Flotation Tests of Complex Sulphide Ore

The complex ore from Osarizawa mine, Taihei Mining Co. was used as the sample. It consists mainly of galena, zincblende, chalcopyrite, pyrite, quartz and barite, and was analysed 1.21% of Cu, 3.27% of Fe, 8.62% of Pb and 12.39% of Zn. The degree of liberation of the constituent minerals in the sample ground 80% of -150 mesh in size was calculated by the theoretical equation after Prof. A. M. Gaudin to be 79.5~89.6% for galena, 87.2~93.7% for zincblende, 65.9~81.9% for chalcopyrite and 31.8~56.3% for pyrite. The quantity of collector to be added in the pulp was determined by measuring the surface tension variation of the pulp before and after the adsorption of the collector on the surface of the ore sample in the pulp. The surface tension of the pulp was measured by a stalagmometer after filtering the solid phase in the pulp. As an example of the collector, Aerofloat No.25 was selected, and it was found that the amount of Aerofloat No.25 adsorbed on the ore would increase gradually and attain to the saturation value of 120~130 g/t as the amount of addition to the pulp increased. The amount of addition corresponding to the saturation of adsorption was found to be 250~300 g/t of ore. As the floatability of zincblende was greater than that of chalcopyrite, the order of collection of minerals in the differential flotations was found to be galena, zincblende, chalcopyrite and pyrite. The reagents used were K isopropyl xanthate, K ethyl xanthate or Aerofloat No. 25 as collector, pine oil or white camphor oil as frother and NaCN, ZnSO_4,CuSO_4,Na_2S, K_4Fe(CN)_6,Na_2CO_3 and Ca(OH)_2 as modifiers. The grade and recovery of concentrates were respectively 59.06% and 77.14% Pb for lead concentrate, 43.35% and 88.01% Zn for zinc concentrate, 8.59% and 72.00% Cu for copper concentrate and 13.01% and 14.23% Fe for iron concentrate. Because of the low degree of liberation and of low floatability of pyrite, both the grade and recovery of iron in the iron concentrate were very low. The selective sulphatization-flotation process developed in the institute was tried for the treatment of the sample. It was found that the lixivation of zinc was over 90% when sulphuric acid 1.5 times that of theoretical was added. The sulphatization residue was treated by differential flotation, with Ca(OH)_2 and Na_2CO_3 as modifiers, Aerofloat No.25 as collector and white comphor oil as frother. The grade and recovery of the concentrates were respectively 26.87% and 82.85% Pb for lead concentrate, 14.40% and 82.34% Cu for copper concentrate, 9.43% and 21.51% Fe for iron concentrate. The lixivation of zinc was 89.87%. It was verified that there was almost no difference in the results obtained by differential flotation and selective sulphatization-flotation for the present sample of ore. For the representation of the metallurgical results obtained in the differential flotation in a graphical form, the block diagram was devised by the authors. In the block diagram the ordinate indicates the ratio of enrichment expressed in terms of the grade of product divided by that of feed, the abscissa indicates the weight per cent of product and the area of the block indicates the recovery of any constituent of the product