APPLICATION OF THE JET SCHEME AND FLOTATION MODE WITH A VAPOR-AIR MIXTURE DURING GOLD-BEARING ORES ENRICHMENT

Increasing the economic efficiency of enrichment of refractory gold-bearing ores is possible by reducing the cost of opening sulphide gold-bearing concentrates by reducing the yield of flotation concentrate sent to pyro- or hydrometallurgical processing. It is important to maintain the achieved level of gold recovery into concentrate with a lower concentrate yield. In order to achieve this goal, the gold content in the main flotation operation is increased by mixing the rough concentrate separated from ½ part of the ore feed with another ½ part of it. Mixing in the flotation operation of products with a high level of wash ability (ability to separate) – initial feed and rough concentrate – is identical to an increase in the content of the extractable component in the original ore. In accordance with the new regime of flotation, the process is carried out in cold slurry with bubbles filled with hot steam. The physical basis of the new flotation regime is the dependence of the surface properties of air bubbles and surface forces that determine the stability of the liquid film separating the bubble and the particle on the temperature, which increases due to the heat of the vapor-liquid phase transition. Radial oscillations of the bubble surface as a result of pressure pulsations during condensation and vapor evaporation and surface thermal flows of liquid in the wetting film are factors that provide an increase in the completeness of gold recovery and the selectivity of flotation adhesion obtained in ore flotation experiments. In laboratory conditions, a comparison was made of the indicators obtained during the dressing of gold-bearing ores according to the factory and new schemes. It was revealed that the use of the developed technology allows, with a lower concentrate yield, to obtain an increase of 7.06% abs. extraction of gold into a concentrate of the best quality. A decrease in gold losses with flotation tailings is a consequence of an increase in the number of particles reaching the bubble surface due to radial oscillations of its surface and a weakening of the stability of the wetting film by the thermo-capillary mechanism. An increase in adhesion selectivity can be interpreted using the concept of surface forces of structural origin – an increase in the forces of hydrophobic attraction and hydrophilic repulsion with increasing temperature. The interaction of vapor bubbles with nanobubbles on the surface of the solid phase (wetting by the Cassie-Baxter mechanism) ensures efficient particle recovery by the coalescence mechanism of flotation.

Gravity-flotation gold-bearing ore concentration

The paper focuses on the study of the gold-bearing ore dressability. According to technological research, the average gold content is 11.88 g/t. The silver content is insignificant – 2.43 g/t. Main ore minerals in the sample are pyrite and pyrrhotite. According to mineralogical and X-ray structural analysis, the average content of these minerals in the ore is about 6 % (in total). Main rock-forming minerals of the original ore are: quartz (60.1 %), quartz-chlorite-mica aggregates (3.8 %), carbonates (7.1 %). According to the study results, it was found that the gold recovery in the GRG test was 72.75 % with a total concentrate yield of 1.34 % and a content of 664.78 g/t. At the same time, the gold content in tailings was 3.29 g/t. A stage test showed that it is advisable to use a two-stage scheme for ore processing by gravity technology only. The first stage is in the grinding cycle with the 60–70 % ore size, and the second stage is with the final classifier overflow size of 90 % –0.071 mm. Centrifugal separation has high performance as a free gold recovery operation in the grinding cycle. A concentrate with a gold content of 2426 g/t was obtained with a yield of 0.31 % and a recovery of 63.74 %. The beneficiation of first stage tailings ground to 90 % –0.071 mm at the KC-CVD concentrator (modeling) made it possible to extract gold into a total gravity concentrate (KC-MD + KC-CVD) of 87.25 % with a concentrate yield of 22.63 %. The gold content in tailings was 1.97 g/t. The results of gravity and flotation concentration of the original ore indicate the feasibility of using a combined gravity-flotation technological scheme. In a closed experiment of the initial ore beneficiation according to the gravity-flotation scheme at a natural pH of the pulp (without adding acid), the following products were obtained: gravity concentrate with a gold content of 2426 g/t at a yield of 0.31 % and recovery of 64.06 %; flotation concentrate (after the II cleaning) with a gold content of 122 g/t at a yield of 2.90 % and recovery of 33.01 %; the total gold recovery in the gravity-flotation concentrate was 94.07 % with a yield of 3.21 % and an Au content of 345.87 g/t, the gold content in the flotation tailings was 0.72 g/t.

Laser-photometric lump separation of gold-bearing ore

This paper establishes the conditions for selecting the optimal lump separation method based on the values of indicator properties of the ore minerals and demonstrates respective pilot testing results for low-grade quartz-type sulfide gold-bearing ores. The studies have shown the high efficiency of laser-photometric processing of ores with low contents of finely disseminated ore components, for which other methods are ineffective. On the example of ore from the Natalka deposit, it has been shown that acceptable lump separation performance requires division of the original ore into high-grade balance ore (the separation of which would be impractical), low-grade balance ore (to be subjected to lump separation), and an off-balance mineralized mass. Lump processing remains efficient up to ore temperatures of minus 15 °C, if rinsed with 4 °C water. The following concentration indicators were achieved for the lean ore of the process sample taken from the Alternativny section at ore temperatures above minus 15 °С: with the concentrate yield of 20 to 28 %, the gold content increases by 3.0–4.5, with an average level of metal losses in the tailings of approximately 13 %. At ore temperatures down to minus 21 °C, when an opaque ice film with the thickness of over 1 mm is formed on the lump during its screening with washing, metal losses in the separation tailings become 2.0–2.5 times higher. The authors are grateful to the employees of TOMRA, CC of CJSC Polyus Zoloto, Natalka FMS department of OJSC RiM, and FSBI VIMS, who assisted in the pre-concentration technology testing for the ore of the Natalka deposit.

BEHAVIOUR OF SULFUR, ARSENIC AND ORGANIC CARBON IN A GRAVITY CONCENTRATION OF GOLD FROM REFRACTORY ORE

The results of assay-gravimetric, chemical, mineralogical analyzes of gold ore are presented. According to the content of sulfide sulfur and the degree of oxidation of sulfur gold-bearing ore is assigned to the poor sulfide type of ore in the primary zone. Gold ore refers to refractory carbonaceous arsenic-containing sulfide ores. The gravity concentration of ore was assessed using a laboratory 3-inch Knelson KC-MD3 centrifugal concentrator. According to the results of the GRG test, the total gold extraction was 39.58% with the total concentrate yield of 3.96%. The extraction of gold in the first stage is lower than in the subsequent stages. This indicates the absence of large gold in the ore. The behavior of sulfur and arsenic during gravity concentration is similar to the behavior of gold. These data confirm that the main amount of gold is associated with arsenopyrite (arsenic pyrite) and pyrite. The behavior of organic carbon during gravity is significantly different. Only 3.39% of organic carbon passes into the gravity concentrate from ore, which reduces the persistence of the concentrate. Most of the organic carbon (96.61%) remains in the tails of gravity. The content of components in the total gravity concentrate was: Au 21.63 g/t, S 6.40%, As 0.82%, C(organic) 0.58%. The extraction of the components in the total concentrate is as follows, %: Au 39.58; S 34.52; As 27.27; C(organic) 3.39.

Investigation of Flotation Separation of Elemental Sulfur and Sulfides from Lead Sulfate from Low-Temperature Leaching of Lead Concentrates

Monometallic ore that is mostly lead found in nature is extremely rare. The main natural raw material for the lead production is sulfide polymetallic ores. In this study the filter cake processing after the low-temperature autoclave leaching of the lead concentrate to produce a sulphide concentrate and lead tailings was investigated The study of component separation was carried out using the methods of mathematical planning of the second order experiment. The following optimal costs of reagents, g/t: 140-200 potassium xanthate, 70-100 foaming agent, 100 copper sulfate; the concentrate yield is 41-43 %; it is extracted to, %: 95 Fe, 49 Cu, 96 Zn, 98 S0, 18-19 Pb. At flotation 18,7% of lead goes into flotation concentrate and 80.5% is lead sulfate. The rectification of the obtained concentrate by flotation did not give acceptable results, since the yield of the foam product in all experiments was 93-96%. For the separation of lead sulphate from sulphur-sulfide concentrate was used in the granulation of sulfur in the following conditions: t = 145 °C, Po2 = 0,0-0,5 MPa, τ = 60 to 120 min. In the granulation process of the flotation concentrate is a division of lead sulfate and elemental sulfur, the sulfate lead content in sulphur-sulfide phase is decreased from 28.44 % to 3.5 %, its recovery in a sulfate filter cake has reached 90.6 %

Improvement of the technological effects of flotation of lean fine disseminated scheelite ores

The paper describes the results of studying ways to improve the contrast of calcite and scheelite technological properties using water glass combined with aluminum, zinc, iron, magnesium sulphate salts, a mixture of water glass and calcium chloride, sodium carboxymethyl cellulose (CMC), combinations of sodium oleate with low-polar compounds (neonol, fatty isoalcohols), liquid phase and oleate ultrasound treatment. The monomineralic fraction of calcite floated by mechanical cell demonstrated that the minimum recovery of calcite is achieved by combining the Fe(II) salt and water glass (3(4) : 1). When f loating lean sheelite ore with a high carbonate modulus on domestic water, the combined use of water glass and CaCl2 reduces the floatability of calcium. Calcium chloride added to water glass on recycling water leads to a certain increase in the rough concentrate yield (13.8 to 14.1 %) with a significant decrease of WO3 recovery to the finished selection concentrate (72.7 to 53.3 %) and a deterioration in the concentrate quality. Replacement of water glass with CMC did not show satisfactory results. Ultrasonic treatment of pulp, liquid phase, collector leads to a certain increase in the calcite floatability, possibly due to the higher liquid phase temperature and increased proportion of the oleate ionic form. The use of neonols in the reagent scheme of flotation of scheelite-containing ore with a high carbonate modulus found no evidence of a decrease in the flotatability of calcite obtained when studying monomineralic calcite fractions unlike fatty isoalcohols that provided better concentrates in the selection cycle in comparison with a single oleate.

Groundwater nanofiltration process efficiency improvement with additional concentrate membrane treatment

Nanofiltration (NF) of waste water originated from nanofiltration drinking water plant stationed in town of Kikinda (Northern Serbia) was investigated. Experiments on removal characteristics of nanofiltration membranes when exposed to influent rich in arsenic and natural organic matter under different flux and transmembrane pressure (TMP) conditions were conducted in order to obtain drinking water from waste water and reduce total amount of waste water. Applied NF membranes showed remarkable removal characteristic. Also, obtained result for concentrate yield, an indicator of reduced concentrate amount, of 8.89% under optimum flux value presents considerable amount of reclamated drinking water. Calculated empirical and theoretical concentration factors quotients were indicated to probable accumulation of some inlet water components at NF membrane surface. Based on the obtained results that nanofiltration membranes remove natural organic matter, arsenic, ammonia and sodium with high efficiencies of 98%, 96%, 80% and 93% respectively, conceptual design of drinking water plant for City of Kikinda was defined and presented.

Experimental Research of Roughing and Cleaning Column Flotation of Difficult-to-Float High Ash Coal

Coal slime in Zhongcheng coal preparation plant in Baotou, Inner Mongolia, was analyzed. It is found that this coal sample is with high inner ash content and very difficult to float. It is hard to obtain low ash concentrate and high ash tailing with primary flotation. The live scale required ash content is 12.00% for total clean coal, 15.00% for flotation concentrate. It is found that the lowest concentrate ash content is 29.24% through exploring flotation experiments, which can not meet the producing requirement. Results of timed release flotation experiment show that when concentrate ash content is 15.00%, the theoretical concentrate yield is 24.36%, with 50.03% of flotation tailing ash content and 35.49% of combustible matter recovery. It can be concluded that the floatable grade of this coal sample is Very Difficult to Float. So flotation column is taken into account. The lowest concentrate ash content of roughing flotation by flotation column is 18.94%, which does not meet the requirement, either. Results of cleaning flotation of roughing flotation concentrate show that when the recycling pressure is 0.14 MPa, the concentrate ash content is 14.15% with a yield of 30.39%, which is satisfactory and can be referred to direct industrial production.

The Magnetic Separation Study on Flotation Concentrate Powder of Magnesite Ore

Iron in magnesite concentrate powder would seriously endanger the quality of refractories. Slon-500 vertical ring pulsating and high gradient magnetic separator and CRIMM DCJB70-200 laboratory electromagnetic splint VPHGMS were used to remove iron from magnesite flotation concentrate powder produced in Haicheng region. The magnetic separation results showed that using CRIMM DCJB70-200 laboratory electromagnetic splint VPHGMS is more suitable, and the most suitable operating conditions is the background field strength 641kA / m and bar spacing 2mm, at which time the magnesite concentrate yield was 91.25%, and the Fe2O3 content was 0.34%.