RU2144429C1 - Method of dressing sulfide copper-and-nickel ores containing inherent minerals of platinum metals and magnetite - Google Patents

Abstract

FIELD: dressing of sulfide copper-and-nickel ores. SUBSTANCE: method consists in separation of inherent minerals of platinum metals into platinum-containing gravitation concentrate before conducting floatation dressing; at mass ratio of sum of sulfides and magnetite to sum of silicon and aluminium oxides in raw ore lesser than 1:2, ore is ground to size 30-65% of class lesser than 74 mcm and separation of inherent minerals of platinum metals is performed at maximum magnitude of centrifugal Froude number equal to 2 to 10 and ratio of this magnitude to pressure of fluidizing water of 0,025÷0,23 kPa^-1; in processing raw ore at mass ratio of sum of sulfides and magnetite to sum of silicon and aluminum oxides more than or equal to 1:2, ore is crushed to size of 60 to 95% of class lesser than 74 mcm; separation of inherent minerals of platinum metals is performed at maximum magnitude of centrifugal Froude number equal to 0.5 to 1.75 and ratio of this magnitude to pressure of fluidizing water ranging from 0,0058÷0,019 kPa^-1. EFFECT: increased extraction of platinum metals into gravity concentrate. 2 cl, 4 dwg, 2 tbl

Description

 The invention relates to the field of mineral processing by technology that combines the methods of hydromechanical separation of minerals in an artificially created force field and foam flotation, in particular to the beneficiation of sulfide copper-nickel ores containing proprietary minerals of platinum metals and magnetite, and can be used in the processing of sulfide platinum-containing polymetallic ores and industrial products in which platinum metals are represented by the actual mineral form.

 A known method of beneficiation of sulfide copper-nickel ores of the Norilsk-1 deposit interspersed in intrusive rocks. In these ores, both the mineral and scattered forms of the platinum metals associated with the sulfide component of the ore are of great importance. Own minerals of platinum metals (PM) are divided into 3 main groups: 1) native Pt-metals and their alloys, as well as alloys of Pt-metals with Fe, Ni, Cu, Co; 2) intermetallic compounds - compounds of Pt metals with Pb, Bi, Sn, Te, As, Sb; 3) sulfides, arsenides and sulfoarsenes of PT metals.

 In the known method, the initial ore is subjected to 4 stage crushing, wet grinding in ball mills operating in a closed cycle with spiral classifiers, and subsequent flotation to obtain a collective copper-nickel concentrate and dump tailings. Copper-nickel concentrate, in which platinum group metals are extracted along with non-ferrous metals, is divided by selective flotation into two sulfide products: copper and total nickel concentrates. Moreover, platinum metals as part of selective flotation concentrates undergo the full technological cycle of copper and nickel branches of the metallurgical redistribution (Polkin S.I., Adamov E.V. Enrichment of non-ferrous metal ores. - M .: Nedra, 1983. - P. 238-245; Genkin A.D., Distler V.V., Gladyshev G.D. et al. Sulphide copper-nickel ores of the Norilsk deposits - M .: Nauka, 1981. - P. 110-134).

The disadvantage of this method is the unacceptably high level of losses of platinum metals (PM) with waste technological products from the beneficiation and metallurgical industry. In this case, the bulk of platinum metals (35-40% of their content in the original ore) are lost with dump tailings. The main reason for the increased transition of PM to tails is the size of their mineral deposits and low flotation activity of non-sulfide mineral forms. In particular, for the Norilsk ores, the size of the PM mineral deposits varies from 1-5 to 150-200 microns, sometimes more. Studies have shown that ~ 7-10% of palladium is in fractions larger than 100 microns, not recoverable by flotation. In addition, the PM technological groups represented by natural alloys and intermetallic compounds, in which the phases based on Pt-Fe, Pt-Pd-Sn, Pd-Pb-Bi, etc., are most developed, due to their high physical density (13-19 g / cm ² ) and sizes for a long time is in circulation in the grinding cycle, where it is subjected to constant impact loads. Due to the high malleability of these minerals, this leads to their flattening and the formation of conglomerates with a high specific surface area and, as a result, to their transfer to the dump product. Uncoiled particles, upon reaching a certain degree of “flattening,” begin to tear and rivet on solid silicate minerals that go into the tails, and it becomes impossible to extract valuable metals into any intermediate product.

 Minerals of the third PM group, whose typical representatives are cuperite, sperrylite, braggite, and vysotskite, are very easily crushed, sludge to sizes <20 μm and are not recovered in the flotation concentrate.

 A study of the distribution of PM minerals in the tailings of enrichment by size classes showed the presence of two maxima: in the size range of ~ 25 μm and (more diffuse peak) in the region of 100-400 μm. Moreover, in a collective copper-nickel concentrate, particles with a size> 70 μm were not found. The data obtained indicate insufficient flotation of their own mineral forms of PM. Therefore, to increase their target extraction, it is necessary to intensify the flotation of thin sludge and to isolate large heavy PM particles by hydromechanical (in particular, gravitational) methods in the grinding cycle of ores and from tailings (Blagodatin Yu.V., Nikolaev Yu.M., Chegodaev V .D. On the possibility of additional extraction of platinum metals from the tailings of the beneficiation of Norilsk copper-nickel ores // Non-ferrous metals. - N 12. - 1995. - S. 58-60).

 A known method of beneficiation of continuous sulfide copper-nickel ores of the Talnakh and Oktyabrsky deposits (JSC "Norilsk Combine"). These ores are ~ 80% pyrrhotite type, which contains up to 60% of the minerals of the pyrrhotite group. In addition to pyrrhotite, solid ores contain chalcopyrite, pentlandite, cubanite, troilite, talnachite, magnetite and non-metallic minerals. The metals of the platinum group in this type of ores are present both in the proper mineral form and in the form of isomorphic impurities in the crystal lattice of the main carrier minerals: chalcopyrite, pentlandite and pyrrhotite. The prevalence of PM minerals in solid ores differs from disseminated ores both in the mineralization spectrum and in the size of the precipitates. In solid ores, PM intermetallides predominate and their sulfides (cuperite, braget and vysotskite), which are characteristic of disseminated and veinlet-disseminated ores, are completely absent. In addition, the separation of PM minerals in solid ores is much thinner than in disseminated ores, which makes them a more complex technological object from the point of view of using PM extraction by hydromechanical methods (Polkin S.I., Adamov E.V. Enrichment of non-ferrous metals. - M: Nedra, 1983. - S. 238-246; Genkin A.D., Distler V.V., Gladyshev G.D. et al. Sulphide copper-nickel ores of the Norilsk deposits. - M.: Nauka, 1981, - P. 110-134; Blagodatin Yu.V., Nikolaev Yu.M., Chegodaev VD On the possibility of additional extraction of platinum metals from dump tailings gascheniya Norilsk copper-nickel ore // Nonferrous metals -. N 12. - 1995. - pp 58-60).

 The known enrichment method includes ore preparation, wet grinding of material in ball mills operating in a closed cycle with spiral classifiers, sequential selective flotation of copper, nickel minerals and pyrrhotite into concentrates of the same name with the transfer of non-metallic minerals to dump tailings (Technological instruction for ore dressing at Talnakhsky processing plant . - TI 0401.14.39.-11-65-85. - The term of introduction is from 01.01.86. - MCM USSR. Soyuznickel. NMMC. - Norilsk: 1985. - 241 p.).

 The disadvantage of this method is the low extraction of PM in the target products of enrichment - copper and Nickel concentrates. The total level of PM extraction in these concentrates is 70-75%. In this case, up to 20-25% of the amount of PM goes into pyrrhotite concentrate and 3-5% into dump tailings. The transition of PM to pyrrhotite concentrate is undesirable, since during its subsequent autoclave-hydrometallurgical processing ~ 30% of platinum metals are lost with iron-hydrated tails, and part of PM in the composition of excess concentrate is sent for long-term storage to pyrrhotine storages, without falling into the scope of production.

 Another disadvantage of this method is that copper and nickel flotation concentrates go through a full cycle of metallurgical processing. This causes additional losses of PM with waste slag and dust from the smelting units, as well as ferruginous iron cakes of hydrometallurgical production.

 The closest to the proposed method for the totality of the characteristics and the achieved result is a method of beneficiation of disseminated sulfide copper-nickel ores from the Norilsk-1 deposit (Norilsk Combine JSC), including ore preparation, wet grinding of the material and its hydraulic classification, flotation separation of nickel sulfides and copper into the same selective flotation concentrates and the subsequent extraction of the PM's own minerals from the flotation tailings by the hydromechanical gravity method into an independent production CT - platinum-containing gravity concentrate. In this case, the separation of PM proper minerals from flotation tailings to a gravity concentrate is carried out on a centrifugal concentrate with a fluidized bed created by water jets in a direction that does not coincide with the centrifugal field strength vector (in the Knelson apparatus). Selective flotation concentrates and a platinum-containing gravity concentrate enter the corresponding cycles of metallurgical processing, and the discharge of the gravity concentrate containing minerals of the host rocks and magnetite is sent to the dump (Ivanov V.A. .- P.36) - prototype.

 The known method has a number of disadvantages.

 A significant disadvantage of this method is the extremely low efficiency of PM extraction from the tailings of the flotation operation, despite the use of such high-intensity gravity concentrators as Knelson centrifugal apparatuses. The extraction of the amount of PM in the gravity concentrate from their content in ore is only 3-4%. Increasing this indicator, ceteris paribus, requires a significant increase in the number of units of used gravity concentrators, which complicates the circuitry of the processing redistribution apparatus and reduces the profitability of processing copper-nickel ores.

 Another serious disadvantage of this method is the low degree of concentration of PM in the gravity concentrate, defined as the ratio of the amount of PM in the gravity concentrate to their content in the original ore. The level of this indicator during the gravity concentration of flotation tailings by a known method does not exceed 80-90.

 Due to these drawbacks, the known method is characterized by low target recovery of PM in commodity metal products, since up to 30-35% of the amount of PM is lost in the enrichment sector with dump flotation tailings and ~ up to 10-15% of the amount of PM goes into dump and difficult to recycle products of metallurgical processing of copper and nickel cycles - waste slag of smelting units, process dust and ferruginous cakes of cobalt production. Due to the low condition of the resulting gravity concentrate in terms of the total PM content (average ~ 500 g / t), its processing technology begins with the head smelting units for the production of electrolyte copper. Moreover, the through extraction of PM from the gravity concentrate into the anode sludge of copper production (target Pt - containing product) does not exceed 90%.

 The reason for the low efficiency of the known method is the peculiarity of the separation of native PM minerals in sulfide copper-nickel ores, the ratio of their density and strength characteristics with the accompanying minerals of sulfide mineralization, magnetite and host rocks, as well as the nature and intensity of mechanical impact on PM minerals in the chain of concentration processing apparatuses .

порядка 1: (2,5-3), имеют крупность до 80% класса минус 74 мкм (Технологическая инструкция. Обогащение руд месторождений Норильск-1, Талнахского и Октябрьского на Норильской обогатительной фабрике. - ТИ 0401.14.52-11-43-97. Введена 28.07.97 г. - С. 45-49 и 136). Вместе с тем, как свидетельствуют результаты исследований, материалы с подобным критериальным отношением требуют более грубого помола (не более 65% класса минус 74 мкм). Данное несоответствие параметров является одной из основных причин низкого извлечения ПМ в гравиоконцентрат при обогащении отвальных хвостов на аппарате конструкции Кнельсона. Средний показатель извлечения ПМ из песковой фракции обогащаемых хвостов в известном способе составляет: 12-15% платины и 0,7-15% палладия, что в пересчете на исходное содержание этих металлов в руде соответствует их извлечению в гравиоконцентрат ~ 3-4 и 0,2-0,4%.In particular, the authors of the present invention have found that for the efficient extraction of PM intrinsic minerals from sulfide copper-nickel ores, the installation point of the gravity concentrator in the ore beneficiation scheme is of fundamental importance. In particular, carrying out the operation of gravel enrichment of the material in the head of the technological scheme, for example, immediately after the first stage of ore grinding and classification of the initial ore, makes it possible to optimize the granulometric composition of the initial feed and prevent overmilling of PM minerals, which, as follows from the table. 2 experimental data, has a decisive influence on the performance of PM extraction in a gravity concentrate. With the release of PM minerals from dump tailings, this possibility is completely excluded and enrichment is carried out outside the optimal range of fineness. So, for example, during the enrichment of disseminated and cuprous copper-nickel ores of the Norilsk deposits, they are subjected to 2 stage grinding. As a result of this, tailings characterized by a criterion relation

about 1: (2.5-3), they have fineness up to 80% of the class minus 74 microns (Technological instruction. Ore dressing of the Norilsk-1, Talnakhsky and Oktyabrsky deposits at the Norilsk concentrator. - TI 0401.14.52-11-43-97 Introduced on July 28, 1997 - pp. 45-49 and 136). At the same time, according to research results, materials with a similar criterion ratio require coarser grinding (not more than 65% of the class minus 74 microns). This mismatch of parameters is one of the main reasons for the low extraction of PM in the gravity concentrate during the enrichment of tailings on a Knelson apparatus. The average PM recovery from the sand fraction of the enriched tailings in the known method is: 12-15% platinum and 0.7-15% palladium, which, in terms of the initial content of these metals in the ore, corresponds to their extraction in a gravity concentrate of ~ 3-4 and 0, 2-0.4%.

 The task of creating gravity-flotation technology for the enrichment of sulfide copper-nickel ores with the separation of the bulk of its own PM minerals into a high-quality gravity concentrate at the stage preceding the flotation operation (to the head of the scheme) has not been set up to date. This was mainly due to 2 circumstances: insufficient study of the quantitative ratio of the forms of the presence of PM in these ores and the lack of effective industrially applicable methods for the separation of PM minerals from polydisperse sulfide materials in which the components to be separated have fairly close density characteristics. It should be noted that sulfide copper-nickel ores containing intrinsic minerals PM are a very complex object for gravity concentration methods. The deposits of these ores include: Canadian deposits in Sudbury, a group of South African deposits and the Norilsk ore region (Borbat V.F. Metallurgy of platinum metals, - M .: Metallurgy, 1977. - P. 30-34).

 When developing gravity-flotation technology for the enrichment of sulfide copper-nickel ores, a number of concentrators were initially tested, based on the principle of separation of a mixture of minerals with different density characteristics in the field of gravitational force. Tests: hydraulic traps, flushing locks, jigging machines, spiral separators and concentrated tables, widely used in the enrichment of gold sands (placers), as well as tin and tin-tungsten placer deposits (Shokhin V.N., Lopatin A.G. Gravity enrichment methods - M .: Nedra, 1980 .-- 400 p.).

 None of the types of the above devices showed sufficient efficiency so that it can be industrially used for the extraction of PM minerals from sulfide copper-nickel ores.

 Wash locks proved to be ineffective for separating finely dispersed particles with a high orientation coefficient resulting from the regrinding of fragile PM minerals (sperrylite, cuperite, braggite, vysotskite, etc.). A critical factor for this separation system was the particle shape of platinum minerals. The presence in the pulp of the crushed ore of relatively large grains of the host rock (100-150 microns) and thin sulfide sludge (10-30 microns) causes clogging of the working elements of the lock, which after 10 minutes leads to a significant decrease in its effectiveness, and after 4 hours In the course of operation, the degree of PM mineral retention in the gateway becomes insignificant.

 Jigging machines have several advantages over flushing locks. In particular, they support (with accurate balancing) a pulsed annealed layer, as a result of which a higher PM content in gravity concentrates is achieved. The efficiency of the jigging machine depends on several factors: the density of the feed stream, the speed of the pulp stream, the pulse frequency, the flow rate of the washing water, the feed rate of the washed concentrate, as well as the composition of the suspension and, in particular, the shape and particle size of the PM's own minerals. This type of apparatus is ineffective for enriching flocculent (uncoiled in a mill) particles of metal platinum and its alloys with other metals. This is due to the fact that the final velocity of such particles is much lower than that of spherical particles of the same mass, and their shape impedes their movement during the seepage phase. In addition, jigging machines do not provide a sufficient level of PM extraction in a gravity concentrate if the particles of extracted minerals are represented by a wide dispersion range and have sizes less than 100 microns. Therefore, for sulfide copper-nickel ores, characterized by the presence of minerals with significantly different density characteristics, as a result of which the crushed material has an increased level of polydispersity, and the bulk of the PM of the third group (sulfides, arsenides, sulfoarsenides) are represented by particles smaller than 74 microns, the use of jigging machines is irrational. As in all systems operating under 1g acceleration (in the field only gravitational forces), the probability of deposition of fine PM particles with a large orientation coefficient is extremely small. Setting the jigging machine to thin classes of PM minerals makes it impossible to achieve a degree of PM concentration at the stage of primary gravity enrichment of more than 100.

 Spiral separators are characterized by relatively high productivity and, in the presence of large mineral deposits of PM in the raw materials, they ensure the production of gravity concentrates with a good degree of enrichment. A serious drawback of this type of apparatus is their low efficiency with respect to polydisperse materials containing fine particles of PM minerals. In this case, it is necessary to use a large number of separation stages, as a result of which the gravity concentration scheme becomes cumbersome, and the enrichment process becomes capital intensive. As a rule, to capture fine particles of an enriched mineral, individual separators are grouped into aggregates (up to 12 devices each), which give a fairly high degree of extraction, but with a low quality of the obtained gravity concentrate (A. Laplant. Use of gravity concentration for gold enrichment. - Part 1. - Efficiency and basic principles. The article is presented at a professional development seminar in the field of small-scale gold processing projects: Mining, Processing. Economics and Politics. Mac University -Gilla. - May 2-6, 1986).

 Concentration tables have the same drawback as spiral separators: they are ineffective in extracting PM minerals from the ore with a grain size of less than 100 microns, therefore this type of apparatus in industry has found application not for primary gravity concentration, but for cleaning rough concentrate.

Thus, all types of devices considered, the enrichment of raw materials in which are carried out in the field of gravitational force, are ineffective for sulfide copper-nickel ores. This is due to the following factors:
- concentration of a significant mass of PM in thin classes (minus 44 microns) of the enriched material;
- the presence in the shared system of magnetite and sulfides of non-ferrous metals having a density close to the density of “light” minerals PM;
- a high coefficient of orientation of particles containing intrinsic PM minerals (brittle PM minerals, such as sperrylite, cuperite, etc.) are represented by acute-angled fragments; while malleable minerals, such as native PM and their alloys, in the crushed ore look like thin flat flakes );
- the presence of PM minerals in intergrowths with less dense minerals of the host rocks.

 For the isolation of Pt - minerals that are in highly refractory forms - highly dispersed, with a high orientation coefficient or contaminated with inclusions with a low density - gravity concentration methods in a centrifugal force field are more promising.

The most famous centrifugal apparatus with high separation factors are hydrocyclones and centrifuges. Creation of a centrifugal field in centrifugal concentrators in principle can be done in two ways:
- tangential flow under pressure into a closed and stationary cylindrical vessel;
- by twisting the freely supplied flow in an open rotating vessel.

 A prerequisite for centrifugal gravitational enrichment is the presence of a transport (flush) flow in a direction that does not coincide with the centrifugal field force vector. In the absence of a flush flow, as well as in the case where the flow direction coincides with the direction of the field, material stratification by density practically does not occur.

Centrifugal concentrators can basically be divided into two types:
- pressure cyclone apparatus for the separation of fine-grained materials;
- non-pressure apparatuses - centrifuges (with a low centrifugal field intensity) for separating both coarse-grained and fine-grained materials.

 The work of the concentrators of the second type, although it resembles the work of a conventional centrifuge, however, differs significantly from it in the presence of elements of a conventional sluice process (Shokhin V.N., Lopatin A.G. Gravity enrichment methods. - M .: Nedra, 1980. - P. 350 -352).

 The use of centrifugal concentrators (cyclone apparatuses and centrifuges) in the operation of primary enrichment of sulfide copper-nickel ores showed higher results in the extraction of PM in a gravity concentrate than when using devices based on the action of gravitational force. However, in this case, too, coarse gravity concentrates were obtained, the degree of enrichment of which in PM did not exceed 200. With an increase in the degree of polydispersity of the starting material, the PM extraction in the gravity concentrate markedly decreased. The extraction of PM at the level of 15% was achieved by combining 7 cyclone apparatuses into a battery, however, this negatively affected the concentration ability of the installation: the degree of enrichment of the gravity concentrate by PM was 55.

 In the study of cyclone-type pressure concentrators in the primary enrichment operation of sulfide copper-nickel ores, it was found that the regulation of bed porosity in the cyclone cone makes it possible to increase the release of PM intrinsic minerals into gravity concentrate. This was achieved by injecting weak streams of water into the lower part of the cone. The supply of additional water led not only to an increase in the quality of the concentrate (PM enrichment increased to 300), but also to an increase in the extraction of mineral forms of PM in it, which in this case reached 15–20%.

 The results presented were the basis for the creation of the present invention.

 The problem solved by the invention is to increase the completeness of the extraction of platinum metals in the gravity concentrate obtained by the hydromechanical separation of their own minerals of platinum metals in the gravity-flotation concentration of sulfide copper-nickel ores, as well as to increase the degree of enrichment of the gravity concentrate for platinum metals by changing the sequence of operations enrichment and flotation in the ore processing scheme and by improving the regime of the hydromechanical concentration stage, igaemogo by optimizing the relation between the process parameters and by regime criterial ratio of mineral ingredients in the composition of the original ore processed.

The essence of the invention lies in the fact that in the method of beneficiation of sulfide copper-nickel ores containing proprietary minerals of platinum metals and magnetite, including ore preparation, wet grinding of the material and its hydraulic classification, separation of non-ferrous sulfides and proprietary minerals of platinum metals from the pulp of the classified material by flotation and gravitational methods into independent products - sulfide (e) flotation concentrate (s) and platinum-containing gravity concentrate, and magnetite and rock nascent minerals are dumped tailings, and platinum native minerals are isolated on a centrifugal concentrator with a fluidized bed created by water jets in a direction that does not coincide with the centrifugal field force vector, according to the invention, platinum native minerals are released into a platinum-containing gravity concentrate before the flotation concentration operation is carried out, at the same time, with a mass ratio of the sum of sulfides and magnetite to the sum of silicon and aluminum oxides in the initial ore, less than 1: 2, I grind the ore t to a particle size of 30-65% of the class less than 74 microns, and the extraction of their own minerals of platinum metals is carried out with a maximum value of the centrifugal Froude criterion equal to 2-10, and the ratio of this value to the pressure of the fluidizing water equal to 0.025-0.23 kPa ^-1 , in the processing of the original ore with a mass ratio of the sum of sulfides and magnetite to the sum of silicon and aluminum oxides greater than or equal to 1: 2, the ore is crushed to a particle size of 60-95% of a class less than 74 microns, and the extraction of intrinsic platinum metal minerals is carried out at a maximum centrifugal value the criteria Froude equal 0,5-1,75 and this value against a pressure fluidizing water equal 0,0058-0,019 kPa ^-1.

 Another difference of the method is that the allocation of their own minerals of platinum metals is carried out from the pulp of the material with a mass ratio W: T equal to (1.5-6.5): 1.

 Studies have shown that the separation of PM from sulfide copper-nickel ores, regardless of the type of ore, its relationship with the host rocks and structural and texture features, most effectively occurs in a centrifugal force field modulated by counter-pressure of the directed flow (jets) of the liquid medium. This condition is provided in various models of centrifugal gravity concentrates of a new generation, of which the Knelson construction apparatus has received the widest application in enrichment practice (Fig. 1).

 Knelson centrifugal gravity concentrators are a centrifuge chamber (Fig. 1) surrounded by a water jacket (2). Food - classified raw material suspension - enters a perforated rotating cone (1) equipped with partitions (corrugations) - (4), where heavy particles of minerals in the form of a wall layer (5) accumulate in the recesses between the corrugations. The feed is loaded into the bottom of the cone through a centrifugal feed tube (3). Compaction of the material between the corrugations is prevented by the injection of water through the holes in the centrifuge body (6) of 2-3 mm in size, which give the jets of water a tangential direction towards the rotation of the cone. Directional jets of water throughout the process maintain the material of the centrifuge wall layer in a fluidized state, allowing denser particles to replace less dense ones. Water injection is the key to a Knelson concentrator. The degree of liquefaction (back pressure of water) allows you to control the bulk density of the concentrate and its porosity, and, therefore, the ability to control which minerals should (or should not) enter the concentration layer. There is a general rule for choosing the operating mode of a concentrator: low water pressure is used for raw materials with low density, and high pressure - for raw materials with high density. When processing coarse raw materials, higher pressure is required, since more water is required to liquefy a layer with high porosity. During the operation of the apparatus, the material rises upward along the wall of the rotating cone. In this case, the pulp of depleted material (tails of gravity concentration) over the edge of the cone is poured into a stationary drum (7) and flows through the discharge pipe (8) into the collection tank. The main controlled parameters of the process are: feedstock fineness, degree of pulp liquefaction, centrifuge rotation speed and fluidizing water back pressure. The installation is semi-continuous: every few hours of operation, it must be stopped to unload the gravity concentrate. The duration of the material accumulation cycle depends on its composition, solid process productivity, requirements for the conditions of the resulting gravity concentrate and a given level of extraction of valuable components. Unloading time for industrial machines is approximately 10 minutes (Laplante AR, Lui L., Cauchon A. Mineralogy and Fiowsheet Changes at the Camchim Mines Inc. Mill, AIME Spring Meeting. - Las Vegas, Fe. 1989 .; Veioo C. Knelson Concentrator Test Program - Executive Summary and Reccommendations // Intenal Westmin report. - Sept., 1991).

 For a clearer understanding of the essence of the invention, it is necessary to consider the main known patterns of operation of centrifugal gravity concentrators and the behavior of mineral particles of different sizes and densities in the field of centrifugal forces modulated by the pressure of the flow (jets) of water oriented in a direction that does not coincide with the centrifugal field force vector.

The separation efficiency of heterogeneous systems in the field of centrifugal force is determined by the separation factor (f), which is the ratio of the acceleration of centrifugal force to the acceleration of gravity:
f = ω ² R / g, (1)
where ω is the angular velocity of rotation of the centrifuge, s ^-1 ;
R is the radius of the centrifuge, m;
g is the acceleration of gravity, m / s ² .

где n - частота вращения центрифуги, с^-1;
D - внутренний диаметр центрифуги, м.The separation factor is a modified (centrifugal) Froude criterion (Fr _c ) and is associated with it by the ratio:

where n is the rotational speed of the centrifuge, s ^-1 ;
D is the inner diameter of the centrifuge, m

 (Romankov P.G., Kurochkina M.I. Hydromechanical processes of chemical technology. - L.: Chemistry, 1982. - P. 150-182; Pavlov K.F., Romankov M.G., Noskov A.A. Examples and tasks on the course of processes and apparatuses of chemical technology. - Chemistry, 1981. - S. 97-98).

 In a Knelson concentrator, in contrast to conventional centrifuges, along with the centrifugal force, the counterpressure of the fluidizing water flow acts simultaneously with the centrifugal force, which simultaneously plays the role of a transport (flushing) stream. In this regard, a device of this type greatly enhances and improves centrifugal factors: it increases the force of gravity tens of times, but at the same time it does not allow compaction of the precipitate obtained.

As in any gravitational enrichment device, larger and denser particles of minerals are concentrated on the lower rings of the centrifuge, while finely dispersed particles of the same minerals, the deposition of which is difficult, are held by the upper rings, which have a higher value of the centrifugal Froude criterion. Due to the fact that the material enclosed in the rings is injected with water in the Knelson construction hub, even very small particles of dense minerals with a high orientation coefficient are able to penetrate and displace less dense particles of the same linear size or volume. This ability of concentrators of this type is extremely important in the enrichment of sulfide copper-nickel ores, in which part of the platinum metal minerals after the ore crushing stage is represented by refractory forms: flat (unchained) particles of metal alloys and slimy particles of brittle minerals (sulfides, arsenides and
The difficulty of gravity concentration of sulfide copper-nickel ores also lies in the fact that they contain minerals with an almost continuous spectrum of density characteristics (Table 1): the minerals of the host rocks have a density of 2.6 to 4.4 kg / dm ³ ; basic sulfides of non-ferrous metals and iron - from 4.0 to 5.0 kg / dm ³ ; magnetite - 5.2 kg / dm ³ ; nickel arsenides - from 7.05 to 8.1 kg / dm ³ . The density of PM minerals accompanying this type of ore varies from 5.0 (Ru, Rh - pentlandite) to 21.5 kg / dm ³ (native platinum), i.e. with its lower limit, a number of PM minerals are in contact with the upper limit of the density range of the accompanying sulfides and magnetite. In addition, the crushed ore is characterized by an extremely high degree of polydispersity of both the main ore-forming minerals and PM minerals: the size of the bulk of those and other particles lies in the range from the first microns to 1-3 mm.

It was shown above that all known types of gravitational apparatuses used in the beneficiation of minerals of ore origin proved to be ineffective at beneficiating such a complex object as sulfide copper-nickel ores. The main advantage of the Knelson concentrator, which fundamentally distinguishes it from centrifugal apparatuses of the previous generation, is the possibility of replacing heavy, but less dense particles, lighter, but having a higher density. It is this property that determined the high efficiency of this type of concentrator for the enrichment of sulfide copper-nickel ores, in which a significant part of the heavy minerals PM (8.4-10.6 kg / dm / ^{3 are} represented by slimy particles (less than 10 microns), and light rock (2.6-4.4 kg / dm ³ ) with large grains up to 3 mm in size The indicated property in the Knelson apparatus is achieved by the fact that water is not only injected from the inside of the material enclosed in the rings, but also tangentially, opposite to the direction moves inside the centrifuge, and so far liquefaction and displacement of the material occurs, and the predominant deposition of dense particles (PM minerals) proceeds. The tangential injection of water in the direction opposite to the rotation of the centrifuge allows the material to be maintained in constant liquid rotation, preventing it from gaining cone speed and occurring. any denser particle than those that are in the rings, infiltrate and replace less dense ones (Knelson B.V. Gravity dressing and separation of precious metal ores. - 17th annual congress of SMR. - Ottawa, 1985 .-- S. 346-357).

 The most critical issue during enrichment in a Knelson concentrator is the balance of forces acting on particles of material from the centrifugal field and the flow of fluidizing water. By changing the value of the centrifugal Froude criterion, it is possible to increase the difference in density between particles of the separated minerals tens and hundreds of times, which ensures a high level of extraction of PM minerals in the resulting precipitate. On the other hand, in order to increase the selectivity of the process in a Knelson concentrator, conditions are created that impede the compaction of the material. This is ensured by the injection of water through the wall of the cone, as a result of which 100% of the operating time of the apparatus, the layer is in a fluidized state.

 This is where the so-called the problem of the “pendulum”, which is that when the injected water is not used enough, the sediment becomes denser, causing a sharp decrease in its enrichment in PM, while at the same time, when the excess water is supplied, all injected material is ejected from the centrifuge into the drain slurry and the gravel enrichment process ceases.

 The density values of the main minerals that form the basis of sulfide copper-nickel ores (Mineralogical tables. Reference book // E.I. Semenov, O.E. Yushko-Zakharova, I.E. Maksimyuk et al. - M .: Nedra, 1981, - 399 s.) Are given in table. 1.

 Thus, each enriched material in a Knelson concentrator corresponds to the optimal value of the centrifugal Froude criterion and the ratio of this criterion to the pressure of the injected water. Moreover, all 100% of the time in the layer should be a continuous exchange of matter (particles with a low density should be exchanged for denser ones), and the liquefaction should be kept constant, balanced and stable, and the sediment should never be at rest.

 A fundamental important design feature of conical centrifugal concentrators is that the radial and axial velocities of the suspension are different at different heights. In the lower section of the apparatus of radius r, the average axial flow velocity is higher than in the annular gap (R-a) at the discharge level (where a is the radius of the loading pipe). At the same time, the centrifugal field strength at the lower level (at ω = const) is less than at the upper level by R / r times. Given the fluid slip, this difference is even greater.

 It follows that the flushing force of the flow decreases sharply as it passes along the wall of the concentrator, while the centrifugal force in the rotating flow increases, i.e. material enrichment in centrifugal concentrators proceeds with a variable parameter that increases along the flow and reflects the ratio of the strength of the flush (fluidizing) flow and the centrifugal Froude criterion.

 A similar phenomenon leads to the fact that the state of the wall layers of the flow turns out to be unequal in height of the centrifuge. In its lower part almost no bed of mineral grains is formed. All grains, with the exception of the heaviest ones, experiencing a strong action of a flushing (fluidizing) flow and a relatively weak action of centrifugal force, are carried away upward. As the flow moves along the generatrix of the cone, the dynamic pressure of the flow weakens, and the centrifugal force and, accordingly, the Froude criterion, on the contrary, increase. Grains from the stream fall onto the wall and form a wall bed, the loosening of which gradually decreases, and the layer thickness and grain size of its constituents increase. This layer of mineral grains serves as a capture coating, similar to how it is carried out at concentration locks. The difference between centrifugal enrichment in a centrifuge-type apparatus and a gravitational airlock is the inconstancy of the “flushing” (“fluidizing”) and “pressing” forces acting in the flow along the length of the apparatus formed and, as a result of this, different size, looseness and roughness of the wall mineral bed ( Shokhin V.N., Lopatin A.G. Gravity methods of enrichment. - M .: Nedra, 1980. - S. 356-357).

 From formula (1) it follows that conical centrifugal concentrators have an upper section (at the discharge level) with respect to hard-to-extract PM minerals, characterized by high dispersion, high orientation coefficient, and located in intergrowths with light minerals of the host rocks. In this section, the separation factor of the apparatus and, accordingly, the Froude centrifugal criterion take maximum values, while the “fluidizing” factor and, accordingly, the force of the wash-off flow become smaller than in lower-level sections. Therefore, when determining the optimal conditions for enrichment on centrifugal concentrators of similar geometry (with a radius of the cross section varying in height), it seems appropriate to set the main parameters of the apparatus for a cross section with a maximum radius. In Knelson design concentrators, this section is the discharge level, where the centrifugal Froude criterion assumes the maximum value.

согласно результатам статистической обработки экспериментального массива данных составляет соответственно 1:2 (где MeS и Mnt - массовое содержание сульфидов и магнетита в исходной руде, %).It was experimentally established that for sulfide copper-nickel ores there is a boundary value for the criterion ratio of the sum of sulfides and magnetite to the sum of silicon and aluminum oxides (hereinafter referred to as the “criterion ratio”), which distinguishes the entire spectrum of ores of this type into 2 subgroups according to the conditions of gravity concentration. Each of these subgroups is characterized by its ranges of optimal values of the centrifugal Froude criterion and the ratio of the Froude criterion to the pressure of fluidizing water. Two points were unexpected: firstly, the indicated ranges are practically isolated (the region of operating parameters separating them is characterized by a significantly lower quality of the obtained gravity concentrates), and secondly, for materials with a high criterion ratio (rich ores) optimal was the work of centrifugal gravity concentrates at lower values of the Froude criterion than for ores with a low criterion ratio (disseminated and poor cuprous ores, as well as mixtures thereof). The boundary value of the criterion relation

according to the results of statistical processing of the experimental data array, it is 1: 2, respectively (where MeS and Mnt are the mass content of sulfides and magnetite in the initial ore,%).

 The need for a qualitative (spasmodic) change in the conditions of centrifugal ore dressing when moving from a subgroup with a low criterion ratio to a subgroup with a high value of this parameter is apparently due not so much to the difference in the size of the mineral deposits of PM, which occurs for disseminated and rich varieties of sulfide ores, but the difference in strength and density characteristics of ore-forming minerals and their associations with the mineral phases of the host rocks.

 In gravitational enrichment, the grain size of light and heavy minerals, their density, as well as the ratio of the concentrations of heavy (recoverable) minerals and intermediate density minerals are most important. The latter in this case include sulfides (copper, nickel, iron) and magnetite. The enrichment of materials containing a significant amount of heavy grains of intermediate density, passes with low efficiency. Therefore, the enrichment of continuous sulfide copper-nickel ores (with a high criterion ratio) requires a higher degree of grinding of raw materials than the enrichment of disseminated ore.

 Coarse grinding of ore with a low criterion ratio (a high proportion of rock) is due to two factors. Firstly, it was established that the minerals of the host rocks have a higher strength than sulfides and arsenides of PM. When grinding ore with a high rock content, the latter has an active effect on the destruction of these mineral forms of PM. The deep grinding of disseminated ore leads to the fact that the minerals of the rock contribute to the sludging of the bulk of brittle minerals PM (sperrylite, cuperite, braggite, etc.), as well as flatten, tear and rivet the malleable minerals of platinum and palladium (polyxene, ferroplatinum, drnopalladite. ), which significantly complicates their subsequent separation into concentrates by gravity-flotation methods. Secondly, it was experimentally established that the extraction of PM minerals in the centrifugal force field depends on the ratio of the size of the recoverable minerals and the size of the grains of gangue. At the same time, the enrichment of material containing coarsely ground rock, ceteris paribus, proceeds much more efficiently than the enrichment of material in which the size of the minerals PM and rock is at the same level.

The process of gravity concentration of sulfide copper-nickel ores and their by-products is very sensitive to the deviation of the value of the centrifugal Froude criterion characterizing the intensity of the centrifugal field from its optimal value. Thus, according to experimental data, the deviation of Fr _c by 20–25% from the optimal level reduces the extraction of platinum metals by ~ 20% abs. Lower optimum values Fr _u characteristic of materials with a high ratio of the criterion, apparently due to the fact that this subgroup is different intermediate density increased content of minerals (sulphides, arsenides, magnetite). In addition, when grinding such a material, grains of sulfide minerals of non-ferrous metals and iron, due to their increased fragility, are commensurate in size with the grains of minerals PM. This requires a finer balance of forces during ore processing (“flushing” and “pressing”) along the height of the centrifuge, which becomes possible with some weakening of the centrifugal field in the central part of the apparatus with respect to the back pressure of the fluidizing water.

The centrifugal force pressing the particles to the wall of the centrifuge tens of times increases their relative weight in comparison with gravity. Therefore, the heavier the material, the ceteris paribus the higher the pressure of the particles on the wall of the centrifuge, and therefore the higher the optimal back pressure of the fluidizing water. Since sulfide minerals and magnetite are 1.2–2 times heavier than gangue minerals, the optimal ratio of Fr _c to the pressure of the fluidizing water is lower for ore dressing with a high criterion ratio than for the enrichment of another subgroup characterized by a low criterion ratio.

 A concentrator-centrifuge is one of the few devices for which large dilutions not only do not interfere, but, on the contrary, contribute to the enrichment process. However, excessive watering of the process complicates the subsequent scheme of the enrichment of the resulting drain. Therefore, experimentally, the optimal dilution area of the initial pulp was found, which, on the one hand, corresponds to high gravel concentration in the entire real range of ore size, and, on the other hand, to the minimum cost for subsequent re-enrichment of the resulting drain by foam flotation.

According to experimental data for sulfide copper-nickel ore or its intermediate product with a criterion ratio of less than 1: 2, the range of optimal values of the centrifugal Froude criterion for gravity concentration is in the range of 2-10, and the ratio of the value of the Froude criterion to the pressure of the fluidizing water corresponds to the optimal range (0.025-0 , 23) kPa ^-1 . Moreover, the highest results are achieved when grinding ore to a particle size of 30-65% class less than 74 microns. The final result of ore dressing is determined by the combined effect of these parameters and far exceeds the result of their additive influence on this process. Outside the ranges of optimal values, enrichment rates are sharply reduced.

 The insufficient degree of grinding of the initial ore (the crushed material contains less than 30% of the class minus 74 microns) does not provide the necessary depth of disclosure of mineral aggregates. PM minerals are in close germination with the rock, as a result of which the mineral contains a large number of intermineral associates of intermediate density, the separation of which is ineffective by hydromechanical methods. When gravel-enrichment of such a material, the resulting concentrates have low conditions, and the ore dressing technology is characterized by a high level of platinum metal losses. A high degree of ore grinding (the crushed material contains more than 65% of the minus 74 micron class) leads to mechanical disruption of the geometric shape of PM minerals, which complicates their subsequent separation by gravity-flotation methods: brittle PM minerals are crushed to a sludge-like state, malleable platinum and palladium minerals are pulled to a flocculent state state, repeatedly increasing its orientation coefficient and forming hard-to-remove "hardened" particles of waste rock. Going beyond the limit in this case sharply reduces the extraction of PM in the gravity concentrate and increases the loss of platinum metals with dump tailings. The resulting gravity concentrate is diluted with waste rock as a result of the co-extraction of the heavier particles with the “hardening” of metallized minerals PM.

 Material enrichment in a centrifugal force field, characterized by low values of the Froude criterion (less than 2), is characterized by a sharp decrease in the extraction of PM minerals in gravity concentrate relative to the level of their extraction, achieved in the range of optimal values. This mainly occurs as a result of an increase in the “breakthrough” of thin hardly enriched particles of acute-angled shape with a high orientation coefficient. At the same time, ore beneficiation in a centrifugal force field with values of the Froude criterion above 10 does not lead to a noticeable increase in the completeness of PM extraction into gravity concentrate. Along with this, the quality of the concentrate decreases noticeably and the wear of the working surface of the centrifugal concentrate sharply accelerates.

It was found that at low values of the ratio of the Froude criterion to the pressure of the fluidizing water (less than 0.025 kPa ^-1 ), the removal of fine and highly oriented particles enriched with PM minerals from the gravity concentrate sharply increases, which reduces their extraction into the concentrate and increases losses with tailings. The quality of the gravity concentrate is improved very slightly. An increase in the ratio of the centrifugal Froude criterion to the fluidizing water pressure above the upper limit of the declared range (more than 0.23 kPm ^-1 ) slightly increases the target PM recovery, but at the same time sharply reduces the quality of the gravity concentrate due to the deterioration of mass transfer conditions in the forming sediment and the weakening of the transport bearing capacity ("flush") flow. Gravity concentrates obtained in this mode, due to the relatively low PM content, are not suitable for processing according to the "short" technology along with Pt-containing electrolyte sludge.

According to the results of experimental studies, for the gravel concentration of sulfide copper-nickel ores with a criterion ratio greater than or equal to 1: 2, the range of optimal values of the centrifugal Froude criterion lies in the range 0.5-1.75. Moreover, the ratio of the value of the Froude criterion to the pressure of the fluidizing water corresponds to the optimum in the range (0.0058-0.019) kPa ^-1 , and the degree of grinding of the ore should ensure that it contains 60-95% of particles of the class minus 74 microns.

 Due to the increased content of minerals of intermediate density (sulfides, arsenides, magnetite) in the ore, an insufficient degree of its refinement (content less than 60% of the class minus 74 microns) leads to the presence in the initial feed of the gravity concentrate of a large number of heavy particles, practically containing or containing in negligible amounts own minerals of platinum metals. The enrichment of such material requires increased pressure of the fluidizing water, which, on the one hand, upsets the balance of forces between the centrifugal field and the wash flow, thereby complicating the design of the apparatus and the process mode, on the other hand, it reduces the extraction of PM in the gravity concentrate and significantly dilutes it non-ferrous metal and iron minerals. At the same time, there is a decrease in the quality of the resulting flotation concentrates and an increase in PM losses with waste tailings. With an excessively high degree of ore grinding (yield class minus 74 μm more than 95%), a large number of PM minerals are crushed and transform into refractory forms: fine sludge and particles with a high orientation coefficient. As a result of this, PM extraction into the gravity concentrate sharply decreases and their losses with dump tailings increase. At the same time, the concentrate is diluted with a sulfide-magnetite mass and its quality is markedly reduced.

Gravel enrichment of the material at extremely low values of the centrifugal Froude criterion (less than 0.5) noticeably reduces the quality of the concentrate, plant productivity and is accompanied by a significant “slip” of PM minerals into the tail suspension. The performance of the proposed method is closer to the results obtained using the prototype method. Material enrichment at extremely high values of the centrifugal Froude criterion (more than 1.75) leads to the need to increase the pressure of the fluidizing water, which complicates the design of centrifugal apparatus and increases operating costs. With increasing Fr values _q in the art reduction gravity concentrate quality occurs at a (slight) increase in recovery of platinum group metals. An assessment of the ratio of these factors indicates that a decrease in the selectivity of the process in this area, which causes the production of low-grade gravity concentrates, does not pay off with the indicated increase in PM extraction in the gravity enrichment cycle. This is due to the fact that the subsequent processing of such concentrates is possible only according to the “complete” scheme and the loss of PM with slags from the head smelting units is much higher than the increase in extraction at the ore concentration stage.

The extremely low ratio of the value of the centrifugal Froude criterion to the pressure of the fluidizing water (less than 0.0058 kPa ^-1 ) leads to an unbalance of forces and, therefore, is accompanied by an increased removal of slimy PM particles into the drain suspension. In this case, the extraction of PM into the gravity concentrate sharply decreases and their losses with dump tailings increase. On the contrary, the prohibitively high ratio of the value of the centrifugal Froude criterion to the pressure of fluidizing water (more than 0.019 kPa ^-1 ) slightly increases the recovery of PM, but at the same time sharply reduces the quality of gravity concentrates. All concentrates obtained in this area in terms of PM content are below the standard level, which determines the possibility of processing platinum-containing raw materials according to a "short" scheme.

 The highest rates of gravity concentration of sulfide copper-nickel ores and their intermediate products, regardless of the ratio of the mineral forms contained in them, are achieved in the pulp dilution range corresponding to the mass ratio W: T - (1.5-6.5): 1. With small ratios W: T (less than 1.5), the separation sharply worsens, up to a complete cessation. The reason for this is the increasing viscosity of the thick pulp and the high cohesion of its grains, which prevent not only centrifugal, but also segregation separation of the material. This feature of the centrifugal apparatus does not allow their use directly in closed grinding cycles. The enrichment of the material in pulps with a high degree of liquefaction (L: T greater than 6.5: 1) only slightly improves the performance of the process compared with the upper limit value of this parameter related to the claimed range. At the same time, the work of gravity concentrators in the area of highly diluted pulps reduces their productivity, significantly increases the specific energy consumption of the process and complicates the scheme for the subsequent processing of depleted drains.

 Information about the prominence of the combination of distinguishing features that provide a cumulative effect in the proposed technical solution during the beneficiation of sulfide copper-nickel ores containing proprietary minerals of platinum metals and magnetite has not been identified in the study of patet and scientific and technical literature.

Known is the use of centrifugal installations with a fluidized bed of the Knelson design for the extraction of free gold in the beneficiation of gold-bearing placer deposits (Knelson B.V. Gravity dressing and separation of precious metal ores. ", 17th Annual Congress of Construction and Installation, Ottawa, 1985. - P. 346-357) At the same time, gold-bearing sands are a much simpler object for the process of gravel dressing than sulfide copper-nickel ores or their intermediate products. This is because the contrast of density characteristics for free gold (19.3 kg / dm ³ ) and host sands (~ 3-3.5 kg / dm /) ~ 2.5-4 times higher than the difference between the density characteristics of a number of platinum-palladium minerals (8, 5-12.5 kg / dm) and the sulfides accompanying them - pentlandite, pyrrhotite, chalcopyrite (4.3-4.9 kg / dm ³ ) and, especially, magnetite (5.2 kg / dm ³ ). Another difference is in that a significant part of the PM minerals in sulfide copper-nickel ores is represented by brittle minerals (arsenides, sulfides, sulfoarsenides), which are sludge and become difficult to process during ore grinding proper geometric shapes with high orientation coefficients. Gold in its free state has a rounded, easy to enrich form and, as a rule, a larger size than the allocation of minerals PM. Therefore, both the very possibility of using the proposed centrifugal concentrators for beneficiation of sulfide copper-nickel ores, and the claimed regime of this process, which is the subject of the present invention and developed for raw materials with specific specific physical and mineralogical characteristics, do not follow explicitly from the achieved level of technology. These features of the known technical solution and the distinguishing features of the present invention are essentially similar.

 Thus, the set of distinctive features of the proposed technical solution explicitly from the existing level of technology does not follow, which indicates the compliance of the claimed object with the criterion of "Inventive step".

 The method is as follows.

 The initial sulfide copper-nickel ore is subjected to ore preparation, including crushing, classification of the material and, if necessary, its preliminary concentration in heavy suspensions and mechanical disintegration to separate from clay inclusions. The crushed material is ground in an aqueous medium using standard grinding units, for example, ball mills operating in a closed cycle with hydraulic (spiral) classifiers and / or classifying hydrocyclones. After the classification stage, the pulp of ore, crushed to a given size, is brought to optimum density with water and subjected to gravity-flotation concentration. The operation of hydromechanical (gravio-) enrichment of the material is carried out in order to isolate the maximum amount of its own minerals of platinum metals into an independent product - enriched by PM for gravity concentrate. In this case, the material is subjected to gravity concentration at the point of the technological scheme preceding the flotation operation. In particular, the pulp of the initial crushed ore, preliminarily ore that has passed the preliminary stages of classification and liquefaction, is subjected to gravity concentration. The resulting gravel dressing tailings (depleted ore pulp) are thickened to the ratio W: T = (1.2-1.4): 1 and sent to the flotation concentration process. In the claimed method, the separation of PM in the gravity concentrate is carried out on a centrifugal concentrator with a fluidized bed created by water jets in a direction that does not coincide with the centrifugal field force vector. As such concentrates, various types of modified centrifuges (vertical, horizontal, inclined) or hydrocyclones with the forced supply of water into the layer of the forming sediment with the purpose of its liquefaction can be used. A specific example of a centrifugal unit of this design, which has now found industrial application, is the Knelson centrifugal separator, the operation scheme of which is shown in FIG. 1.

 For gravity concentration of sulfide copper-nickel ores in centrifugal concentrators, two different modes have been developed that provide maximum process performance under certain boundary conditions.

 In each of the modes, the pulp of the material before gravity concentration is adjusted with water to the required degree of dilution, which, depending on the specific characteristics of the initial ore, varies in the range W: T = (1.5-6.5): 1.

In the gravitational enrichment of the initial ore, characterized by the mass ratio of the sum of sulfides and magnetite to the sum of silicon and aluminum oxides less than 1: 2 (less than 0.5), the ore is ground to a particle size of 30-65% of the class less than 74 microns, and the extraction of their own minerals PM is carried out at the maximum value of the centrifugal Froude criterion equal to 2-10 and the ratio of the value of the Froude criterion to the pressure of the fluidizing water equal to (0.025-0.23) kPa ^-1 . This class of materials includes: disseminated varieties of sulfide copper-nickel ores in intrusive rocks (including normally disseminated and low-sulfide varieties of this type of ore), vein-disseminated and breccia ores in rocks of endocontact and exocontact of intrusions, including low-sulfide mineral types of copper ores. The same class of materials includes chamber intermediate products of flotation concentration of all types of ore, depleted in sulfide content as a result of their separation into the corresponding flotation concentrates.

In the case of gravity concentration of ores with a high ratio of the sum of sulfides and magnetite to the sum of silicon and aluminum oxides greater than or equal to 1: 2 (≥0.5), the ore is ground to a particle size of 60-95% of the class less than 74 microns, and the extraction of its own platinum minerals metals are conducted at a maximum value of the centrifugal Froude criterion equal to 0.5-1.75, and the ratio of the value of the Froude criterion to the pressure of the fluidizing water equal to (0.0058-0.019) kPa ^-1 . This gravity concentration mode is effective for continuous (rich) ores (endo and exocontacts of stratified intrusions) regardless of their mineral varieties.

 The ratio of these parameters in each particular case is selected experimentally, focusing on the regulatory requirements for the quality of the obtained grave concentrates and the completeness of extraction of PM in them.

 Gravel concentrates are preferably processed in a separate cycle to produce salable platinum group metals.

 Gravity dressing tailings (PM lean ore pulp) are subjected to flotation concentration operations carried out according to standard technological schemes: for example, collectively selective (disseminated and cuprous ores) and selective (rich ores). In the process of flotation enrichment of the material receive sulfide concentrates, sent for further processing to the appropriate metallurgical processing, and waste rock-containing tailings that are discharged into the dump.

 The resulting gravity concentrates, if necessary, can be brought to a higher standard by concentrating or hydrometallurgical methods. This allows to a certain extent to reduce the quality requirements of concentrates and, thereby, increase the direct extraction of platinum metals in them. For refinement of gravity concentrates, among other known beneficiation devices (flushing locks, jigging machines, spiral separators, concentration tables), centrifugal units with a fluidized bed operating at low values of the ratio of the centrifugal Froude criterion to the pressure of the fluidizing water can also be used.

 Gravity and flotation concentration products are subjected to volumetric and weight measurements, they are tested and analyzed. Based on the results of analyzes and measurements, the material balance of each of the stages of the process and the overall balance of gravity-flotation enrichment technology are calculated.

 The proposed method is described in specific examples and its result is shown in table. 2.

 Gravel concentration experiments were carried out on an industrial scale on various types of sulfide copper-nickel ore of the Norilsk and Talnakh-Oktyabrsk deposits of Norilsk Combine JSC, which enters the industrial flotation process after the 1st grinding stage.

 In the experiments according to the proposed method, as well as in experiments with exorbitant values of distinctive features (op. 2-12, 14-24, 26-36 and 40), the pulp of the crushed ore or a mixture of various types of ore was selected in a continuous stream with a constant output in the form of a discharge hydraulic spiral classifier, working in a closed cycle with a ball mill, and sent to the process of gravity enrichment. The tail pulp of the gravity concentrator was concentrated to the ratio W: T = 1.3: 1, divided into weights and subjected to laboratory flotation. Flotation was carried out according to standard methods worked out for different types of the studied ore, with the production of sulfide flotation concentrates and dump tailings.

 In the experiments carried out according to the prototype method (op. 1, 13, 25, 37 and 39), industrial flotation tails obtained by beneficiation of the same types of sulfide copper-nickel ore were subjected to gravity concentration. The dump rock-containing tails in these experiments were obtained in the form of a discharge of the process of gravity enrichment.

 In all experiments, the gravity concentration of materials (initial ore and tailings) was carried out in a centrifugal gravity concentrator with a fluidized bed of a Knelson concentrate design. A size apparatus of 7.5 '' (inches) was used with a maximum centrifuge diameter (in the discharge section) of 190.5 mm. The design of the apparatus included a vibrating sieve with a hole size of 6 mm to hold the scrap and wood chips. The main elements of a Knelson concentrator and the principle of its operation are shown in FIG. 1. The apparatus of the used standard size operates in a semi-continuous mode with a periodic stop to unload the accumulated gravity concentrate. The duration of the accumulation cycle of the gravity concentrate was established from the conditions for the collection of its mass in the amount of 1.3-2.0 kg, which corresponds to the output for Knelson apparatuses of this standard size. The solid load was set taking into account the passport capacity of the apparatus and for various materials varied within 300-460 kg / h.

 The efficiency of the gravity concentration mode and the method as a whole were evaluated by the quality of the obtained gravity concentrate, the degree of its enrichment (by the amount of PM), the level of PM extraction in the gravity concentrate, as well as the total PM extraction into gravity and flotation concentrates and the PM loss rate with tailings.

 Example 1 (experiment 1 of table 2) is the implementation of the prototype method.

The experiment was conducted on salable sulfide copper-nickel ore of the Norilsk-1 deposit, in which the host rocks were represented by picrite (60%), taxitic (30%) and contact (10%) gabbro dolerites with uniform and uneven (in taxitic) interspersed sulfides of three morphological varieties: fine-grained, medium-grained and teardrop-shaped, as well as coarse-fingered. The mineral composition of sulfides,%: chalcopyrite - 1.6; pentlandite - 1.4; pyrrhotite - 9.0. The mass fraction of non-metallic was 88.0%, including,%: SiO ₂ - 40.1; CaO - 7.7; MgO - 13.5; Al ₂ O ₃ - 11.3; magnetite - 10.4. The ore content of platinum group metals, g / t: platinum - 2.10; palladium - 4.90; rhodium - 0.035 (the amount of PM - 7.35 g / t). Mineral composition of platinum: Braggite (Pt _0.65 Pd _0.4 S) - 50%; ferroplatinum (PtFe) - 20%; sperrylite (PtAs ₂ ) - 20% and the remaining 10% - cuperite (PtS), palladium platinum, etc. The bulk of palladium was represented as solid solutions in sulfides and about 6-8% stannopalladinite (Pd ₃ Sn ₂ Cu). Rich PM solid solutions were associated with chalcopyrite (up to 30%), pyrrhotite (up to 20%), pentlandite (up to 25%) and nonmetallic (up to 25%). The mineral forms of PM were represented by two technological varieties: coarse-grained, relatively easily recoverable by gravity methods, and extremely finely dispersed (first units of microns) in association with sulfides contained in large silicate formations, and biotite. The mass ratio of the sum of sulfides and magnetite to the sum of oxides of silicon and aluminum in the initial ore averaged 1: 2.3.

 The initial ore was continuously crushed (to a fineness of -16 + 0 mm), crushed in a ball mill operating in a closed cycle with a spiral classifier, and the discharge of the classifier, the solid phase of which contained 45% of particles of particle size less than 74 μm, was subjected to flotation enrichment according to standard 2-stage scheme for 1 collective section of the Norilsk Concentrating Plant (NOF) of JSC Norilsk Combine, (Technological instruction // Ore dressing of the Norilsk-1, Talnakhsky and Oktyabrsky deposits at the Norilsk Concentrating Plant factory.-ty 0401.14.52-11-43-97. Approved by the chief engineer of the PF from 07.28.97 - Norilsk, 1997. - P. 45-47). In the flotation process, a collective flotation concentrate and rock-containing tailings were obtained.

 The tail pulp from the flotation enrichment operation, the solid phase of which contained 75% of particles of particle size less than 74 microns (a typical particle size for this product), was sent to the gravity concentration process, carried out in a centrifugal gravity concentrator with a fluidized bed created by water jets injected under pressure in the direction counter to the rotational movement of the pulp stream. As a centrifugal concentrator, a Knelson apparatus of a standard size of 7.5 inches was used (the diameter of the largest cross section of the centrifuge was 190.5 mm).

где n - скорость вращения центрифуги, с^-1;
D_m - максимальный диаметр центрифуги (D_m = 190,5 мм), м;
g - ускорение силы тяжести (g = 9,81 м/с²).The tail pulp with a mass ratio of W: T equal to 5.0: 1, by gravity entered the receiving vat with a stirrer. From here, it was pumped in a continuous flow with constant flow rate to a pulp splitter with a capacity of 150 dm ³ and then the pulp was gravity-fed through the flow slot in the receiving tray of the Knelson apparatus. Excess power through the side pocket was dumped into the discharge sump. The productivity of the process of gravity concentration on solid was 460 kg / h and was adjusted using a valve installed on the line of the feed pump. In this experiment, the centrifuge rotation speed by selecting the appropriate pulley was set equal to 17.6 s ^-1 , which ensured the maximum value of the Froude centrifugal criterion (in the uppermost section of the centrifuge) at level 6 (centrifugal acceleration ~ 120 g). By adjusting the flow rate of water supplied to the jacket of the apparatus, the pressure of the fluidizing water was set at 46.2 kPa according to the pressure gauge. By adjusting these parameters, the ratio of the maximum value of the centrifugal Froude criterion (Fr _c ) to the pressure of the fluidizing water (P _c ) equal to 0.13 kPa ^{-1 was} ensured. In this case, the maximum value of the centrifugal Froude criterion was calculated by the formula:

where n is the speed of rotation of the centrifuge, s ^-1 ;
D _m - the maximum diameter of the centrifuge (D _m = 190.5 mm), m;
g is the acceleration of gravity (g = 9.81 m / s ² ).

 The gravity concentrate accumulated in the apparatus was discharged at the end of the control time determined by the results of preliminary experiments. In this experiment, the accumulation time of material in Knelson was 30 hours, which ensured the production of a gravity concentrate in an amount of 1352 g per dry weight. The discharge of the gravel dressing process, which is the tailings of the sulfide copper-nickel ore processing technology, was continuously discharged into the dump.

 The results of the experiment are given in table. 2. The resulting gravity concentrate contained 640 g / t of the amount (Pt + Pd + Rh) and was characterized by the enrichment degree in the amount of PM equal to 87. Extraction of PM in the gravity concentrate was,%: platinum - 5.1; palladium - 2.8; rhodium - 0.1; the amount of PM - 3.2. In total, the gravity concentrate and collective flotation concentrate were recovered,%: platinum -65.7; palladium - 71.9; rhodium - 69.2; the amount of PM - 69.8. Accordingly, with waste tailings, enrichment is lost,%: platinum - 34.3; palladium - 28.1; rhodium - 30.8; the amount of PM - 30.2.

 The indicators obtained in this experiment reflect the real state of the industrial process for the dissemination of disseminated ores of the Norilsk and Talnakh ore clusters, which is characterized by a low yield and quality of the obtained gravity concentrates and an unacceptably high level of PM losses with dump flotation tailings.

 Example 2 (experiment 2 of table 2) - the proposed method.

The composition of the initial ore, the equipment and the mode of the process of gravity concentration are the same as in example 1. The difference is that in this example the process of gravity concentration was not subjected to flotation tailings, but the initial ground ore, selected at the stage preceding the flotation operation. As a gravity concentrate, the same apparatus of the Knelson design of the standard size of 7.5 inches was used as in Example 1. To make it possible to vary the value of the centrifugal Froude criterion (Fr _c ), which determines the efficiency of centrifugal concentration of platinum group minerals, the Knelson apparatus was equipped with a set of pulleys of different diameters, providing a step change in the speed of rotation of the centrifuge from 272 to 1491 min ^-1 . A variable parameter of gravity concentration was also the counterpressure of fluidizing water, which in various experiments varied in the range from 35 to 105 kPa. The set value of this parameter was set by valve adjustment, monitoring according to the pressure gauge.

The ore, crushed in an industrial ball mill to a grade of 45% grade less than 74 microns, as part of a hydraulic classifier discharge with a W: T mass ratio of 4.0: 1, was fed to a receiving vat with a mixer, from where it was fed by continuous flow to the apparatus by hydrotransport Knelson. The productivity of the process of gravity concentration on solid was 460 kg / h. In this experiment, the centrifuge rotation speed was also set equal to 17.6 s ^-1 , which corresponded to the maximum value of the centrifugal Froude criterion at level 6 (centrifugal acceleration ~ 120 g). The pressure of the water supplied to the jacket of the apparatus was 46.2 kPa. Adjusting the parameters provided the ratio of the maximum value of the centrifugal Froude criterion (Fr _c ) to the pressure of the fluidizing water (Fr _c ) equal to 0.13 kPa ^-1 .

 In this experiment, the material accumulation cycle was 8 hours, and the mass of the obtained gravity concentrate was 1401 g (dry weight).

где n - скорость вращения центрифуги, с^-1;
D_m- максимальный диаметр центрифуги (D_m= 190,5 мм), м;
g, - ускорение силы тяжести (g = 9,81 м/с²).From here, it was pumped in a continuous flow with constant flow rate to a pulp splitter with a capacity of 150 dm ³ and then the pulp was gravity-fed through the flow slot in the receiving tray of the Knelson apparatus. Excess power through the side pocket was dumped into the discharge sump. The productivity of the process of gravity concentration on solid was 460 kg / h and was adjusted using a valve installed on the line of the feed pump. In this experiment, the centrifuge rotation speed by selecting the appropriate pulley was set equal to 17.6 s ^-1 , which ensured the maximum value of the Froude centrifugal criterion (in the uppermost section of the centrifuge) at level 6 (centrifugal acceleration ~ 120 g). By adjusting the flow rate of water supplied to the jacket of the apparatus, the pressure of the fluidizing water was set at 46.2 kPa according to the pressure gauge. By adjusting these parameters, the ratio of the maximum value of the centrifugal Froude criterion (Fr _c ) to the pressure of the fluidizing water (P _c ) equal to 0.13 kPa ^{-1 was} ensured. In this case, the maximum value of the centrifugal Froude criterion was calculated by the formula:

where n is the speed of rotation of the centrifuge, s ^-1 ;
D _m - the maximum diameter of the centrifuge (D _m = 190.5 mm), m;
g, is the acceleration of gravity (g = 9.81 m / s ² ).

 The gravity concentrate accumulated in the apparatus was discharged at the end of the control time determined by the results of preliminary experiments. In this experiment, the material accumulation cycle was 8 hours, which corresponded to a mass of gravity concentrate of 1401 g (dry weight).

 The tailings of the gravel dressing process (pulp of the combined ore) were concentrated to the ratio W: T = 1.3: 1 and subjected to flotation enrichment carried out on laboratory grinding and flotation equipment.

 Flotation was carried out according to a collective scheme (Fig. 2) with intermediate regrinding of the chamber product of the 1st control flotation (up to 70% of the class less than 74 microns) and rough collective concentrate (up to 85% of the class less than 45 microns). The adopted enrichment scheme corresponds to the technological scheme of flotation enrichment of disseminated sulfide copper-nickel ores, industrially used at the Norilsk processing plant of JSC Norilsk Combine. The final products of laboratory flotation were: collective sulfide copper-nickel concentrate and dump rock-containing tailings.

 The results of the experiment are shown in table 2.

 The obtained gravity concentrate in this experiment, in contrast to the prototype method, was significantly enriched in platinum group metals. The content of PM in it amounted to 4132 g / t, with an enrichment rate of 562. Such a rich gravel concentrate can be sent for processing according to a “short” scheme, that is, directly to platinum production. Knelson concentrate was able to extract a significant amount of platinum - 47.6%, as well as 11.9% of palladium. The extraction of rhodium was only 1.1%, which is due to the negligible content of this satellite of platinum in the form of its own mineral forms, and its presence, mainly, in the form of an isomorphic impurity in the crystal lattice of sulfide minerals - carriers: pentlandite and pyrrhotite. The extraction of the amount of PM in the gravity concentrate amounted to 21.4% of their content in the original ore, which is ~ 6.7 times higher than in the prototype method. This experiment also has a high total PM content in gravity and flotation concentrates,%: platinum - 89.9; palladium - 92.6; rhodium -88.5; the amount of PM - 91.6. PM losses with tailings amounted to,%: platinum - 10.1; palladium - 7.4; rhodium -11.5; the amount of PM - 8.4. This, on average, is approximately 3.6 times lower than in the prototype method.

 An important positive factor of the proposed technical solution, first identified during the formulation of this series of experiments, was the effect of the activating effect of the gravitational concentration process in the Knelson apparatus on the subsequent flotation of sulfide copper-nickel ores. This was manifested primarily in the fact that the increase in the total extraction of palladium and, especially, rhodium in gravel and flotation concentrates relative to the base level (op. 1) turned out to be significantly higher for these metals than would be expected only due to their extraction in gravity concentrate. This fact cannot be explained otherwise than by an increase in the flotation activity of sulfide minerals - carriers of PM, which is especially clearly seen when studying the behavior of rhodium: the distribution of this metal during flotation, since it practically does not have its own minerals, is mainly determined by the flotation activity of sulfides. A comparison of the results of op.1 and op.2 shows that when extracting rhodium in a gravity concentrate equal to 1.1%, the increase in its extraction in the amount of 2 concentrates compared to the prototype method was 19.3% abs. Thus, the hydromechanical treatment of crushed ore in a rotating pulp stream penetrated by directed high-energy water jets provides an additional effect - it increases the completeness of PM extraction during subsequent flotation of the material into the foam product. The observed effect of ore activation in the Knelson apparatus can be explained by two possible reasons: mechanical peeling of oxide films from the surface of the particles as a result of their cramped intensive movement in the near-wall centrifuge layer, which occurs under the conditions of abrasive action of solid rock-forming minerals, as well as washing the surface of sulfide particles from gel-like clay deposits screening particles from flotation reagents and air bubbles, while dispersing clay formations and their peptides nation. Subsequently, when the tails of the Knelson apparatus are thickened, clay particles in the form of the finest suspension are discharged with discharge, which ensures material freeing and improves the conditions for the formation of flotation complexes, and therefore, the selectivity of the flotation process of sulfides hydrophobized by the addition of a sulfhydryl reagent-collector.

 Example 3 (experiment 14 of table 2) - the proposed method.

The gravity and flotation equipment used is the same as in Example 2. The difference is that in this example copper sulfide ore of the chalcopyrite mineral type of the Talnakh - Oktyabrsky deposit was used, which is mainly a brexitic texture variety. The main rock-forming minerals in this type of ore were: pyroxenes, feldspars, garnets, serpentines, chlorite, calcite, anhydrite. The mineral composition of sulfides,%: chalcopyrite - 15.1; pentlandite - 2.7; pyrrhotite - 7.4; valerite and millerite are single grains. Mass fraction of non-metallic - 74.8%, including,%: SiO ₂ - 42.6; CaO - 6.9; MgO - 3.7; Al ₂ O ₃ - 11.4; magnetite - 1.8. The ore content of platinum group metals, g / t: platinum - 2.14; palladium - 8.98; rhodium - 0.57 (the amount of PM - 11.69 g / t). The ratio of intrinsic mineral forms of PM in cuprous ore differs from the mineralization of disseminated ore of the intrusive type. The main form of the presence of platinum and palladium in this ore was its own minerals of these elements, forming grains ranging in size from 0.02 to 0.5 mm. PM minerals were present in close association with each other, forming polymineral intergrowths among sulfides, as well as at the border of sulfides with magnetite or silicates. Platinum was represented by rather large (up to 0.5 mm) grains of sperrylite (PtAs ₂ ), couperite (PtS) and braggite (Pt _0.65 Pd _0.4 S), the total share of which was about 95%. A small amount of platinum (~ 5%) was represented by compounds of the type (Pd, Pt) _3-x S _n and ferroplatinum (PtFe). Palladium minerals were 60% represented by intermetallic compounds with tin, lead, bismuth, arsenic, nickel, copper, antimony, tellurium and ~ 35% sulfide minerals, of which vysotskite (Pd _0.7 Ni _0.3 ) S occupied the main place . In addition, in small amounts, palladium was present as an isomorphic impurity to pentlandite. Rhodium and other platinum moons were in dispersed form in the main ore-forming sulfides: chalcopyrite, pentlandite, and pyrrhotite. The mass ratio of the sum of sulfides and magnetite to the sum of oxides of silicon and aluminum in the initial ore was 1: 2.

Another difference is that the initial ore was ground to a content of 80% of particles of size class less than 74 microns and subjected to gravity-flotation concentration according to the technological scheme shown in FIG. 3. At the same time, the productivity of gravel concentration in ore was 300 kg / h, and the pulp density in the feed of the gravity concentrator corresponded to the mass ratio W: T equal to 4.0: 1. Gravel concentrate was accumulated in the Knelson apparatus for 11 hours, which ensured that the concentrate was obtained in an amount of 1493 g (dry weight). The speed of rotation of the centrifuge using a pulley was set equal to 7.6 s ^-1 , and the pressure of the fluidizing water in the jacket of the apparatus was stabilized at a damage of 93.3 kPa. This ensured the maximum value of Fe _{c of} the centrifugal force field at the level of 1.12 (centrifugal acceleration ~ 22.4 g), and the ratio of Fe _c to the pressure of the fluidizing water was 0.012 kPa ^-1 .

 Flotation enrichment of the tailings obtained in the operation of gravity concentration was carried out on laboratory grinding and flotation equipment according to the selective-collective scheme shown in FIG. 3. At the same time, three products were obtained: selective copper concentrate, collective copper-nickel concentrate, and dump rock-containing tailings.

 The results of the experiment are presented in table 2.

The concentrate obtained by gravity concentration of ore contained 2352 g / t of PM amount and was characterized by the degree of enrichment for platinum metals equal to 201. This is a significantly higher result than in the experiment by the prototype method (op. 13), where, all things being equal, under the same conditions A gravel concentrate containing only 415 g / t PM amount was obtained at ore with a degree of enrichment of ~ 5.7 times lower. The extraction of PM in the gravity concentrate was,%: platinum - 25.3; palladium - 5.8; rhodium - 0.5; the amount of PM - 9.1. The total extraction of PM in the target products of gravity-flotation concentration (gravity concentrate, selective copper and collective
copper-nickel flotation concentrate) indicate the high efficiency of the proposed method. These concentrates were recovered,%: platinum - 94.8; palladium - 96.7; rhodium -92.9; the amount of PM - 96.2. In the prototype method (op.13) in the total target product, the extraction of the amount of PM was 92.6%, i.e. by 3.6% abs. lower than in the proposed method, ceteris paribus. In this example, as in example 2, the activating effect of the hydromechanical processing of ore in the Knelson apparatus on the indicators of the flotation concentration process is traced. In particular, with the extraction of rhodium in a gravity concentrate equal to 0.5%, the increase in its extraction in the combined gravity-flotation product compared to the prototype method (op.13) amounted to 3.1% abs. With dump tailings, enrichment is lost,%: platinum - 5.2; palladium - 3.7; rhodium - 7.1; the amount of PM - 3.8. This, on average, is 1.9 times lower than in the prototype method (op.13).

 Example 4 (experiment 26 of table 2) - the proposed method.

The technological scheme of gravity-flotation ore dressing, equipment and process conditions are the same as in example 3. The difference is that in this example we used a technological mixture of rich ore of pyrrhotite type (chalcopyrite - pyrrhotite mineral variety) and cuprous ore (pyrrhotite mineral varieties) of the Talnakh-Oktyabrsky deposit. The mineral composition of sulfides,%: chalcopyrite - 6.5; pentlandite - 4.2; cubanite - 1.6; pyrrhotite - 16.7. Mass fraction of non-metallic - 64.6%, including,%: SiO ₂ - 37.1; CaO - 7.9; MgO - 5.2; Al ₂ O ₃ - 11.2; magnetite - 3.2. The metal content of the platinum group in the ore mixture, g / t: platinum - 2.17; palladium - 8.95; rhodium - 0.63; the amount of PM - 11.75. A feature of the forms of the presence of PM was the almost complete absence of their sulfide minerals (cuperite, braggite, and vysotskite). Platinum was represented by the smallest (0.03-0.06 mm) grains of mainly two minerals: sperrylite (PtAs ₂ ) - 60% and compounds of the type (Pd, Pt) _3-x S _n - 30%. An insignificant part of platinum (~ 5%) was contained in the crystal lattices of pyrrhotite and pentlandite, the rest was ferroplatinum, niggliite (PtSn), and others. Palladium in the ore mixture was ~ 80% dispersed in pentlandite, into the crystal lattice of which it entered as isomorphic impurities. A small part of palladium (~ 10%) was present in the form of stannopalladinite (Pd ₃ Sn ₂ Cu) and other mineral forms such as Pd ₂ B (where B-Sn, As, Sb). A small portion of palladium was in dispersed form in pyrrhotite. Rhodium was mainly dispersed in the main ore-forming sulfides, mainly in pyrrhotite. The mass ratio of the sum of sulfides and magnetite to the sum of oxides of silicon and aluminum in the initial ore was 1: 1.5.

The initial ore was ground to a content of 80% of particles of size class less than 74 microns and subjected to gravity-flotation concentration according to the technological scheme shown in figure 4. The duration of the process of gravity concentration in ore was 300 kg / h; the pulp density in the diet corresponded to a W: T ratio of 4.0: 1. The accumulation of gravity concentrate in Knelson was carried out for 11 hours. The mass of the concentrate in terms of dry matter was 1404 g. The centrifuge rotational speed was set equal to 7.6 s ^-1 , the pressure of the fluidizing water in the Knelson jacket was 93.3 kPa. This ensured the maximum value of Fe _{c of} the centrifugal force field in the rotating cone at the level of 1.12 (centrifugal acceleration ~ 22.4 g), and the ratio of Fe _c to the pressure of the fluidizing water was 0.012 kPa ^-1 .

 Knelson tails were subjected to flotation enrichment in a laboratory setup according to the direct selective flotation scheme (Fig. 4). In this case, 3 foam products (copper, nickel and pyrrhotite concentrates) and dump tailings were obtained.

 The results of the experiment are shown in table 2.

 The gravity concentrate obtained in this experiment contained 2210 g / t of the PM amount, which corresponded to the degree of enrichment equal to 188. In the experiment by the prototype method (op. 25), the results are significantly lower: the content in the gravity concentrate of the PM amount is 396 g / t; the degree of enrichment - 34. The extraction of PM in the gravity concentrate by the proposed method was,%: platinum - 21.9; palladium - 5.2; rhodium - 0.4; the amount of PM - 8.0. In a parallel experiment conducted by the prototype method (op. 25), the extraction of the amount of PM was 6.2 times lower. Including platinum recovery is 9.1 times less. The total extraction of PM in the gravity concentrate and combined flotation concentrates was characterized by a high level,%: platinum -92.5; palladium - 95.3; rhodium - 90.7; the amount of PM - 94.5. In the corresponding prototype method (op.25), the extraction of the amount of PM in the total target product of enrichment was 91.8%, which is 2.7% abs. below. Loss PM with tailings tailings in the considered example amounted to,%: platinum - 7.5; palladium - 4.7; rhodium - 9.3; the amount of PM - 5.5. This is approximately 1.5 times lower than under similar conditions when using the prototype method (op.25).

 Table 2 shows examples that differ in the composition of the initial sulfide copper-nickel ore, ore preparation conditions (the degree of ore grinding and liquefaction of the resulting ore suspension), as well as the place of installation of the gravity concentrator in the beneficiation scheme and its operation mode: productivity, time of accumulation of the gravity concentrate and specific separation factors of mineral components of the ore, characteristic of a centrifugal Knelson concentrator. The latter include: the maximum value of the centrifugal Froude criterion, the back pressure of the fluidizing water and the ratio of the maximum value of the centrifugal Froude criterion to the pressure of the fluidizing water.

 According to the obtained experimental results (experiments 2-4, 14-16, 22-28, 38 and 40), the proposed method of gravity-flotation concentration of sulfide copper-nickel ores, based on the use of a centrifugal gravity concentrator with a fluidized bed of the concentrated material, when installed in " the head "of the technological scheme (up to the flotation concentration stage) and certain values of operating parameters, ensures high extraction of the platinum metals' own minerals, mainly platinum, into rich gravel centrate and low losses of platinum metals with tailings enrichment. Depending on the type of ore being mined and operating parameters of the process, the extraction of platinum metals in a gravity concentrate in the proposed method is,%: platinum - 16.7-57.0 palladium - 4.0-12.5; rhodium - 0.3-1.2; the amount of PM - 6.1-23.4. The resulting gravity concentrates are characterized by a significant degree of concentration of platinum metals, which determines a very high level of their total content in this product -1986-4132 g / t of concentrate. This level of platinum metals allows you to process the gravity concentrates obtained by the proposed method directly in the production of metallurgical finishing of rich platinum raw materials along with anode sludge and, thus, bypass the whole chain of redistribution of copper and nickel production, which platinum metals are part of selective ore flotation concentrates and gravity concentrates isolated by the prototype method. Shortening the technological cycle of processing gravel concentrates eliminates PM losses with dump metallurgical products: slag from smelting units, process dusts and ferrous cakes.

 The indicators of gravity-flotation concentration of similar types of ore obtained using the prototype method will significantly tire of the indicators achieved by the proposed technical solution. The extraction of the amount of PM in the gravity concentrate from ore by the prototype method (op. 1, 13, 25, 37 and 39), on average, 5.7-7.9 times lower than in the proposed method, and the resulting gravity concentrates even in optimal conditions contain only 415-640 g / t PM amount. Such concentrates are not suitable for processing according to the “short” scheme and are processed together with ore flotation concentrates at the head pyrometallurgical redistributions of copper production, which inevitably causes PM losses with the above dump products.

 Comparison of the proposed method with the prototype method in terms of PM losses with waste tailings indicates a significant advantage of the first. The most significant reduction in PM losses with tailings in the proposed method is achieved in the processing of poor types of sulfide copper-nickel ore (with a ratio of the sum of sulfides and magnetite to the sum of silicon and aluminum oxides less than 1: 2). In experiments with the enrichment of such ore, the loss of PM amount with tailings amounted to: in the proposed method (op. 2.38) - 7.6-8.4%; in the prototype method (op. 1.37) - 26.0-30.2%. This trend, although with a lower contrast of results, also occurs in the concentration of ores with a high content of valuable components (cuprous, rich and their mixtures). When processing such ores according to the proposed method (op. 14, 26, 40), the loss of the amount of PM with waste tailings amounted to 3.8-6.1%, while in parallel experiments using the prototype method (op. 13, 25 , 39) 7.4-9.0% of the amount of the PM was lost.

 Therefore, the proposed technical solution in comparison with the prototype method provides a double positive effect with respect to the target extraction of platinum metals: it reduces PM losses with tailings and eliminates PM losses with waste products in metallurgical processes in the non-ferrous metal production cycle. At the same time, an improvement in the extraction rate during ore dressing in the proposed method is provided not only for PM present in their own mineral forms, but also for PM contained in the ore in the form of an isomorphic impurity in the crystal lattice of sulfide carrier minerals: chalcopyrite, cubanite, pentlandite and pyrrhotite . This is achieved due to the side activating effect of hydromechanical processing, which takes place in a gravity concentrator with a fluidized bed of material, on the surface of PM sulfide minerals-carriers, increasing their flotation activity and, thereby, contributing to a more complete transition during flotation into a foam product.

 When the values of the parameters of gravity concentration are outside the stated range, the main indicators of the gravity-flotation process of ore dressing significantly worsen, approaching the results obtained by the prototype method or taking even lower values.

 An analysis of the data given in Table 2 shows that the technology for the enrichment of sulfide copper-nickel ores with a mass ratio of the sum of sulfides and magnetite to the sum of oxides of silicon and aluminum in the feed of the gravity concentration process less than 1: 2 (op. 2-12, 38) The following regularities are characteristic.

 In particular, in op. 5, when the mass content in the crushed ore of 70% of the particle size class is less than 74 microns (extremely high degree of grinding of the ore), the content of PM in the gravity concentrate decreased from 3564-4132 g / t (in the claimed range - op.2-4) to 1121 g / t with a simultaneous decrease in the extraction of platinum in the gravity concentrate by 31.2% abs., and the amount of PM - by 14.1% abs. in comparison with op.2. In addition, this mode is characterized by a high level of PM losses with waste tailings,%: platinum - 31.5; palladium - 2.94; the amount of PM - 30.1, approaching in this indicator to the prototype method (op. 1), in which the data loss of the PM is 34.3, respectively; 28.1 and 30.2%.

 When grinding the initial ore to a particle size of 25% of the class less than 74 microns (op. 6), i.e., with an extremely low degree of grinding of the ore, the quality of gravity concentrate sharply decreases (PM content is 935 g / t) and by 5.9% abs. the extraction of the amount of PM into it is reduced. Another serious drawback of this regime is the high level of irretrievable losses of all PM with waste tailings (platinum - 28.7; palladium - 26.6; rhodium - 24.8; the amount of PM - 27.3), which is comparable with the level of PM losses in prototype method (op.1) and is explained by insufficient depth of disclosure of mineral splices.

Gravel enrichment of the material with a maximum value of the centrifugal Froude criterion equal to 12.0 (op. 7, with an extremely high value of the maximum Froude criterion) reveals a tendency to a decrease in the amount of PM in the gravity concentrate to 2270 g / t compared with the declared range of values (2.0- 10.0), in which concentrates containing 3564-4132 g / t of PM amount were obtained (op. 3564-4132 g / t of PM amount (op. 2-4). In this case, PM extraction into the gravity concentrate is kept at a high level, %: platinum - 48.1; palladium - 11.7; rhodium - 1.1; PM amounts -21.5, and PM losses with tailings tailings eniya correspond to the level of losses, corresponding to the range of optimal values of Fr _n. Thus, maintaining gravioobogascheniya process at prohibitively high values Fr _u is impractical because it leads to lower quality of gravity concentrate and simultaneously increases the speed of the abrasive working surface wear of the centrifuge that reduces the turnaround time operation gravioapparata and unnecessarily increases operating costs.

With a prohibitively low value of the centrifugal Froude criterion of 1.0 (op. 8), the quality of the gravity concentrate remains at a fairly high level (3961 g / t of the total PM), however, the extraction of PM into the gravity concentrate drops sharply compared to the regime with the optimal value of Fr _c (op.2) platinum - 3.7 times; palladium - 3.8 times; rhodium - 3 times; PM amounts - by 3.7 times and losses of PM with tailings (platinum - by 9.5% abs.; palladium - by 9.8% abs.; rhodium - by 9.4% abs.; amounts of PM - increase significantly by 9.7% abs.).

In experiment 9, the ratio of the maximum value of the centrifugal Froude criterion to the pressure of the fluidizing water in the jacket of the gravity concentrate was 0.25 kPa ^-1 , which corresponded to a prohibitively high value of this parameter, the optimal range for which was determined to be equal to 0.025-0.23 kPa ^-1 . In this mode, the degree of PM concentration during gravel ore beneficiation decreased by about 4 times as compared to option 2, as a result of which a low-quality gravity concentrate (1046 g / t PM amount) was obtained, which is not suitable for direct processing of platinum sludge in the metallurgical finishing cycle according to the "short "scheme for the production of commercial platinum concentrates. By the level of PM extraction into the gravity concentrate (platinum - 49.5%; PM amounts - 22.0%) and their losses with tailings (platinum - 9.2%; PM amounts - 7.5%), this regime did not practically differ from mode in op. 2 and 3 (by the proposed method).

The extremely low ratio of the maximum value of the centrifugal Froude criterion to the pressure of the fluidizing water - 0.022 kPa ^-1 (op. 10) allows to obtain a high-quality gravity concentrate containing 4243 g / t of PM amount. However, at the same time, there is a sharp decrease in the extraction of all PM in the gravity concentrate, which compared with the proposed regime (option 2) is: platinum - 5.8 times; palladium - at 6.0; rhodium - 5.5 times; the amount of PM - 5.9 times. Another disadvantage of the regime is the high level of PM losses with tailings,%: platinum - 21.3; palladium - 18.9; rhodium - 22.6; the amount of PM - 19.8, which, on average, ~ 2.4 times higher than in the claimed mode (op.2).

 Gravel enrichment of the material with a mass ratio W: T of feed equal to 7.0 (op. 11), characterized by a prohibitively high proportion of the liquid phase, allows one to obtain high-quality gravity concentrate (4271 g / t of total PM) and at the same time a high level of PM extraction into it,%: platinum - 48.0; palladium - 11.7; rhodium - 1.1; the amount of PM - 21.4. These indicators slightly differ from the results obtained in experiments on the proposed method (op. 2-4). However, the high level of pulp liquefaction in this mode during its industrial implementation will require the installation of additional units of dewatering equipment (thickeners, hydrocyclones, etc.), which will significantly complicate the circuit diagram of the enrichment redistribution apparatus, increase capital investments and increase the level of operating costs. Therefore, the gravitational enrichment of the material using excessively liquefied food (L: T greater than 6.5: 1) is economically unjustified.

 However, the prohibitively high density of nutrition (op. 12) leads to a sharp deterioration in all indicators of gravity enrichment. In particular, in this experiment, when the ratio of W: T supply is 1: 1, the quality of the obtained gravity concentrate in terms of the amount of PM (372 g / t) is lower than in the prototype method (op.1). Due to the increased viscosity of the treated pulp, this regime is characterized by a very low level of PM extraction into the gravity concentrate,%: platinum - 6.4; palladium - 1.6; rhodium - 0.2; the amount of PM - 2.8, which approximately coincides with the level of extraction of PM in the prototype method. At the same time, PM losses with tailings of enrichment are,%: platinum - 25.8; palladium - 23.6; rhodium - 27.0; the amount of PM - 24.4, on average, ~ 2.9 times higher than in the proposed method (op.2).

 Experiments 14-24, 26, 36, and 40 illustrate the results of gravity-flotation concentration of sulfide copper-nickel ore when a material enters the process of gravity concentration in which the mass ratio of the sum of sulfides and magnetite to the sum of silicon and aluminum oxides is less than or equal to 1: 2. In these experiments, regardless of the type and mineral variety of the initial ore, when the regime parameters deviate from the declared range of values, the same tendencies in the change of indicators appear as described above for materials with a criterion ratio greater than 1: 2.

 In general, an analysis of the results shows that the use of the proposed method for the beneficiation of sulfide copper-nickel ores in comparison with the prototype allows for optimal gravel concentration parameters to significantly increase the technical and economic indicators for the extraction of platinum metals from all types and main mineral varieties of this category of ores containing their own minerals PM and magnetite.

 The proposed method provides the greatest efficiency in the enrichment of poor disseminated ores, including when processing their low-sulfide variety. Using the proposed method for the enrichment of this type of ores, characterized by a criterion ratio of less than 1: 2, provides gravity concentrates with a total PM content of 6.5-7.3 times higher (op. 2 and 38) than in the corresponding experiments carried out by the method prototype (op. 1 and 37). The level of extraction of PM in the gravity concentrate in the known method is significantly higher than in the prototype: platinum - 42.5-50.7% abs .; palladium - by 9.0-9.1% abs. ; rhodium - at 0.9% abs .; the amount of PM - by 18.2-19.7% abs. at the same time a higher productivity of the process for the initial nutrition and other equal conditions for its pulp preparation. Another important advantage of the proposed method is the low level of PM losses with dump tailings, which for disseminated ore is (op. 2 and 38),%: platinum - 8.7-10.1; palladium - 6.9-7.4; rhodium - 11.5-12.5; the amount of PM - 7.6-8.4. This, on average, is 3.4-3.6 times less than the PM loss when using the prototype method. And finally, the serious advantages of the proposed method, of course, include the possibility of processing the resulting gravity concentrates (up to 4 kg / t PM amount) according to the "short" scheme, which avoids PM losses in the non-ferrous metal production cycle.

Claims (2)

1. A method of beneficiation of sulfide copper-nickel ores containing proprietary minerals of platinum metals and magnetite, including ore preparation, wet grinding of the material and its hydraulic classification, separation of non-ferrous sulfides and proprietary minerals of platinum metals from the pulp of the classified material by flotation and gravity methods into independent products - sulfide (e) flotation concentrate (s) and a platinum-containing gravity concentrate, and magnetite and rock-forming minerals into dump tailings, and own platinum metal minerals are isolated on a centrifugal concentrator with a fluidized bed created by water jets in a direction that does not coincide with the centrifugal field force vector, characterized in that the platinum native minerals are isolated in a platinum-containing gravity concentrate before the flotation enrichment operation, while the mass ratio of the sum sulfides and magnetite and the sum of silicon and aluminum oxides in the original ore, less than 1: 2, the ore is crushed to a particle size of 30 - 65% of the class less than 74 microns, and division own minerals platinum metals are at the maximum centrifugal criterion Froude 2 - 10 and the ratio of this value to the pressure fluidizing water 0.025 - 0.23 kPa ^-1 when processing the original ore weight ratio of sulfides and amounts of magnetite and the amount of silicon and aluminum oxides greater than or equal to 1: 2, the ore is crushed to a particle size of 60 - 95% of the class less than 74 microns, and the extraction of native platinum minerals is carried out with a maximum value of the centrifugal Froude criterion of 0.5 - 1.75 and the ratio of this value to pressure fluidizing water 0.0058 - 0.019 kPa ^-1.  2. The method according to claim 1, characterized in that the allocation of their own minerals of platinum metals is carried out from the pulp of the material with a mass ratio W: T (1.5 - 6.5): 1.

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