Reductive-oxidative pretreatment of a stibnite flotation concentrate: Thermodynamic and kinetic considerations

1998 ◽  
Vol 11 (6) ◽  
pp. 563-580 ◽  
Author(s):  
F.P Gudyanga ◽  
T Mahlangu ◽  
J Chifamba ◽  
D.J Simbi
Keyword(s):  
Flotation Concentrate

Metering device for flotation concentrate

1971 ◽  
Vol 7 (2) ◽  
pp. 107-108
Author(s):  
B. S. Modestov ◽  
V. L. Burmaka ◽  
Yu. V. Minaev ◽  
E. P. Moskalev
Keyword(s):  
Flotation Concentrate

Improvement of zinc bioleaching from a zinc flotation concentrate using mechanical activation

2021 ◽  
Vol 163 ◽  
pp. 106793
Author(s):  
Ali Zahiri ◽  
Ali Ahmadi ◽  
Abdolrahim Foroutan ◽  
Mahdi Ghadiri
Keyword(s):  
Mechanical Activation ◽  
Flotation Concentrate

Gravity-flotation gold-bearing ore concentration

2021 ◽  
Vol 1 (1) ◽  
pp. 4-15
Author(s):  
P. K. Fedotov ◽  
A. E. Senchenko ◽  
K. V. Fedotov ◽  
A. E. Burdonov
Keyword(s):  
High Performance ◽  
Gold Recovery ◽  
Gold Content ◽  
Flotation Concentrate ◽  
Stage Scheme ◽  
Study Results ◽  
Ore Minerals ◽  
Concentrate Yield ◽  
Flotation Concentration ◽  
Gold Bearing

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.

scholarly journals The Use of Heap Bioleaching as a Pre-Treatment for Platinum Group Metal Leaching

2017 ◽  
Vol 262 ◽  
pp. 151-154
Author(s):  
James M. Mwase ◽  
Jochen Petersen
Keyword(s):  
Platinum Group Metal ◽  
Metal Leaching ◽  
Flotation Concentrate ◽  
Group Metal ◽  
Two Samples ◽  
Heap Bioleaching ◽  
Mineral Recovery ◽  
The Difference ◽  
Pre Treatment

Two samples, a Platreef flotation concentrate and coarse ore (<6 mm), were column bioleached at 65°C using a culture dominated by Metallosphaera hakonensis. Based on solution assays, extractions in excess of 90% Cu and Ni were achieved from the flotation concentrate, while from the coarse ore 96% Cu and 67% Ni extractions were achieved. The difference in extraction levels and leaching patterns despite identical conditions used for both samples is discussed, as is the performance of the samples during a follow-up leach step using cyanide to extract the PGMs in a separate column leach experiment. While the recovery of Pd and Au was excellent during these steps, recovery of Pt was limited to 35% after 45 days for the concentrate and 56% after 60 days for the whole ore material, primarily due to the presence of a refractory Pt mineral. Recovery from a concentrate without pre-treatment was substantially lower.

scholarly journals Gold recovery from flotation concentrate from gold mine tailings using dissolve smelting

2019 ◽  
Vol 55 (3) ◽  
pp. 343-349
Author(s):  
U. Erdenebold ◽  
C.-M. Sung ◽  
J.-P. Wang
Keyword(s):  
Molten Slag ◽  
Precious Metals ◽  
Copper Slag ◽  
Molten Copper ◽  
Flotation Concentrate ◽  
Sulphide Minerals ◽  
Oxidative Roasting ◽  
Valuable Metals ◽  
Gold Mine Tailings ◽  
High Concentrations

Gold flotation concentrate may contain relatively high concentrations of valuable metals such as iron, copper, and zinc, and occasionally, even precious metals such as gold. The major components of the concentrate are SiO2, Fe2O3, and Al2O3, but it also contains reactive sulphide minerals such as pyrite. The sulphides in the flotation concentrate are fully converted into an oxide form during oxidative roasting, therefore, the chemical composition of the roasted concentrate is considered to be a copper slag. High temperature smelting with additives to dissolve Au from the gold concentrate into a molten copper was used in the research. Gold greatly dissolved at 1600?? under a CaO/SiO2 ratio of 1.25, suggesting the increase in the dissolution of gold into molten copper with decreasing viscosity of the molten slag-like concentrate at high temperatures.

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

2020 ◽  
Vol 989 ◽  
pp. 537-542
Author(s):  
Kirill A. Karimov ◽  
Aleksei V. Kritskii ◽  
Sergey E. Polygalov
Keyword(s):  
Low Temperature ◽  
Filter Cake ◽  
Elemental Sulfur ◽  
Lead Sulfate ◽  
Raw Material ◽  
Granulation Process ◽  
Flotation Concentrate ◽  
Autoclave Leaching ◽  
Sulfate Lead ◽  
Concentrate Yield

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 %

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