a pH regulator, and a frother to the rougher concentrate slurry, and performing a primary concentration to obtain a primary concentrated concentrate and a primary concentrated tailings;

[0060] (2) forming a primary concentrated concentrate slurry from the primary concentrated concentrate, sequentially adding water glass, the salt described above, a pH regulator, and a frother to the primary concentrated concentrate slurry, and performing secondary concentrating to obtain a secondary concentrated concentrate and a secondary concentrated tailings;

[0061] (3) The secondary concentrated concentrate is formed into a secondary concentrated concentrate slurry, water glass, the salt as described above, a pH regulator and a frother are sequentially added to the secondary concentrated concentrate slurry, and three concentrations are performed to obtain a final concentrated concentrate and a final concentrated tailings.

[0062] In step (1), the roughing concentrate slurry is first heated to 40-70° C., preferably 50-70° C., and more preferably 60-70° C. The pH value of the slurry formed after adding the salt of the compound of the present invention can be 7-10, preferably 8.5-10, and more preferably 9.1 or 9.2. Based on the total mass of the mineral to be selected, the amount of water glass can be 0.5-2 kg / t, preferably 0.53-1 kg / t. Based on the total mass of the mineral to be selected, the amount of the salt of the compound of the present invention can be 0.001-1.5 kg / t, preferably 0.16-0.75 kg / t. Based on the total mass of the mineral to be selected, the amount of the foaming agent can be 0.01-0.2 kg / t, preferably 0.02-0.15 kg / t.

[0063] In step (2), the primary concentrated ore pulp is heated to 40-70° C., preferably 50-70° C., more preferably 60-70° C. The pH value of the pulp after adding the salt of the compound of the present invention can be 7-10, preferably 8.5-10, more preferably 9.1 or 9.2. Based on the total mass of the mineral to be selected, the amount of water glass can be 0.25-1 kg / t, preferably 0.26-0.8 kg / t. Based on the total mass of the mineral to be selected, the amount of the salt of the compound of the present invention can be 0.001-1.5 kg / t, preferably 0.13-0.7 kg / t. Based on the total mass of the mineral to be selected, the amount of the foaming agent can be 0.01-0.1 kg / t, preferably 0.04-0.1 kg / t.

[0064] In step (3), the secondary concentrated ore slurry is heated to 40-70° C., preferably 50-70° C., more preferably 60-70° C. The pH value of the slurry after adding the salt of the compound of the present invention can be 7-10, preferably 8.5-10, more preferably 9.1 or 9.2. Based on the total mass of the mineral to be selected, the amount of water glass can be 0.25-1 kg / t, preferably 0.26-0.7 kg / t. Based on the total mass of the mineral to be selected, the amount of the salt of the compound of the present invention can be 0.001-1.5 kg / t, preferably 0.08-0.5 kg / t. Based on the total mass of the mineral to be selected, the amount of the foaming agent can be 0.01-0.1 g / t, preferably 0.02-0.05 kg / t.

[0065] Compared to existing single-component collectors, the salts of the compounds of the present invention, when used as rare earth mineral flotation collectors, can significantly improve REO recovery in roughing and significantly increase the grade of the resulting rare earth concentrate in fine separation, resulting in significantly better overall flotation performance. Furthermore, compared to existing composite collectors, the salts of the compounds of the present invention, when used as rare earth mineral flotation collectors, can simplify the process flow and complex operations during use, reducing production costs; avoid interference from composite components, ensure reagent stability; and facilitate storage and management, reducing operational risks and maintenance costs. BRIEF DESCRIPTION OF THE DRAWINGS

[0066] Figure 1 This is the infrared spectrum (FT-IR) of the product 5-allyl-3-methoxysalicylic acid collector of Example 1.

[0067] Figure 2 This is the hydrogen nuclear magnetic resonance spectrum of the product 5-allyl-3-methoxysalicylic acid collector of Example 1. DETAILED DESCRIPTION

[0068] The present invention will be further described below with reference to specific embodiments, but the protection scope of the present invention is not limited thereto.

[0069] Other implementation methods based on routine experiments and operations performed by technicians in this field without creative work should also be considered as part of the present invention and included in its scope of protection.

[0070] Unless otherwise specified, all raw materials, reagents, instruments, and equipment involved in the present invention can be purchased through commercial channels or prepared according to existing technical methods. The embodiments of the present invention are merely exemplary descriptions and should not be construed as restrictive requirements. Any equivalent alternatives or technical solutions should comply with the basic concept of the present invention and fall within the scope of protection of the claims.

[0071] <Ingredients>

[0072] The sodium carbonate solution used in the following examples has a concentration of 12 wt%.

[0073] Foaming agent 2# oil: pine oil.

[0074] <Term Explanation>

[0075] Grade: The total mass of all rare earth oxides (REO) in an ore is expressed as a percentage of the total mass of the ore.

[0076] Example 1 Preparation of 5-allyl-3-methoxysalicylic acid

[0077] 8.0 g of 98% pure methyl 5-allyl-3-methoxysalicylate was added to a reaction flask, followed by 4.0 g of hydroxylamine hydrochloride and 120 mL of anhydrous methanol. The reaction mixture was stirred until fully dissolved. Once completely dissolved, the pH of the solution was adjusted to 12.5 with 2.5 mol / L sodium hydroxide solution. The temperature was then raised to 50°C and the reaction was refluxed at this temperature for 6 hours.

[0078] After the reaction is completed, it is cooled to room temperature and the pH value of the reaction system is adjusted to 6.5 with 15wt% hydrochloric acid. Then, ethyl acetate is used as an extractant for extraction, and three extractions are performed (the amount of ethyl acetate used for each extraction is 100mL). The organic phases are combined after extraction. The combined organic phases are distilled under reduced pressure to remove the solvent. After the distillation is completed, 10°C cold water is added to the distillation concentrate to precipitate a solid product, which is filtered with suction to obtain a filter cake. The filter cake is dried in an oven at 45°C for 12h to obtain 5-allyl-3-methoxysalicylic acid. It is a yellow powder with a yield of 80%.

[0079] The infrared spectrum and nuclear magnetic resonance spectrum of the obtained product 5-allyl-3-methoxy salicylic acid are shown in Figure 1 and Figure 2 .

[0080] Example 2

[0081] Add 4.0 g of 98% pure methyl 5-allyl-3-methoxysalicylate to a reaction flask, followed by 1.5 g of hydroxylamine hydrochloride and 140 mL of ultrapure water. Stir and heat in a water bath to 50°C to fully dissolve the reaction materials. Once completely dissolved, adjust the pH of the solution to 12.5 with 2.5 mol / L sodium hydroxide solution and reflux at 50°C with stirring for 6 h.

[0082] After the reaction is completed, the temperature is lowered to room temperature and the pH value of the system is adjusted to 6.5 with 15wt% hydrochloric acid. Then, ethyl acetate is used as an extraction agent and extracted three times (the amount of ethyl acetate used during each extraction is 100mL). The organic phases are combined after extraction. The combined organic phases are desolvated by reduced pressure distillation. After the distillation is complete, 10°C cold water is added to the distillation concentrate to separate out a solid product, which is filtered with suction to obtain a filter cake. The filter cake is dried in a 45°C baking oven for 12h to obtain 5-allyl-3-methoxy salicylic acid as a yellow powder with a yield of 65%.

[0083] Application Example 1

[0084] The 5-allyl-3-methoxysalicylic hydroxamic acid obtained in Example 1 was added to water, and then saponified with sodium hydroxide to obtain a collector.

[0085] The raw ore was selected from the Bayan Obo iron ore tailings. The raw ore contained 9.08wt% REO, 27.00wt% CaO, 17.12wt% SiO2, and 12.77wt% F. Minerals with a mesh size below 200 mesh accounted for 81.68wt%.

[0086] The flotation experiment was carried out using a 500 mL single-cell flotation machine with an aeration rate of 0.1 L / min.

[0087] 0.375 kg of iron ore tailings were mixed with 300 mL of tap water to form a slurry (the slurry concentration was about 55.56 wt%). Under the stirring speed of 1980 r / min, the mixture was heated to 60°C and kept at a constant temperature for 4 minutes. Industrial water glass (the concentration of water glass was 9 wt%, and the amount of water glass added was 3.2 kg / t) was added to the slurry and reacted for 3 minutes. Then, the above-mentioned flotation collector (the amount of collector added was 0.8 kg / t) was added and reacted for 1 minute. The pH value of the slurry was adjusted to 9.1 using sodium carbonate solution, and then the reaction was continued for 3 minutes. Then, the foaming agent 2# oil was added (the amount of foaming agent added was 0.14 kg / t). The reaction was maintained for 3 minutes to ensure that the foaming agent was completely dissolved in the slurry and successfully reacted with the collector. Then, ventilation was started to start the roughing step to obtain roughing concentrate and roughing tailings respectively.

[0088] The REO grade of the obtained rougher concentrate was 21.44 wt%, the REO grade of the rougher tailings was 1.60 wt%, and the REO recovery rate was 91.88 wt%.

[0089] The rougher concentrate is then further refined, and the refinery experimental steps are consistent with the rougher experiment. The specific steps are as follows:

[0090] Primary concentration: The obtained roughing concentrate is subjected to primary concentration. The roughing concentrate slurry is heated to 60°C at a stirring speed of 1980r / min and kept at a constant temperature for 4 minutes. Water glass with a concentration of 9wt% is added to the slurry (the amount of water glass added is 0.53kg / t) and reacted for 3 minutes. Then, the above-mentioned flotation collector is added (the amount of collector added is 0.16kg / t) and reacted for 1 minute. The pH value of the slurry is adjusted to 9.1 using sodium carbonate solution and the reaction is continued for 3 minutes. Then, the foaming agent 2# oil is added (the amount of foaming agent added is 0.035kg / t), the reaction is continued for 3 minutes, and then aeration is performed for flotation. After flotation is completed, a primary concentrated concentrate and a primary concentrated tailings are obtained.

[0091] Secondary concentration: further concentrate the primary concentrate. Heat the primary concentrate slurry to 60°C at a stirring speed of 1980r / min and keep the temperature constant for 4 minutes. Add water glass with a concentration of 9wt% to the slurry (the amount of water glass added is 0.265kg / t) and react for 3 minutes. Then add the above-mentioned flotation collector (the amount of collector added is 0.134kg / t) and react for 1 minute. Use sodium carbonate solution to adjust the pH value of the slurry to 9.1 and continue to react for 3 minutes. Then, add the foaming agent 2# oil (the amount of foaming agent added is 0.07kg / t), continue to react for 3 minutes, and then ventilate for flotation. After flotation is completed, secondary concentrated concentrate and secondary concentrated tailings are obtained.

[0092] Tertiary concentration: Continue to concentrate the secondary concentrated concentrate. Heat the secondary concentrated concentrate slurry to 60°C at a stirring speed of 1980r / min and keep the temperature constant for 4 minutes. Add water glass with a concentration of 9wt% to the slurry (the amount of water glass added is 0.265kg / t) and react for 3 minutes. Then add the above-mentioned flotation collector (the amount of collector added is 0.08kg / t) and react for 1 minute. Use sodium carbonate solution to adjust the pH value of the slurry to 9.1 and continue to react for 3 minutes. Then, add the foaming agent 2# oil (the amount of foaming agent added is 0.035kg / t), continue to react for 3 minutes, and then ventilate for flotation. After flotation is completed, the final concentrated concentrate and the final concentrated tailings are obtained. The final concentrated concentrate is the final rare earth concentrate.

[0093] Comparative Application Example 1

[0094] Benzohydroxamic acid was used to replace the collector in Application Example 1. Except for the following process parameters, the remaining steps and conditions were the same as those in Application Example 1:

[0095] In the roughing step, the dosage of benzohydroxamic acid collector is 0.8 kg / t, the dosage of water glass is 3 kg / t, the pH value of flotation pulp is 9.2, and the dosage of frother 2# oil is 0.14 kg / t.

[0096] Table 1. Comparison of collector flotation effects on rare earth ores

[0097]

[0098] Note: Grade and recovery rate refer to the grade and recovery rate of REO.

[0099] In summary, under the same beneficiation conditions, the 5-allyl-3-methoxysalicylic acid of the present invention improves both the grade and recovery of the selected ore during the roughing process, compared to benzohydroxamic acid, and also increases the final concentrate grade during the beneficiation process. Therefore, the compound of the present invention is more effective in the flotation of rare earth minerals.

[0100] The present invention is not limited to the above-mentioned embodiments. Any modification, improvement, or substitution that can be conceived by those skilled in the art without departing from the essential content of the present invention shall fall within the scope of the present invention.

Claims

1. A compound or a salt thereof, characterized in that: The chemical structure of the compound is shown in formula (I): In formula (I), R1 and R2 are each independently selected from hydrogen, a C1-C6 alkyl group or a C2-C6 monoolefin.

2. The compound or salt thereof according to claim 1, characterized in that The salt is potassium salt, sodium salt or ammonium salt.

3. The method for preparing the compound or its salt according to claim 1, characterized in that The steps include: (1) subjecting the compound represented by formula (II) to a reflux reaction with hydroxylamine under alkaline conditions to obtain a hydroxamate, which is then acidified to obtain an acidified product; In formula (II), R1 and R2 are independently selected from hydrogen, C1-C6 alkyl or C2-C6 monoolefin, and R is independently selected from C1-C6 alkyl; (2) extracting the acidified product with an organic solvent, combining the extracted organic phases, and distilling to obtain a distillation concentrate; adding water to the distillation concentrate to precipitate a solid, filtering the obtained solid, and drying to obtain the compound.

4. The preparation method according to claim 3, characterized in that R1 and R2 are each independently selected from hydrogen, and R is methyl or ethyl.

5. The preparation method according to claim 3, characterized in that The mass ratio of the compound represented by formula (II) to hydroxylamine is 1.5 to 3:1; the solvent used in the reaction is C1 to C4 alkyl alcohol or water.

6. The preparation method according to claim 5, characterized in that The mass ratio of the compound represented by formula (II) to hydroxylamine is 2-2.3:1; and the solvent used in the reaction is C1-C4 alkyl alcohol.

7. The preparation method according to claim 6, characterized in that The specific steps include: (1) Add the compound represented by formula (II) to a reaction vessel, then add hydroxylamine hydrochloride and anhydrous methanol, and fully dissolve the reaction materials under stirring; after complete dissolution, adjust the pH value to 11-13 with a sodium hydroxide aqueous solution; then raise the temperature to 35-50°C and reflux the reaction at this temperature for 3-8 hours to obtain a hydroxamate; after the reaction is completed, reduce the temperature to 20-30°C, and adjust the pH value to 4-7 with hydrochloric acid to obtain an acidified product; In formula (II), R1 and R2 are independently selected from hydrogen, C1-C6 alkyl or C2-C6 monoolefin, and R is independently selected from C1-C6 alkyl; (2) extracting the acidified product multiple times with ethyl acetate, combining the extracted organic phases, and distilling to obtain a distillation concentrate; adding 0-10°C water to the distillation concentrate to precipitate a solid, filtering the obtained solid, and drying it at 35-55°C for 6-18 hours to obtain the compound.

8. Use of the compound according to claim 1 or a salt thereof as a collector in the flotation of rare earth minerals.

9. A flotation method for rare earth minerals, characterized in that: The following steps are involved: (a) forming a slurry from the mineral to be processed; (b) adding water glass, the salt according to claim 2, a pH regulator and a foaming agent to the ore pulp in sequence, and performing roughing to obtain a roughing concentrate and a roughing tailing; (c) forming a rougher concentrate slurry from the rougher concentrate, and performing two or more rounds of ...

10. The flotation method according to claim 9, characterized in that: Step (c) includes the following process: (1) forming a rougher concentrate slurry from a rougher concentrate, sequentially adding water glass, the salt according to claim 2, a pH regulator, and a foaming agent to the rougher concentrate slurry, and performing a primary concentration to obtain a primary concentrated concentrate and a primary concentrated tailings; (2) forming a primary concentrated concentrate slurry from the primary concentrated concentrate, sequentially adding water glass, the salt according to claim 2, a pH regulator, and a foaming agent to the primary concentrated concentrate slurry, and performing secondary concentrating to obtain a secondary concentrated concentrate and a secondary concentrated tailings; (3) The secondary concentrated concentrate is formed into a secondary concentrated concentrate slurry, water glass, the salt according to claim 2, a pH regulator and a foaming agent are added to the secondary concentrated concentrate slurry in sequence, and three concentrations are performed to obtain a final concentrated concentrate and a final concentrated tailings.

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