166results about How to "High enrichment ratio" patented technology

The invention belongs to the technical field of hydrometallurgy, and particularly relates to a method for individually processing high-iron zinc sulfide concentrate. The method comprises a step of subjecting the high-iron zinc sulfide concentrate to calcination in a fluidized bed combustion boiler to obtain zinc calcine; a step of subjecting the zinc calcine to neutral leaching to produce neutral leaching solution and neutral leaching residue; a step of, after the neutral leaching residue and the high-iron zinc sulfide concentrate are mixed, successively performing reduction leaching and oxidation leaching, and circulating oxidation leaching solution to the reduction leaching to produce reduction leaching solution and silver-rich sulfur residue; a step of replacing the reduction leaching solution by using iron powder to precipitate copper and to produce copper-rich slag and solution after copper precipitation; a step of subjecting the solution after copper precipitation to pre-neutralization by using the zinc calcine, and then replacing by using zinc powder to precipitate indium and to produce indium-rich slag and solution after indium precipitation; and a step of bubbling oxygen into the solution after indium precipitation, heating and removing iron to obtain iron removal solution and hematite slag. The hematite slag can be utilized as a raw material for ironmaking. The method has strong pertinence, short technological process and high metal recovery yield, and the method is clean, efficient, energy-saving and environmental friendly. Separation and comprehensive utilization of zinc, indium, copper and iron are achieved.

The invention provides a method for treating arsenic-containing waste copper slag. The method comprises the following steps of: adding alkali and arsenic fixation roasting on the arsenic-containing waste copper slag from a copper electrolysis purification procedure to convert the arsenic into low-toxicity, water-soluble and nonvolatile arsenate, leaching the roasting slag into water to remove the arsenic, recovering copper and enriching antimony bismuth silver by virtue of acid leaching, and embedding calcium arsenate sediment converted from the arsenic in the water leaching solution; and comprehensively recovering valued metals such as copper, silver, antimony and bismuth in the black copper slag. The method is a safe and effective wet smelting method for recovering the valued metals in the black copper slag, and the arsenic and the copper are leached separately; the recovery rate of the copper in the black copper slag reaches 99.6 percent, and the removal rate of the arsenic reaches over 98 percent; over 95 percent of antimony and over 98 percent of bismuth enter the slag, and over 98 percent of silver also enters the acid leaching slag by adding trace chlorine radicals during acid leaching, so that the antimony, the bismuth and the silver are comprehensively recovered; and the method has the advantages of low equipment investment, short flow, low running cost and safe and reliable operating environment.

The invention discloses a rutile roughing technology for consisting of selective ore grinding, coarse particle gravity separation and fine particle floatation, belonging to the technical field of mineral processing engineering. According to the characteristics of complex mineral composition of rutile, containing of a large quantity of flaky rocky minerals in raw ore, small crystal size of rutile,non-uniform distribution, remarkable selective grinding phenomenon in an ore grinding process and the like, the technology comprises the following steps of: strengthening selective ore grinding by taking a steel forging as an ore grinding medium; pre-separating and discarding coarse particle tailings with a high-frequency fine sieve; narrowing the particle composition undersize minerals; recovering large-particle rutile by performing gravity separation, wherein rutile in gravity-separated tailings is distributed at a fine particle scale; sieving and discarding coarse-particle tailings; recovering small-particle rutile from minus sieve by suppressing rutile reverse flotation desliming and activating rutile positive floatation; and combining gravity-separated rough concentrate with floated rough concentrate, concentrating and purifying to obtain a high-quality rutile concentrate product. The technology has the characteristics of simple process, low ore dressing cost, high recovery rate,high concentration ratio, and the like.

The invention relates to a method for the floatation of potassium chloride in a sylvite mine, and a method for producing the product of chemical industry, in particular to a novel floatation method for extracting the potassium chloride from the sylvite mine. The method of the invention mainly comprises the following steps: a. the sylvite mine is pulverized and graded, the particle size of the pulverized mine rages from 3 to 8mm, and the pulverized mine is processed by the operation of cooling decomposition by being added with sweet water; b. mine slurry obtained in the step a is led into a mineralization slot and is added with the saturated mother solution of potassium chloride, sodium chloride and magnesium chloride as well as a positive floating agent; and the mine slurry, the saturated mother solution and the positive floating agent are completely mixed to control the concentration of the mine slurry to be from 20 to 40 percent; c. the mine slurry obtained in the step b is led into a flotation separation device to obtain the potassium chloride by the positive floatation; and d. particles fallen to the bottom of a flotation column are milltailings, particles which are discharged from the flotation column and are caused to float to the top end of the flotation column are potassium chloride particles, and the particles are collected to be concentrated, filtered and dried. The method has the advantages of simple and stable process, easier automatic control, favorable flotation effect, low agent consumption and production cost, etc.

The invention provides a beneficiation method for a high-carbon molybdenum nickel ore, which adopts floatation decarbonization so as to eliminate the influence of high-content carbon substances on subsequent nickel floatation. Furthermore, the beneficiation process adopts sieve classification with jigging and tail throwing so as to enhance the floatation and selected grade, then rough concentrates are selected again and treated by floatation after being ground, and the ore concentrates are selected again with tail throwing to enhance the floatation and selected grade without ore grinding. The invention can greatly reduce the floatation quantity, has good beneficiation index and provides a new way for developing a low-grade high-carbon molybdenum nickel ore.

A mineral dressing technique used for enriching nickel and/or cobalt is characterized in that limonite type ore and serpentine type ore are separated for mineral dressing, and adopt a sieve to be processed by classification for at least once after being processed by ore washing; mineral dressing is carried out on the products with granulometric classes, and the target concentrated ore is obtainedby the techniques of reelection and magnetic classification. According to the invention, after the limonite type ore and the serpentine type ore are treated by sieve classification, coarse fraction materials directly enter subsequent hydrometallurgy operation, so that the handling capacity of the selected crude ore is reduced, and the sorting conditions are improved; materials with medium fraction and fine fraction adopt a combined process flow of reelection and magnetic classification for sorting to obtain the target concentrated ore, so that the mineral quantity entering the process flow ofhydrometallurgy is reduced. Mineral dressing is respectively carried out on the limonite type ore and the serpentine type ore in lateritic nickel, thus obtaining the concentrated ore having higher quality, recovery rate and concentration ratio.

The invention discloses a method for enriching precious metal by dissolving multi-metal alloy material through nitric acid. The method comprises the steps that the multi-metal alloy material is mixed with the nitric acid, placed in a titanium reaction kettle to be leached, heated to 85 DEG C-95 DEG C and kept for a period of time, when the multi-metal alloy material is concentrated to be in a sticky state, heating is stopped, water with the weight ratio being two to six times of that of the multi-metal alloy material is added at the same time so as to reduce the acidity, and the pH value of an end point is increased by one to three times or so, so that filtrating and separating are conveniently conducted between main metal lead, nickel, copper and the like and the precious metal; after filtering and washing are conducted, leaching liquid and leaching residue are obtained, the leaching liquid contains the lead, the copper and the like and serves as a raw material for recycling the lead and the copper, and the leaching residue contains bismuth, antimony, gold, silver, platinum, palladium and the like; and the antimony is removed through pressurizing and alkali leaching, alkali leaching liquid and alkali leaching residue are obtained after filtering and washing are conducted, and the precious metal enters the alkali leaching residue and effectively enriched. The method for enriching the precious metal by dissolving the multi-metal alloy material through the nitric acid is simple in process, easy to operate, mature in involved production equipment, easy to industrialize, environmentally friendly, low in production cost, high in precious metal enrichment ratio and good in industrialized application prospect.

The invention discloses a method for separating and enriching valuable metals from copper anode mud. The method comprises the following steps that (1) the copper anode mud and a smelting agent are mixed evenly and are smelted at the temperature of 350-700 DEG C, and a smelting product is obtained; (2) the smelting product is crushed and immersed in water, alkaline leaching residues and alkaline leaching liquid are obtained, and selenium and arsenic are recycled from the alkaline leaching liquid; (3) acid and sodium chloride are added into the alkaline leaching residues, ozone is introduced in the alkaline leaching residues for ozone strengthening acid leaching, and acid leaching residues and acid leaching liquid are obtained; and copper and tellurium are recycled from the acid leaching liquid, and lead, stibium and precious metals are recycled from the acid leaching residues. According to the process, the method has the advantages that the selenium and arsenic removal rate is high; the moving directions of the valuable metals are more reasonable and concentrated; the enrichment ratio of the precious metals is high; the direct recovery rate of all elements is high; the comprehensive recovery benefits are good; the problem that tellurium and arsenic in the copper anode mud are dispersed seriously in a traditional process is solved; operation is safe; the labor intensity is low; the treatment time is short; and the operation environment is good.

The invention discloses a technology for floatation of lead and zinc sulfide ores through inflatable floatation columns. The floatation columns are fully utilized, and the method includes the steps of ore grinding, mixing of ore pulp and flotation reagents, rougher flotation operation, concentration operation, scavenging operation and the like. The method is based on the flotation column three-phase flow mechanics theory, through process mineralogy research and technological process research of low-grade, complex and refractory lead and zinc sulfide ores of mines and research and improvement on an inflatable flotation column device, the flotation columns are successfully applied to the flotation technology of low-grade, complex and refractory lead and zinc sulfide ores, and the technological process that the floatation columns are adopted in all the flotation operations of rougher flotation, concentration and scavenging of lead and zinc sulfide ores is achieved. The technology has the advantages that the automation degree is high, labor intensity of workers is low, ore flotation efficiency is high, the technological process is simple, production indexes are stable, maintenance is convenient, building investment and production cost are low, consumption of the reagents is saved by about 10%, electricity consumption is saved by about 20 %, and water consumption is saved by about 20 %.

The invention relates to a lead-zinc oxide ore flotation inhibitor and application thereof, and belongs to the technical field of mineral processing. The designed inhibitor comprises the following components in parts by mass: 30-80 parts of lignosulfonate, 10-40 parts of assistant A, and 10-30 parts of soluble starch. The application scheme comprises the following steps: a raw ore is finely grinded to particles with a particle size not larger than 0.074 mm; after the particles are more than 85% of total mass of the raw ore, the size mixing is performed to obtain preselected slurry; after the pH value of the preselected slurry is adjusted to 8-11, the lead-zinc oxide ore flotation inhibitor is added for uniformly stirring; and then, a capturing agent is added to obtain a rough concentration of zinc oxide. The inhibitor is reasonable in component design, has dual attributes of dispersion and inhibition, has such advantages as high selectivity, high inhibition capacity, no poison and pollution and convenience for operation and management, and is convenient for large-scale industrial application.

A process for removing Mg from phosphorus ore in order to prepare phosphoric acid by wet method features use of static microbubble floatation column, and includes such steps as breaking the crude phosphorus ore, grinding, classifying, mineralizing by mixing the reverse floatation chemical with ore sludge, and floatating in said static microbubble floatation column while conditioning the ore sludge to obtain the bubble-phase tailings and low-Mg phosphorus ore concentrate from the column bottom.

The invention discloses a flotation cell without transmission, which comprises a cell body and a pulp distributer; wherein, a feeding pipe arranged above the pulp distributer, and at least one venturi pipe is arranged under the pulp distributer; the venturi pipe is connected with a mineralized pipe which is connected with the pulp distributer, the bottom end of which is provided with at least onepiece of steel pipe, and a pulp ejector device is arranged at the bottom end of the steel pipe and is provided with a plurality of nozzles; the steel pipe is installed in the cell body, and a foam collecting groove is installed at the upper part of the cell body; the cell body is connected with a middling ore box by a pipeline; the flotation cell has high separation efficiency and low energy consumption, can meet the requirements of aluminum-silicon separation process of alumyte flotation desiliconization with different fractions, and has wide application scope.

The present invention discloses a complex force field spiral chute for ore dressing, and fine particles can be effectively prevented from entering into a tailing zone. The spiral chute comprises a channel steel support, an ore-feeding device, magnetic poles, a pulsating pressure water tank, a spiral blade, an ore dividing bucket, an ore intercepting bucket and an ore receiving bucket. The permanent magnetic fixed magnetic poles are installed at an external portion of a spiral blade of an existing spiral chute, so that magnetic fine particles in magnetic mineral can be forced by an inward-radial magnetic retentive force to enter into a concentrate zone under the function of an applied magnetic field. The pulsating water tank is installed at the bottom of a spiral surface, ore particles are prevented from rubbing with the spiral surface, so that the ore particles can be well layered and separated conveniently, and equipment abrasion can be eliminated conveniently. Concentrate grades can be improved, concentration ratio can be increased, mineral recovery rates can be improved, an existing spiral ore dressing technology can be simplified, and equipment maintenance rates and equipment power consumption can be reduced.

The invention discloses a method and a device for mineral processing of micro-granular magnetic minerals. The method is as follows: the gradient and the rotating speed of a sorting and revolving cylinder are set according to the components of mineral materials; after being put into the sorting and revolving cylinder, the mineral materials are absorbed on the sorting and revolving cylinder and rotate along with the sorting and revolving cylinder; the mineral materials are washed by rinse water and non-magnetic minerals and weak magnetic minerals in the mineral materials are separated to move towards the lower end of the sorting and revolving cylinder to a mine tailing discharge hopper; and under the action of the relative movement of the sorting and revolving cylinder and a spiral magnetic system, the magnetic minerals are moved to the high end of the sorting and revolving cylinder spirally, and are discharged into a concentrate discharge hopper through a spiral mineral discharging plate after the magnetic minerals break away from a magnetic field. The device comprises a stander, a support, a rinsing device, a feeding hopper, a spiral magnetic system, a magnetic conducting outer cylinder, the sorting and revolving cylinder and a transmission device. The sorting and revolving cylinder is supported on the support through a carrier roller; one end of the sorting and revolving cylinder is provided with the mine tailing discharge hopper, while the other end is provided with the concentrate discharge hopper and the spiral mineral discharging plate; and the support is hinged with the stander at the mine tailing discharge end, and is supported on the stander at the concentrate discharge end through a gradient regulating device. The method has large handling capacity and high concentration ratio, and the device has strong handling ability and high coefficient of recovery.

A technology for dressing the primary ilmenite includes removing Fe component, coarse separating for enriching TiO2 to more than 20%, drying, and fine dressing by permanent-magnet drum-type separator to enrich TiO2 to more than 47%.

The present invention relates to sulfur slag vacuum volatilizing process for enriching noble metal, and belongs to the field of vacuum fire metallurgy process. Sulfur slag material is distilled in a vacuum smelting furnace at pressure of 10-30 Pa and temperature of 250-400 deg.c for 50-120 min to volatilize sulfur and the sulfur vapor is cooled to form liquid and exhausted from the vacuum smelting furnace, so as to obtain metal sulfide powder with enriched Cu, Ni, Os, Ag, Au, etc. The process is simple, and has high noble metal enriching ratio, no environmental pollution, low power consumption and low production cost.

The invention discloses a selective flocculation flotation method for extremely fine ilmenite. The selective flocculation flotation method comprises the following steps of stirring the extremely fine ilmenite with a dispersing agent, an activating agent, collector and an auxiliary collector into slurry, and then, performing rougher flotation by adopting a flotation column to obtain rougher concentrate and rougher tailings; adding the collector to the obtained rougher tailings and stirring, and performing secondary scavenging by adopting the flotation column to obtain secondary scavenging concentrate and secondary scavenging tailings, wherein the secondary scavenging tailings are final tailings; adding the dispersing agent and the activating agent to the obtained rougher concentrate and stirring; performing triple concentration by adopting the flotation column to obtain triple concentration concentrate and triple concentration tailings, wherein the triple concentration concentrate is final concentrate. The current situation that the extremely fine ilmenite is directly thrown away as slurry can be solved by the selective flocculation flotation method disclosed by the invention; moreover, the selective flocculation flotation method is high in flotation efficiency and high in stability, and has good indexes.

The invention belongs to the technical field of nonferrous metallurgy, and particularly relates to a selective oxygen pressure leaching method for the enrichment of precious metals from high-iron, high-copper and nickel-containing materials. The method comprises the following steps of purification and impurity removal, normal-pressure leaching, oxygen-pressure deironing, oxygen-pressure nickel immersion and oxygen-pressure copper immersion. The method is short in technological process and high in direct yield, and can be used for selectively leaching nickel, cobalt, copper and iron and respectively realizing open circuit; and the method is wide in raw material adaptability, suitable for treating various high-iron high-copper nickel-containing materials such as intermediate product low-nickel matte generated in the conventional nickel sulfide ore smelting process, primary alloys, fine grain alloys, cobalt-rich low-nickel matte and cobalt-rich matte, and is remarkable in benefit. The nickel sulfate cobalt sulfate solution obtained by adopting the method is high in the product quality, and the requirements of multi-variety production can be met at the same time; the noble metal can be totally enriched in the oxygen pressure leaching copper slag and can be used as a noble metal concentrate, the direct yield and the enrichment ratio are high, and iron is directly used in a product form; and copper is leached to obtain a copper sulfate solution, and crystallized copper sulfate or electrodeposition copper can be produced.

The invention provides a printing-dyeing wastewater treatment process, which belongs to the technical field of wastewater treatment. The treatment process uses a foam separation technique together with Fenton oxidation to treat printing-dyeing wastewater containing more than two dyes, and concretely comprises the following: a step of collecting and filtering the printing-dyeing wastewater containing more than two dyes; a step of detecting and pretreating, which is to detect the pigment composition, chemical oxygen consumption, color, pH value and suspended solid content of the dyes contained in the printing-dyeing wastewater and to regulate the pH value of the wastewater to between 6.0 and 8.0 by use of hydrochloric acid or sodium hydroxide; a step of separating foam and decolorizing, which is to add Dodecyl trimethyl ammonium bromide as surfactant to the wastewater treated through the previous step and then to decolorize by use of a bubbling tower; and a step of carrying out Fenton oxidation to the wastewater obtained in the previous step. The COD of the treated wastewater meets the national emission standard. The process overcomes the disadvantage that the prior printing-dyeing wastewater treatment process is large in occupied area, causes secondary pollution, is low in color removal rate, great in treatment effect fluctuation, complex in process and high in cost, or can not achieve ideal effects.

The invention discloses a method for enrichment of platinum in a fluorine-containing spent platinum catalyst. The method comprises the following steps of mixing a fluorine-containing spent platinum catalyst, a fluoride retention agent, a platinum trapping agent, a slagging agent, a reducer and a binder to obtain a uniform mixture, carrying out granulation by a granulator to obtain granules with 5cm, carrying out drying, carrying out melting by an electric-arc furnace at a temperature of 1300-1400 for 1-2h to obtain platinum alloy and melting slag, adding fluorine into the melting slag in melting, melting the platinum alloy by an intermediate frequency furnace, carrying out atomization powder injection to obtain fine platinum alloy particles, carrying out selective leaching of iron in the platinum alloy particles, and carrying out filtration and washing to obtain platinum richment which is platinum concentrate. The platinum concentrate has platinum content greater than 30%, a platinum enrichment ratio 25-35 times the original ratio and a platinum yield greater than 99.0% and is a high-quality raw material for platinum purification. The method has simple processes, a high platinum yield, a high enrichment ratio, environmental friendliness, a low cost and a good industrialization prospect.

The invention discloses a method for efficiently enriching precious metal from electronic waste. Particularly, anode slime is obtained from the electronic waste which is subjected to copper smelting capture and then subjected to electrolysis. The method comprises the steps that the electronic waste is crushed, added with copper, a reducing agent and a slag former and placed in an electric arc furnace to be smelted, so that crude copper is obtained, and the precious metal is effectively captured; the crude copper is then subjected to electrolysis, the precious metal enters copper anode slime, and thus preliminary enrichment of the precious metal is achieved; the copper anode slime material is mixed with sulfuric acid and hydrogen peroxide and placed in a reaction kettle to be leached, so that acid leaching residue and acidic leaching liquid are obtained; the acid leaching residue is added with sodium hydroxide and sodium carbonate and subjected to alkali leaching, so that alkali leaching residue and alkali leaching liquid are obtained; and the alkali leaching residue is added with nitric acid to be leached, and lead enters the solution, so that the enriched precious metal is obtained after filtering and washing are conducted. The enriched precious metal is obtained from the electronic waste. The method for efficiently enriching the precious metal from the electronic waste is easy to operate, environmentally friendly, low in production cost, easy to industrialize, high in precious metal yield and good in application prospect.

The invention discloses a method used for improving the recovery of the molybdenum processing by fine sweeping section; the method comprises five selection operations such as fine sweeping I, II, III, IV and V; wherein, fine sweeping I foam and fine sweeping foam are returned to a slurry pond 2 together; fine sweeping II gangue is classified by a hydrocyclone and the overflow thereof is used as fine I given ore; fine sweeping I gangue enters a fine sweeping II flotation machine; the fine sweeping II flotation machine foam enters fine sweeping; fine sweeping II gangue enters a fine sweeping III flotation machine and is classified; the fine sweeping III gangue is re-classified by fine sweeping IV and used as final fine sweeping gangue; the foam thereof is re-returned to the fine sweeping III flotation machine to be classified; after the selection operation of the fine sweeping IV, a floatation column is provided so as to lead the fine sweeping IV gangue to directly enter a newly increased flotation column (fine sweeping V) and to be classified; subsequently, the gangue is directly used as final fine sweeping gangue tail and the foam as well as the fine sweeping III foam and fine sweeping gangue is returned to the fine sweeping II flotation machine and re-classified. The method leads the original fine gangue to be re-classified, reduces the grade of the final gangue and improves the recovery at the same time, and increases the yield of molybdenum concentrates.