RU2245740C1 - Method of heavy metals deposits ore dressing - Google Patents

Abstract

FIELD: mining; production of heavy metals.

SUBSTANCE: the invention concerns to the field of heavy metals deposits and placers ore and rock dressing with the purpose of extraction of heavy metals, including small-sized and the thin heavy metals, for example, gold or platinum being in chemically free state. The technical result is an increase of efficiency of deposits and placer heavy metals rock ore dressing. The method provides for removal: with the help of the calibration equipment of buckstones of the size exceeding the maximum size of particles of extracted metal, determined by laboratory analyses and experimental treatment using the equipment of centrifugal separation; material of initial mined rock with consequent sizing of the received product using a hydraulic sizing screen and a hydrogravitational concentration of the under screen product of the hydraulic sifter in the slant sluice pipe with continuous withdrawal from the slant sluice pipe through the slot of a base splitter of dressed product with its consequent control gauging using an additional hydraulic sifter and the hydrogravitational concentration of the under screen product of the additional hydraulic sifter in the additional slant sluice pipe with continuous withdrawal from the slant sluice pipe through the slot of a base splitter of dressed product of the secondary concentrate with consequent continuous treatment of the latter using centrifugal concentrators for consequent extraction of large particles of extracted metal by the finishing equipment. At that all the drains of the slant sluice pipes are jointly stage by stage treated in the gravitational settler and a thin-layer thickener for settling in the gravitational settler of the of the draw rock particles. At that the thickened in the thin-layer thickener sediment is treated by additional centrifugal classifiers for extraction of a small particles of the extracted metal, and the waste products of centrifugal separation are treated by the methods of cyanidation or a heap leaching.

EFFECT: the invention ensures increased efficiency of deposits and placer heavy metals rock ore dressing.

9 tbl, 11 dwg

Description

The invention relates to the field of enrichment of the rock mass of placer, ore and man-made deposits in order to extract heavy metals, including small and thin ones, for example gold or platinum, which are in a free, chemically unbound state.

The invention can be used for the enrichment and separation of heavy metals from the rock, if its bulk density of particles of the extracted metal exceeds the density of the particles of rock containing this metal, two or more times.

The invention can be used to extract metals or to obtain their rich concentrates from the rock mass of placer, ore and man-made deposits of mass or waste of its primary technological processing, both independently and as part of the technological lines of mining and processing enterprises.

The invention can be used to create laboratory, pilot and industrial plants.

The well-known “Method of processing mineral-containing rock mass” (Patent RU No. 2144430), according to which the exclusion from the initial rock mass of waste rock particles is carried out by dry or wet screening, followed by stepwise screening of the undergrate product on high-performance hydraulic screens with subsequent thickening and finishing in a concentrator. In this case, the entire liquid phase of the hydraulic mixture is passed through a thickener in order to isolate fine and fine particles of the allocated mineral.

The disadvantage of this method is that when processing large volumes of rock mass, it is necessary to pass a large volume of the liquid phase of the hydraulic mixture through the sump and thickener, which leads to the creation of large (in size) sump and thickener, the complexity of installation and operation of such equipment directly at the field, and also to the use of energy-intensive centrifugal concentrators of high productivity, which due to their design features are not able to extract concentrate from the processed s, especially from concentrates processing crushed ore mass, particles with a particle size of less than 50 microns.

Closest to the proposed one is the “Method for the enrichment of heavy small-fraction concentrates” (Patent RU No. 2174448), according to which the exclusion from the initial rock mass of gangue particles of placer deposits, crushed rock mass of ore deposits and their primary processing products (tailings, dumps, etc. .) produced by wet screening on the sieve of the hydrovasherd with subsequent step-by-step screening of the under-sieve product on several high-performance hydraulic screens, the sub-sieve product from the previous hydraulic screen is fed into the next one, and the resulting sublattice from the last hydraulic screen is subjected to hydrogravity enrichment in a slanted honeycomb pulp conduit with the concentrate obtained through enrichment through the bottom cutters for subsequent thickening in a thin-layer thickener and finishing of the condensed material at a finishing concentrator, centrifugal magnetic and magnetically-liquid separators) in order to obtain highly enriched concentrate of mineral (metal) for chemical cleaning methods refining.

The disadvantage of this method is that the sequential calibration of large volumes of rock mass, especially from concentrates of crushed ore mass, at several hydraulic screens requires the sequential pumping of the entire, almost constant, volume of the liquid phase of the hydraulic mixture with stepwise decreasing volumes of the sieve product from the previous hydraulic screen in subsequent, which requires the use of several pumps of high energy intensity.

In addition, the technological scheme used does not provide for technological operations for the extraction of mass of “tails” from particles discharged into the dump through the end sections of the slurry conduit containing particles of the emitted heavy mineral (metal) of less than 44 microns in size, especially lamellar particles that, according to scientific studies, cannot precipitate in a moving stream of pulp and, therefore, cannot be distinguished by gravitational methods.

In addition, the technological scheme used does not solve the problem of determining the width of the slit of the bottom cutter by mathematical methods, which leads to a loss of time when selecting the dimensions of the slit during commissioning.

The aim of the invention is to remedy the above disadvantages.

Achievable technical result - increasing the efficiency of enrichment of the rock mass of heavy metal deposits through the use of a highly efficient technological method of rock dressing, based on the phased reduction of the volume of processed rock mass due to the gradual elimination of waste rock particles from the processed volumes and the allocation of consolidated, chemically unrelated particles of recoverable metal methods of hydrogravity enrichment and thin-layer thickening according to a special technique.

The technical result is achieved in that in the enrichment method, which consists in eliminating waste rock on calibration equipment that exceeds the maximum particle size of the recoverable metal, determined by laboratory analyzes and experimental processing on centrifugal separation equipment, of the material of the initial rock mass with subsequent calibration of the resulting product on a hydraulic rumble and hydrogravity enrichment of the under-sieve product of a hydraulic screen in an inclined slurry conduit with continuous removal from the inclined slurry line through the slot of the bottom cut-off of the enriched product with its subsequent calibration on an additional hydraulic screen and hydrogravity enrichment of the undergrate product of an additional hydraulic screen in an additional inclined slurry line with the constant removal of the last secondary concentrate from the additional inclined slurry line through the slot of the bottom cut-off with the secondary concentrate centrifugal concentrators for subsequent the extraction of large particles of recoverable metal on the finishing equipment, while the effluents of the inclined slurry conduits are jointly processed in stages in a gravity settler and a thin-layer thickener for sedimentation of gangue particles in a gravity settler, and the precipitate thickened in a thin-layer thickener is treated with additional centrifugal classifiers for extracting metal particles and centrifugal separation waste is treated with cyanidation or heap leaching methods.

This result is achieved using more accurate methods for determining the parameters of enrichment processes based on the use of experimentally confirmed mathematical and empirical formulas in the calculations of technological equipment in order to increase the efficiency of extraction of free, chemically unbound particles of heavy metals, especially precious metals, by the methods of hydrogravity enrichment and thin-layer thickening.

The specified result is achieved by the fact that in order to clarify the technical parameters of the enrichment process of the rock mass of heavy metal deposits by the proposed method, the following laboratory analyzes are carried out:

- a complete mineralogical analysis of the initial rock mass with a description of the particle shapes of all the minerals that make up the processed rock mass, and primarily the particles of recoverable metals, with the determination of clay and other binders in the composition of the rock mass, with the determination of dry bulk weight;

- granulometric analysis of the initial rock mass according to the largest possible number of granulometric classes with determination of the recoverable metal content in each class, and it is necessary to carefully carry out the above studies in classes from 0 to 44 microns and from 44 to 74 microns.

Based on the data of the granulometric composition of the processed initial rock mass according to standard granulometric classes, especially when processing the rock mass of placer deposits, the number of stages of reduction of the initial rock mass using the used technological equipment used in the proposed method is determined, and the boundary parameters obtained by reducing the particle size distribution are numerically determined classes at each stage both in rock mass and in recoverable metal, namely:

- on equipment for preliminary calibration of the initial rock mass;

- in a hydraulic screen of the first stage of enrichment;

- in the pulp line of the first stage of enrichment;

- hydraulic screen of the second stage of enrichment;

- in the pulp line of the second stage of enrichment;

- in the inclined channels of the gravity settler of a thin-layer thickener;

- in the inclined channels of the thin layer thickening unit of the thin layer thickener.

Based on the analysis of the content of particles of recoverable metal in each particle size class, experimental processing of the processed rock is performed on finishing equipment (centrifugal concentrators of large and low capacity, on magnetic and magneto-liquid separators) to determine the boundaries of the stepwise reduction of the processed material on finishing equipment, as well as to determine the minimum particle size of recoverable metal in order to determine economic feasibility osleduyuschego application of technological operations and additional extraction of small metal fine particles recovered, subject health measures and environmental safety, chemical methods for extracting the object, for example, gold or cyanidation heap leaching, especially against rock mass crushed ore processing.

On any type of calibration equipment, gangue particles are excluded from the original rock mass, exceeding in size the maximum particle size of the recoverable metal, determined by laboratory analyzes of the initial rock mass.

The calibrated rock mass is subjected to secondary calibration on a hydraulic screen of the first stage of enrichment to a particle size corresponding to that determined earlier by the calculation of the particle size classification at the stage of calibration on the screen of the first stage of enrichment for rock mass and recoverable metal, with the processed rock being removed from the screen in the form of under-sieve and over-sieve products, both of these products being removed from the screen in the form of hydraulic mixtures with different ratios of solid and liquid phases (T: G).

The sublattice product removed from the hydraulic screen of the first enrichment stage is enriched by the method of hydrogravity classification in an inclined slurry conduit of the first enrichment to the particle size corresponding to the granulometric class determined earlier by calculation in the inclined slurry conduit of the first enrichment both in rock mass and in rock mass recoverable mineral (metal), with the output from the slurry line through the bottom cut-off of the primary concentrate and the “tails” discharged from the slurry line Erez end section, both of these products are derived from the screen in the form of hydraulic mixtures with different ratios of solid and liquid phases (S: L).

The primary concentrate discharged through the bottom cut-off of the pulp line of the first enrichment stage is subjected to secondary calibration on a hydraulic screen of the second enrichment stage to a particle size corresponding to the particle size distribution determined earlier by the calculation at the stage of calibration on the screen of the second enrichment stage according to rock mass and recoverable metal, with output from rumble of processed rock mass in the form of an under-sieve product and an over-sieve product, both of which are removed from the screen in the form of hydra ble of mixtures with different ratios of solid and liquid phases (S: L).

The sublattice product removed from the hydraulic screen of the second enrichment stage is enriched by the method of hydrogravity classification in an inclined slurry conduit of the second enrichment to the particle size corresponding to the granulometric class determined earlier in the calculation at the calibration stage in the inclined slurry conduit of the first enrichment both in rock mass and in rock mass recoverable metal, with the output from the slurry line through the bottom cut-off of the primary concentrate and the “tails” discharged from the slurry line through the end -hand section, both of these products are derived from the screen in the form of hydraulic mixtures with different ratios of solid and liquid phases (S: L).

The analysis of the solid phase of hydraulic mixtures, derived from hydraulic screens of the first and second stages of enrichment in the form of oversize products, is carried out to determine the presence of particles of recoverable metal that have got into the oversize product as a result of demolition by the flow of the hydraulic mixture.

In the presence of particles of recoverable metal in the test products, the over-sieve product is sent for retrieval, for example, to a fine-filling sluice, jigging machine, etc.

In the absence of particles of recoverable mineral or metal in their composition, it is sent to the dump.

The secondary concentrate discharged through the bottom cutoff of the pulp line of the second enrichment stage in the form of a hydraulic mixture is subjected to processing at a centrifugal concentrator of calculated capacity in order to extract free, chemically unbound metal particles, with laboratory analysis of the material of the “tailings” of the centrifugal separation for the presence of particles of emitted metal .

If there are free, chemically unbound particles of metal released in the “tailings” of the concentrator, they are subjected to repeated single or double processing (purification), for example, at a low-capacity centrifugal concentrator.

If there are proprietary, chemically bonded particles of emitted metal in the “tailings” of the concentrator, the “tailings” are subjected, subject to sanitary and environmental safety measures, to further technological processing for the recovery of, for example, gold by cyanidation or heap leaching.

A centrifugal concentration concentrate, for example, a gold-containing concentrate obtained by processing a secondary concentrate on a centrifugal concentrator, is dried in any type of drying ovens (furnaces), the dry material is processed on a magnetic separator to separate it into magnetic and non-magnetic fractions, while the non-magnetic fraction is processed on magneto-liquid separator in order to exclude from the processed material particles of waste rock with a density close to gold, and the magnetic fraction is sent to thwarted.

This result is achieved by the fact that the hydraulic mixture flowing from the end sections of the slurry lines of the first and second enrichment stages (“tails”) and containing particles, including recoverable metal, the size of which corresponds to the particle size class determined earlier by the calculation at the calibration stage in inclined slurry pipelines of the first and second stages of enrichment, both for rock mass and for recoverable metal, is sent for further technological processing into a thin-layer thickener.

The thin-layer thickener is structurally divided into two technological units connected to each other - the gravity sedimentation unit and the thin-layer thickening unit, and in each unit the solid phase of the hydraulic mixture is divided into two fractions by size.

The hydraulic mixture arriving from the end sections of the slurry pipelines, containing particles of gangue and recoverable metal, enters the gravity sedimentation unit, where it is calibrated in inclined channels to a size corresponding to the particle size classification previously determined by calculation at the calibration stage in the gravitational settler of a thin-layer thickener.

The cross-sectional area, the length, the angle of inclination to the horizontal plane and the number of inclined channels installed in the upper part of the gravity settler are determined, with the aim of sedimenting the condensed deposited mass from the particles of only waste rock in the collection tank of the gravitational settler and transferring the particles of the extracted metal to the thin-layer thickening unit.

The fraction obtained in the form of a condensed sediment in the gravity settler of a thin-layer thickener is sent to the dump.

The hydraulic mixture emerging from the gravitational thickener, containing particles of gangue and recoverable metal, enters the thin-layer thickening unit, where it is calibrated in inclined channels to a size corresponding to the particle size classification previously determined by the calculation at the calibration stage in the thin-layer thickening thickening thickening thickening block.

The cross-sectional area, the length, the angle of inclination to the horizontal plane and the number of inclined channels installed in the upper part of the thin-layer thickening unit of the thin-layer thickener are determined so that a mass containing large particles of emitted metal and gangue particles is deposited in the collection tank of the thin-layer thickener and the hydraulic mixture leaving the thin-layer thickening unit contained in its composition particles of gangue and small particles of recoverable metal, the sizes of which were There are the granulometric class defined earlier by the calculation at the calibration stage in the thin-layer thickening unit.

The fine fraction, concentrated in the form of a condensed precipitate in the collecting tank of a thin-layer thickener, is subjected to single or double processing at a centrifugal concentrator of calculated capacity in order to extract free, chemically unbound metal particles.

In the case of the presence of free, chemically unbound particles of emitted metal in the “tailings” of the concentrator, the “tailings” are subjected to repeated single or double processing (cleaning), for example, at a low-capacity centrifugal concentrator.

If there are proprietary, chemically bonded particles of emitted metal in the “tailings” of the concentrator, the “tailings” are processed, subject to sanitary and environmental safety measures, for the extraction of, for example, gold, cyanidation or heap leaching methods.

If there are no particles of recoverable metal in the “tailings” of the concentrator, the “tailings” are sent to the dump without processing.

If the clarified hydraulic mixture flowing from the thin-layer thickening unit meets the technical requirements for the water transfer pumps used in the proposed technological scheme, in terms of particle size distribution and particle concentration in the pumped liquid (water), these drains can be used as a source of reverse water supply in the proposed technological scheme, and a thin-layer thickener can be used as a device for clarification (purification) of water from heavy mechanical impurities.

The method of enriching the rock mass of heavy metal deposits is as follows and in the following sequence.

1. On calibration equipment of any type (on a cradle, on a scrubber butyram, etc.), waste rock particles exceeding the maximum particle size of the extracted metal, determined by laboratory analyzes of the initial rock mass, are excluded from the original rock mass.

2. Calibrated rock mass according to any transport scheme (bulldozer, front-end loader, road transport, etc.) is fed to the metering hopper.

3. The rock mass mixed with water in the form of a hydraulic mixture (pulp) enters the receiving chamber of the jet pump (ejector) mounted in the lower part of the mixer hopper.

4. The pressure flow of water supplied by the pump to the working chamber of the jet pump, due to the ejection effect, captures the pulp and feeds it into the pressure pulp conduit, the design of which provides a gravity self-trapping device for trapping and accumulating large native particles of recoverable metal.

5. The number of pipes of the slurry line of the first enrichment stage and their inner diameter are determined based on the capacity of the installation. In this case, the speed of movement of the pulp stream through the slurry duct should be greater than the minimum velocity of the movement of particles of the solid phase of the hydraulic mixture along the bottom of the slurry duct Vmin calculated by hydraulic formulas, i.e. the rate at which particles are not yet deposited on the bottom of the slurry line.

The minimum speed Vmin is determined by the formula:

where: W ₀ - hydraulic fineness (deposition rate of particles of diameter d in still water);

d _sr is the average diameter of suspended particles, m;

p is the percentage (by weight) of particles with a particle size d≥ 0.25 mm;

R is the hydraulic radius of the slurry line, m;

n is the roughness coefficient of the inner walls of the slurry line.

6. The deposition of particulate matter in an inclined slurry conduit, in contrast to the deposition rate of a single particle in an unlimited space in a statically stationary system, occurs in a moving flow of a hydraulic mixture, and the deposition rate of particles depends on the transporting ability of the flow, the nature of the mode of motion, concentration and shape of the solid deposited particles, the viscosity of the liquid and its density, as well as reducing (shading) the actual cross section of the flow due to the space occupied by solid particles.

The particle deposition rate under such conditions is called the constrained deposition rate W _st and is determined by the formula:

W _st = W ₀ × K ₁ × K ₂ × K ₃ × K ₄

The deposition rate W _{0 of} particles with a particle size of 0 to 0.12 mm (120 μm), provided that a single particle is deposited in unlimited space in a statically stationary system, is determined by the Stokes formula:

where: W ₀ is the particle deposition rate;

q _t is the density of particles of the solid phase of the hydraulic mixture;

q _v is the density of the liquid (water);

d _t is the particle size of the solid phase of the hydraulic mixture;

μ is the dynamic coefficient of viscosity of water;

g = 9.81.

K ₁ - coefficient taking into account the decrease in the actual cross section of the fluid flow due to the space occupied by solid particles, is determined by the formula:

K ₁ = 1-γ ^2/3

where: γ - part of the volume of the flow of the hydraulic mixture, which is occupied by the solid phase (solid particles);

K ₂ - coefficient taking into account the change in the average difference in specific gravity between solid particles and liquid by a value of 1-γ

K ₂ = 1-γ;

K ₃ - coefficient taking into account the correction for viscosity;

K ₃ = 1-2.5γ;

K ₄ - coefficient taking into account the shape of the particles.

The value of the form factor K ₄ is selected depending on the shape of the particles from the hydraulic reference books.

7. Together with the particles of the metal being released, particles of gangue are also involved in the deposition process, the size difference between these particles is determined by the coefficient of equidistance K _p .

The coefficient of equidivision for particles of different particle size distribution d _t is determined by the following formulas:

for small particles: d _t <0.1 mm

where: q _p is the particle density of the inclusive rock;

q _sm is the density of the hydraulic mixture;

q _m is the particle density of the recoverable metal;

d _t is the particle size;

for medium particles: 0.1 <d _t <0.1 mm

for large particles: dt> 1 mm

The density of the hydraulic mixture qsm is determined by the formula:

where: q _p is the density of particles of rock mass;

q _v is the density of the liquid (water);

V _p - the volume of particles of rock mass;

V _v - the volume of the working medium (water).

8. Studies of laboratory samples of primary concentrate discharged from the slurry line through the bottom cutter with varying values of the density of the hydraulic mixture (pulp), width and underestimation of the slit of the bottom cutter of the slurry line of the first enrichment stage, determine the ratios of the solid and liquid phases of the hydraulic mixture, i.e. the ratio of T: W.

Based on the above T: Ж values obtained, a part of the cross-sectional area of the slurry duct F _{p is} mathematically determined, which is occupied by a moving bottom stream with a cross-sectional area f _k geometrically representing a segment of the corner sector of a circle with a diameter Dp with a central angle of 2α, a chord b _s and an arrow h ₁ .

The cross-sectional area of the bottom stream F _f = b _f × h ₂ output from the slurry line of the first enrichment stage through the bottom cutter is selected equal to the cross-sectional area of the moving bottom stream f _k .

9. The flow of the hydraulic mixture, entering the inclined slurry line, passes a certain distance L _st , called the stabilization section. Experimental studies have proved that the practical formation of the velocity field in a turbulent flow ends at the length of the slurry conduit equal to

L _st = 40 × D _p

where: D _p is the diameter of the slurry line.

10. The process of deposition of solid particles in an inclined slurry line 15 is carried out in a moving stream of pulp (see Figure 4), i.e. simultaneously with the deposition under the influence of gravitational force, solid particles 46 participate in the process of their transfer by a pulp stream moving along the slurry conduit, and the picture of the velocity distribution in the flow section is a rotation paraboloid.

In the turbulent mode of movement of the hydraulic mixture flow along the slurry line, an increase in the flow velocity (in the direction from the slurry wall walls to the center) from the minimum value V _min on the pipe wall to the maximum value V _max on the longitudinal axis of the slurry pipe is described by the equation of the cubic parabola.

The thickness of the wall laminar layer 5, which includes solid particles in the process of turbulent motion and in which they are locked, as in a moving hydraulic trap, is comparable with the size of solid particles moving in a laminar layer with a thickness of 5, is determined by the formula:

where: δ is the thickness of the parietal laminar layer;

D _p is the inner diameter of the slurry line;

Re - Reynolds criterion;

λ is the Darcy coefficient

where: V _plp is the average speed of the hydraulic mixture flow along an inclined slurry line;

ξ is the kinematic viscosity coefficient.

Solid particles moving in a hydraulic trap are affected by a gravitational force, which forces the particles to move downward along the arc of the wall of the slurry duct, while the movement of these particles along the walls of the slurry duct in the axial direction is due to layers of the hydraulic mixture moving at high speeds (located closer to the axis of the slurry duct).

As a result of such movements, due to the interaction of two forces, the solid particles moving along the arc of the wall of the slurry duct pass in the vertical direction the calculated distance of the vertical precipitation of the solid particle h _t , while the particles are displaced in the horizontal direction relative to the longitudinal axis of the slurry duct and enter the bottom flow zone, whose width is limited by the calculated value of the chord b _s .

11. Structurally, the width of the slit of the bottom cutter b _{f of the} slurry conduit of the first enrichment stage is determined from the ratio of the equality of angles:

α _f = θ

where: α _f -1/2 of the Central angle of the angular sector of the circle with a diameter equal to the inner diameter of the slurry conduit Dp, based on the chord, equal to the width of the slit b _f bottom cutter;

θ is the angle of repose of the rock material diluted with water.

12. The underestimation of the slit of the bottom cutter h ₂ is determined from the calculation of the equality F _K = F _f and is determined by the formula:

where: F _k is the area of the moving bottom flow;

D _p is the inner diameter of the slurry line;

2α _f - the Central angle of the angular sector of the circle with a diameter equal to the inner diameter of the pulp duct D _p , based on the chord, equal to the width of the slit b _{f of the} bottom cutter;

π = 3.14

The area of the moving bottom flow f _k is determined by experimental measurements of the ratio T: W in the pulp discharged through the slit of the bottom cutoff of the slurry line.

Geometrically, the area of the moving bottom flow f _k is determined by the formula:

The minimum value of the understatement of the slit of the bottom cutter h _2min is selected in accordance with the equality:

h _2min = d _tmax +1 mm

where: d _tmax is the maximum particle size of the recoverable metal.

The maximum value of the understatement of the slit of the bottom cutter h _2max , is selected in accordance with the equality:

h _2max = 2 × d _tm

13. The length of the slit of the bottom cutter L _f is determined by the formula:

where: h ₂ - understatement of the slit of the bottom cutter;

6 ° - half the maximum opening angle of the diffuser, at which there is no disruption of the fluid flow when moving along the diffuser.

14. The calculation of the time of deposition of a solid particle T _{t is} made according to the value of the minimum vertical precipitation h _t solid particles of size d _t according to the formula:

h _t = D _p -h ₁

where: D _p is the inner diameter of the slurry line;

h ₁ - the value of the arrow of the arc segment with an area of f _k .

where: h _t is the distance of the vertical precipitation of a solid particle;

W _st - the rate of constrained deposition of a solid particle.

15. The average speed of the hydraulic mixture flow V _plp along an inclined slurry _conduit can be practically (when setting up the equipment) based on the equations of the flow velocity, the center line of the jet and the distance of departure of the jet:

where: V _plp is the average speed of the hydraulic mixture flow along an inclined slurry line;

X is the range of departure of the jet;

Y is the excess of the center of the end section of the end section of the inclined slurry conduit above the point of contact of the center of the departing jet with a horizontal plane;

N is the pressure above the center of gravity of the outflow opening;

φ is the velocity coefficient, φ = 0.97-0.98;

g = 9.81.

16. The length of the pipe L _p1 pulp duct of the first stage of enrichment is determined by the formula:

The length of the pulp duct of the first enrichment stage is calculated from the following reasoning: a solid particle of size d _t in the cavity of the pulp duct should descend at a vertical sedimentation speed W _st into the bottom of the pulp duct by the vertical precipitation distance h _t during the settling time T _t , at the same time the particle should move along the pulp duct together with pulp flow with a flow velocity V _plp at the estimated distance L _p to the slit of the bottom cutter. This dependence is expressed by the following mathematical equality:

where: h _t is the distance of the vertical precipitation of a solid particle;

W _st - the rate of constrained deposition of a solid particle;

V _plp is the average velocity of the hydraulic mixture flow along an inclined slurry line;

L _p - the horizontal distance by which the deposited particle moves along the axis of the slurry conduit during the vertical movement by the distance of the vertical sedimentation of the particle h _t ;

L _p1 is the length of the slurry conduit of the first enrichment stage (the distance from the vertical axis of the hydraulic screen of the first enrichment stage to the slit of the bottom cutoff of the slurry conduit);

T _t is the deposition time of a solid particle of size d _t by the calculated value of the vertical precipitation h _t .

17. The process of deposition of solid particles in an inclined slurry conduit is carried out in a moving pulp stream, i.e. simultaneously with the deposition, solid particles participate in the process of their transfer by a pulp stream moving along the pulp line, and the picture of the velocity distribution in the flow section is a rotation paraboloid.

In the turbulent mode of movement of the pulp stream along the slurry duct, an increase in the flow velocity (in the direction from the walls of the slurry duct to the center) from the minimum value V _min on the pipe wall to the maximum value V _max on the longitudinal axis of the slurry duct is described by the equation of the cubic parabola.

18. The flow rate of the pulp q _p1 through one pipe of the pulp line of the first enrichment stage is determined by the formula:

q _p1 = 0.25 × V _plp × π × D $\genfrac{}{}{0ex}{}{\text{2}}{\text{p}}$

where: q _p1 is the flow rate of the hydraulic mixture through one pipe of the slurry line of the first enrichment stage;

V _plp is the average velocity of the hydraulic mixture flow along an inclined slurry line;

D _p is the inner diameter of the slurry line;

π = 3.14159

19. The number of pipes of the slurry line of the first enrichment stage is determined based on the accepted design capacity of the installation according to the composition and volume of the hydraulic mixture and is determined by the formula:

where: q _gr is the productivity of the hydraulic screen of the first stage of enrichment in terms of the volume of the hydraulic mixture;

q _p1 - flow rate of the hydraulic mixture through one pipe of the slurry line of the first enrichment stage.

20. The calculation of the parameters of the slurry conduit of the second enrichment stage in terms of determining the productivity of the screen of the second enrichment slurry, the number and design length of the pipes of the inclined slurry conduit, the speed of movement of the pulp stream through the slurry conduit, as well as the width and understatement of the slit of the bottom cutter, are similar to the above calculation of the parameters of the screen and slurry conduit of the first stage enrichment.

21. The calculation of the cross-sectional area of the inclined channel, its length and the number of inclined channels installed in the upper part of the gravity settler of a thin-layer thickener, is reduced to determining the parameters of the above elements, in which only gangue particles will be deposited in the collection tank of the gravity settler, and particles of recoverable metal as part of the waste (“tailings”) must be carried out by the flow of the hydraulic mixture into the block of thin-layer thickening of a thin-layer thickener.

22. The flow around a particle during its deposition in an inclined channel, when the flow of the hydraulic mixture moves up the channel, is characterized by a dimensionless parameter Re _w , similar to the Reynolds number.

The numerical value of the parameter Re _w is determined by the formula:

where: ξ is the kinematic viscosity coefficient of the liquid (water);

W ₀ is the particle deposition rate in still water;

d _t is the particle size (diameter).

23. The value of W ₀ for solid particles with a particle size from 0 to 0.12 mm (120 μm) is determined by the Stokes formula, subject to the deposition of a single particle in unlimited space in a statically stationary system:

where: W ₀ is the particle deposition rate under the condition of deposition of a single particle in unlimited space in a statically stationary system;

q _t is the density of particles of the solid phase of the hydraulic mixture;

q _v is the density of the liquid (water);

d _t is the size (diameter) of the particles of the solid phase of the hydraulic mixture;

μ is the dynamic coefficient of viscosity of water;

g = 9.81.

24. The deposition of solid particles in inclined channels located in the upper part of the gravity sedimentation tank, in contrast to the deposition of a single particle in infinite space in a statically stationary system, occurs in an upward flowing hydraulic mixture, and the particle deposition rate depends on the nature of the particle deposition mode and its forms.

The particle deposition rate under such conditions is called the constrained deposition rate W _st and is determined by the formula:

W _st1 = W ₀ × K _R × K ₄

where: W _st1 - speed constrained deposition of particles;

K ₄ - coefficient taking into account the shape of the particle;

S - concentration of the hydraulic mixture - the ratio of the volume of solid particles deposited in the inclined channel to the volume of the hydraulic mixture moving along the inclined channel;

W ₀ is the particle deposition rate provided that a single particle is deposited in unlimited space in a statically stationary system;

k _r - coefficient taking into account the mode of deposition of the particle during tight deposition.

The value of the form factor K ₄ is selected depending on the shape of the particles from the hydraulic reference books.

25. The value of k _r is determined by the formula

When Re _w <2, the value of k _r is determined by the formula:

where: S is the concentration of the mixture;

k _r - coefficient taking into account the mode of deposition of the particle during tight deposition.

Re _w is a dimensionless parameter.

The numerical value of the parameter Re _w is determined by the formula:

where: ξ is the kinematic viscosity coefficient of the liquid (water);

W ₀ is the particle deposition rate in still water;

d _t is the particle size (diameter).

26. The length of the inclined channels l _k1 located in the upper part of the gravity settler, their height h _K1 and the cross-sectional area Fk _{1 are} determined from mathematical equality, which is a necessary condition for the operation of a thin-layer settler:

The channel length Lк ₁ is determined by the formula:

where: Lv ₁ - the distance of the vertical precipitation of a solid particle;

W _st1 - speed constrained deposition of a solid particle;

v _g1 is the flow velocity of the hydraulic mixture up the inclined channel;

l _k1 is the length of the inclined channel;

T _t1 is the deposition time of a solid particle of size d _t1 by the calculated value of the vertical precipitation L _v1 .

h _k1 - the height of the inclined channel in the section perpendicular to the longitudinal axis of the channel;

φ is the angle of inclination of the channel of the horizontal plane.

Analyzing the formula for determining the length of the inclined channel, we can say that during the precipitation of the solid particle of the extracted mineral (metal) to the bottom of the inclined channel, it, the precipitated particle, should not be carried out by the flow of a hydraulic mixture moving up the inclined channel with a speed of V _G1 .

The distance of the vertical precipitation of a solid particle L _V1 is determined by the formula:

where: h _k1 is the height of the inclined channel in the section perpendicular to the longitudinal axis of the channel;

φ is the angle of inclination of the channel of the horizontal plane.

27. The flow of the hydraulic mixture, passing through the inclined channels, goes up and has a free surface. To improve the conditions of deposition in inclined channels and to create a stable hydrodynamic structure of the upward flow over the entire cross section of the gravity settler tank, a vertical stabilization grating made of rectangular rectangular channels is installed over the inclined channels, which divides the upward flow of the hydraulic mixture into many separate vertical flows with alignment vertical velocities over the entire cross section of the capacity of the gravity sedimentation tank and eliminates the possibility The probability of the occurrence of horizontal flows in the upper part of the capacity of the gravity sedimentation tank.

The total area of the cross-sectional capacity of the gravitational sump tank shaded by the grating partitions should not exceed 5-7% of the total cross-sectional area of the gravitational sump tank, and the grate height is selected within (5-10) × h _k1 , the area limited by one grate cell is selected within (20-50) × F _k1 , where F _k1 is the cross-sectional area of the inclined channel, and h _k1 is the height of the inclined channel.

28. The second prerequisite for the calculation of thin-layer sedimentation tanks is the presence of a laminar regime of movement of the hydraulic mixture along the inclined channel with the prerequisite for maintaining flow stability, ie the values of the Reynolds and Froude criteria must correspond to the parameters of the following equalities:

where: ξ is the kinematic viscosity coefficient of the liquid (water);

v _g1 is the flow velocity of the hydraulic mixture up the inclined channel;

R is the hydraulic radius of the inclined channel;

Re is the Reynolds criterion.

v _g1 is the flow velocity of the hydraulic mixture up the inclined channel;

R ₁ is the hydraulic radius of the inclined channel;

Fr — Froude criterion;

g = 9.81

The value of the hydraulic radius R ₁ for a rectangular channel is determined by the formula:

where: R ₁ is the hydraulic radius;

b _k1 is the width of the inclined channel;

h _k1 - the height of the inclined channel.

29. The specific load on one inclined channel Z ₁ is determined by the formula:

Z ₁ = b _K1 × h _k1 × v _g1

where: Z ₁ - specific load on one inclined channel;

d _k1 is the width of the inclined channel;

h _k1 - the height of the inclined channel;

v _g1 is the flow velocity of the hydraulic mixture up the inclined channel;

30. The number of inclined channels N ₁ is determined by the formula:

where: N ₁ - the number of inclined channels;

Q ₁ - the capacity of the sump in the hydraulic mixture;

Z ₁ - the specific load of the hydraulic mixture per channel.

31. The calculation of the cross-sectional area of the inclined channel, its length and the number of inclined channels installed in the upper part of the thin-layer thickening unit of the thin-layer thickener, is reduced to determining the parameters of the above elements, in which small particles of emitted metal that do not precipitate in the collection tank of the thin-layer thickening unit were isolated as a result of hydrogravitational enrichment of rock mass in inclined slurry conduits of the first and second stages of enrichment, in the gravity settler and Does the unit thin-layer thickening.

32. The calculation of the parameters of the block of thin-layer thickening is carried out according to the formulas specified in paragraphs 21-30 with the use of the notation of the values used in the calculation of the parameters of the block of the gravity sedimentation tank.

33. If the clarified hydraulic mixture (water) flowing from the block of thin-layer thickening complies with the technical requirements for water transfer pumps used in the proposed technological scheme in terms of particle size and particle concentration in the water pumped by the pump, these drains can be used as a source of reverse water supply in the proposed technological scheme, and a thin-layer thickener can be used as a separate device for clarification (purification) of water from heavy mechanical impurities in various industrial fields.

The proposed method for enriching the rock mass of heavy metal deposits consists in combining and combining the enrichment methods given and the equipment used.

The essential features of the above features of the proposed method exhibits new properties in such a way that the proposed solution meets the criterion of "inventive step".

The technological scheme of the proposed method is shown in Fig.1.

A cross section of an inclined slurry line is shown in FIG. 2.

A cross section of the bottom cutter of an inclined slurry duct is shown in FIG. 3.

A plan view of the slit of the bottom cutter and the plot of the flow rates of the hydraulic mixture is shown in Figure 4.

A fragment of the cross section of the slurry line and a diagram of the forces acting on the deposited solid particle with reference to the vertical and horizontal axes of the cross section of the slurry line are shown in FIG.

A fragment of a longitudinal section of the slurry duct with the particle path along the slurry wall in the wall laminar layer with reference to the axes of the slurry duct is shown in FIG. 6.

A fragment of the upper part of the conical gravity settler with the image of inclined channels and the grating stabilization of the upward flow of the hydraulic mixture is shown in Fig.7.

The technological scheme of processing the secondary concentrate on the finishing equipment is shown in Fig. 8.

A fragment of the upper part of the conical gravity settler with the image of inclined channels, the grating stabilization of the upward flow of the hydraulic mixture and the collar is shown in Fig.9.

A fragment of the upper part of the block of thin-layer thickening of a thin-layer thickener with the image of inclined channels is shown in FIG. 10.

The technological scheme of processing the thickened mass of sediment block thin-layer thickening by chemical methods is shown in Fig.11.

The method is as follows.

1. The full mineralogical and lithological compositions of the processed rock mass, screening of the initial rock mass by granulometric classes with determination of the total content of recoverable metal, as well as the content of recoverable metal in a free, chemically unbound form, in each granulometric class, especially in grades from 0 to 44 microns and from 44 to 74 microns.

2. The number of stages of reduction of the initial rock mass on various technological equipment used in the proposed method is determined, and the boundary parameters obtained by reducing the particle size distribution at each concentration stage both by rock mass and by recoverable metal are numerically determined.

3. Based on the analysis of the content of particles of recoverable metal in each particle size class, experimental processing of the processed rock is performed on finishing equipment (centrifugal concentrators of large and low capacity, on magnetic and magneto-liquid separators) to determine the limits of stepwise reduction of the processed material on finishing equipment, as well as for determine the minimum particle size of the recoverable metal in order to determine the economic advisable ti subsequent application of technological operations for additional extraction of small and fine particles recovered metal by chemical means - or cyanidation heap leaching, especially regarding the processing of the rock mass finely divided ore processing.

4. On any type of calibration equipment, gangue particles that exceed the maximum particle size of the recoverable metal, determined by laboratory analyzes of the initial rock mass, are excluded from the source rock mass.

5. The calibrated rock mass (see Figure 1) is fed to any type 1 dispenser by any transport scheme, from which it enters into a receiving chamber-mixer 2 in metered volumes, where it is mixed by means of barbation with water supplied by a pumping unit 4.

6. The rock mass mixed with water in the form of pulp enters the suction port of the pump 3 for pumping any type of hydraulic mixture (jet, sand, soil, etc.) mounted in the lower part of the mixing chamber, which feeds the pulp through a pressure pulp 6.

In the design of the pressure slurry conduit 6, a self-trapping device 7 is made, which is installed in the pulp conduit during processing of rock mass containing native particles of heavy precious metals, for example, Au. Native particles from the self-trapping device 7 are periodically discharged into a transport container 8 for transportation to a storage location or further processing.

7. The flow of pulp moving through the discharge pulp through the inlet pipe 9 enters a hydraulic screen 10, in which the particles of the solid phase of the hydraulic mixture are calibrated on a conical sieve installed in the conical part 12 of the screen-separator 10, until the calibration parameters of the particles of the solid phase of the hydraulic mixture are determined by calculation on a hydraulic screen of the first stage of enrichment, dividing into oversize and under-sieve products.

8. Oversize product in the form of large particles of waste rock that have not passed through the calibration sieve is discharged from the conical part 12 of the hydraulic screen 10 through the side pipe 11 in the form of a hydraulic mixture (pulp).

In the process of calibrating the rock mass on a conical sieve, it is possible to accidentally flush some part of the recoverable metal and take it out together with the gangue particles out of the hydraulic screen by the flow of the liquid phase of the hydraulic mixture through the side pipe 11. To correct this phenomenon, the flow of the hydraulic mixture from the pipe 11 is directed to recover particles emitted metal in finishing equipment of any type 42 (fine-filling gateway, jigging machine, etc.).

9. The sublattice product in the form of small particles of gangue and recoverable metal, passed through a calibration sieve, is removed from the conical part of the hydraulic screen through a divider 13 and enters in the form of a pulp into an inclined slurry line 14 made of previously determined one or more pipes.

Moving by gravity in the composition of the pulp along the inclined pulp line 14 (see Figure 2), all particles of the extracted metal and particles of waste rock are exposed to gravitational force, under the influence of which they are deposited in the bottom of the slurry line 15 and move along the bottom of the slurry line in the form of an flooded layer of solid particles with an area of f _k , while the density of the flooded layer depends on the mineralogical, lithological and granulometric compositions of the particles of the processed rock mass, as well as the shape of the particles.

The calculation of the parameters of the slurry conduit of the first enrichment stage (the length of the slurry conduit, the width and the underestimation of the slit of the bottom cutter and the length of the end section of the slurry conduit) is carried out in accordance with the previously determined particle size distribution of the particles of recoverable metal removed from the slurry conduit of the first enrichment stage through the bottom cutter.

The process of deposition of solid particles in an inclined slurry conduit 15 is carried out in a moving pulp stream (see Figure 4), i.e. simultaneously with the deposition under the influence of gravitational force, solid particles 46 participate in the process of their transfer by a pulp stream moving along the slurry conduit, and the picture of the velocity distribution in the flow section is a rotation paraboloid.

In the turbulent mode of movement of the hydraulic mixture flow along the slurry duct, an increase in the flow velocity (in the direction from the walls of the slurry duct to the center) increases from the minimum value V _min on the pipe wall to the maximum value V _max on the longitudinal axis of the slurry line. The pattern of increasing speed is expressed by the equation of the parabola.

9. As part of the pulp in the stabilization section of the pulp line of the first enrichment stage 14, solid particles 46 move along the pulp line 15 in an unsteady flow regime characterized by random particle movement in axial and transverse directions with respect to the direction of flow of the pulp.

As a result of the randomness of the movement of particles in the space of the slurry conduit 15, particles 46, solid particles (see Figure 4), fall into the wall ring laminar layer with a thickness of 5.

Particles of gangue and particles of recoverable metal, the sizes of which are smaller than the thickness of the annular laminar layer 6, are “locked” in the thin laminar layer by moving pulp layers located closer to the axis of the pulp duct and, accordingly, having high speeds of movement, continuing to move with speed V _t in the axial direction along the slurry conduit “locked” in a moving hydraulic trap.

Particles of waste rock and particles of recoverable metal, the dimensions of which are larger than the thickness of the annular laminar layer 5, are deposited in the pulp stream moving along the pulp duct under the action of gravitational force.

In the annular gap of the laminar wall layer, the solid particle 46 (see Fig. 4, 6) moves along the trajectory FF ₁ -F ₂ -F _n in the axial direction with a speed V _t .

Under the influence of gravitational force, the solid particle 46 slides down the wall of the slurry duct 15 by a distance S _t , passing a vertical distance h _t and moving horizontally in the transverse direction by a distance b _t , in the axial direction the solid particle is transferred by a distance L _x , and the particle transfer It is produced by pulp layers located closer to the axis of the slurry line and moving at a higher speed.

10. Laboratory studies of the pulp volumes discharged through the bottom cutter 16 of the slurry line of the first enrichment stage (see Figure 2) determine the ratio T: G in the test pulp volumes, based on which the volume of the bottom layer F “is mathematically determined, layer thickness hi, width layer along the upper edge at points L and T (chord b _s ) and the central angle 2a, based on the chord b _s formed by the radii LO and TO, drawn from the center at point O of the cross section of the slurry line 15, to the points L and T of the intersection of the circle with chord.

11. A solid particle 46 (see Figure 5) located on the inner wall of the pipe of the slurry duct 15 and moving in the wall laminar layer of thickness δ is affected by the gravitational force Q, which decomposes into the normal pressure force G _f , whose vector is perpendicular to the plane passing through the point of contact by the particle of the wall of the slurry conduit at point F, and the rolling force C _f directed tangentially to the indicated plane.

From the analysis of the graphical construction (see Figure 5), it can be seen that the angle α _f formed by the normal pressure force vectors C _f and the gravitational force vector Q is half the central angle 2α _f , and the angle β formed by the vectors of the gravitational force Q and rolling force C _f , is the angle of repose Q for the material of the deposited particle 46 of the extracted mineral (metal). Data on the angles of repose of materials 9 in dry and wet are given in the special technical literature or are determined experimentally for each type of rock mass processed specifically.

12. Based on the established patterns (see Figure 2) and based on the fact that the solid particle 46 of the extracted metal will slide down the arc (along the inner wall of the slurry duct 15) until the angle β becomes equal to or less than the angle the natural slope θ for the material of the deposited particle in the wet state, while the chord value, on which the central angle 2α rests, determines the size of the slit width b _s at which the particle can still slide down the inner wall of the slurry duct.

We take the angle α _f = θ, which guarantees the movement of a solid particle down the inner wall of the slurry duct 15, then based on the half value of the central angle 2α _f and the inner diameter of the slurry duct D _p , the slit width of the bottom cutter (chord value b _f ) is determined by the geometric formula:

13. The bottom cutoff of the slurry duct of the first enrichment stage (see Figure 4) is structurally a tape slit 43 of length L _f and width b _f , made in the bottom of the slurry duct 15. The bottom of the slit is closed by a housing 44, in which the valve 45 is pivotally mounted, overlapping the cavity of the slit 43 from the bottom, when the valve 45 is fully pressed to the slurry pipe 15.

The position of the valve 45 in the housing 44 is determined by an adjustable distance h ₂ , which can vary from 0 to a value of h ₁ (depth of the housing 44), thereby allowing you to adjust the volume of the hydraulic fluid discharged from the slurry line through the bottom cutoff.

The maximum distance of the vertical precipitation h _t particles (see Figure 2) to the entrance to the gap zone of the cutter is determined by the formula:

h _t = D _p + 0.5 × D _p -h ₁

where: h ₁ - the distance of the vertical precipitation of a solid particle;

D _p is the inner diameter of the slurry line;

h ₁ - the value of the arrow of the arc segment with an area of F _to , determined by the formula:

where: 2α _f is the central angle of the angular sector of the circle with a diameter equal to the inner diameter of the slurry conduit D _p based on the chord equal to the width of the slit b _{f of the} bottom cutter.

14. The pulp in the form of a primary concentrate (see FIG. 1) discharged through the bottom slurry cutter of the first stage of enrichment of pulp enters the receiving chamber-mixer 18, in the lower part of which the receiving port of any type of pulp transfer pump is structurally made (jet, sand, soil and .t.) 19, which feeds the pulp through the pressure pulp line 20, while the water in the pump 19 is supplied by the pumping unit 4.

15. The flow of pulp moving through the discharge slurry 20 through the inlet pipe 21 enters a hydraulic screen 22, in which the solid particles of the pulp are calibrated on any type of conical sieve installed in the conical part 24 of the screen-separator 22, up to the particle calibration parameters determined by the calculation of the hydraulic screen the second stage of enrichment, dividing into oversize and sublattice products.

16. Oversize product in the form of large particles of gangue that did not pass through the calibration sieve is discharged from the conical part 24 of the hydraulic screen 22 through the side pipe 23 in the form of a pulp.

In the process of calibrating the rock mass on a conical sieve, it is possible to accidentally flush some part of the recoverable metal and take it out together with the gangue particles out of the hydraulic screen by the pulp stream through the side pipe 23. To correct this phenomenon, the pulp stream from the pipe 23 is directed to further extract the particles of the metal to be emitted into finishing equipment of any type 42 (fine-filling gateway, jigging machine, etc.).

17. The sublattice product in the form of small particles of gangue and recoverable metal passing through the calibration sieve is discharged from the conical part of the hydraulic screen and through the divider 25 it enters in the form of a pulp into an inclined slurry line 26 made of one or more pipes previously determined by calculation.

18. The calculation of the parameters of the slurry conduit of the second enrichment stage (length of the slurry conduit, the width and the underestimation of the slit of the bottom cutter 28, the length of the end section of the slurry conduit) is performed in accordance with the previously determined particle size distribution of particles of recoverable metal removed from the slurry conduit of the first enrichment stage through the bottom cutter, and according to the methodology and the sequence of calculations of the pulp line parameters of the second enrichment stage in accordance with paragraphs 9-13.

19. Particles of recoverable metal and particles of waste rock, removed from the slurry duct 27 through the slit of the bottom cutter 28, are fed by gravity to further technological processing at the finishing equipment complex, which carries out a number of sequential technological operations united in the “Process I” group (see Figure 1 , Fig. 8).

20. The pulp from the end sections 17 and 29 (see Figure 1) of the pulp lines of the first and second enrichment stages enters the conical capacity of the gravity settler 33, which is structurally part of the thin-layer thickener 39.

The flow of pulp in the composition of solid particles of waste rock, recoverable mineral (metal) and liquid (water), falling into the tank of the gravity sump 33 (see Figure 7), loses speed and disperses in the space of the tank in horizontal and vertical directions in the process of spatial diffusion .

21. The pulp flow due to the closed capacity of the gravity sedimentation tank 33 rises upward and enters the block of inclined, isolated from each other channels 34 installed at an angle φ ₁ to the horizontal plane, length L _K1 , height h _k1 and cross-sectional area F _k1 which are calculated so that the bottoms down the inclined channels with velocity V _t1 moving large particles only gangue 63 precipitating in the channel cavity with hindered settling velocity W _st1, but upward at a speed V _g1 together with a stream of liquid particles move isolated emogo metal and fine particles of gangue 64, the difference in the granulometric size of which is determined ravnopadaemosti coefficient.

22. Particles of gangue 63 under the influence of gravitational force fall down and accumulate in the form of a condensed sediment in the tank of the gravity sump 33, from which, as they accumulate, they are removed by a jet pump 40 located in the bottom of the gravity sump 38. The thickened sediment in the form of a mixture of gangue particles and water supplied by the pumping unit 4 is sent to the hydraulic dump 41.

23. Leaving the block of inclined channels 34, the pulp stream (see Fig. 7) at a speed of V _g1 enters the inter-partition spaces of the vertical stabilization lattice 32 made of rectangular rectangular channels of size b _r × L _r and height h _r .

Being poured over the edges of the stabilization grating (see Fig. 9), the pulp stream with solid particles 64 enters the open collar type collecting tray 65, the cross section of which allows the outgoing pulp stream through the channel of the collection tray of width b _{v to} flow by gravity into the receiving funnel 35 of a thin-layer block thickening 36 (see Figure 1).

23. Entering the capacity of the block of thin-layer thickening 36, the pulp stream, due to the closed capacity of the block of thin-layer thickening 36, rises upward and enters the block of inclined, isolated from each other channels 37 installed at an angle φ ₂ to the horizontal plane, length l _k2 , height h _k2 and cross-sectional area F _K2 are calculated in such a manner, so that a collecting container of a thin thickener precipitated mass comprising larger particles emitted metal and gangue particles, the difference in particle-size p zmerah ravnopadaemosti coefficient is determined, and the block leaving the thin layer of condensation hydraulic mixture would contain in its structure particles of gangue and the metal being recovered, which corresponds to a specific size previously granulometric class calculation step for calibration at block thin-layer thin-film condensation thickener.

Large particles 66 of waste rock and emitted metal move down the inclined bottoms of the channels with a speed V _t2 , deposited in the cavity of the channel with a constrained deposition velocity W _st2 , and small particles 67 of the released metal and empty rock particles move together with the fluid flow with a velocity V _g2 . The difference in particle size distribution of particles is determined by the coefficient of equidistance.

24. Particles of recoverable metal and particles of gangue 66 under the influence of gravitational force fall down and accumulate in the form of condensed sediment in the collection tank of a thin-layer thickener 36, from which they are discharged through an outlet pipe located in the bottom of the tank for further processing at the finishing equipment complexes, carrying out a number of sequential technological operations, united in the groups “Process I” and “Process II” (see Figure 1, Figure 8).

The essence of technological operations, united in the group “Process I”, is as follows.

The concentrated condensed sediment of particles in the collection tank of a thin-layer thickener 36 is processed in a concentrator 47, the concentrate of the concentrator 50 obtained by processing is dried in any type of drying equipment 51, and 52 is processed in a dry form on a magneto-liquid separator 53. The separation concentrate obtained by treatment on a magneto-liquid separator 54 is stored in a collection tank 55 and sent for chemical cleaning by refining methods. The “tails” of the magneto-liquid separation 56 are sent to the dump.

In the case of the presence in the “tailings” of the concentrator 48, stored in the collection tank 49, of free, chemically unbound particles of the metal to be released, they are subjected to repeated single or double processing (purification), for example, at a low-capacity centrifugal concentrator.

If there are 48 non-free, chemically bonded particles of emitted metal in the “tailings” of the concentrator, they are subject to further technological processing for the extraction of gold, for example, gold, subject to sanitary and environmental safety measures, at a finishing equipment complex that implements a series of sequential technological operations grouped “Process II” (see Figure 1, Figure 11), in the absence of particles of recoverable mineral in the “tails” - is sent without processing to the dump.

25. The essence of technological operations, united in the group “Process II”, is as follows.

Concentrated condensed sediment of particles in the collection tank of a thin-layer thickener 36 or tailings of concentration 48 from the collection tank 49 in the form of a technological product 58 are processed by heap leaching in a technological tank 57 or cyanidation in a technological tank 61.

Obtained as a result of processing in one way or another chemical processing, the solution of the recoverable metal 60 is poured into the collection tank 61 and goes to further processing in order to extract from the metal solution. The waste rock in the form of “tails” 59, 62 is sent to the dump.

26. The pulp emerging from the thin-layer thickening unit through the pipe 39, containing particles of gangue and recoverable metal, the sizes of which correspond to the particle size classification determined earlier by the calculation at the stage of calibration of solid particles in the thin-layer thickening unit of the thin-layer thickener, is sent to the settler 5.

If the clarified hydraulic mixture flowing from the block of thin-layer thickening flows meets the technical requirements for the water transfer pumps used in the proposed technological scheme in terms of particle size distribution and particle concentration in the liquid pumped by the pump (water), these drains can be used as a source of circulating water water supply in the proposed technological scheme.

Calculation example.

It is proposed to calculate the main parameters of the installation for extracting free, chemically unrelated gold particles from the rock mass of the placer deposit.

The rock mass of the placer deposit by lithological composition consists of river sand with boulder-pebble inclusions characteristic of the indicated type of deposits, sand particles up to 3.0 mm in size have a rounded shape.

The density of the rock mass of the deposit is 2.4 t / m ³ , the average gold content is 2.88 g / m ³ .

The particle size distribution of the particles of the host rock mass, gold particles and the distribution of gold particles by particle size classes are presented in Table 1.

Analyzing the data given in Table 1, we can say that the task of extracting gold particles is reduced to the deposition of metal from the host rock mass, represented by particles of a particle size class - 3.0 mm.

Gold in the grading classes presented in Table 1 is represented by particles of various shapes, the shape of the particles is shown in Table 2.

Experimental processing of the initial rock mass at centrifugal concentrators with a capacity of 1 t / h showed that the recoverability of gold particles of round and angular shape of particle size classes - 3 + 0.074 mm is within the economic feasibility of using these devices in terms of extracting composite, chemically unrelated gold particles, and experimental processing of the initial rock mass at centrifugal concentrators with a capacity of 100 kg / h showed that the recoverability of elongated gold particles to the face ometricheskih grades - 0.074 + 0.044 mm is also within the economic feasibility of using these devices in terms of extraction of summary, not chemically bound gold particles.

Experimental processing of the initial rock mass at centrifugal concentrators with a capacity of 100 kg / h showed that the recoverability of gold particles of oblong and lamellar shape of particle size classes - 0.044 + 0.022 mm is about 68% of their initial content in the processed rock mass.

Experimental processing of the initial rock mass showed that there is no extraction of gold particles of a plate-shaped granulometric class - 0.022 mm when processed on centrifugal concentrators with a capacity of 100 kg / h. Therefore, the extraction of gold particles of this shape and such particle size distribution is possible only using cyanidation or heap leaching methods.

The use of such chemical extraction methods is possible with strict observance of sanitary and environmental safety measures.

Extraction of such an amount of metal (14.4 mg: 2880 mg × 100% = 0.5%) by chemical methods is not economically feasible, therefore, a fraction of 0.022 mm will not be processed by cyanidation or heap leaching and will be disposed of.

Based on the data given in Table 1, we will disaggregate the stages of calibration of the initial rock mass by granulometric classes in relation to the equipment used for reduction in terms of 1 m ^{3 of the} processed mass and summarize the data in Table 3.

In a moving pulp stream in the pulp line of the first enrichment stage, particles of recoverable metal and 3 mm grade rock will be deposited.

An analysis of the data in Table 3 shows that when processing the original rock mass of the above particle size distribution by the proposed method, 46.4% of the rock mass (from the initial volume) calibrated to a size of 3 mm enters the slurry line of the first enrichment stage.

We calculate the parameters of the installation equipment with a capacity of 25 m ³ / h based on the initial mass (2400 × 25 = 60,000 kg = 60 t). 25 × 0.434 = 10.85 m ³ / h (10.85 × 2.4 = 26.04 t / h) of calibrated rock mass will enter the pulp line of the first enrichment stage per hour.

Pulp line to make of pipes 159 × 4.5 mm, the cross-sectional area of one pipe:

Ft _p = 0.25 × π × D $\genfrac{}{}{0ex}{}{\text{2}}{\text{p}}$ = 0.25 × 3.14 × 15 ² = 176.71458. Ft _p = 176.71458 cm ²

We perform the slurry line of the first stage of enrichment with a cellular one - consisting of four pipes D _p = 150 mm, then the total area of the slurry line will be equal to:

F = 4 × Ft _p = 176.71458 × 4 = 706.85832 cm ²

Let us define the ratio T: L = 1: 20 in the pulp flow moving along the pulp line. Pulp consumption Q _p along the pulp line:

Q _p = 10.85 × 20 + 10.85 = 227.85 (m ³ / h) = 227.85: 3600 = 0.063291666 (m ³ / s)

Q _p = 63,291.666 cm ³ / s

The average speed V _plp of the flow of the hydraulic mixture through the pipes of the slurry pipe:

V _plp = Q _p : F = 63,291.666: 706.85832 = 89.539394

V _plp = 95.728753 cm / s = 0.895393 m / s

The temperature of the water supplied to the units of the installation is + 12 ° C.

The density of the flow of the hydraulic mixture q _sm moving through the pipes:

The weighted average size d _{sr of} particles moving along the slurry duct for the indicated particle size classes by the formula:

The value of W ₀ for solid particles with a particle size of d _sr = 0.9163318 mm:

Minimum travel speed V _min :

V _min = 52.271338 cm / s

Experimental measurement in the product (pulp), discharged through the slit of the bottom cutter, determined T: W = 1: 5.

The volume of solid particles moves along one pipe of the slurry duct:

The flow of pulp through the pulp line moves at a speed of 89.5393 cm / s, then the area of the arc segment f _to occupied by the volume of solid particles is:

f _kt = 753.47222: 89.5393 = 8.41498. f _kt = 8.41498 cm ²

The area of the moving bottom stream f _k equal to the area of the arc segment occupied by the moving bottom stream of pulp with the ratio T: W = 1: 5 is determined by the formula:

f _k = 8.41498 × 6 = 50.48988 (cm ² )

According to the tables of geometric relationships between the area of the arc segment, the diameter of the slurry duct D _p and the central angle 2α _f, we find that at 2α _f = 140 ° the area of the arc segment f _k = 50.643922 cm ² .

Geometrically, the area of the moving bottom flow:

The underestimation of the slit of the bottom cutter h ₂ is determined from the equality f _k = F _f :

h ₂ = 2.7635074 mm

With such an underestimation of the slit of the bottom cutoff h ₂ = 2.7635074 mm along the arc (in the bottom of the pipe) of the slurry conduit of 140 °, the entire moving bottom flow F _K leaves the slurry conduit, i.e. reduction of the volume of output mass of solid particles in relation to their original volume will not occur.

The bulk of the particles removed from the slurry line is represented by flooded sand particles, the angle of repose 9 of which is 25 °.

We take the angle α _f = 25 °, which guarantees the movement of a solid particle down the inner wall of the slurry duct, then based on half the central angle 2α _f = 50 ° and the inner diameter of the slurry duct D _p = 15 cm, determine the width of the slit of the bottom cutter (chord value b _f ) by the geometric formula:

b _f = D _p : Sinα _f = 15: Sin 25 ° = 6.4 b _f = 6.4 cm

The amount of understatement of the slit of the bottom cutter h _2min with a slit width b _f = 6.4 cm is selected in accordance with the equality:

With such an underestimation of the slit of the bottom cutoff h ₂ = 14.29731 mm along the arc (in the bottom of the pipe) of the slurry conduit at 50 °, the entire moving bottom stream f _k leaves the slurry conduit, i.e. reduction of the volume of output mass of solid particles in relation to their original volume will not occur.

The minimum value of the understatement of the slit of the bottom cutter h _{2min is more} rational to choose depending on the size of the output particles d _tm :

h _2min = d _tm +0.5 mm = 3 + 0.5 = 3.5 h _2min = 3.5 mm

where: d _tm is the maximum particle size of the recoverable metal.

Accepting the width of the slit of the bottom cutter b _f = 6.4 cm, and the underestimation of the slit h _2min = 3.5 mm, we determine the cross-sectional area f _k1 of the pulp flow cross section through the slot from the slurry duct:

f _k1 = b _f × h _2min = 3.5 × 6.4 = 22.4. f _k1 = 22.4 cm ²

The reduction in the volume of particles removed from the slurry line at f _k1 = 22.4 cm ² with respect to the volume of particles entering the slurry line is determined by the ratio:

f _k : f _k1 = 50.643922: 22.4 = 2.26088

13. The minimum length of the slit of the bottom cutter L _fmin :

14. The rate of cramped deposition in the slurry line of the first enrichment stage is performed for the deposited quartz particles and gold particles with a size of 0.074 mm. Quartz particle size:

d _q = d _Au × K _p

d _q = 0.074 × 3.171861 = 0.234717; d _sq = 0.234717 mm

Equivalence coefficient for particles with a size of 0.234717 mm:

d _q = 0.074 × 3.171861 = 0.234717; d _sq = 0.234717 mm

The value of W ₀ for t particles d _q = 0.234717 mm:

The constrained deposition rate W _st and is determined by the formula:

W _st = W ₀ × K ₁ × K ₂ × K ₃ × K ₄

K ₁ - coefficient taking into account the decrease in the actual cross section of the fluid flow due to the space occupied by solid particles, is determined by the formula :.

K _i = 1-γ ^2/3

where: γ is the part of the volume of the hydraulic mixture flow that is occupied by the solid phase (solid particles): γ = 11.6 m ³ : 243.6 m ³ = 0.047619

K ₁ = 1-0.047619 ^2/3 = 0.868623

K ₂ - coefficient taking into account the change in the average difference in specific gravity between solid particles and liquid by a value of 1-γ.

K ₂ = 1-γ = 1-0.047619 = 0.952381

K ₃ - coefficient taking into account the correction for viscosity.

K ₃ = 1-2.5γ = 1-2.5 × 0.047619 = 0.88

To ₄ - coefficient taking into account the shape of the particles.

K ₁ × K ₂ × K ₃ = 0.868623 × 0.952381 × 0.880952 = 0.728776

The value of the form factor K ₄ is selected, depending on the shape of the particles, from hydraulic references:

The calculation of the time of deposition of a solid particle Tt is carried out according to the formula:

T _t = h _t : W _st

where: h _t is the distance of the vertical sedimentation of the particle, h _t = 13.95195 cm;

W _st - the rate of constrained deposition of a solid particle, cm / s

We take the length of the flow stabilization section to be equal to 40 pulp duct diameters, i.e. l _article = 0.15 × 40 = 6.0. L _st = 6 m

The length of the pulp line L _p = L _article + L _oc

Let us determine the value of the constrained deposition rate W _st for different particle shapes and the sedimentation rate of particles of different shapes in the slurry conduit of the first enrichment stage. The results are summarized in table 4.

When processing 25 m ³ / h of the initial rock mass, q _pk1 = 4.7975 m ³ / h of solid particles, which are calibrated into the hydraulic screen of the second stage of enrichment, are discharged through the bottom cutoffs of the slurry line of the first enrichment stage.

In the hydraulic screen of the second enrichment stage, the pulp solid particles are calibrated to a size of 3 mm and the pulp in the form of an under-sieve product is fed to hydrogravity enrichment in the slurry line of the second enrichment stage, with 0.239875 m ³ / h of solid particles being carried into the side pipe into the dump ( 5%).

Let us set the ratio T: L = 1: 20 in the mixture entering the pulp line of the second enrichment stage, the total volume of pulp flowing through the pulp line is: Q _total = 4.557625 × 20 + 4.557625 = 95.710125 (m ³ / h )

Secondary pulp consumption q _p :

q _p = Q _total : 3600 = 95.710125: 3600 = 26586.145 (cm ³ / s).

The pulp line of the second enrichment stage is made of two pipes with an inner diameter of 150 mm, the cross-sectional area of the slurry line is:

F ₂ = 2 × 0.25 × π × D $\genfrac{}{}{0ex}{}{\text{2}}{\text{p}}$ = 353.42916 (cm ² )

The velocity of the pulp stream V _pp2 along the pulp line of the second enrichment stage:

V _pp2 = q _p : F ₂ = 26586.145: 353.42916 = 75.223405 (cm / s)

We will calculate the parameters of the slurry line of the second enrichment stage using the calculation formulas for the slurry line of the first enrichment stage, since we will take into account all the boundary conditions and parameters that were used in the calculation of the slurry line of the first enrichment stage.

The length of the slurry conduit of the second enrichment stage is determined taking into account the change in the velocity of the pulp stream in the slurry conduit with respect to the value of the pulp flow rate in the slurry conduit of the first enrichment stage. The calculation is summarized in table 5.

The volume of the secondary concentrate q _pk2 , _discharged through the bottom cutoffs of two pipes of the slurry line of the second enrichment stage:

q _pk2 = q _pk1 : K = 4.557625: 2.26088 = 2.0158632 (m ³ / h) = 559.962 (cm ³ / s).

where: K is the coefficient of volume reduction, K = 2,26088.

When the ratio T: W = 1: 5, the volumetric water content in the pulp q _in , discharged through the bottom cutter:

q _in = q _pk2 × 5 = 559.962 × 5 = 2799.81 (cm ³ / s) = 10.079316 (m ³ / h)

Summarize the results of the enrichment of the rock mass in the slurry conduit of the second stage of enrichment and summarize them in table 6.

We summarize the results obtained on the enrichment of 25 m ^{3 of} rock mass in the equipment of the proposed method and summarize them in table 7.

Calculation of the parameters of the inclined channels installed in the upper part of the gravity sump settles to the fact that in the collecting tank of the gravity sump should concentrate the gangue particles of the class - 3 + 0.074 mm, and the gold particles of the particle size class - 0.074 mm and differ in size by the coefficient of equidistance of the particle waste rock should be removed from the block of inclined channels by a pulp stream. The flow density of the hydraulic mixture q _sm :

Equivalence coefficient for particles: d _t <0.1 mm:

A particle size of sand (quartz) having the same deposition rate with a gold particle of 0.044 mm in size:

d _p = q _au × K _p = 0.044 × 3.196651 = 0.14065264 (mm)

W ₀ for solid particles with a particle size d _q = 0.140652 mm:

Calculation of the parameters of the channels will be made by the geometric and physical parameters of the sand particles with a size of d _q = 0.14065264 mm.

The numerical value of the parameter Re _w for a particle of size d _sq :

The concentration of the hydraulic mixture (pulp) of mixture S in the inclined channels of the gravity settler:

The constrained deposition rate W _{st of} solid particles in inclined channels located in the upper part of the gravity settler:

W _st1 = W ₀ × K _R × K ₄

When Re _w <2, the value of k _r is determined by the formula:

We choose the channel cross section equal to h × b = 10 × 15 mm = 1 × 1.5 cm, installed at an angle of 60 ° to the horizontal plane along the long side b, then the value of the particle settling distance is determined by the formula:

L _v1 = 1: Cos60 ° = 1: 0.5 = 2 (cm).

Hydraulic radius of the channel section:

The speed of the constrained movement of the flow of the hydraulic mixture up the inclined channel v _g1 , Re = 485,04446:

Single Channel Z Performance:

Z = h × b × v _gt = 1 × 1.5 × 5 = 7.5 cm ³ / s

The number of inclined channels

Take N = 11644 = 108 × 108. The block of such channels is a rectangle 200 × 150 cm in size.

Stabilization grate height (5-10) × h _k1 = 10 × 2 = 20 (mm)

Area of one lattice cell (20-50) × F _k1 = 50 × 1.5 = 7.5 (cm ² )

We determine the value of the constrained deposition rate of particles of various shapes, determine the time of deposition of particles on the bottom of the channel and the length of the channel, the calculations are summarized in Table 8.

7.46921 m ³ / h of solid particles of gold fraction - 0.074 mm and particles of gangue, the size of which is larger than the gold particles by the coefficient of equidistance, enters the inclined channels of the block of thin-layer thickening.

0.425 m ³ / h of gold particles of the fraction - 0.022 mm and particles of gangue, the size of which is larger than the gold particles by the coefficient of equidistance, leaves the inclined channels.

7.044261 m ³ / h of gold particles - 0.074 + 0.022 mm in size and gangue particles, the size of which is larger than the gold particles by the equidistance coefficient, precipitates.

The calculation will be carried out according to the parameters of quartz particles. We apply channels of the same size as in the gravity settler.

The flow density of the hydraulic mixture q _sm :

Equivalence coefficient for particles: d _t <0.1 mm:

The particle size of sand (quartz) having the same deposition rate with a gold particle of 0.022 mm in size:

d _p = d _Au × K _p = 0.02 × 3.196651 = 0.070326 (mm)

W ₀ for solid particles with a particle size d _q = 0,070326 mm:

Calculation of the parameters of the channels will be performed by the geometric and physical parameters of sand particles with a size of d _q = 0.070326 mm.

The numerical value of the parameter Re _w for a particle of size d _sq :

The concentration of the hydraulic mixture (pulp) S in the inclined channels of the gravity sump:

The constrained deposition rate W _{st of} solid particles in inclined channels located in the upper part of the gravity settler:

W _st1 = W ₀ × K _R × K ₄

Single Channel Z Performance:

Z = h × b × v _g1 = 1 × 1.5 × 5 = 7.5 cm ³ / s

The number of inclined channels

Take N = 11644 = 108 × 108. The block of such channels is a rectangle 200 × 150 cm in size.

We determine the value of the constrained deposition rate of particles of various shapes, determine the time of deposition of particles on the bottom of the channel and the length of the channel, the calculations are summarized in Table 9.

Claims (1)

The enrichment method, which consists in excluding waste rock on calibration equipment that exceeds the maximum particle size of the recoverable metal, determined by laboratory analyzes and experimental processing on centrifugal separation equipment, of the material of the initial rock mass, followed by calibration of the resulting product on a hydraulic screen and hydrogravity enrichment of the sublattice hydraulic product screening in an inclined slurry conduit with a constant output from an inclined slurry conduit through the slit of the bottom cutter of the enriched product, followed by its calibration calibration on an additional hydraulic screen and hydrogravity enrichment of the sublattice product of an additional hydraulic screen in an additional inclined slurry line with the continuous removal of the secondary concentrate from the slant slurry line through the slot of the bottom cutter with subsequent continuous processing of the latter on centrifugal concentrator extraction of large particles of recoverable metal lavage equipment, while the effluent of sloped slurry pipelines is jointly stage-by-stage processed in a gravity settler and a thin-layer thickener for sedimentation of gangue particles in a gravity settler, and the sediment condensed in a thin-layer thickener is treated with additional centrifugal classifiers to extract fine particles of recoverable metal, and treated with cyanidation or heap leaching methods.

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