4579 results about "Ore concentrate" patented technology

The invention relates to a screen method for vanadium titano-magnetite with high quality, belonging to the ore screen field. The method adopts three stage-grinding and stage-concentration process for magnetic separation, wherein the field intensity of one-stage magnetic separation is 3000-4000 Gs; the field intensity of two-stage magnetic separation is 1800-2200 Gs; and the field intensity of three-stage magnetic separation is 1300-1700 Gs. The recovery rate of the screened iron ore concentrate and the titanium ore concentrate is high and the screened cost of the screened iron ore concentrate and the titanium ore concentrate is low, thus providing a new selection for the low grade vanadium titano-magnetite resource and having a broad application prospect.

The invention discloses a preselecting method of low-grade magnetic iron ore, comprising the steps of crushing the raw ore of the magnetic iron ore by coarse crushing, intermediate crushing and fine crushing operations, conveying to subject to ultra-fine crushing operation, and then crushing by a high-pressure roller mill, scattering the crushed product by a wet cylindrical sieve, conveying the ore on the cylindrical sieve to the vibrating sieve by a belt to perform wet gradation, conveying the ore on the vibrating sieve to a dry magnetic separator by the belt to discard the waste, returning the ore concentrate of the dry magnetic separator to the high-pressure roller mill to crush the ore concentrate again, perform magnetic separation and discarding tailings of the ore concentrate below the vibrating sieve and the ore concentrate below the cylindrical sieve in a wet coarse grain intermediate-field intensity magnetic separator, and selecting the coarse ore concentrate of the intermediate-field intensity magnetic separator by a grinding separation technology. The invention has the advantages that the invention selects quantities of qualified mine tailings by wet magnetic separation before grinding ore, reduces ore grinding work, reduces ore grinding energy consumption and has high system processing capacity; the invention can be widely used in magnetic ore dressing plant, and in particular suitable for treatment of preselecting low-grade magnetic iron ore with the iron grade lower than or equal to 22%.

The invention discloses a ferric powder composite block preparing technology, which is characterized by the following: separating fine ferric ore individually; preparing ball; blending rough particle with ferric powder and other raw materials; paving on the traditional sintering machine; inserting high-alkaline sintering ore of acid ball ore.

The invention discloses a beneficiation process of a molybdenum-tungsten oxide ore. The process of which is that: the liberation degree of ore crushing is higher than or equal to 70%; the flotation of molybdenum sulfide is preferred; the floatation tailing is classified and then a separation of slime and sand is carried out, wherein a slime and sand system takes alkali as the conditioning agent, sodium silicate as the depressant, and emulsified fatty acid as the collecting agent. The rough concentrate obtained from the separation of slime and sand is concentrated after being blended to obtain molybdenum-tungsten oxide ore collective concentrate. The collective concentrate is grounded again and added with the depressant, then a secondary concentrate can be gained through the concentration; the secondary concentrate is concentrated by means of Peter Roph method to obtain scheelite ore and oxidized molybdenum-tungsten beneficiation concentrate. The invention has the advantages that: 1. the invention discloses a beneficiation process flow for treating molybdenum-tungsten oxide ore; 2. the invention solves the difficulty of lowe technical specification of molybdenum - tungsten oxide ore beneficiation; 3. the invention completes laboratory tests and researches and fulfills industrialized implementation for the first time in China.

A process for preparing Ti-enriched material from V-Ti magnetite through preseparating to obtain Ti-Fe ore concentrate, proportionally mixing it with V-Ti-Fe core concentrate, adhesive and carbon reducer, smelting to obtain high-Ti slag and semi-steel, blowing V-Cr to melten alloy iron to obtain V-Cr contained steel slag, separating V and Cr, fire metallurge of high-Ti slag to obtain artificial rutile and microcrystal glass, ball grinding of artificial rutile and coal while adding adhesive to obtain C-contained Ti particles, calcining, cooling and classifying.

The invention relates to a method for reducing concentrate magnesium oxide content in copper-nickel sulfide ore flotation. The method includes steps of firstly adding complexing agent oxalic acid, citric acid, tartaric acid, quadrol and EDTA (ethylene diamine tetraacetic acid), cleaning active ions of copper and nickel on the surface of magnesium-contained gangue mineral such as serpentine, removing flotation activity of sulfide ore collector such as xanthogenate, and adding composition agent to suppress gangue mineral. Ore pulp is treated by complexing agent, the flotation environment is improved, and floatability difference between copper-nickel sulfide mineral and magnesium-contained gangue mineral is increased. By combined inhibitor, serpentine and talc gangue mineral are suppressed, and magnesium-reducing effect is better; nickel recovery rate is increased to some extent as compared with that of actual production. Meanwhile, content of magnesium oxide in concentrate is reduced, and melting cost is lowered.

The invention discloses a method for smelting and extracting platinum metal from an alumina-supported petrochemical catalyst, and the method comprises the following steps: mixing the alumina-supported petrochemical catalyst with a trapping agent, a reductant and a slag forming agent and then smelting to trap the platinum metal from the alumina-supported petrochemical catalyst to obtain subsidiary product and alloy containing the platinum metal; using a sulfuric acid solution to dissolve the alloy containing the platinum metal to separate noble metal and base metal so as to obtain noble metal ore; and refining the noble metal ore to obtain the platinum metal product. The method disclosed by the invention has the characteristics that the process design does not produce wastewater, the wastewater treatment and filtration difficulty in a wet treatment process of the alumina-supported petrochemical catalyst are avoided, the recycling rate is higher than 98%, and the product purity is 99.95%.

The phosphate rock floating process is one phosphate rock grading forward-backward floating process, in which the raw ore pulp is first separated into two parts of different fineness and then forward-backward floated in two separated branches. The coarse ore pulp is made to separated in high concentration so as to save floating chemical; while the fine ore pulp is floated through adding dispersant to raise the floating efficiency. The grading floating process can avoid the mutual interference between coarse and fine grains, and thus has raised separating efficiency, raised floated concentrate quality, lowered floating chemical consumption and raised technological and economic indexes.

The invention provides a method of performing chloridizing roasting to synchronously reduce and recover gold and iron from gold concentrate cyanide tailings, and belongs to the technical field of comprehensive utilization of resources. The method comprises the following steps: performing reverse flotation desilicication preconcentration on the gold concentrate cyanide tailings which comprise 0.5-0.95% of phosphorus, 0.9-1.4% of sulfur, 20-25% of iron, and 3.0-4.5 g / t of gold, drying the obtained concentrate, then mixing the concentrate with a chlorinating agent, a reducing agent, a dephosphorization agent and a desulfurizing agent, pelletizing, performing high temperature chlorination to volatilize gold and synchronously performing deep reduction to remove sulfur and phosphorus, collecting the exhaust gas to recover gold, and then performing crushing, ore grinding, magnetic separation on the roasted pellets to recover iron. With application of the method, a gold product and a sponge iron product can be obtained, wherein the gold recovery rate of the gold product is 92-96%; in the sponge iron product, the iron grade is 90-92.5%, the sulfur content is less than 0.06%, the phosphorus content is less than 0.07%, and the iron recovery rate is 88-93%. Through the reverse flotation preconcentration, the method can reduce the energy consumption of the follow-up treatment process and realize recycling of the low-grade gold and iron in the gold concentrate cyanide tailings.

The invention relates to the technical field of black metal ore exploitation and beneficiation, in particular to a beneficiation process of low-grade magnetite and specularite mixed ore, which is especially suitable for associated magnetite and specularite mixed ore of Zhouyoufang iron ore. The novel beneficiation process is realized according to an operation procedure and the process proceduressuch as ore crushing, one-stage grinding and grading, reselecting, one-stage weak magnetic separation, one-stage strong magnetic separation, two-stage grinding and grading (crude concentrate re-grinding), two-stage weak magnetic separation, two-stage strong magnetic separation, reverse floating and the like. The iron ore is comprehensively recycled, and a high-quality iron concentrate product is produced. The process has the characteristics of energy saving, high beneficiation efficiency and low beneficiation cost.

The invention discloses a method for treating low-grade tungsten concentrate and tungsten slag. The method comprises the steps of mixing low-grade tungsten concentrate or tungsten slag and coal powder with a reduction roasting enhancer, and then performing reduction roasting; grinding the product obtained through the reduction roasting, then performing neutral leaching to obtain a tungstate solution and leaching residues, and performing magnetic separation on the leaching residues by adopting a magnetic field so as to obtain iron concentrate and valuable metal tailings; removing silicon from the valuable metal tailings by using dilute hydrochloric acid and leaching out manganese from the valuable metal tailings by using concentrated hydrochloric acid in sequence, and then leaching out tantalum and niobium by using hydrofluoric acid to prepare tantalum and niobium products. By adopting the method, valuable elements including tungsten, iron, copper, manganese, bismuth, cobalt, tantalum, niobium and the like which are difficult to extract in the low-grade tungsten concentrate and tungsten slag can be efficiently enriched, separated and recovered effectively, so that the resource comprehensive utilization of the valuable elements in the low-grade tungsten concentrate or tungsten slag is realized; moreover, the method is simple in equipment, short in flow, simple and convenient in operation, is economical and reliable, and is favorable for industrial production.

The invention relates to a novel industrialized method for comprehensively utilizing vanadium titanium iron ore concentrate. Raw material, an additive and a solid reducing agent are made into a higher-density round ingot through a full automatic press after being mixed evenly by an edge mill; the mixed material is charged into a gradient alloy can and is directly reduced in a tunnel kiln; and after being crushed, the reduced ingot is ground and is magnetically separated to obtain high-quality sponge iron with 92 to 96 percent of TFe and material containing abundant vanadium and titanium with 2.8 to 3.4 percent of V2O5 and 48 to 49 percent of TiO2. After being cold formed, the sponge iron is used for electric steelmaking or is sold as iron powder through deep processing. The material containing abundant vanadium and titanium can be used for producing V2O5 with the purity more than or equal to 98.5 percent through a simplified wet method vanadium extraction process; and vanadium slag with the TiO2 content more than or equal to 48 percent can be used as sulfuric acid method titanium white or other industrial raw materials. The method can greatly improve the capacity of the tunnel kiln and the service life of a charging vessel, has high comprehensive utilization degree, reliable technology, investment conservation and low production cost, energy conservation and emission reduction, and little pollution, and creates conditions for industrialization.

The invention discloses a dry-grinding and dry-separation method of magnetite, which comprises the following steps of: feeding the crude ore of magnetite into a crusher for coarse crushing, and then feeding the crushed magnetite into a vibrating sieve for sieving; controlling the sieve meshes within a range of 35-75 mm; feeding the oversize product to the crusher for intermediate crushing and then feeding the crushed oversize product into a high-pressure roller mill along with the undersize product for fine crushing; breaking and grading the fine-crushed ore by a breaking-grading machine, thereby obtaining the products of three grain grades, i.e. coarse grains, fine grains and fine powder; returning the coarse-grain grade product to the high-pressure roller mill through an elevator, and feeding the fine-grain grade product to a dry type magnetic separator for magnetic separation and discarding tailings, wherein the magnetic field intensity is controlled within a range of 1200-3000 Gs; directly dryly stacking the discarded fine-grain grade tailings, and returning the fine-grain grade concentrate to the high-pressure roller mill through a belt conveyor and the elevator to form a closed loop; and feeding the fine-powder grade product into a cyclone magnetic separator, wherein the magnetic field intensity is controlled within a range of 600-1200 Gs, thereby obtaining concentrate or semi-concentrate. The dry-grinding and dry-separation method is low in power consumption, steel consumption and water consumption, and the tailings are easy to be disposed of.

The cooled agglomerated carbon-contained pellet prepared with vanadium titano magnetite concentrate and opposite direct reduction with former pellet comprises: mixing the concentrate, reducer and bonding agent, drying, grinding, and pelletizing to obtain the cooled agglomerated carbon-contained pellet drying and sieving into the reduction furnace with reducer for reduction atmosphere; loading the reduced pellet into electric furnace to melt and separate the iron liquid, V, Ti and Cr. This invention is fit to the pellet on wide application temperature and low energy consumption, has well efficiency and low running cost, and can also extract V, Ti Fe and Cr from electric furnace with high yield and obvious environmental protection.

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 relates to a method for extracting precious metal from a disabled auto-exhaust catalyst, which comprises the following steps: 1. lixiviating precious metal from the disabled auto-exhaust catalyst by a hydrometallurgy process and obtaining precious metal concentrates after permuting lixivium; 2. lixiviating slag, collecting precious metal of the slag by a pyrometallurgy process to obtain a precious metal phase and selectively lixiviating base metal in the precious metal phase to obtain precious metal concentrates; and 3 combining the precious metal concentrates obtained in the first two steps and refining the precious metal concentrates to produce platinum, palladium and rhodium products. The invention compensates the deficiency that the percent recovery of the precious metal is low by simply treating the disabled auto-exhaust catalyst with the hydrometallurgy process and has the advantages that the contents of platinum, palladium and rhodium in the waste slag are smaller than 1g / t and the product purity reaches 99.95 percent.

The invention discloses a method for extracting high arsenic complicated golden ore concentrate multielement. The method comprises the following steps that: the multielement complicated golden ore concentrate with extra-high arsenic is subjected to the roast dearsenification technique treatment, part of sulfur in the multielement high arsenic golden ore concentrate converted into sulfur dioxide and the arsenic converted into arsenic trioxide enter flue gas which is subjected to dust and arsenic collection, and then enter an acid making system to produce sulphuric acid; the multielement smelting slag and multielement low arsenic golden ore concentrate obtained by roast dearsenification, return products and flux are mixed to obtain copper matte regulus, and the like by adopting oxygen enrichment bottom blowing matte smelting gold collecting process to carry out the smelting process; the copper matte regulus is subjected to converting and refining to cast an anode plate which is sent to an electrolysis system to obtain an electrolytic copper product by refining; anode mud generated through electrolysis is sent to a noble metal refine system to produce gold and sliver, and the like; smelting slag and converting slag are subjected to floatation treatment to obtain copper and iron ore concentrate; and the generated gangue belongs to harmless slag and can be used. The method has the characteristics of wide range of the raw material application, high synthesized recovery efficiency of valuable element, and obvious economic and environment benefits.

The invention discloses a method for separating manganese and silver of manganese-silver ore and purifying manganese dipped solution thereof. The method comprises the following steps of: 1) pre-degrading and saccharifying plant byproducts comprising straws, hulls and slag; 2) reacting the product obtained in the step 1) with crude manganese-silver ore or enriched mixed concentrate dipped manganese; 3) separating, neutralizing and adsorbing the manganese dipped solution in the step 2) to obtain solution for preparing manganese sulfate and the like by further purification and crystallization; and 4) obtaining manganese dipped slag in the step 3), wherein the manganese dipped slag has high reaction activity, and the leaching time is short and the energy consumption is low during silver leaching treatment. The purifying method is also suitable for purifying the manganese dipped solution by adopting a reducing agent similar to rice bran and the like to treat pyrolusite or ferromanganese ore, and the treated manganese solution can be used for the production of manganese sulfate, electrolytic manganese and the like. The method has the characteristics of wide separation raw material source, reasonable process route, low equipment investment, low treatment cost, stable product performance and the like; and the manganese sulfate prepared by the method can meet the standard requirements of the industries such as chemical engineering, feed, agriculture and the like.

The invention discloses a beneficiation method for cu-pb-zn polymetallic ores, and belongs to the technical field of non-ferrous metal beneficiation. According to the technical scheme, the characteristic of complex cu-pb-zn-sulfur polymetallic ores with embedded fine-grained carbon level is utilized, a technological process of carbon-removing, cu-pb partial bulk flotation, cu-pb separation and zn-and-sulfur choosing in tailings is adopted, a concentration and fine grinding mode is adopted to achieve deep reagent removal, the influence of slurry on flotation is effectively eliminated, efficient collectors A8 and depressors T721 for cu-pb separation are used in the cu-pb partial bulk flotation process, separation and recovery of copper, lead, zinc and sulfur are effectively achieved, and concentrate mixing is obviously lowered.

The invention relates to a mineral separation process, in particular to a mineral separation process of complex fluorite difficult to separate. The process includes utilizing sodium hydroxide to treat oleic acid to obtain modified oleic acid, utilizing concentrated sulfuric acid to treat water glass and obtain acidized water glass, conducting ore grinding on fluorite ores according to the prior art, conducting coarse separation under the temperature of 25-30 DEG C, then conducting seven times fine separation on coarse separation foams and adding the acidized water glass into the foams in the first four times of fine separation to prepare fluorite concentrate. The process can improve separation efficiency of calcium fluoride, silica and calcite, improves quality and recycle rate of the fluorite ores, obtains high quality acid grade fluorite ores and resolves the problem of difficulty in separation of complex fluorite flotation, is remarkable in application effect in separation of the complex fluorite difficult to separate, has repeatability, simultaneously reduces requirements of fluorite mineral flotation for ore temperature, reduces energy consumption, reduces usage of collectingagent oleic acid, reduces environment pollution caused by exhaust of a large amount of waste water containing oleic acid and has better economical and social benefits.

The invention discloses a beneficiation method for a pure magnetic iron ore. A crude ore is sequentially treated according to the steps such as first crushing, first-stage dry-type magnetic dressing, medium crushing, second-stage dry-type magnetic dressing, fine crushing and third-stage dry-type magnetic dressing, fine ore is obtained after a great amount of gangues are thrown by the third-stage dry-type magnetic dressing and sent to a grinding head sieve to be griddled and separated, granules griddled by the grinding head sieve are treated in a first-section magnetic dressing process, rough ore griddled by the grinding head sieve is treated in a first-section ball grinding and then enters the first-section magnetic dressing process; ore aggregates obtained from the first-section magnetic dressing process are treated in a second-section ball grinding after discharging gangue sands till the granularity of ore powder is -200meshes being more than or equal to 60 percent, ore powder generated from the second-section ball grinder are treated in second-section magnetic dressing, ore aggregates generated from the second-section magnetic dressing are treated in a high-frequency sieve after discharging gangue sands, products positioned on the high-frequency sieve are filtered to obtain ore concentrate, and products positioned below the high-frequency sieve are treated with desliming and then treated with floatation. The invention adopts two sections of grinding ore open-way processes and distinguishes the ore aggregates during the ore dressing process, thereby greatly reducing the energy consumption of a system, lowering the production cost and enhancing the utility ratio of equipment.

The invention discloses a beneficiation method for low-grade mixed copper ore containing molybdenum. According to the properties and characteristics of the low-grade mixed copper ore containing molybdenum, the copper and molybdenum bulk floatation and re-separation principle technological process is adopted, and high-efficiency concentration of minerals is achieved by using a combination of butyl xanthate and a high-efficiency collector M201 during copper and molybdenum bulk floatation operation; high-efficiency reagent removal of bulk concentrates is achieved through a concentration dewatering and fine grinding method during copper and molybdenum separation operation so that the impact of a large number of reagents existing in copper-lead bulk concentrates on copper and molybdenum separation can be reduced; and moreover, copper minerals are suppressed by means of suppressant F-6, so that separation of copper and molybdenum is effectively realized, and the copper content and molybdenum content of the concentrates are effectively reduced. Meanwhile, in order to stabilize the floatation indexes, the floatation process is optimized for site production, namely, branching and streaming floatation; accordingly, the purpose of comprehensive recovery and utilization of mineral resources is achieved.

The invention discloses a method for recycling valuable elements from vanadium-titanium magnetite ore, which comprises the following steps of: crushing the ore or concentrate, adding sodium salt, performing oxidizing roasting, converting vanadium and chromium into water-soluble sodium vanadate and sodium chromate, performing water leaching in solution, and separating the vanadium and chromium from the solution to obtain vanadium pentoxide and chromium sesquioxide products; and adding coal dust into the leached residue for pelletizing, reducing in a rotary hearth furnace, magnetically separating iron and titanium, using the obtained magnetic iron powder as a raw material for powder metallurgy or steelmaking, and using a nonmagnetic product containing more than 50 percent of TiO2 as a raw material for extracting the titanium; or reducing iron from the leached residue in an electric furnace, using the obtained molten iron as a raw material for steelmaking and using electric furnace slag containing more than 50 percent of TiO2 as a raw material for extracting the titanium. The method is short in process flow and economical; and the recovery rate of the vanadium, chromium, titanium and iron is high.

The invention discloses a fine flaky-cryptocrystalline mixed graphite separation technique which comprises the following steps: primary rough grinding and rough separation, quintic regrinding and sextic concentrating, and middling centralized processing. According to the technique, in the rough separation step, the grinding fineness is -0.074mm, and the content is 90-95%; in the concentrating step, the regrinding fineness is -0.045mm, and the content is 90-95%; and thus, the graphite monomer is sufficiently dissociated, thereby sufficiently ensuring the fixed carbon content and recovery rate of the concentrated concentrate. The technique adopts gradually decreased low concentrating concentrations in multiple concentrating process to ensure the dispersity of ore slurry, thereby solving thesevere problem of entrapped impurities in floatation foam; and thus, in the final concentrate, the fixed carbon content is 91.35%, and the recovery rate is 91.30%, thereby solving the problems of lowtechnical indexes, low grade, low recovery rate and high cost in the fine flaky-cryptocrystalline mixed graphite concentrate separation, so that the fine flaky-cryptocrystalline mixed graphite ores are utilized in an efficient, economical and reasonable way.

Provided is a classified-desliming pressurized-feeding three-product dense medium cyclone coal preparation process. The process comprises the steps that wet classification and primary desliming are carried out on raw coal after prescreening, gangue predischarging and breaking are carried out on the raw coal, secondary desliming is carried out on the screen underflow of the raw coal after the screen underflow is concentrated, the oversize products of the secondary desliming and the screen underflow of the primary desliming are together fed in a pressurized three-product cyclone, three products are obtained after separation, and clean coal, middle coal and gangue are obtained after medium separation and dehydration are carried out on the three products. Part of qualified suspension liquid and dilute suspension liquid are obtained when the medium separation is carried out on the three products, magnetic concentrate is obtained after the dilute suspension liquid is recovered by a magnetic separator, and the magnetic concentrate and the qualified suspension liquid are directly fed in a medium blending barrel. The magnetic tailings of the dilute suspension liquid and the undersize mixtures of pre-desliming, the overflow of concentration classifying equipment and a secondary desliming screen enter a coal slime recovery processing system together. The process is high in separation efficiency, flexible in product structure, low in producing and operating cost and prone to achieving the large-scale single system, the modularization and the integrating intellectualization of plant layout.

The invention relates to a new technology for sorting a lean red magnetite ore containing iron carbonate, which is characterized by comprising the following steps of: enriching and discarding tailings by adopting a 'technology for stage grinding and strong magnetic discarding tailings', thereby obtaining a mixed magnetic concentrate; sorting the mixed magnetic concentrate by adopting an 'open-circuit reverse flotation technology with one-stage roughing, one-stage fining and two-stage scavenging', thereby obtaining a high-grade hematite and magnetite concentrate; adopting a 'close-circuit centrifuge technology with one-stage roughing and one-stage fining' for sorting a high iron carbonate ore enriched in mine obtained by adopting the 'open-circuit reverse flotation technology', thereby obtaining a siderite concentrate; feeding the centrifuge tailings into a weak magnetic machine and sorting; and recycling fine fraction magnetic concentrates. The new technology provided by the inventionhas the advantages that the lean hematite and magnetite ore containing iron carbonate is sorted by adopting a 'combined sorting process flow with stage grinding, strong magnetic discarding tailings, reverse flotation-centrifuge', so that the sorting indexes of the whole iron grade of the obtained comprehensive concentrate being more than 65% and the metal recovery rate being more than 65% are achieved and the maximization of utilization benefit of iron ore resources is realized.

The invention relates to a technique for efficiently removing iron from kaolin by magnetic separation, which is characterized by comprising the following steps: (1) purifying kaolin raw ore to obtain a magnetic separation raw material; (2) adding water to the magnetic separation raw material obtained in the step (1) to prepare a mixed slurry with the solid content of 10-20 wt%, adding sodium hexametaphosphate accounting for 0.2-0.6 wt% of the magnetic separation raw material, slurrying, stirring for 20 minutes, and passing through a 325-mesh wet sieve; (3) carrying out high-gradient magnetic separation on the undersize slurry of the 325-mesh sieve in the step (2): while controlling the magnetic field intensity of the high-gradient magnetic separator at 1.0-1.8T and the slurry flow rate at 0.8-1.8cm / s, the slurry is subjected to magnetic separation by the high-gradient magnetic separator to obtain a concentrate slurry; (4) adding alum accounting for 1-6 wt% of the concentrate slurry into the concentrate slurry obtained in the step (3), flocculating, carrying out pressure filtration, drying, and packaging to obtain a magnetic concentrate; and (5) calcining the magnetic concentrate obtained in the step (4) to obtain the calcined kaolin. The technique has the advantages of high iron removal efficiency and obvious whitening effect.

The invention discloses a method using mineral dressing backwater to float and recover gold, silver, lead and zinc in cyanidation tailings of a gold mine. The method comprises the following working procedures: (1) pulp pre-processing stage: adding mineral dressing backwater in cyaniding tailings, adding concentrated sulfuric acid to stir for 5 minutes, reacting the concentrated sulfuric acid withthe pulp until smoke is completely generated, adding activated carbon to remove reagent for 55 minutes, and then pumping to a flotation workshop section; and (2) bulk flotation stage: pumping the pretreated pulp into a 3-meter stirring buffer slot, stirring for 30 minutes continuously, flowing into a medicine injecting slot 1 and a medicine injecting slot 2 by proper motion, sequentially adding copper sulfate and diethyldithiocarbamate, stirring fully, then entering a floatation unit to perform roughing on the pulp, further carrying out selection stage by stage on the rough concentrate obtained by roughing to obtain qualified lead and zinc bulk concentrate containing gold and silver, and further carrying out scavenging stage by stage on the roughed tailing to obtain floated tailing which is the sulphur concentrate. The invention realizes comprehensive utilization of mineral dressing tailings and zero exhaust of the tailings, thus realizing cleaner production of the mineral dressing process.