RU2530945C2 - Method of three-stage technological parameters optimisation of centrifugal enrichment for recovery of precious metals in mineral form from ores, tailings from processing of embedded copper-nickel ores of norilsk deposits - Google Patents

Abstract

FIELD: mining.

SUBSTANCE: invention relates to the mining industry and can be used to increase the extraction of valuable elements from ores and products of their processing, in particular for the recovery of precious metals in mineral form and in part of sulphides of copper, nickel, iron from old tailings of the preserved tailing dump located in the Norilsk industrial district. The method of three-stage process optimisation of the parameters of the centrifugal enrichment for the recovery of precious metals in the mineral form from ores, tailings during processing of embedded copper-nickel ores of Norilsk deposits comprises selection of its own minerals of platinum group metals in the gravity concentrate before the flotation enrichment operation at a weight ratio of the amount of sulphides and magnetite and the amount of oxides of silicon and aluminium in the initial ore or tailings less than 1:2, the size of 30-65% class less than 74 microns. Isolation of own minerals of platinum metals is carried out at the value of the centrifugal Froude number of 11.75 and ratio of this value to the fluidising water pressure of 0.085 kPa. The method of optimising the parameters of centrifugal enrichment comprises the sequence of operations on the centrifugal separators, designed for continuous operation in industrial environments. At the first step an optimum accumulation time of the concentrate is determined. At the second step the optimum speed of jets of water or optimal water flow through the openings to the separator bowl inter-ripple space is selected. At the third stage the speed of jets of water or water flow through the openings to the separator bowl inter-ripple space is stepwise increased, starting with the optimal water flow, defined in the second step and the optimum time interval of accumulation of concentrate defined in the first step.

EFFECT: improving the efficiency of the recovery of precious metals from ores, tailings from the processing of ores of Norilsk deposits, as well as improving the efficiency of the optimisation parameters of centrifugal enrichment.

3 cl, 2 dwg, 4 tbl

Description

Method of three-stage technological optimization of centrifugal beneficiation parameters for extraction of precious metals in mineral form from ores, tailings from processing disseminated copper-nickel ores of Norilsk deposits.

The invention relates to the mining industry and can be used to increase the extraction of valuable elements from ores and products of their processing, in particular to extract precious metals in mineral form and partially sulfides of copper, nickel, iron from the stale tailings of a preserved tailing dump located in the Norilsk industrial region.

To ensure the achievement of high performance in centrifugal enrichment, it is necessary to take into account the mineral composition of the initial product, the form of the noble and non-ferrous metals, their physical properties, as well as the technical specifics of centrifugal separators. The invention also relates to optimizing the operation parameters of centrifugal separators in order to increase the content of precious metals in the concentrate.

The tails of the mothballed tailings (storage completed in 1975) have a complex mineral composition identical to the ore grades of industrially processed ores of the Norilsk-1 deposit from 1949 to the present. Noble metal metals are in two forms: mineral and scattered. (Yushko-Zakharova O.E. Platinum content of ore deposits. M., Nedra, 1975, pl. 52, 53, 57). According to the microprobe analysis, more than 20 noble metal minerals were found in the tailings. The size of the mineral emissions of precious metals is from 1-5 microns to 150-200 microns or more. Noble metal minerals (MBM) with a particle size of more than 70 μm are not extracted during flotation enrichment (Kovalenko L.N., Blagodatin Yu.V., Golubeva TD, Lomteva L.L. Form of the presence of noble metal minerals in flotated enrichment products sulfide ores of the Norilsk group of deposits // Mineral processing No. 1-2. 1993, p. 18-25). Natural alloys of noble metals, solid and malleable, magnetic and non-magnetic with a density of 13-19 kg / dm ³ , sulfides and arsenides of noble metals, hard and brittle with a density of 9-13 kg / dm ³ , are lost with dump tails in the form of free, large riveted particles 100-400 microns in size and over-crushed, slurred particles less than 25 microns in size.

A known method of enrichment of the material of the mothballed tailings from the enrichment of sulfide normally disseminated ores of the Norilsk-1 deposit. In these tails, MBM are in mineral and dispersed forms.

In the known method, the tails are subjected to two-stage screening according to particle size classes of 15 and 1.5 mm, de-slamming in battery hydrocyclones with a diameter of 250 mm according to particle size less than 20 microns. The discharge of hydrocyclones with a particle size of less than 20 microns is sent for storage in the tailing dump. A class larger than 20 microns (hydrocyclone sands) is sent for centrifugal enrichment to Knelson-48 separators for concentration of MBM larger than 70 microns. The tailings of centrifugal separators are subjected to flotation with a purification of the concentrate to extract MBM with a particle size of less than 70 microns, intergrowths of MBM with sulfides and free sulfides of copper, nickel, and iron. The concentrate of centrifugal and flotation concentration is sent together to the hydrotransport system of the copper or nickel concentrate of the processing plant and then to the pyrometallurgical redistribution (Blagodatin Yu.V., Yatsenko A.A., Zakharov B.A., Chegodaev V.D., Alekseeva L.I. Involvement in the processing of new raw materials of non-ferrous and precious metals. // Non-ferrous metals. 2003 G. No. 8-9. P. 28-29).

The disadvantage of this method is the low content of noble metals in the concentrate, and as a result, their low recovery. In this case, according to our research, in contrast to the known method, noble metals are transferred to secondary tails (after ore dressing and tailings from the tailings by centrifugal separation and flotation), which occur in classes larger than 50 μm and less than 50 μm (table one).

Known designs of separators in which the separation of fine-grained material by density occurs in a centrifugal field, which acts on particles of different density and size in the liquid phase, located in a rotating working body. To prevent particle compaction, under the action of centrifugal forces, water is supplied into the inter-groove space through a large number of small-diameter holes, under pressure, in the form of thin jets to create a fluidized bed that facilitates mixing of particles and removes low-density particles from the separation zone. The most widely used are separators of the Canadian company “KNELSON” (Tsarkov V. A. Foreign apparatus for centrifugal enrichment. // Mining Journal. 1999, No. 3, p. 76-79). The main disadvantage of this type of separator is that when it enriches fine-grained materials, heavy particles of a certain size penetrate intensively through openings through which fluidizing water under pressure is supplied (Bogdanovich A.V. Intensification of gravity concentration processes in centrifugal fields. // Ore concentration . 1999, No. 1-2. P. 33-35). Our experience has established another drawback, it is driving holes with large heavy particles of minerals, which leads to the termination of liquefaction and compaction of the concentrate in the inter-groove space. Ultimately, the quality of the concentrate is reduced in terms of the content of precious metals and, as a result, a decrease in their extraction. To clean the holes, the bowl must be dismantled and mechanically cleaned approximately every 240 hours. The next significant drawback is the long time that the centrifugal separators set up for optimal operating conditions - from several weeks to several months (L. Mitin, On Some Paradoxes in the Hydrogravity Process of Mineral Processing. Kolyma. 2003, No. 1. p. 26-30 ) This drawback is also confirmed by our experience.

Known designs and domestic centrifugal separators of this type. For example, patent RU 2278735 (or patent RU 2196004, where a separator with a varying angle of inclination of the rotor is proposed), in which, unlike the KNELSON separator, water is supplied to the inter-space through diffusers having a cross-section in the shape of rectangles, the water supply to the inter-coil the space is in the form of a pulsating flow using a pulsation source. The disadvantage of this separator is the complexity of the design due to the additional ripple mechanism and the lack of industrial testing data on the advantages of this type of separator.

There is also a method for concentrating heavy minerals by creating mineral particles in a precipitating spiral flow of a pulsating radial component of motion by creating concentration grooves with a variable radius from the center of rotation of the bowl on the inner surface of the bowl; in this separator, water is not supplied to the inter-groove to create a loosening layer (patent RU 2423183), and heavy grains of minerals are held by the rotational component of the spiral flow. The disadvantages of this design of the separator include the following: it is unclear how the concentrate is unloaded, the complexity of manufacturing a bowl with a changing radius of the concentration grooves on the inner surface of the bowl, and also the lack of industrial test data on the advantages of this type of separator.

A common drawback in the above descriptions of the designs of centrifugal separators and methods of enrichment (separation) is the lack of methods for optimizing technological parameters in industrial conditions, with the aim of accelerated tuning and further operation of the separators in optimal conditions.

Closest to the proposed method according to some criteria and the achieved result is a method of beneficiation of sulfide copper-nickel ores containing their own minerals of platinum metals and magnetite. Prototype (patent RU 2144429).

A method of extracting valuable elements from copper-nickel ores containing proprietary minerals of platinum metals and magnetite, including ore preparation, grinding of ore and its hydraulic classification, separation of proprietary minerals of platinum metals by centrifugal separation with a fluidized bed into an independent concentrate before flotation. The result is a platinum-containing gravity concentrate, sulfide flotation concentrates, tailings containing waste rock and magnetite. In this method, optimal centrifugal concentration conditions are determined for two types of ores with different ratios of the sum of sulfides and magnetite to the sum of oxides of silicon and aluminum. When the value of this indicator is less than 1: 2, the ore is crushed to a particle size of 30-65% of the class less than 74 microns, centrifugal enrichment is carried out at a maximum value of the Froude parameter of 2-10 and the ratio of this value to the pressure of the fluidizing water is 0.025-0.23 kPa ^-1 . With a value of this indicator equal to or greater than 1: 2, the ore must be crushed to a particle size of 60-95% of the class less than 74 microns and centrifugal enrichment must be carried out with a maximum value of the Froude parameter of 0.5-1.75 and the ratio of this value to the pressure of the fluidizing water 0.0058-0.19 kPa ^-1 .

A serious disadvantage of the prototype is that one of the optimizing criteria is a parameter - water pressure, which depends on the state of the holes (clogged or free) through which water flows to create a fluidized bed. If the holes are clogged and the bed is pressed into the concentration grooves, the water pressure will be set or higher than in the regime, and the water flow will decrease and the enrichment rate will decrease.

The next disadvantage of the prototype is the use in experiments of a laboratory model of a Knelson centrifugal separator (manual unloading) with a bowl diameter of 190.5 mm (7.5 inches). According to literature data (Bocharov V.A., Gurikov A.V., Gurikov V.V. Analysis of the processes of separation of gold-containing products in concentrators KNELSON and FALKON SB. // Mineral processing. 2002, No. 2, p. 19) and Our experience has established that the performance of the enrichment of finely dispersed minerals obtained on a laboratory separator is significantly higher than the industrial ones, and the optimum conditions obtained are not modeled on industrial models of separators.

The next disadvantage of the prototype is the use as a parameter of the ratio of the sum of sulfides and magnetite to the sum of silicon and aluminum oxides in the feed when beneficiating ore and industrial raw materials, which is almost impossible to quickly analyze due to the duration of the analysis and to quickly change the operation mode of centrifugal separators when beneficiating common tails from these types of ores. This parameter, the ratio of the sum of sulfides and magnetite to the sum of oxides of silicon and aluminum will change in the total stored tailings with a change in the volume of ore processing.

The problem solved by the invention is to optimize the parameters of centrifugal enrichment to increase the extraction of precious metals from the tailings from ore processing of the Norilsk deposits, reducing the time to bring the process parameters to optimal conditions.

The technical result achieved by the implementation of the invention is to obtain a centrifugal concentration concentrate, in which a high content of precious metals (increase in their extraction) is achieved due to the operation of centrifugal separators in optimal conditions.

The problem is solved in that in the proposed method for the technological optimization of centrifugal enrichment parameters, according to the invention, the optimal parameters are purposefully selected in three stages under industrial conditions: at the first stage, the accumulation time of the concentrate is selected; on the second, the speed of the water jets (water flow) is selected through the holes in the inter-groove space of the separator bowl to create a fluidized bed of particles; in the third step, the speed of the jets (water flow) is increased in an optimal time period for the accumulation of concentrate, which is determined in the first stage, in order to free clogged holes from particles with a size close to the diameter of the holes through which fluidizing water is supplied.

The efficiency of separation of grains of various densities in a rotating stream of an industrial centrifugal separator depends on a large number of factors acting on the particle:

1. centrifugal force (bowl diameter and number of revolutions);

2. The vertical velocity of the pulp flow (volumetric capacity);

3. pulp density and particle size in the separator feed;

4. the time of accumulation of the concentrate;

5. flow rate (speed of water jets) of fluidizing water through openings into the inter-groove space in the enrichment zone.

All of these parameters are variable and their regulation limits are laid down in the design of the separator, with the exception of fineness and density of minerals in the initial power of the separators. Based on the literature data, it can be concluded that when enriching finely divided raw materials in centrifugal separators, the magnitude of the centrifugal force or separation factor is not a determining criterion for evaluating work efficiency (L. Mitin On some paradoxes in the hydrogravity process of mineral processing. Kolyma. 2003 G. № 1. S. 29), and the magnitude of the centrifugal force of 60 g is optimal (V. Tsarkov. Foreign apparatus for centrifugal enrichment. // Mining Journal. 1999 G. No. 3, S. 77). During the tests, in order to determine the optimal parameters and to avoid distortion of the parameters of subsequent technological operations, it is necessary to stably maintain the volumetric capacity and power density in accordance with the technical characteristics of the separator and the requirement of the technological instruction for enrichment for this type of raw material.

Thus, the task of optimizing the parameters of centrifugal enrichment in an industrial environment for finely divided raw materials containing noble metals in mineral form is reduced to determining the accumulation time of the concentrate and the flow rate (speed of the water jets) of fluidizing water into the interfloor space.

The goal to be solved in the first and second stages is to obtain a concentrate by the content of the amount of precious metals close to their content in copper or nickel concentrates obtained in the ore mining cycle and delivered to the corresponding metallurgical plants, under conditions of optimal concentrate accumulation time and flow rate (stream velocity) of fluidizing water.

The goal at the third stage is a further significant increase in the content of the amount of precious metals in the centrifugal concentration concentrate: by increasing the speed of the jets in the inter-groove space and freeing the holes from large particles of minerals, several times exceeding the maximum permissible sizes of feed particles (falling into the material, possibly from due to non-observance of technological parameters), recovery of liquefaction and reduction of compaction of concentrate particles in the inter-groove space of the separator bowl. (Fedotov KV Theory and practice of enrichment of gold-containing raw materials in centrifugal concentrators. Abstract of dissertation for the degree of Doctor of Technical Sciences. Irkutsk. ISU. 2000, p. 21).

It is known that during centrifugal enrichment of finely dispersed raw materials, the effect of antisegregation forces increases, which changes the looseness of particles in the inter-groove space, and they are pressed. With a decrease in the size of the particles to be separated, the effect of friction forces increases. When the particle size is tens of micrometers, the friction forces are greater than the separation forces, which leads to a decrease in enrichment indicators (L. Mitin, On some paradoxes in the hydrogravity process of mineral processing. Kolyma. 2003, No. 1. P. 29).

In our case, raw materials are enriched with a fineness class content of less than 0.074 mm in 47%, which also causes blocking of the holes for supplying water to the inter-groove space. An increase in the amount of water (speed of the jets) will lead to a decrease in the number of compacted zones formed due to the friction forces of particles in the inter-groove space, and will favorably affect the technological parameters.

In centrifugal enrichment at the point of entry of the jet into the inter-groove space, oppositely directed forces act on the particle: centrifugal (Fс) and the pressure force of the water jet (Fst). If the centrifugal force acting on the particle, the size of which is commensurate with the size of the holes for supplying water for liquefaction in the inter-groove space, exceeds the pressure force of the water jet at the point of entry into the inter-groove space, then the holes become clogged, the liquefaction stops, the friction forces increase and the particles are compressed in the inter-groove space. The centrifugal force acting on the particle is calculated by the formula:

- угловая скорость вращения чаши, рад/сек; Vт - объем частицы, м³; (для упрощения расчета форма частицы принята шарообразной); r - радиус вращения частицы, м (в расчете - средний радиус чаши сепаратора); ρ_т - плотность частиц, кг/м³ (кварц, пирротин, тетраферроплатина); ρ_ж - плотность жидкости, кг/м³ (вода).where: Фц - centrifugal force, (H);

- angular velocity of rotation of the bowl, rad / sec; Vt is the particle volume, m ³ ; (to simplify the calculation, the particle shape is adopted spherical); r is the radius of rotation of the particle, m (in the calculation - the average radius of the separator bowl); ρ _t - particle density, kg / m ³ (quartz, pyrrhotite, tetraferroplatinum); ρ _W - the density of the liquid, kg / m ³ (water).

, где:

- количество оборотов чаши сепаратора, 316 об/мин; π - 3,14), так как пристенные слои жидкости вместе с частицами вращаются с той же угловой скоростью, что и стенка межрифельного пространства чаши, у внутренней поверхности чаши окружное смещение жидкости составляет всего несколько процентов величиной, которой можно пренебречь. (Лопатин А.Г. Центробежное обогащение руд и песков. М., Недра, 1987 г. С.179).In the calculation, the particle rotation speed is equal to the rotation speed of the separator bowl (

where:

- the number of revolutions of the separator bowl, 316 rpm; π - 3.14), since the wall layers of the liquid together with the particles rotate at the same angular velocity as the wall of the inter-rip space of the bowl, at the inner surface of the bowl the circumferential fluid displacement is only a few percent, which can be neglected. (Lopatin A.G. Centrifugal beneficiation of ores and sand. M., Nedra, 1987, p. 179).

where: Fst is the force of the jet per particle, H; ρ _W - the density of the liquid, kg / m ³ (water); S is the cross-sectional area of the jet, m ² ; V is the jet velocity, m / s. (Chugaev PP Hydraulics. Energy Publishing, 1982, p. 123).

The separator bowl has six storage channels (inter-groove space) for the concentrate, which includes 1780 openings with a diameter of 1.8 mm for supplying water for liquefaction.

The kinematic viscosity coefficient of water is small, it is not involved in the calculation (Fedotov K.V., Romanchenko A.A., Senchenko A.E. Calculation of the velocities of hydrodynamic flows in a centrifugal concentrator. // Mining Journal. 1998, No. 5 , p.23).

Table 2 shows the calculation data of the comparative values of the centrifugal force, the force of the jet of water for liquefaction at the point of entry of the jet into the inter-groove space of the centrifugal separator for particles of quartz, pyrrhotite and tetraferroplatinum.

The tailing dump was formed from ore processing at the Norilsk-1 deposit. The concentration of mineralization is in picritic and taxitic gabbro-dolerites (containing rocks) (Genkin AD, Distler VV, Gladyshev GD and others. Sulfide copper-nickel ores of the Norilsk deposits. M., Nauka, 1981 . Page 39). The hard-to-grind minerals that make up their composition have a hardness of 6.5-7 and after grinding are in large classes, they also include quartz, light and hard, also hard to grind. In the Norilsk ores, the minerals of the pyrrhotite group are the most abundant, platinum is represented by ferruginous forms, of which tetraferroplatinum, ferroplatinum, and the most abundant (Genkin AD, Distler VV Gladyshev GD, etc. Sulfide copper-nickel ores of the Norilsk deposits M. Science. 1981, p. 63, 113).

Table 2 shows that the centrifugal force acting on particles of quartz or the host rocks of gabbro-dolerite with a grain size of up to 1.8 mm exceeds the force of the jet at the point of its entry into the inter-groove space at a jet velocity of 1.39; 1.52; 1.67 m / s At a jet velocity of 1.81 m / s, the jet force already exceeds the centrifugal force and may prevent the penetration of particles into the holes or release already clogged ones.

Minerals of the pyrrhotite group (or other sulfides of chalcopyrite, pentlandite, cubanite) can also block inlet openings for supplying fluidizing water. Sulfides are brittle (Doroshenko MV, Bashlykova TV Technological properties of minerals. M. Heat engineering. 2007, pp. 24, 75, 78, 115) compared with the host rocks and their penetration by unmilled particles with a particle size of more than 1.5 mm unlikely.

Minerals of platinum (glandular forms) are malleable and can, after grinding, come out riveted in the form of plates, flakes. Their size in the deposit is up to 0.2 mm (Genkin A.D., Distler V.V. Gladyshev G.D. et al. Sulphide copper-nickel ores of the Norilsk deposits. M. Nauka. 1981, p. 113), therefore very low probability of these particles blocking fluidizing water inlets.

Information on the use of the method of three-stage technological optimization of centrifugal enrichment parameters for finely divided raw materials containing noble metals, including a sequence of operations under industrial conditions: at the first stage, the accumulation time of the concentrate is determined, at the second stage, the speed of the water jets (water flow) is selected through the openings in the inter-groove space of the separator bowl , at the third stage, the speed of the jets (water flow) is stepwise increased, starting with the optimal water flow defined at the second stage, and in the optimal period of time the accumulation of the concentrate, defined at the first stage, is not revealed in the patent and scientific literature. There is also no information about the fame of the distinguishing features of the proposed method in the application of a stepped mode of water flow (jet velocity) upward, starting from the optimal and in the optimal period of time the accumulation of concentrate, when used in industrial and laboratory research. Therefore, the claimed method meets the criterion of "Inventive step". The effectiveness of the proposed method is the result of the combined action of a three-stage technological optimization of the parameters of centrifugal enrichment.

The method is as follows. The tailings tailing factory for the mothballed tailings (SG Cheburashkin, NI Geonya. Analysis of the implementation of a multi-purpose industrial investment project for the enrichment of the platinum deposit. // Gold and Technology. 2010 No. 2, p. 39) works according to the following scheme. The pulp of tails from the quarry field is classified according to the class of 1.8 mm on vibrating screens. The under-sieve product is sent for de-slamming into two hydrocyclone batteries of 16 pcs. in each battery. The diameter of hydrocyclones is 250 mm. Sands of hydrocyclones will be tested with a cross sampler (PRO-6M), a private sample is cut off by a divider (GDP-4P) and dehydrated (OP-3M). Density of hydrocyclone sands is controlled by a radioisotope densitometer (Berthold LB-444-01), the volume of sands is controlled by an electromagnetic flow meter (ABB FSM4000 / SE41F) and pumped to a pressure pulp splitter and then to 4 centrifugal separators (KC-XD48) operating according to a given time program, excluding moments of their simultaneous unloading of the concentrate. The concentrate will also be tested with a cross sampler, density control with a radioisotope densitometer, pulp volume measurement with an electromagnetic flow meter. The tails of centrifugal separators are treated with a collector (butyl xanthate), a blowing agent (pine oil) and the main flotation is carried out with the following two cleanings. The centrifugal concentration concentrate and the second purification concentrate are combined and pumped to the hydrotransport system of concentrates obtained from ore raw materials. The tailings of the main flotation together with the drains of the hydrocyclones and the plus product of the screens are sent to the existing tailing dump.

At the first stage, in industrial conditions, the optimal time of accumulation of the concentrate is determined at 10, 16, 20, 24, 30, 44, 60, 120 minutes. During the specified accumulation time, all four separators operate, with the exception of the accumulation time at 10 and 16 minutes, where two separators operate at this time to avoid simultaneous unloading. In determining the optimal accumulation time, the flow rate of fluidizing water (jet velocity) is the same for all time modes and is 22.71 m ³ / h (1.39 m / s). Consider the procedure for an accumulation time of 24 minutes. All four KC-XD48 separators are in operation. On the interface of the control panel of each separator, the accumulation time of the concentrate is 24 minutes. All four are launched at 5-minute intervals to prevent simultaneous unloading. The sampler goes into continuous operation. The readings of densitometers, flow meters installed on the sands of hydrocyclones (feeding separators KC-XD48) and centrifugal concentration concentrate are recorded. In centrifugal enrichment of finely divided raw materials, the parameter determining the extraction of precious metals is their content in the concentrate, the mass of the concentrate varies slightly. During the experiments, for the reliability of the obtained indicators, it is necessary that each separator produce three unloadings. At the end of the experiment, the readings of densitometers, flow meters are again recorded. Samples of the initial food and concentrate are taken from the dehydrator, they are dried, weighed, reduced, divided for sieve and chemical analysis into non-ferrous and noble metals. Similarly, actions are performed at a different time of accumulation of the concentrate. Based on the results of measurements of the contents of noble metals, a graph is constructed according to which the optimal accumulation time of the concentrate is determined.

In the second stage, also under industrial conditions, the optimum water flow rate (jet velocity) is determined for liquefaction in the inter-groove space at the optimal accumulation time determined in the first stage. The procedure for performing the second stage of optimization is similar to the first. Four separators operate in all modes. At the control panel interface, the required fluidizing water flow is set for all four separators. Water consumption (speed of the jets) varies according to experiments, m ³ / h (m / s): 22.71 (1.39); 24.98 (1.52); 27.25 (1.67); 29, 52 (1.81); 31.79 (1.94); 36.34 (2.07). Based on the results of measurements of the contents of noble metals, a graph is constructed according to which the optimal water flow (jet velocity) is determined.

At the third stage, also under industrial conditions, four experiments are carried out to obtain reliable results.

The first experiment is standard, lasting 240 hours, is carried out in the optimal parameters mode obtained at the first and second stages of optimization with a constant flow of fluidizing water, the procedure is similar to the first stage of optimization, only all four separators are in operation. Samplers are switched to balance sampling mode with a cutoff interval of a private sample of 6 minutes. The second experience, also 240 hours. The procedure is similar to the first experiment. The accumulation time of the concentrate is 24 minutes, the step-by-step flow rate of fluidizing water is 24.98 m ³ / h (jet velocity 1.52 m / s) for 10 minutes, the next 10 minutes the flow rate of fluidizing water is 27.52 m ³ / h (jet velocity 1, 67 m / s) and 4 min water consumption 29.52 m ³ / h (jet velocity 1.81 m / s), at which the jet force exceeds centrifugal. The third experience is 240 hours. The procedure is similar to the second experiment. The accumulation time of the concentrate is 20 minutes, the step-by-step flow rate of fluidizing water is 24.98 m ³ / h (jet velocity 1.52 m / s) for 10 minutes, the next 6 minutes the flow rate of fluidizing water is 27.52 m ³ / h (jet velocity 1, 67 m / s) and 4 min water consumption 29.52 m ³ / h (jet velocity 1.81 m / s). The fourth experience is 240 hours. The procedure is similar to the second experiment. The accumulation time of the concentrate is 18 minutes, the step-by-step flow rate of fluidizing water is 24.98 m ³ / h (jet velocity 1.52 m / s) for 10 minutes, the next 6 minutes the flow rate of fluidizing water is 27.52 m ³ / h (jet velocity 1, 67 m / s) and 2 min water consumption 29.52 m ³ / h (jet velocity 1.81 m / s).

An example of a specific use of the proposed method in an industrial environment is given below in the text, figures and tables.

1. The first stage of optimization. The proposed method. Determination of the optimal accumulation time at a constant flow rate of fluidizing water (jet velocity). The flow rate of fluidizing water is 22.71 m ³ / h (1.39 m / s). The time of experiments on the accumulation of concentrate 10, 16, 20, 24, 30, 44, 60, 120 minutes Based on the results of the analyzes, a graph was built of the dependence of the content of the amount of precious metals (platinum, palladium, gold) in a centrifugal concentration concentrate (Fig. 1).

According to the graph, the optimal time of accumulation of the concentrate is in the range of 20-24 min with a total precious metals content of 25.7-26.1 g / t.

2. The second stage of optimization. The proposed method. Determination of the optimal flow rate of fluidizing water (the speed of the jets) into the inter-groove space at the optimal accumulation time determined in the first stage. Water consumption (speed of the jets) varies according to experiments, m ³ / h (m / s): 22.71 (1.39); 24.98 (1.52); 27.25 (1.67); 29, 52 (1.81); 31.79 (1.94); 36.34 (2.07). Based on the results of measurements of the contents of noble metals, a graph is plotted for the content of the amount of noble metals (platinum, palladium, gold) in a centrifugal concentration concentrate, which determines the optimal flow rate (speed of jets) of fluidizing water (Fig. 2). The graph shows that the optimal water flow is in the range of 22.7-27.25 m ³ / h (1.39-1.67 m / s). The concentrate has a content of noble metals in the range of 26.34-26.85 g / t.

3. The third stage of optimization. The proposed method. Four experiments are conducted under industrial conditions of 240 hours each, to increase the reliability of the results (table 3).

The first experiment is standard, with a concentrate accumulation time of 24 minutes and a constant flow rate of 24.98 m ³ / h (jet velocity 1.52 m / s), which are defined as optimal in the first and second stages of optimization. This experience can be attributed to the prototype. The mineral composition of hydrocyclone sands (initial feed of centrifugal enrichment) for all four experiments: pentlandite - 0.3%; chalcopyrite - 0.2%; pyrrhotite - 4%; magnetite - 5%; SiO ₂ - 40%; Al ₂ O ₃ - 10%. The content of the particle size class less than 74 microns is 50%. The ratio of the sum of sulfides and magnetite to the sum of oxides of silicon and aluminum is 0.19; Froude criterion (calculation according to patent No. 2144429, prototype) for an industrial separator KC-XD48 is 11.75. The ratio of the Froude criterion to the pressure of the fluidizing water before unloading the concentrate (in the experiments the final pressure of the fluidizing water before unloading the concentrate is 116.7-137.4 kPa) is 0.0855 kPa at a pressure of 137.4 kPa.

The technological indicators of the first experiment are given for non-ferrous and platinum metals in table 3. The content of the amount of platinum and palladium in the concentrate is 17.1 g / t, the degree of enrichment is 14.5 (the ratio of the total amount of platinum metals in the separator's feed to their content in the concentrate), extraction the amount of 1.94%. The indicators are low, despite the fact that the parameters of centrifugal enrichment are close or are within the scope of the claims of the patent of the prototype for the initial feed of centrifugal enrichment with the ratio of the sum of sulfides and magnetite to the sum of oxides of silicon and aluminum less than 1: 2.

Second experience. The accumulation time of the concentrate is 24 minutes Water consumption (speed of jets) for liquefaction in the inter-groove space is stepwise. For 10 minutes, the water flow rate (stream velocity) is 24.98 m ³ / min (1.52 m / s), the next 10 minutes 27.52 m ³ / min (1.67 m / s) and 4 min flow rate 29, 52 m ³ / min (1.81 m / s). The content of the amount of platinum and palladium is 62 g / t, the degree of enrichment is 46.6, the recovery is 6.53%. From literature data it is known that platinum in this technogenic deposit is represented by mineral forms and 87% of it is in coarseness classes -0.14 + 0.05 mm, palladium is also represented by mineral forms and 74.8% of it is in the same coarseness class. (Dodin D.A., Isoitoko V.A. Super-large technogenic deposits of platinum metals. // Mineral processing. 2006 No. 6, p.20).

Thus, it can be stated that in the second experiment with a step-by-step flow of water, an increase in this size class productive for platinum metals should occur.

The granulometric composition of centrifugal concentration concentrates with constant and stepwise water consumption (experiments 1 and 2) is shown in table 4, where it is shown that in the second experiment there was an increase in the content of the productive size class for platinum metals by 5.55%, due to the use of stepwise water consumption (jet velocity) and the release of clogged holes from mineral particles with a size close to the diameter of the holes through which fluidizing water is supplied, reducing the number of sealed zones formed due to the friction forces of the particles in the interfacial th space, which favorably influenced technological indicators. In the proposed method of three-stage technological optimization, real optimal operating conditions of centrifugal separators in industrial conditions are obtained, ready for use, in a short period of time, without preliminary laboratory studies with unrealistic indicators.

The third experience. The accumulation time of the concentrate is 20 minutes Water consumption (speed of jets) for liquefaction in the inter-groove space is stepwise. Within 10 minutes, the water flow rate (stream velocity) 24.98 m ³ / min (1.52 m / s), the next 6 minutes 27.52 m ³ / min (1.67 m / s) and 4 min flow rate 29, 52 m ³ / min (1.81 m / s). The content of platinum and palladium is 49.2 g / t, the concentration is 42.4, and the recovery is 6.4%. The content of the amount of platinum and palladium in the concentrate is slightly lower than in the two experiments, due to a decrease in the accumulation time of the concentrate, but the recovery remained at the same level.

Fourth experience. The accumulation time of the concentrate is 18 minutes Water consumption (speed of jets) for liquefaction in the inter-groove space is stepwise. Within 10 minutes, the water flow rate (jet velocity) is 24.98 m ³ / min (1.52 m / s), the next 6 minutes 27.52 m ³ / min (1.67 m / s) and 2 min flow rate 29, 52 m ³ / min (1.81 m / s). The content of the amount of platinum and palladium is 31 g / t, the degree of enrichment is 29.2, the recovery is 4.09%. The content of the sum of platinum and palladium in the concentrate is less in comparison with the second and third experiments, due to a decrease in the total accumulation time of the concentrate, a decrease in operating time at a jet speed of 1.81 m / s, when clogged holes are released to liquefy the separator interframe space and increase the quantity compacted zones formed due to the forces of friction of particles in the inter-groove space.

Table 1 The distribution of metals by size classes in the tailings after ore dressing and tailings dressing from the tailings by centrifugal separation and flotation. No. p / p Size class (mm) Exit, % Content,%, g / t Distribution% Ni Cu Pt Pd Pt + pd Ni Cu Pt Pd Pt + pd one +0.2 0.74 0.09 0.10 0.68 0.85 1,53 0.53 1.18 0.66 0.85 0.75 2 -0.2 + 0.14 14.2 0,07 0.04 0.39 0.6 0.99 7.92 9.05 7.28 11.49 9.36 3 -0.14 + 0.09 21.05 0.08 0.05 0.71 0.49 1,2 13,42 16.79 19.65 13.90 16.81 four -0.09 + 0.071 16.95 0.11 0.05 1.33 0.85 2.18 14.86 13.50 29.64 19.42 24.59 5 -0.071 + 0.05 12.58 0.11 0.05 0.64 0.71 1.35 11.05 10.03 10.58 12.04 11.30 6 -0.05 + 0.00 34.48 0.19 0.09 0.71 0.91 1,62 52.22 49.45 32.19 42.30 37.19 7 Source product one hundred 0.12 0,063 0.76 0.74 1,5 one hundred one hundred one hundred one hundred one hundred

table 2 Estimated comparative values for particles of quartz, pyrrhotite, tetraferroplatinum, centrifugal force and water jet force for liquefaction at the point where the jet enters the inter-groove space of the centrifugal separator KC-XD48 * No. p / p Particle size (mm) Vst = 1.3 9 m / s (22.71 m ³ / h) Vst = 1.52 m / s (24.98 m ³ / h) Vst = 1.67 m / s (27.25 m3 / h) Vst = 1.81 m / s (29.52 m ³ / h) Fc (N) × 10 ^-3 Fc (H) × 10 ^-3

$$\frac{Fc}{Fc}$$

$$\frac{F\mathrm{from}}{Fc}$$

Fc (N) × 10 ^-3 Fc (H) × 10 ^-3 $$\frac{Fc}{Fc}$$

$$\frac{F\mathrm{from}}{Fc}$$

Fc (N) × 10 ^-3 Fc (H) × 10 ^-3 $$\frac{Fc}{Fc}$$

$$\frac{F\mathrm{from}}{Fc}$$

Fc (N) × 10 ^-3 Fc (H) × 10 ^-3 $$\frac{Fc}{Fc}$$

$$\frac{F\mathrm{from}}{Fc}$$

Quartz, density - 2.7 g / cm ³ , hardness - 7. one 1.8 6.21 4.91 1.26 6.21 5.87 1.06 6.21 7.1 1.14 6.21 8.33 1.34 2 1,5 3,59 4.91 1.38 3,59 5.87 1,63 3,59 7.1 1.37 3 1,0 1.06 4.91 4.63 1.06 5.87 5.54 1.06 7.1 2.06 four 0.5 0.13 0.13 5.87 45.15 0.133 7.1 4.11 Pyrrhotite, density - 4.7 g / cm ³ , hardness - 4. 1.8 13.53 4.91 2.75 13.53 5.87 2,30 13.53 7.1 1.9 13.53 8.33 1,62 2 1,5 7.83 4.91 1,59 7.83 5.87 1.33 7.83 7.1 1,1 7.83 8.33 1.06 3 1,0 2,3 4.91 2.13 2,3 5.87 2,55 2,3 7.1 3.09 2,3 8.33 3.6 four 0.5 0.29 4.91 16.9 0.29 5.87 20.54 0.29 7.1 24.48 0.29 8.33 28.7 Tetraferroplatinum, density - 15 g / cm ³ , malleable. one 1.8 51.18 4.91 10.42 51.18 5.87 8.72 51.18 7.1 7.21 51.18 8.33 6.14 2 1,5 29.62 4.91 6.03 29.62 5.87 5.04 29.62 7.1 4.17 29.62 8.33 3,55 3 1,0 8.78 4.91 1.79 8.78 5.87 1.49 8.78 7.1 1.24 8.78 8.33 1.05 four 0.5 1,1 4.91 4.46 1,1 5.87 5.34 1,1 7.1 6.45 1,1 8.33 7.57 * Designations: Vst is the velocity of the fluidizing water jet at the inlet of the annular channel (inter-groove space) of the separator;
Fц - centrifugal force acting on the particle;
Fc — force of a stream of fluidizing water acting on a spherical particle;
$$\frac{Fc}{Fc}$$

- the ratio of the centrifugal force to the force of the jet;
$$\frac{F\mathrm{from}}{Fc}$$

is the coefficient of the ratio of the force of the water jet to the centrifugal force.

Table 3 Industrial comparative performance of four centrifugal separators KC-XD48, at a constant and stepwise flow rate (jet velocity) of fluidizing water through openings in the inter-groove space of the working cone. Optimal and constant conditions for each experiment: operating time - 240 hours, bowl rotation speed of 316 rpm (acceleration -60g), solid content in the initial feed 34-36%. Experience number Product name Exit, % Content,%, g / t Recovery% Test conditions Ni Cu Pt Pd Pt + pd Ni Cu Pt Pd Pt + pd one Concentrate 0.134 0.3 0.15 14,4 2.7 17.1 0.31 0.21 5.22 0.44 1.94 Standard experience. Tails 99,866 99.69 99.79 94.78 99.56 98.06 1. The amount of water (W) for 24 minutes is 24.98 m ³ / h. Source 100.0 0.127 0,093 0.37 0.81 1.18 one hundred one hundred one hundred one hundred one hundred 2. The performance of one separator (Qhard). - 68.3 t / h 3. The speed of the jets of water (V) through the holes - 1.52 m / s 2 Concentrate 0.14 0.37 0.41 47 fifteen 62 0.4 0.6 14.0 2.44 6.53 Stepped. Tails 99.86 99.6 99,4 86.0 97.56 93.47 1.W for 10 min - 24.98 m ³ / h, 10 min - 27.52 m ³ / h, 4 min - 29.52 m ³ / h Source 100.0 0.13 0.09 0.47 0.86 1.33 one hundred one hundred one hundred one hundred one hundred 2. Qhard. - 68.0 t / h 3.V - 10 min - 1.52 m / s V - 10 min - 1.67 m / s V - 4 min - 1.81 m / s 3 Concentrate 0.151 0.44 0.38 43 6.2 49.2 0.44 0.63 15.09 1.28 6.4 Stepped. Tails 99,849 99.56 99.37 84.91 98.72 93.60 1. W for 10 minutes - 24.98 m ³ / h, 6 min - 27.52 m ³ / h, 4 min - 29.52 m ³ / h Source 100.0 0.13 0.09 0.43 0.73 1.16 one hundred one hundred one hundred one hundred one hundred 2. Qhard. - 64.7.0 t / h 3.V - 10 min - 1.52 m / s V - 6 min - 1.67 m / s V - 4 min - 1.81 m / s

four Concentrate 0.14 0.46 0.40 24.6 6.4 31 0.48 0.7 10.42 1.22 4.09 Stepped. Tails 99.86 99.52 99.3 89.58 98.78 95.91 1. W for 10 minutes - 24.98 m ³ / h, 6 min. - 27.52 m ³ / h, 2 min. - 29.52 m ³ / h Source 100.0 0.134 0.08 0.33 0.73 1.06 one hundred one hundred one hundred one hundred one hundred 2. Qhard. - 66.7.0 t / h 3. V - 10 min - 1.52 m / s V - 6 min - 1.67 m / s V - 2 min - 1.81 m / s

Table 4 Granulometric composition of centrifugal concentration concentrates (experiments 1 and 2) experience number Product name Size classes, mm,% -1.8 + 0.14 -0.14 + 0.05 -0.05 one Centrifugal Concentrate Concentrate (Standard) 26.87 53.77 19.36 2 Centrifugal Concentrate 25.11 59.32 15,57

Claims (3)

1. A method of three-stage technological optimization of the parameters of centrifugal beneficiation for the extraction of precious metals in mineral form from ores and tailings during the processing of disseminated copper-nickel ores of the Norilsk deposits and comprising the separation of native platinum metal minerals into a gravity concentrate prior to the flotation concentration operation with a mass ratio of the sum of sulfides and magnetite and the sum of silicon and aluminum oxides in the initial ore or tails is less than 1: 2, with a particle size of 30-65% of the class less than 74 microns This is due to the fact that the isolation of native minerals of platinum metals is carried out at a value of the centrifugal Froude criterion of 11.75 and the ratio of this value to the pressure of the fluidizing water of 0.085 kPa, as well as by the method of three-stage technological optimization of the parameters of centrifugal enrichment for finely divided raw materials containing precious metals in mineral form, which includes a sequence of operations on centrifugal separators designed for continuous operation in an industrial environment: at the first stage, opti the total accumulation time of the concentrate, at the second stage, the optimal speed of the water jets or the optimal water flow rate through the openings into the inter-groove space of the separator bowl is selected; in the third stage, the speed of the jets or water flow through the openings into the inter-groove space of the separator bowl is gradually increased, starting from the optimal water flow rate determined on the second stage and in the optimal period of time the accumulation of the concentrate, defined in the first stage. 2. The method according to claim 1, characterized in that in the enrichment of finely divided raw materials containing precious metals in mineral form in centrifugal separators, the flow rate of fluidizing water or the speed of the jets in the inter-groove space of the separator stepwise increases at time intervals: for 10 minutes and subsequent 10 min and 4 min, for 10 min and the next 6 min and 4 min, for 10 min and the next 6 min and 2 min, with an optimal accumulation time. 3. The method according to claim 1, characterized in that in the enrichment of finely divided raw materials containing noble metals in mineral form in centrifugal separators, the flow rate of fluidizing water is increased stepwise or the speed of the jets in the inter-groove space of the separator bowl is increased, starting from the minimum optimum and ending with the water flow rate or a jet velocity at which the jet force exceeds the centrifugal force for the particles of the enclosing rocks, commensurate with the diameter of the holes for supplying fluidizing water.

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