Manganese extraction by reduction–acid leaching from low-grade manganese oxide ores using CaS as reductant

2015 ◽  
Vol 25 (5) ◽  
pp. 1677-1684 ◽  
Author(s):  
Chang-xin LI ◽  
Hong ZHONG ◽  
Shuai WANG ◽  
Jian-rong XUE ◽  
Fang-fang WU ◽  
...  
2013 ◽  
Vol 49 (1) ◽  
pp. 97-106 ◽  
Author(s):  
D. Hariprasad ◽  
M. Mohapatra ◽  
S. Anand

Low and medium grade land as well sea based manganese ores were used for manganese extraction in H2SO4 - NH3NH2HSO4 (hydrazine sulphate) medium For land based Mn ores, only Mn recovery is important but for sea nodules which contain substantial amounts Co, Ni, and Cu, their recovery is equally important. In the present studies four samples used were: Indian ocean manganese nodules, medium and low grade Mn ores of Gujarat, and low grade Mn ore of Orissa, India. The Mn content of these ores varied from 15 to 39%. The objective of this work is to establish a reductant which can be used for leaching Mn from all types of ores. The optimum conditions established for nodules by varying parameters such as time, temperature, pulp density, H2SO4 and NH3NH2HSO4 concentrations were: pulp density 10%, time 0.5h, temperature 110?C, NH3NH2HSO4 3.25 g/10g, H2SO4 2.0% (v/v) for 96.9% Mn, 85.25% Cu, 92.58% Ni and 76.5% Co extractions. More than 92% Mn could be leached from different types of ores by varying amount of reductant and acid concentration at 35?C. Depending on Mn content 1.0 to 1.2 times stochiometric amount of reductant and 1.5 to 1.8 times sulphuric acid were required for >92% Mn extraction.


2010 ◽  
Vol 49 (3) ◽  
pp. 219-226 ◽  
Author(s):  
F. R. Carrillo-Pedroza ◽  
M. A. Sánchez-Castillo ◽  
M. J. Soria-Aguilar ◽  
A. Martínez-Luévanos ◽  
E. C. Gutiérrez

2014 ◽  
pp. 105-109 ◽  
Author(s):  
Alexander G. Suss ◽  
Alexander A. Damaskin ◽  
Alexander S. Senyuta ◽  
Andrey V. Panov ◽  
Andrey A. Smirnov

Author(s):  
Yuanbo Zhang ◽  
Daoxian Duan ◽  
Zhixiong You ◽  
Guanghui Li ◽  
Xiaohui Fan ◽  
...  

2020 ◽  
Vol 989 ◽  
pp. 559-563
Author(s):  
Ashimkhan T. Kanayev ◽  
Khussain Valiyev ◽  
Aleksandr Bulaev

The goal of the present work was to perform bioleaching of uranium from low grade ore from Vostok deposit (Republic of Kazakhstan), which was previously subjected to long-term acid leaching. The ore initially contained from 0.15 to 0.20% of uranium in the form of uraninite, but ore samples used in the study contained about 0.05% of uranium, as it was exhausted during acid leaching, and uranium was partially leached. Representative samples of ore were processed in 1 m columns, leach solutions containing 5, 10, 20 g/L of sulfuric acid and bacterial cells (about 104) were percolated through the ore. Leaching was performed at ambient temperature for 70 days. In one of the percolators, the leaching was performed with leaching solution containing 10 g/L of H2SO4, cells of A. ferrooxidans, and 0.5 g/L of formaldehyde. Leaching with the solution containing 5, 10, and 20 g/L of sulfuric acid made it possible to extract 50, 53, and 58% of uranium. Addition of formaldehyde in leach solution led to the decrease in uranium extraction extent down to 37%. Thus, the results of the present work demonstrated that uranium ore exhausted during long-term acid leaching may be successfully subjected to bioleaching, that allows extracting residual quantities of uranium. Leaching rate of uranium from exhausted ore depended on both sulfuric acid concentration and microbial activity of bacteria isolated from acid mine drainage, formed on uranium deposit. In the same time, acid mine drainage may be used as a source of inoculate, to start bioleaching process.


Processes ◽  
2019 ◽  
Vol 7 (6) ◽  
pp. 376 ◽  
Author(s):  
Qian Zhang ◽  
Qicheng Feng ◽  
Shuming Wen ◽  
Chuanfa Cui ◽  
Junbo Liu

In this work, oxidizing roasting was combined with leaching to separate copper, lead, and zinc from a concentrate obtained by bulk flotation of a low-grade ore sourced from the Jiama mining area of Tibet. The flotation concentrate contained 7.79% Cu, 22.00% Pb, 4.81% Zn, 8.24% S, and 12.15% CaO; copper sulfide accounted for 76.97% of the copper, lead sulfide for 25.55% of the lead, and zinc sulfide for 67.66% of the zinc. After oxidizing roasting of the flotation concentrate, the S content in the roasting slag decreased to 0.22%, indicating that most sulfide in the concentrate was transformed to oxide, which was beneficial to leaching. The calcine was subjected to sulfuric acid leaching for separation of copper, lead, and zinc; i.e., copper and zinc were leached, and lead was retained in the residue. The optimum parameters of the leaching process were: a leaching temperature of 55 °C; sulfuric acid added at 828 kg/t calcine; a liquid:solid ratio of 3:1; and a leaching time of 1.5 h. Under these conditions, the extents of leaching of copper and zinc were 87.43% and 64.38%, respectively. Copper and zinc in the leaching solution could be further separated by electrowinning. The effects of leaching parameters on the extents of leaching of copper and zinc were further revealed by X-ray diffraction and scanning electron microscopy analysis.


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