CN108940563B - Carbon silicon collophanite cation reverse flotation desilication method based on foam regulation - Google Patents

Carbon silicon collophanite cation reverse flotation desilication method based on foam regulation Download PDF

Info

Publication number
CN108940563B
CN108940563B CN201810480343.0A CN201810480343A CN108940563B CN 108940563 B CN108940563 B CN 108940563B CN 201810480343 A CN201810480343 A CN 201810480343A CN 108940563 B CN108940563 B CN 108940563B
Authority
CN
China
Prior art keywords
flotation
foam
collophanite
pulp
ore
Prior art date
Legal status (The legal status is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the status listed.)
Active
Application number
CN201810480343.0A
Other languages
Chinese (zh)
Other versions
CN108940563A (en
Inventor
余俊
甘顺鹏
季荣
胡勇
蒋世鹏
谢超
郑贤福
Current Assignee (The listed assignees may be inaccurate. Google has not performed a legal analysis and makes no representation or warranty as to the accuracy of the list.)
China Bluestar Chonfar Engineering and Technology Co Ltd
Original Assignee
China Bluestar Chonfar Engineering and Technology Co Ltd
Priority date (The priority date is an assumption and is not a legal conclusion. Google has not performed a legal analysis and makes no representation as to the accuracy of the date listed.)
Filing date
Publication date
Application filed by China Bluestar Chonfar Engineering and Technology Co Ltd filed Critical China Bluestar Chonfar Engineering and Technology Co Ltd
Priority to CN201810480343.0A priority Critical patent/CN108940563B/en
Publication of CN108940563A publication Critical patent/CN108940563A/en
Application granted granted Critical
Publication of CN108940563B publication Critical patent/CN108940563B/en
Active legal-status Critical Current
Anticipated expiration legal-status Critical

Links

Classifications

    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03B—SEPARATING SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS
    • B03B7/00—Combinations of wet processes or apparatus with other processes or apparatus, e.g. for dressing ores or garbage
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B01—PHYSICAL OR CHEMICAL PROCESSES OR APPARATUS IN GENERAL
    • B01D—SEPARATION
    • B01D19/00—Degasification of liquids
    • B01D19/02—Foam dispersion or prevention
    • B01D19/04—Foam dispersion or prevention by addition of chemical substances
    • B01D19/0404—Foam dispersion or prevention by addition of chemical substances characterised by the nature of the chemical substance
    • B01D19/0418—Foam dispersion or prevention by addition of chemical substances characterised by the nature of the chemical substance compounds containing P-atoms
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03B—SEPARATING SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS
    • B03B9/00—General arrangement of separating plant, e.g. flow sheets
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D1/00—Flotation
    • B03D1/001—Flotation agents
    • B03D1/002—Inorganic compounds
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D1/00—Flotation
    • B03D1/001—Flotation agents
    • B03D1/004—Organic compounds
    • B03D1/008—Organic compounds containing oxygen
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D1/00—Flotation
    • B03D1/001—Flotation agents
    • B03D1/004—Organic compounds
    • B03D1/01—Organic compounds containing nitrogen
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D2201/00—Specified effects produced by the flotation agents
    • B03D2201/007—Modifying reagents for adjusting pH or conductivity
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D2201/00—Specified effects produced by the flotation agents
    • B03D2201/02—Collectors
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D2201/00—Specified effects produced by the flotation agents
    • B03D2201/04—Frothers
    • B—PERFORMING OPERATIONS; TRANSPORTING
    • B03—SEPARATION OF SOLID MATERIALS USING LIQUIDS OR USING PNEUMATIC TABLES OR JIGS; MAGNETIC OR ELECTROSTATIC SEPARATION OF SOLID MATERIALS FROM SOLID MATERIALS OR FLUIDS; SEPARATION BY HIGH-VOLTAGE ELECTRIC FIELDS
    • B03D—FLOTATION; DIFFERENTIAL SEDIMENTATION
    • B03D2203/00—Specified materials treated by the flotation agents; specified applications
    • B03D2203/02—Ores
    • B03D2203/04—Non-sulfide ores
    • B03D2203/06—Phosphate ores

Abstract

The carbon silicon collophanite cation reverse flotation desilication method based on foam regulation comprises the following steps: (1) decarburization and flotation: wet grinding the carbon-silicon collophanite raw ore, adding water for size mixing, adding a foaming agent, stirring, performing air flotation, and throwing the flotation foam as tailings to obtain decarburization flotation ore pulp; (2) desliming and flotation: firstly, carrying out cyclone classification desliming, taking an overflow part as tailings to discard tails, adding a collecting agent into bottom flow, stirring, carrying out air flotation, and taking flotation foam as tailings to discard tails to obtain desliming flotation pulp; (3) desiliconization reverse flotation: and adjusting the pH value, adding a collecting agent, stirring, adding a defoaming agent, stirring, performing air flotation, scavenging flotation foam, scavenging tailings and discarding tails to obtain the phosphate concentrate pulp. The method effectively reduces the viscosity and stability of the foam, increases the brittleness of the foam, has the defoaming efficiency of more than or equal to 81 percent, and has the foam dynamic stability coefficient of less than or equal to 0.25 min; the phosphate concentrate obtained by defoaming by the method has high grade and less impurities.

Description

Carbon silicon collophanite cation reverse flotation desilication method based on foam regulation
Technical Field
The invention relates to a carbon-silicon collophanite cation reverse flotation desilication method, in particular to a carbon-silicon collophanite cation reverse flotation desilication method based on foam regulation and control.
Background
The phosphorite resources in China are abundant but not rich, and more than 90 percent of the phosphorite reserves are middle-low grade ores, P2O5The average grade is only 17 percent, wherein collophanite is taken as the main material and accounts for about 80 percent of the total reserves of the whole country, and the ore particles are fine, the embedding is compact and the separation difficulty is high. In addition, the phosphorus ore is divided into siliceous, calcareous and silico-calcic phosphorus ores according to the types and contents of gangue minerals contained in the phosphorus ores, wherein the silico-calcic phosphorus ores account for more than 85 percent of the total reserves of the national phosphorus ores. At present, high-magnesium calcium phosphorite is developed and utilized in large quantity due to mature mineral separation technology and low cost, but the requirement of economic growth on phosphorite resources is difficult to meet due to limited resource reserves. Therefore, the development and utilization of medium-low grade siliceous collophanite (such as carbon-silicon collophanite) are very important.
The flotation method is the preferred scheme for separating collophanite and silicate gangue minerals, and comprises an anion direct flotation desilication process and a cation reverse flotation desilication process. Because the anion collecting agent has low activity under the normal temperature condition, the process needs heating and has large energy consumption. In addition, the process has the disadvantages of complex medicament system, large consumption and non-ideal economic benefit. The cationic collector has strong low-temperature resistance, so that the energy consumption is saved; the cation collecting agent has high adsorption speed on the surface of the silicate mineral, and the process is simplified. Therefore, the cation reverse flotation desilication process has better development potential than the anion forward flotation process. However, when the cation reverse flotation desilication process is used for treating medium and low grade siliceous collophanite, the cation collecting agent is sensitive to micro-fine particle minerals, so that a large amount of over-stable foam accumulation is easy to occur, defoaming is difficult to realize, the flotation effect and subsequent operation are seriously influenced, the grade of the concentrate is reduced, and the impurity content of the concentrate is increased.
In the current collophanite flotation process, desilication synergist and defoamer are common means for improving cation reverse flotation desilication foam. The research shows that: kerosene, octanol, industrial fusel Zc and 800# (petroleum by-product, main component is polyalcohol) can effectively improve the foaming property and selectivity of aliphatic amine and ether amine collecting agent. Although our country has made a certain progress in research on improving the cation flotation froth, in the flotation industry, the problems of large flotation three-phase froth amount, stickiness, difficult dispersion and the like are still obvious, and the application of the cation reverse flotation process in the phosphorite beneficiation in our country is restricted.
CN 103285625A discloses a defoaming method for phosphorite reverse flotation desilication, which is to hang a spray pipe on a reverse flotation desilication flotation foam tank and spray an aqueous solution containing a defoaming agent on the foam to realize the defoaming purpose. However, the defoamer tributyl phosphate used in the method is difficult to dissolve in water, so that the method is difficult to realize; meanwhile, the volatilization of the defoaming agent in the spraying process can also influence the production air environment; in addition, for collophanite with high fine particle content, the purpose of defoaming is difficult to realize only by adopting a defoaming agent.
CN 105013617A discloses a mineral processing technique of collophanite, which utilizes a flotation column to float phosphate ore to remove MgO and Al in the phosphate ore2O3And Fe2O3. However, the process is a phosphorite direct flotation desilication process, so that the chemical system is complex, the chemical consumption is large, and the cost is high.
CN 105750090A discloses a method for sorting silico-calcium collophanite, which comprises the steps of firstly, preselecting ores by a cyclone, and respectively carrying out anion positive flotation desilicication and anion reverse flotation demagnetisation on the preselected ores. Although the method can obtain high-quality phosphate concentrate, the process is long and the production investment cost is high; meanwhile, the desiliconization process also adopts anion positive flotation, the medicament system is complex, and the medicament cost is higher.
Therefore, the development of a method for reducing the stability of the desiliconized foam of the collophanite cation reverse flotation and improving the brittleness and the fluidity of the foam is urgently needed.
Disclosure of Invention
The invention aims to solve the technical problem of overcoming the defects in the prior art and provide the foam regulation-based carbon silicon collophanite cation reverse flotation desilication method which can effectively reduce the viscosity and stability of the carbon silicon collophanite cation reverse flotation desilication foam, increase the foam brittleness, enable the flotation process flow to operate stably and efficiently, and enable the obtained phosphate concentrate after defoaming to have high grade and less impurities.
The technical scheme adopted by the invention for solving the technical problems is as follows: the carbon silicon collophanite cation reverse flotation desilication method based on foam regulation comprises the following steps:
(1) decarburization and flotation: wet grinding the carbon-silicon collophanite raw ore, adding water for size mixing, adding a foaming agent into the obtained ore pulp, stirring, performing air flotation, and discarding the flotation foam as tailings to obtain decarburization flotation ore pulp;
(2) desliming and flotation: carrying out cyclone classification desliming on the decarburized ore pulp obtained in the step (1), using an overflow part as tailings for tailing discarding, adding a collecting agent into bottom flow, stirring, carrying out air flotation, and using flotation foam as tailings for tailing discarding to obtain desliming flotation ore pulp;
(3) desiliconization reverse flotation: and (3) adjusting the pH value of the desliming flotation pulp obtained in the step (2), adding a collecting agent, stirring, adding a defoaming agent, stirring, performing air flotation, scavenging flotation foam, and discarding tailings of the flotation foam to obtain the phosphate concentrate pulp.
Preferably, in step (1), P is in the carbon-silicon collophanite2O518-25% by mass, 1-8% by mass of carbon, 22-32% by mass of CaO, and SiO2The mass percentage of the component (A) is 20-35%.
Preferably, in the step (1), the amount of water used in the wet grinding is 0.5-0.8 times of the mass of the carbon-silicon collophanite.
Preferably, in the step (1), the ore is ground by a wet method until the mass of the ore under the screen of 0.074mm is less than or equal to 85 percent of the mass of the total raw ore. Grinding to the particle size is more conducive to dissociating the valuable mineral from the gangue mineral monomers.
Preferably, in the step (1), water is added for size mixing until the mass concentration of the ore pulp is 20-35%.
Preferably, in the step (1), the amount of the foaming agent is 100-600 g of foaming agent per ton of carbon silicon collophanite raw ore (more preferably 200-400 g of foaming agent per ton of carbon silicon collophanite raw ore). The carbon in the ore has the effect of enhancing the stability of the foam, and the step uses a foaming agent to remove the carbon in the ore so as to reduce the stability of the foam; if the amount of the blowing agent is too small, the decarburization is not complete, and if the amount of the blowing agent is too large, the blowing agent remaining in the solution rather enhances the stability of the subsequent desiliconized foam.
Preferably, in the step (1), the foaming agent is one or more of terpineol, methyl isobutyl carbinol, diethyl phthalate and the like. The foaming agent has strong foaming capacity and is particularly suitable for carbon flotation.
Preferably, in the step (1), the stirring time is 1-3 min.
Preferably, in the step (2), the feeding speed of the cyclone classification desliming is 1-4 m/s, and the feeding mass concentration is 10-25%. 50-75% of minerals with-10 mu m (under 10 mu m sieve) size fraction can be removed through cyclone classification desliming.
Preferably, in the step (2), the amount of the collector is 50-200 g per ton of the carbon-silicon collophanite raw ore (more preferably 100-150 g per ton of the carbon-silicon collophanite raw ore). 15-25% of-10 mu m (under 10 mu m sieve) size fraction minerals can be removed by the desliming of the collecting agent.
In the step (2) of the method, the slime content in the ore can be reduced through desliming, so that the stability of foam can be reduced, and the using amount of the collecting agent can be reduced.
Preferably, in step (2), the collector is ether amine acetate and/or ether polyamine acetate. The collecting agent has strong collecting capacity, good foam fluidity and easy defoaming.
Preferably, in the step (2), the stirring time is 1-3 min.
Preferably, in the step (3), the pH value of the desliming flotation pulp is adjusted to 7.5-8.5. Within the pH value range, the collecting performance of the collecting agent is strongest, and foam is easier to defoam.
Preferably, in the step (3), the pH value regulator for regulating the pH value is sodium carbonate and/or sodium hydroxide.
Preferably, in the step (3), the amount of the collector is 100-500 g per ton of the carbon-silicon collophanite raw ore (more preferably 200-400 g per ton of the carbon-silicon collophanite raw ore). The step (3) of the method is used for removing silicate gangue minerals in ores, the step (1) of decarburization is carried out, and the step (2) of cyclone classification desliming and flotation desliming serve the step, so that the stability of flotation foam in the step (3) is reduced, the moderate foam stability of desiliconization operation is ensured, and the whole process can be stably operated.
Preferably, in step (3), the collector is ether amine acetate and/or ether polyamine acetate. The collecting agent has strong collecting capacity, good foam fluidity and easy defoaming.
Preferably, in the step (3), the stirring time is 1-3 min.
Preferably, in the step (3), the amount of the defoaming agent is 50-150 g defoaming agent/t carbon silicon collophanite raw ore. If the using amount of the defoaming agent is too small, the defoaming effect is not obvious, the process cannot be stably operated, and if the using amount of the defoaming agent is too large, the reverse flotation desilication is not easy to foam, so that the desilication operation is difficult to complete.
Preferably, in the step (3), the defoaming agent is tributyl phosphate and/or trioctyl phosphate. The addition of the defoaming agent can reduce the viscosity of the cation reverse flotation desilication foam, reduce the water content and improve the brittleness of the foam, thereby reducing the stability of the foam, reducing the mechanical entrainment rate of useful minerals and improving the concentrate grade and the recovery rate.
Preferably, in the step (3), the stirring time is 1-3 min.
Preferably, in the step (3), the scavenged middlings are returned to the step for desilication and reverse flotation.
The method has the following beneficial effects:
(1) according to the method, a decarbonization process is adopted to remove carbon with a stable foam effect in the ore, a desliming process eliminates the influence of fine-grained minerals on the foam stability, a collecting agent with good foam brittleness is adopted in the desiliconization process and is used in combination with a defoaming agent, so that combined defoaming is realized, the viscosity and stability of foam are effectively reduced, the foam brittleness is increased, the defoaming efficiency is more than or equal to 81%, the dynamic stability coefficient of the foam is less than or equal to 0.25min, the fluidity of a foam product is improved, defoaming is easy, a large amount of washing water is not needed, water resources are saved, the next flotation operation can be smoothly carried out, and the flotation process flow can stably and efficiently operate;
(2) in the phosphate concentrate obtained by flotation according to the method of the invention, P2O5The grade is more than or equal to 30 percent, the recovery rate is more than or equal to 70 percent, the mass content of carbon is less than or equal to 0.8 percent, and SiO2The mass content of the phosphate concentrate is less than or equal to 14 percent, which shows that the phosphate concentrate obtained after defoaming has high grade and less impurities.
Detailed Description
The present invention will be further described with reference to the following examples.
Ether amine acetate and ether polyamine acetate used in the embodiment of the invention are purchased from Yueyangxing group member company of Changyuan refining in Hunan; the ore starting materials or chemicals used in the examples of the present invention, unless otherwise specified, are obtained by conventional commercial means.
Example 1
(1) Decarburization and flotation: in 0.3kg of carbon-silicon collophanite raw ore (P)2O520.43% by mass, 3.58% by mass of carbonaceous material, 28.80% by mass of CaO, and SiO229.30 percent) of the total ore pulp, adding 0.18kg of water, grinding the ore by a wet method until the mass of the ore below a sieve of 0.074mm is 80.23 percent of the total raw ore mass, adding water to adjust the pulp until the mass concentration of the pulp is 22 percent, adding a foaming agent, namely pine oil, into the obtained pulp according to the dosage of 350g of the foaming agent/t of the carbon silicon collophanite raw ore, stirring for 1.5min, performing air flotation, and throwing flotation foam as tailings to obtain decarburization flotation pulp;
(2) desliming and flotation: adding the decarburized ore pulp obtained in the step (1) into a cyclone classifier at a feeding speed of 2m/s and a feeding mass concentration of 25%, performing cyclone classification desliming, taking an overflow part as tailings and discarding tailings (-10 mu m ore removal rate of 65.50%), adding a collecting agent ether amine acetate in the amount of 125g of the collecting agent/t of carbon silicon collophanite raw ore in underflow, stirring for 2.5min, performing air flotation, and taking flotation foam as tailings and discarding tailings (-10 mu m ore removal rate of 18.60%), thereby obtaining desliming flotation ore pulp;
(3) desiliconization reverse flotation: and (3) adjusting the pH value of the desliming flotation pulp obtained in the step (2) to 7.5 by using a pH value regulator sodium hydroxide, adding a collecting agent ether amine acetate according to the using amount of 215g of collecting agent/t of carbon silicon collophanite raw ore, stirring for 1.5min, adding a defoaming agent tributyl phosphate according to the using amount of 80g of defoaming agent/t of carbon silicon collophanite raw ore, stirring for 2.5min, performing air flotation, scavenging flotation foam, using the flotation foam as tailings to discard the tailings, and scavenging middlings to perform desilicication reverse flotation so as to obtain phosphate concentrate pulp.
Through detection, the dynamic stability coefficient of the flotation foam after scavenging in the step (3) is 0.20min, the viscosity and stability of the foam are effectively reduced, the foam brittleness is increased, the defoaming efficiency is not less than 83%, the fluidity of the foam product is improved, defoaming is easy, the foam product can smoothly enter the next flotation operation, and the flotation process flow can stably and efficiently run.
In the phosphorus concentrate obtained by flotation in the embodiment of the invention, P2O530.56% grade, 72.68% recovery rate, 0.76% carbonaceous mass content, 33.58% CaO mass content, and SiO2The mass content of the phosphate concentrate is 12.08%, which shows that the phosphate concentrate obtained after foam regulation has high grade and less impurities.
Example 2
(1) Decarburization and flotation: in 0.3kg of carbon-silicon collophanite raw ore (P)2O522.15% by mass, 2.26% by mass of carbonaceous material, 30.21% by mass of CaO, and SiO227.05 percent) of the raw ore, adding 0.2kg of water, grinding the ore by a wet method until the mass of the ore under a sieve of 0.074mm is 82.50 percent of the total mass of the raw ore, adding water to adjust the pulp until the mass concentration of the pulp is 28 percent, adding a foaming agent methyl isobutyl carbinol into the obtained pulp according to the dosage of 225g of the foaming agent/t of the raw ore of the carbon silicon collophanite, stirring for 2min, aerating, and throwing flotation foam as tailings to obtain the de-tailingC, flotation of ore pulp;
(2) desliming and flotation: adding the decarburized ore pulp obtained in the step (1) into a cyclone classifier at a feeding speed of 3m/s and a feeding mass concentration of 20%, performing cyclone classification desliming, taking an overflow part as tailings and discarding tailings (-10 mu m ore removal rate of 68.80%), adding a collecting agent ether amine acetate in the amount of 100g of the collecting agent/t of carbon silicon collophanite raw ore in underflow, stirring for 2min, performing air flotation, and taking flotation foam as tailings and discarding tailings (-10 mu m ore removal rate of 18.46%), thereby obtaining desliming flotation ore pulp;
(3) desiliconization reverse flotation: and (3) adjusting the pH value of the desliming flotation pulp obtained in the step (2) to 8.4 by using a pH value regulator sodium carbonate, adding a collecting agent ether amine acetate according to the using amount of 250g of collecting agent/t of carbon silicon collophanite raw ore, stirring for 2min, adding a defoaming agent trioctyl phosphate according to the using amount of 65g of defoaming agent/t of carbon silicon collophanite raw ore, stirring for 2min, aerating, scavenging flotation foam, taking the flotation foam as tailings to discard the tailings, and returning the scavenged middlings to the step for desiliconization and reverse flotation to obtain the phosphate concentrate pulp.
Through detection, the dynamic stability coefficient of the flotation foam after scavenging in the step (3) is 0.25min, the viscosity and stability of the foam are effectively reduced, the foam brittleness is increased, the defoaming efficiency is not less than 81%, the fluidity of the foam product is improved, defoaming is easy, the foam product can smoothly enter the next flotation operation, and the flotation process flow can stably and efficiently run.
In the phosphorus concentrate obtained by flotation in the embodiment of the invention, P2O530.70% grade (C), 71.42% recovery rate, 0.54% carbonaceous mass content, 33.88% CaO mass content, and SiO2The mass content of the phosphate concentrate is 13.55%, which shows that the phosphate concentrate obtained after foam regulation has high grade and less impurities.
Example 3
(1) Decarburization and flotation: in 0.3kg of carbon-silicon collophanite raw ore (P)2O518.98% by mass, 4.50% by mass of carbonaceous material, 26.10% by mass of CaO, and SiO234.22 percent) of the raw materials are added with 0.21kg of water, and the wet grinding is carried out until the ore with the size of 0.074mm is sievedThe mass of the ore pulp is 83.70 percent of the total raw ore mass, water is added to adjust the pulp until the mass concentration of the ore pulp is 32 percent, in the obtained ore pulp, a foaming agent diethyl phthalate is added according to the using amount of 400g of foaming agent/t of carbon silicon collophanite raw ore, the mixture is stirred for 2.5min, air flotation is carried out, flotation foam is used as tailings to be thrown, and decarburized flotation ore pulp is obtained;
(2) desliming and flotation: adding the decarburized ore pulp obtained in the step (1) into a cyclone classifier at a feeding speed of 4m/s and a feeding mass concentration of 15%, performing cyclone classification desliming, taking an overflow part as tailings and discarding tailings (-the removal rate of the minerals with the size of 10 mu m is 64.50%), adding a collecting agent ether polyamine acetate in the amount of 150 g/t of carbon silicon collophanite raw ore into underflow, stirring for 1.5min, performing air flotation, and taking flotation foam as tailings and discarding tailings (the removal rate of the minerals with the size of 10 mu m is 20.15%) to obtain desliming flotation ore pulp;
(3) desiliconization reverse flotation: and (3) adjusting the pH value of the desliming flotation pulp obtained in the step (2) to 8.4 by using a pH value regulator sodium carbonate, adding a collecting agent ether polyamine acetate according to the using amount of 325g of collecting agent/t of carbon silicon collophanite raw ore, stirring for 2.5min, adding a defoaming agent tributyl phosphate according to the using amount of 125g of defoaming agent/t of carbon silicon collophanite raw ore, stirring for 1.5min, performing air flotation, scavenging flotation foam, using the flotation foam as tailings to discard the tailings, and returning the scavenged middlings to the step for desiliconization reverse flotation to obtain the phosphate concentrate pulp.
Through detection, the dynamic stability coefficient of the flotation foam after scavenging in the step (3) is 0.18min, the viscosity and stability of the foam are effectively reduced, the foam brittleness is increased, the defoaming efficiency is not less than 85%, the fluidity of the foam product is improved, defoaming is easy, the foam product can smoothly enter the next flotation operation, and the flotation process flow can stably and efficiently run.
In the phosphorus concentrate obtained by flotation in the embodiment of the invention, P2O530.15% grade (C), 70.12% recovery rate, 0.65% carbonaceous mass content, 32.87% CaO mass content, and SiO2The mass content of the phosphate concentrate is 12.89%, which shows that the phosphate concentrate obtained after foam regulation has high grade and less impurities.
Comparative example 1
(1) Desliming and flotation: in 0.3kg of carbon-silicon collophanite raw ore (P)2O520.43% by mass, 3.58% by mass of carbonaceous material, 28.80% by mass of CaO, and SiO229.30 percent) of the raw ore, adding 0.18kg of water, grinding the raw ore by a wet method until the mass of the ore under a screen of 0.074mm is 80.23 percent of the total raw ore mass, adding water to adjust the mass concentration of the ore pulp to be 22 percent, adding the obtained ore pulp into a cyclone classifier at the feeding speed of 2m/s and the feeding mass concentration of 25 percent, performing cyclone classification desliming firstly, throwing the overflow part as tailings to throw off the tailings (-10 mu m of the ore at the removal rate of 65.50 percent), adding the collector ether amine acetate into the underflow according to the dosage of 125g of the collector/t of the carbon silicagel phosphorite raw ore, stirring for 2.5min, performing air flotation, and throwing the flotation foam as tailings to throw off the tailings (-10 mu m of the ore at the removal rate of 18.6 percent) to obtain desliming flotation ore pulp;
(2) desiliconization reverse flotation: and (2) adjusting the pH value of the desliming flotation pulp obtained in the step (1) to 7.5 by using a pH value regulator sodium hydroxide, adding a collecting agent ether amine acetate according to the using amount of 215g of collecting agent/t of carbon silicon collophanite raw ore, stirring for 1.5min, adding a defoaming agent tributyl phosphate according to the using amount of 80g of defoaming agent/t of carbon silicon collophanite raw ore, stirring for 2.5min, performing inflation flotation, scavenging flotation foam, using the flotation foam as tailings to discard the tailings, and returning the scavenged middlings to the step for desiliconization reverse flotation to obtain the phosphate concentrate pulp.
Through detection, the dynamic stability coefficient of the flotation foam after scavenging in the step (2) is 1.28min, the foam viscosity is high, the stability is high, the foam brittleness is small, the defoaming efficiency is only 41.82%, the flowability of a foam product is poor, defoaming is not easy to occur, next flotation operation is difficult to enter, and the flotation process flow is hindered.
In the phosphate concentrate obtained by the flotation of the comparative example, P2O5Has a grade of 28.23%, a recovery rate of 65.23%, a carbonaceous mass content of 1.55%, a CaO mass content of 31.43%, and SiO2The mass content of the phosphate concentrate is 15.78 percent, which shows that the phosphate concentrate obtained after defoaming has low grade and more impurities.
Comparative example 2
(1) Decarburization and flotation: in the presence of 0.3kg of carbon-silicon collophaniteRaw ore (P)2O522.15% by mass, 2.26% by mass of carbonaceous material, 30.21% by mass of CaO, and SiO227.05 percent) of the total ore pulp, adding 0.2kg of water, grinding the ore by a wet method until the mass of the ore below a sieve of 0.074mm is 82.50 percent of the total raw ore mass, adding water to adjust the pulp until the mass concentration of the pulp is 28 percent, adding a foaming agent, namely pine oil, into the obtained pulp according to the dosage of 225g of the foaming agent/t of the carbon silicon collophanite raw ore, stirring for 2min, performing air flotation, and throwing flotation foam as tailings to obtain decarburization flotation pulp;
(2) desiliconization reverse flotation: and (2) adjusting the pH value of the decarbonization flotation pulp obtained in the step (1) to 8.4 by using a pH value regulator sodium carbonate, adding a collecting agent ether amine acetate according to the using amount of 250g of collecting agent/t of carbon silicon collophanite raw ore, stirring for 2min, adding a defoaming agent tributyl phosphate according to the using amount of 65g of defoaming agent/t of carbon silicon collophanite raw ore, stirring for 2min, aerating, scavenging flotation foam, taking the flotation foam as tailings to discard the tailings, and returning the scavenged middlings to the step for desiliconization reverse flotation to obtain the phosphate concentrate pulp.
Through detection, the dynamic stability coefficient of the flotation foam after scavenging in the step (2) is 2.58min, the foam viscosity is high, the stability is high, the foam brittleness is small, a large amount of foam is accumulated, the defoaming efficiency is only 15%, the flowability of a foam product is poor, defoaming is not easy to occur, next flotation operation is difficult to enter, and the flotation process flow is hindered.
In the phosphate concentrate obtained by the flotation of the comparative example, P2O529.42% grade (C), 61.58% recovery rate, 0.50% carbonaceous mass content, 32.64% CaO mass content, and SiO2The mass content of (A) is 13.22%, which shows that the obtained phosphate concentrate P2O5Although the grade reaches the standard, the phosphorus recovery rate is low, and the flotation froth is not suitable for industrial production.
Comparative example 3
(1) Decarburization and flotation: in 0.3kg of carbon-silicon collophanite raw ore (P)2O518.98% by mass, 4.50% by mass of carbonaceous material, 26.10% by mass of CaO, and SiO234.22 percent) is added with 021kg of water, grinding the ore by a wet method until the mass of the undersize ore of 0.074mm is equal to 83.70 percent of the total mass percent of the ore, adding water to adjust the pulp until the mass concentration of the pulp is 32 percent, adding a foaming agent, namely pine oil, into the obtained pulp according to the dosage of 400g of the foaming agent per t of the carbon silicon collophanite raw ore, stirring for 2.5min, performing air flotation, and throwing flotation foam as tailings to obtain decarburization flotation pulp;
(2) desliming and flotation: adding the decarburized ore pulp obtained in the step (1) into a cyclone classifier at a feeding speed of 4m/s and a feeding mass concentration of 15%, performing cyclone classification desliming, taking an overflow part as tailings and discarding tailings (-the removal rate of the minerals with the size of 10 mu m is 64.50%), adding a collecting agent ether polyamine acetate in the amount of 150 g/t of carbon silicon collophanite raw ore into underflow, stirring for 1.5min, performing air flotation, and taking flotation foam as tailings and discarding tailings (the removal rate of the minerals with the size of 10 mu m is 17.98%) to obtain desliming flotation ore pulp;
(3) desiliconization reverse flotation: and (3) adjusting the pH value of the desliming flotation pulp obtained in the step (2) to 8.4 by using a pH value regulator sodium carbonate, adding a collecting agent ether polyamine acetate according to the using amount of 325g of collecting agent/t of carbon silicalite collophanite raw ore, stirring for 2.5min, performing air flotation, scavenging flotation foam, discarding tailings by using the flotation foam, returning the scavenged middling to the step for desiliconization reverse flotation, and obtaining the phosphate concentrate pulp.
Through detection, the dynamic stability coefficient of the flotation foam after scavenging in the step (3) is 1.88min, the foam viscosity is high, the stability is high, the foam brittleness is small, a large amount of foam is accumulated, the defoaming efficiency is only 37.42%, the flowability of a foam product is poor, defoaming is not easy to occur, next flotation operation is difficult to enter, and the flotation process flow is hindered.
In the phosphate concentrate obtained by the flotation of the comparative example, P2O5Has a grade of 29.15%, a recovery rate of 66.58%, a carbonaceous content of 0.55%, a CaO content of 32.18%, and SiO2The mass content of (A) is 13.56%, which shows that the obtained phosphate concentrate P2O5Although the grade reaches the standard, the foam is too stable and is not suitable for industrial production.
In conclusion, in comparative example 1, when decarburization flotation is not carried out, the obtained phosphate concentrate has low grade,Impurities are more, and the defoaming efficiency is only 41.82%; comparative example 2 when desliming flotation was not performed, the resulting phosphate concentrate P was2O5Although the grade reaches the standard, the phosphorus recovery rate is low, the flotation foam is not suitable for industrial production, and the defoaming efficiency is only 15%; comparative example 3 desilication flotation step without addition of antifoaming agent, phosphorus concentrate P was obtained2O5Although the grade reaches the standard, the foam is over-stable and is not suitable for industrial production, and the defoaming efficiency is only 37.42%. Therefore, the desliming process has the largest influence on the defoaming efficiency of foam, the next defoaming agent in the desiliconization flotation process has the smallest influence on the decarburization process, but the decarburization process has the largest influence on the quality of the phosphate concentrate. Therefore, for the carbon-silicon collophanite, the decarburization process, the desliming process and the defoaming agent must be combined for ensuring the quality of the cation desilication phosphate concentrate and the stable operation of the process.

Claims (4)

1. A carbon silicon collophanite cation reverse flotation desilication method based on foam regulation is characterized by comprising the following steps:
(1) decarburization and flotation: wet grinding the carbon-silicon collophanite raw ore, adding water for size mixing, adding a foaming agent into the obtained ore pulp, stirring, performing air flotation, and discarding the flotation foam as tailings to obtain decarburization flotation ore pulp;
(2) desliming and flotation: carrying out cyclone classification desliming on the decarburized ore pulp obtained in the step (1), using an overflow part as tailings for tailing discarding, adding a collecting agent into bottom flow, stirring, carrying out air flotation, and using flotation foam as tailings for tailing discarding to obtain desliming flotation ore pulp;
(3) desiliconization reverse flotation: adjusting the pH value of the desliming flotation pulp obtained in the step (2), adding a collecting agent, stirring, adding a defoaming agent, stirring, performing air flotation, scavenging flotation foam, and discarding tailings as flotation foam to obtain phosphate concentrate pulp;
in the step (2), the amount of the collecting agent is 50-200 g of collecting agent per t of carbon silicon collophanite raw ore; the collecting agent is ether amine acetate and/or ether polyamine acetate; the stirring time is 1-3 min;
in the step (3), adjusting the pH value of the desliming flotation pulp to 7.5-8.5; the pH value regulator for regulating the pH value is sodium carbonate and/or sodium hydroxide; the using amount of the collecting agent is 100-500 g of collecting agent per t of carbon silicon collophanite raw ore; the collecting agent is ether amine acetate and/or ether polyamine acetate; the stirring time is 1-3 min; the dosage of the defoaming agent is 50-150 g of defoaming agent per t of carbon silicon collophanite raw ore; the defoaming agent is trioctyl phosphate; the stirring time is 1-3 min.
2. The method for desilication of carbon silicon collophanite cation reverse flotation based on foam regulation and control as claimed in claim 1 is characterized in that: in the step (1), in the carbon-silicon collophanite, P is2O518-25% by mass, 1-8% by mass of carbon, 22-32% by mass of CaO, and SiO2The mass percentage of the components is 20-35%; in the wet grinding, the amount of water is 0.5-0.8 times of the mass of the carbon silicon collophanite; wet grinding until the mass of ore under a sieve of 0.074mm is less than or equal to 85 percent of the total mass of the raw ore; adding water to mix pulp until the mass concentration of the pulp is 20-35%.
3. The carbon-silicon collophanite cation reverse flotation desilication method based on foam regulation and control as claimed in claim 1 or 2, is characterized in that: in the step (1), the amount of the foaming agent is 100-600 g of foaming agent per t of carbon silicon collophanite raw ore; the foaming agent is one or more of pinitol oil, methyl isobutyl carbinol or diethyl phthalate; the stirring time is 1-3 min.
4. The method for desilication of carbon silicon collophanite cation reverse flotation based on foam regulation and control as claimed in claim 1 is characterized in that: in the step (2), the feeding speed of the cyclone classification desliming is 1-4 m/s, and the feeding mass concentration is 10-25%.
CN201810480343.0A 2018-05-18 2018-05-18 Carbon silicon collophanite cation reverse flotation desilication method based on foam regulation Active CN108940563B (en)

Priority Applications (1)

Application Number Priority Date Filing Date Title
CN201810480343.0A CN108940563B (en) 2018-05-18 2018-05-18 Carbon silicon collophanite cation reverse flotation desilication method based on foam regulation

Applications Claiming Priority (1)

Application Number Priority Date Filing Date Title
CN201810480343.0A CN108940563B (en) 2018-05-18 2018-05-18 Carbon silicon collophanite cation reverse flotation desilication method based on foam regulation

Publications (2)

Publication Number Publication Date
CN108940563A CN108940563A (en) 2018-12-07
CN108940563B true CN108940563B (en) 2020-11-06

Family

ID=64499232

Family Applications (1)

Application Number Title Priority Date Filing Date
CN201810480343.0A Active CN108940563B (en) 2018-05-18 2018-05-18 Carbon silicon collophanite cation reverse flotation desilication method based on foam regulation

Country Status (1)

Country Link
CN (1) CN108940563B (en)

Families Citing this family (4)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN110124870B (en) * 2019-05-31 2021-04-30 北京矿冶科技集团有限公司 Mineral separation and impurity removal method for carbon-containing magnesium-containing high sesquioxide collophanite
CN110369152B (en) * 2019-07-15 2021-12-03 宜都兴发化工有限公司 Flotation process for micro-fine particle phosphorite
CN111298979A (en) * 2020-03-27 2020-06-19 云南铁峰矿业化工新技术有限公司 Foaming agent
CN113800540B (en) * 2021-09-30 2023-06-02 浙江镇洋发展股份有限公司 Method for removing silicon aluminum by one-time refining of ionic membrane caustic soda salt water

Family Cites Families (5)

* Cited by examiner, † Cited by third party
Publication number Priority date Publication date Assignee Title
CN102441498B (en) * 2011-10-31 2013-07-24 中蓝连海设计研究院 Phosphorite double-reverse flotation process
CN103285625B (en) * 2013-06-13 2015-08-26 北京矿冶研究总院 Defoaming method for reverse flotation desilication of phosphorite
CN103909017B (en) * 2014-04-18 2017-02-15 武汉工程大学 Flotation process for silicon-calcium mixed collophanite containing organic carbon
CN104841569B (en) * 2015-06-12 2017-10-10 中蓝连海设计研究院 A kind of middle-low grade silico-calcium matter Collophanite flotation technique
CN105597939A (en) * 2015-12-25 2016-05-25 湘西自治州鑫汇金属新材料有限公司 Process for mineral processing of low-grade silicon calcium collophanite

Also Published As

Publication number Publication date
CN108940563A (en) 2018-12-07

Similar Documents

Publication Publication Date Title
CN108940563B (en) Carbon silicon collophanite cation reverse flotation desilication method based on foam regulation
CN102744151B (en) Branch flotation technology for silicon calcium collophanite
CN101972710B (en) Double reverse flotation process of middle-low grade phosphorite
CN1911527A (en) Collector for bauxite flotation
CN101121151A (en) Bauxite surface preprocessing-reverse floatation desiliconizing method
CN109382213A (en) A kind of beneficiation method of gibbsitic bauxite
CN105880032B (en) A kind of mid-low grade collophanite ore floats combined sorting method again
CN104511373A (en) Mineral separation method for high-oxidative molybdenum ore
CN104984835A (en) Selective flocculation-column flotation recovery method and system of micro-fine particle molybdenum cleaner tailings
CN111389598B (en) Method for recovering mica and feldspar quartz from rare metal ore dressing tailings
CN105750089B (en) A kind of magnesia collophane method for separating
CN104174504A (en) Direct-flotation branching floatation method of low-and-medium-grade mixed type refractory collophane
CN111468302B (en) Beneficiation inhibitor and purification method of molybdenum rough concentrate
CN107029896B (en) Flotation process for separating and enriching apatite, dolomite and quartz in phosphorite
CN107433230A (en) A kind of Scheelite Flotation inhibitor and its preparation technology
CN112474065A (en) Method for selecting phosphorus from low-grade vanadium titano-magnetite tailings
CN106076607B (en) A kind of two step desliming reverse floatation process handling high alumina-silica collophane
CN111298978B (en) Method for flotation of lepidolite without desliming
CN113976304A (en) Beneficiation method for comprehensively recovering tungsten tin bismuth silicon from skarn type tin tailings
CN106733220B (en) A kind of zinc oxide ore slurry bubble sorting ore-dressing technique
CN105327784B (en) It is a kind of to select separation method of the association rich in fluorite and low-grade cassiterite in Tungsten tailing
CN112844818B (en) Beneficiation separation method for copper-zinc sulfide ore
US3259326A (en) Method of slime beneficiation
CN114749281B (en) Calcium gangue zinc oxide ore mixed collector, preparation method and use method thereof
CN113877721B (en) Method for deeply removing micro-fine black-white mica from granite type metal ore tailings

Legal Events

Date Code Title Description
PB01 Publication
PB01 Publication
SE01 Entry into force of request for substantive examination
SE01 Entry into force of request for substantive examination
GR01 Patent grant
GR01 Patent grant