404results about How to "Reduce the cost of beneficiation" patented technology

The invention relates the mineral separation process for low-grade collophane. The method comprises the following steps: carrying out dense-media dressing ore for low-grade collophane, getting the finished ore, then carrying out double reverse flotation, removing magnesium and silicon, and getting the phosphor concentrate. The method transforms the low-grade cellophane to high-grade concentrate, and recovery ratio is more than 80%, solving the problem that the low-grade cellophane is hard to utilize in industry.

The invention belongs to the technical field of efficient development of ultralow-grade ore resources and comprehensive utilization of associated minerals, and in particular relates to a combined beneficiation method and a combined beneficiation system for comprehensive recovery of associated iron-phosphate minerals. The combined beneficiation method for the comprehensive recovery of the associated iron-phosphate minerals comprises the following steps: sequentially performing three-section and two-closed-loop crushing and screening flow on raw ores, and performing dry separation to obtain fine ores; performing two-section closed-loop ore-grinding flow on the fine ores, performing fine magnetic separation for three times, and concentrating by using a wash mill to separate iron ore concentrate and discharge tailings; performing one-time roughing, one-time scavenging and three-time concentrating on the tailings at the temperature of 10-15 DEG C to obtain phosphate ore concentrate by floatation. Compared with the prior art, the invention provides a comprehensive recovery scheme for magnetic iron ore resources and phosphate ore resources in ultralow-grade associated iron-phosphate minerals. According to the scheme, the separation cost is low, the efficiency is high, energy is saved, the space occupation is low, and the comprehensive recovery of the magnetic iron ores and the phosphate ores can be realized.

The invention discloses a technology of enhanced-dispersion partial selective and bulk flotation of lead and zinc sulfide ores under low and high alkalinity. During grading of the lead and zinc sulfide ores, sodium hexametahposphate is taken as a dispersion agent and directly added into a ball mill, the ores are ground till the ores with the size of 0.074mm account for 67%-70%, lime is taken as an adjusting agent, zinc sulfate and sodium sulfite are taken as inhibitors, dithiophosphate BA and diethyldithiocarbamate are taken as collectors, and selective flotation of part of lead minerals with good floatability can be performed under low alkalinity; then xanthate and the diethyldithiocarbamate are taken as collectors, and the flotation of the lead minerals is further performed under high alkalinity; copper sulfate is added in lead flotation tailings for activation, butyl xanthate is further taken as the collector for flotation of zinc blende and part of pyrite, and zinc-sulfur separation flotation is further performed on zinc-sulfur mixed concentrate; and sulfuric acid is added in zinc flotation tailings for activation, and the xanthate is taken as the collector for flotation of the remaining pyrite. By adopting the technology, the lead-silver recovery rate can be improved, the using amount of lime and sulfuric acid can be reduced, the circulating amount of middlings can be reduced, the ore dressing cost can be reduced and the grade of the concentrate can be improved.

The invention discloses a beneficiation method for separating fluorite, white tungsten (black tungsten) and gangue minerals in complex polymetallic ore through flotation. Non-magnetic products of tailings subjected to sulphide ore flotation or tailings subjected to sulphide ore flotation and then subjected to strong magnetic separation for recycling black tungsten serve as two kinds of samples. The pH value is controlled by adding regulators, white tungsten, black tungsten, other gangue combined inhibitors and collectors are added, fluorite is subjected to differential flotation, then, tungsten minerals are subjected to flotation, and therefore fluorite and tungsten can be efficiently recycled. Loss of fluorite in tungsten flotation concentrate is avoided when tungsten minerals are subjected to differential flotation, and the defect that when tungsten flotation is performed, fluorite is difficult to recycle through flotation because the inhibitors have strong inhibition on fluorite, and beneficiation efficiency is low is avoided. Compared with a current beneficiation situation, the fluorite recycling rates of tests on the two different samples are increased by 39% to 48 % and 9% respectively, and the tungsten recycling efficiency is improved by 3%.

The invention belongs to the field of complicated copper-lead-zinc sulphide ore beneficiation separation and provides a copper-lead separation flotation inhibitor and applications thereof. The copper-lead separation flotation inhibitor is formed by mixing polymeric ferric sulfate solution with a mass concentration of 5.0%-6.0% and glycerin solution with a mass concentration of 0.8%-1.2%, wherein the mass ratio of the polymeric ferric sulfate solution to the glycerin solution is 1:1-4. Addition of the inhibitor in copper-lead separation operation can remarkably improve copper and lead concentrate qualities, evidently lower mutual containing of copper and lead, and improve flotation recovery rate of copper and lead minerals. The copper-lead separation flotation inhibitor and the applications thereof have the advantages that the copper-lead separation flotation inhibitor is environment-friendly, efficient in sorting, low in reagent cost, easy to operate and the like and applicable to the complicated copper-lead-zinc sulphide ore beneficiation separation.

The invention relates to a coarse grain pre-selection and magnetic-floating separation process for mixed ores with embedded micro-fine particles. The coarse grain pre-selection and magnetic-floating separation process is characterized in that a 'half-self-grinding, wet-type pre-selection, continuous ore grinding, weak magnetic-strong magnetic fine screening and re-grinding and strong magnetic reverse flotation' operating system is adopted to sort the mixed ores with embedded micro-fine particles, and iron ore concentrates with grade of 65%-65.5% and yield of 35%-40% are obtained. The coarse grain pre-selection and magnetic-floating separation process has the advantages that 15%-20% large-particle qualified tailings are obtained from the ores through the wet-type pre-selection, the magnetic ores with yield of 15%-20% and grade of higher than 65% can be separated in advance through magnetic-floating separation, the ore feeding amount of separation operation is reduced, the number of flotation machines is decreased, agent using amount is decreased, and ore separation cost is reduced.

The invention relates to a rutile ore part size fraction benefication joint production process. The joint production process comprises the steps of part size fraction magnetic separation, reselection, magnetic separation, desliming and grading, and flotation in sequence, so that the limit of only recovering rutile or comprehensively recovering garnet and rutile can be broken, and three main minerals including rutile, garnet and omphacite in the ore can be recovered comprehensively. By adopting the process, the rutile concentrate with TiO2 being in 90% grade and garnet and omphacite concentrate with purity being 88% can be obtained, the recovery rates of TiO2, the rutile, garnet and omphacite are 67.44%, 77.95%, 81.50 and 71.18%. Test results show that the main mineral resources in raw ores can be reasonably and effectively utilized with low flotation cost.

The invention relates to a refining process for lepidolite in tantalum-niobium ore waste rocks. A combined method with magnetic separation, high gradient magnetic separation, dense medium separation and flotation is adopted. The flotation adopts the process of rough flotation, dual fine flotations and sweeping. The process for refining the lepidolite from the granite-type tantalum-niobium ore waste rocks is provided. The tantalum-niobium ore waste rocks are crushed, screened, ground and subjected to high-frequency screening, spiral grading, permanent magnet magnetic separation for iron removal, high gradient magnetic separation, dense medium separation and flotation, so that lepidolite concentrates are obtained. By means of the method, the lepidolite concentrates are separated from the tantalum-niobium ore waste rocks. The lepidolite concentrates contain at least 4.5% of Li2O, at least 8.0% of K2O and Na2O and at most 0.10% of TiO2 and Fe2O3, the recovery rate of the lepidolite concentrates is larger than 80%, the qualities of the lepidolite concentrates meet the requirement of national standards, and the recovery rate of the lepidolite concentrates exceeds the industry level.

The invention provides an ore dressing method for lead-zinc-silver poly-metallic ores. Lead-silver concentrates and zinc concentrates are finally obtained through the steps of raw ore grinding, lead fast roughing, lead slow roughing, lead scavenging, lead fast concentration, lead slow concentration, zinc fast roughing, zinc slow roughing, zinc scavenging, zinc fast concentration and zinc slow concentration. The lead-zinc-silver poly-metallic ores sorted by the ore dressing method provided by the invention are good in sorting index and high in separation efficiency; the ore dressing method is a lead-zinc-silver ore dressing method which is stable, efficient, strong in adaptive capacity, good in sorting effect, high in comprehensive recovery index and low in production cost and is suitable for popularization and application.

The invention relates to a rutile beneficiation method comprising magnetic separation, tailing discarding reselection, desliming and fine particle floatation. According to the method, raw rutile is crushed and screened to have a certain feed size of ore grinding; obtained rutile particles are ground to a certain particle size through an ore grinding machine to obtain ore pulp with a certain concentration through preparation; the ore pulp is led to an intensity magnetic separator for separation to obtain magnetic minerals and non-magnetic minerals; the non-magnetic minerals are subjected to the reselection for desliming to obtain ore concentrate, and then the ore concentrate is dewatered to a certain concentration; the dewatered ore concentrate is subjected to the floatation operation, gangue mineral inhibitors, rutile activators, collectors and foaming agents are added in the dewatered ore concentrate, the rutile is subjected to the flotation, and after rough separation, 1-2 times of fine separation and scavenging are carried out in sequence, rough concentrate of the flotation rutile is obtained; the rough concentrate of the flotation rutile is subjected to the magnetic separation to eliminate magnetic impurities and then eliminate sulphur and phosphorus impurities through roasting and acid pickling, and finally high-grade rutile concentrate is obtained. The method has the advantages that the process procedure is simple, beneficiation cost is low, the recovery rate is high, and the concentration ratio is high.

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.

Disclosed is an ore dressing method of micro-fine particle tantalum-niobium ores. The ore dressing method is characterized by including steps of firstly, performing desliming pre-treatment to obtain set sand and micro-mud; secondly, adding an adjusting agent of sodium carbonate, an inhibiting agent of water glass, an activating agent of lead nitrate and collectors of benzohydroxamic acid and hydroxyamino acid to the set sand for once roughing; adding the benzohydroxamic acid and the hydroxyamino acid for once or twice scavenging; adding the water glass for once or twice concentration or blank concentration to obtain tantalum-niobium rough concentrates and flotation tailings; thirdly, performing de-reagent and de-foaming: adding sulfuric acid to the tantalum-niobium rough concentrates, and stirring the mixture until foams are disappeared; and fourthly, performing gravity concentration: subjecting the tantalum-niobium rough concentrates to de-reagent and de-foaming, and subjecting the tantalum-niobium rough concentrates to gravity concentration to obtain tantalum-niobium concentrates and gravity concentration tailings. According to the ore dressing method of the micro-fine particle tantalum-niobium ores, the recovery rate of the micro-fine particle tantalum-niobium ores is increased by over 30%, and the ore dressing method is applicable to primary mud and secondary mud with sizes of -0.037mm in tantalum-niobium ores.

The invention discloses a carbonate phophorite reverse flotation combined collecting agent and a preparation method thereof. The collecting agent is a compound anion collecting agent and is formed by mixing basic substance with water, wherein the basic substance mainly comprises the following components in parts by weight: 0.2-0.5 part of higher fatty acid sulfated soap containing unsaturated fatty acid, 0.5-0.8 part of higher fatty acid saponified matter with the fatty acid content of being not more than 50% and 0.01-0.1 part of foaming agent, and the ratio of the basic substance to the water in parts by weight is 1: (20-70). The combined collecting agent has the advantages of strong collecting capability, high selectivity, low dosage, low temperature resistance and the like and is applied to carbonate phophorite reverse flotation with the P2O5 grade of being 25-26%, the P2O5 grade of the obtained concentrate is more than 34%, MgO content is less than 1.2%, and P2O5 recovery rate of the concentrate is more than 90%. The invention has the characteristics of simple preparation process, easy operation, high concentrate P2O5 recovery rate, low MgO content in the concentrate and low flotation cost and is applicable to flotation production of carbonate phophorite and various phosphate ores.

The invention relates to the technical field of ore beneficiation, and in particular relates to an extremely lean magnetite beneficiating process. The process is characterized by comprising the following steps of: 1) carrying out wet pre-beneficiation on raw ores with grain sizes being 0-5mm and concentration being 35-45%; 2) processing the concentrates obtained through wet pre-beneficiation with a primary closed-circuit milling system formed by primary ball milling and primary classifying and discarding the tailings obtained through wet pre-beneficiation; 3) carrying out first-stage magnetic separation on overflow obtained through primary classifying; 4) processing the concentrates obtained through first-stage magnetic separation with a secondary closed-circuit milling system formed by secondary ball milling and secondary classifying; 5) carrying out second-stage magnetic separation on overflow obtained through secondary classifying; 6) carrying out third-stage magnetic separation on concentrates obtained through second-stage magnetic separation; 7) taking the concentrates obtained through third-stage magnetic separation as the final concentrates with grades being 63-64%; and 8) discarding the final tailings formed by the tailings obtained through first-stage magnetic separation, second-stage magnetic separation and third-stage magnetic separation. The process has the following characteristics that: 30% of tailings are discarded by adopting the pre-beneficiation method, thus improving the beneficiation feed grade of the ores to 16-18%, reducing the quantity of the ores entering the ball mill by 30% and obviously lowering the ore beneficiation cost; and the energy is saved and the economic benefits are obvious.

The invention relates to a method for recycling metal iron resources from waste residues of the iron and steel industry, particularly relates to a method for recycling metal iron from steel slag of a converter, and belongs to the technical field of benefication and comprehensive utilization of resources. The method includes recycling the metal iron from the slag of the converter by a cascade separation technological method, a dry magnetic separation technological method for discarding tailings and a wet concentration technological method, and is characterized in that +70mm large high-grade steel slag of the steel slag of the converter is in cascade separation and recycling, -70mm large high-grade steel slag is in classification dry magnetic separation for discarding tailings, rough concentration is further in full-grain wet ore grinding for separating slag from iron, cascade screening and cascade magnetic separation, and finally the metal iron with different grain sizes and magnetic iron concentration are obtained. The method has the advantages of high recycling rate and the grade of the metal iron, simplicity in technological process, simplicity and convenience in operation, low production cost, applicability to processing different types of steel slag of converters, and the like.

The invention discloses a beneficiation method for comprehensively using vanadium-bearing titanomagnetite. The method comprises the following steps of: performing mixed exploiting of ore above cut-off grade and ore below cut-off grade; crushing the mixed ore into ore with the granularity of 10mm completely, performing weak magnetic rough flotation or medium magnetic rough flotation on the crushedproduct to select out mineral aggregates of which the main component is the anadium-bearing titanomagnetite, performing strong magnetic scavenging on rougher tailings to select out the mineral aggregates of which the main component is ilmenite, combining the mineral aggregates into a pre-sorted concentrate, and discarding gangue of which the yield is 31.88 to 33.20 percent; performing I section coarse grinding the concentrate into concentrate with the granularity of 1mm completely, performing weak, medium and strong magnetic separation, discarding coarse gangue, performing II section fine grinding on the obtained coarse mixed concentrate to grind the obtained concentrate into concentrate with the granularity of 0.15mm completely, adding a modifier, a yellow catching agent and a foaming agent into the product of the II section ore grinding to perform flotation for a sulfide ore and obtain high-quality linnaeite; performing weak magnetic separation on titanomagnetite on tailings of the flotation for the sulfide ore to obtain high-quality iron ore concentrate; and performing strong magnetic titanium separation on tailings of iron separation, further gathering the ilmenite to obtain strong magnetic separation concentrate, and adding the modifier, sodium aliphatate and an auxiliary catching agent into the strong magnetic separation concentrate to perform flotation for the ilmenite and obtain high-quality titanium concentrate.

The invention discloses a magnetic ore rescreening process for mixed iron ore. The process comprises the following process steps: performing primary ore grinding, performing primary grading, performing primary low intensity magnetic separation, performing primary strong magnetic separation, rescreening, performing high frequency screening and performing secondary ore grinding and selecting. According to the invention, partial qualified iron ore concentrate and partial tailings are thrown out preferentially under the condition that the dissociation degree of useful mineral monomers is insufficient after the primary ore grinding operation, the amount of materials subjected to secondary regrinding and reselecting is reduced, and the aim of reducing the ore dressing cost is achieved.

The invention belongs to the technical field of lean hematite processing, in particular to a lean hematite processing technology which comprises the steps of stage grinding, coarse-fine separation, coarse fraction gravity separation, fine-fraction magnetic separation and centrifuge beneficiation. The lean hematite processing technology disclosed by the invention adopts centrifuge beneficiation to replace reverse flotation beneficiation in the original technology, so that the whole production technological flow is optimized and becomes shorter, and the loss of fine fraction high-grade hematite particles in tailings is reduced; besides, the technological course is more adaptive to properties of the hematite, so that the influence of over-high content of iron carbonate over reverse flotation in the original technological course is effectively solved; and furthermore, due to the adoption of centrifuge beneficiation, reagents does not need to be used and hematite pulp does not need to be heated, so that expenses of reagents and coal are effectively saved and the hematite processing cost is reduced.

The invention relates to a mineral separation method for recycling iron, rare earth, fluorite and niobium from iron tailings of associated multi-metal minerals, and belongs to the fields of mineral process engineering and comprehensive recycling of resources. The mineral separation method for recycling iron, rare earth, fluorite and niobium from iron tailings of associated multi-metal minerals comprises the following steps: carrying out low intensity magnetic separation and high intensity magnetic separation pre-concentration by taking bayan obo tailings as raw materials, then carrying out fluid bed roasting on obtained pre-concentrated concentrates, and carrying out low intensity magnetic separation operation on roasted minerals obtained by roasting so as to obtain weak-magnetism concentrates and rare-earth-containing weak-magnetism tailings; and carrying out rare earth flotation operation on the weak-magnetism tailings to obtain rare earth flotation concentrates and rare earth flotation tailings finally, then carrying out acid leaching process on the rare earth flotation tailings to obtain niobium-enriched slag and acid pickling tailings, and meanwhile, carrying out fluorite flotation operation on pre-concentrated strong-magnetism tailings to obtain fluorite flotation concentrates and fluorite flotation tailings. By the method, fluorite concentrates, iron core concentrates, rare earth concentrates and the niobium-enriched slag are obtained finally, and thus, the bayan obo tailings are utilized comprehensively.

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.

The invention relates to a method for dehydrating micro- particle slurry, in particular to a method for dehydrating the micro-particle products, which is suitable for ore-dressing practice. The method comprises the processes of grading, sedimentating and filtering. The method is characterized in that the micro-particle slurry is graded firstly in the dehydration process to obtain the particle slurry and fine particle slurry, wherein the coarse particle slurry is sedimentated via a thickening machine, and then is filtered by using a micropore material or device. After the fine particle slurry is sedimentated, a membrane chamber-type filter press or a centrifugal hydroextractor is used to dehydrate and filter. The method is easy and simple, solves the problems of low productive capacity of filtering the micro-sized particle slurry and enhances the sedimentating and filtering performance of the micro-sized particle slurry and the moisture of a filter mass so as to greatly reduce the energy consumption of metallurgy and the mineral separation cost.

The invention provides a mineral separation process and system for an ultra-lean magnetite ore. The mineral separation process for the ultra-lean magnetite ore comprises the following steps of sequentially carrying out coarse crushing, intermediate crushing and fine crushing on a raw magnetite ore to obtain a fine-particle ore; carrying out dry magnetic separation on the fine-particle ore, and throwing tailings to obtain a dry concentrate; and carrying out wet grinding and separation on the dry concentrate. The dry pre-separation step is additionally arranged in the mineral separation process, so that the technical problems of high ore grinding and magnetic separation energy consumption, high cost, high mineral separation water consumption, tailing pond construction and the like are solved.

The invention relates to the technical field of ore dressing, in particular to an ultra-lean hematite dressing process. The process is characterized by comprising the following steps of: performing primary grinding on the ultra-lean hematite with grade of 18 to 22 percent and granularity of 0 to 12 millimeters till the 200-mesh granule content of primary graded overflow reaches 60 percent, applying primary medium magnetism and primary strong magnetism to the primary graded overflow, abandoning the obtained tailings, performing secondary closed circuit grinding on the concentrate till the 200-mesh granule content of secondary graded overflow reaches 90 percent, performing primary concentration operation on the secondary graded overflow, then applying secondary medium magnetism and secondary strong magnetism to the secondary graded overflow, abandoning the tailings, performing secondary concentration operation on the mixed concentrate with grade of 40 to 50 percent, then performing rough flotation, fine flotation and third scavenging flotation, selecting final concentrate with grade of 67 to 68 percent, and abandoning the flotation tailings with grade of 12 to 13 percent. The process can develop and utilize the ultra-lean hematite with grade of 18 to 22 percent and reduce the dressing cost.

The invention relates to a magnetic-floating separation technology of micro-fine particle dissemination magnet-hematite mixed ore. The processes of wet-type semi-autogenous grinding, stage grinding, phase-flux weakening-strong magnetism, and strong magnetism concentrate reverse flotation are adopted, the anion reverse flotation operation adopts one-section flotation and roughing, one-section flotation and concentration and a three-section flotation and scavenging ratio closed circuit system to sort flux weakening concentrate and flotation concentrate and combine the flux weakening concentrate and flotation concentrate into the final concentrate, and the grade of the final concentrate is 65-65.5%, and the yield of the final concentrate is 35-40%; one-section strong tailings, two-section strong magnetism tailings and three-section scavenging operation tailings are combined into the final tailings, and the grade of the final tailings is 13-13.5%, and the yield of the final tailings is 60-65%. The magnetic-floating separation technology has the following advantages: the wet-type semi-autogenous grinding process is adopted, the moderate and fine crushing processes are reduced, the cost is decreased, stage grinding and stage separation are realized, magnetic ore is recycled in advance by adopting the single magnetic separation method, 30% of feeding capacity of the flotation operation is reduced, and the chemical dosage is decreased, and the effects of cost decreasing and benefit increasing are realized.