<Original Papers>A Research on the Treatment of Complex Sulphide Ores. (IV) : Differential Flotation of Sulphatization Residue

The selective sulphatization-flotation process for the treatment of complex sulphide ores containing Pb, Zn, Cu, Fe, etc., was developed in the institute. The process is an application of the sulphatization process after Gromov and Derkachev. The bulk sulphide concentrate is ground and treated with selective sulphatization. In this treatment, galena and sphalerite are converted into sulphates, but chalcopyrite and pyrite remain practically unaffected. Zinc sulphate is extracted with water from the resultant product. Chalcopyrite, pyrite and lead sulphate are recovered from the residue with differential flotation. The advantage of this process from the standpoint of flotation may be said to be as follows : (a) The liberation of minerals will be promoted chemically, (b) the surface of mineral grains will be cleaned, and (c) the separation of minerals will be easy, because sphalerite is dissolved and galena is converted into lead sulphate. The tests on differential flotation of the sulphatization residue of the bulk sulphide concentrate from the Hanaoka mine proved the applicability of this new process. The samples of sulphatization residue were analyzed 4.3-5.7% of Cu, 14.5-20.0% of Fe, 6.2-13.9% of Pb, and 2.1-3.2% of Zn, and were 100% of -150 mesh and 59.3% of -200 mesh in size. The main problem in this process was the floatability of lead sulphate produced by the sulphatization treatment and the degree of regrinding necessary for the liberation of chalcopyrite and pyrite. From the fundamental flotation tests of the chemically pure lead sulphate it was verified that the lead sulphate was depressed at pH of below 5.8 or above 6.2 when Aerofloat No.25 was used as a collector and that the lead sulphate was floated at pH of 8 to 9 when it was sulphidized with the predetermined quantity of sodium sulphide and potassium amyl xanthate was used as a collector. The sulphatization temperature suitable for the separation of chalcopyrite and pyrite from lead sulphate from the standpoint of flotation was found to be 160℃ or 170℃ in the temperature range of 130℃ to 170℃. In the separation of chalcopyrite and pyrite from lead sulphate, the recoveries were 92.9% for Cu and 96.2% for Fe and the rejection was 83.5% for Pb with 0.2 lb/t of white camphor oil and 0.25 lb/t of Aerofloat No.25 and at the pH of 6. In the recovery of lead sulphate, the lead concentrate 36.7% pb was obtained with the pb recovery of 89.8% with 1.2 lb/t of white camphor oil and 2,0 lb/t of potassium amyl xanthate and at the pH of 9.2. In the separation of chalcopyrite from pyrite, three stages of regrinding, one stage of roughing and two stages of scavenging were required to obtain the copper concentrate of 9.9% Cu with the Cu recovery of 95.2% and the pyrite concentrate of 41.4% Fe with the Fe recovery of 61.4% with 1.4 lb/t of white camphor oil, 1.6 lb/t of Aerofloat No.25 and 120 lb/t of slaked lime and at the pH of 10.6. The size of the flotation products was very minute when microscopically examined, namely, 72.5% of -20 micron for the copper concentrate, 71.8% of -20 micron for the lead concentrate, 40.2% of -20 micron for the pyrite concentrate and 83.3% of +20 micron for the tailing. The highest grade of concentrates was relatively low, namely, 49.2% of Pb for the lead concentrate, 11.7% of Cu for the copper concentrate, and 41.4% of Fe for the pyrite concentrate. It was, however, considered that the higher grade could not be obtained, because the degree of grinding of the sample was near the limit in flotation. The combined recoveries in the selective sulphatization and differential flotation was 97% for zinc, 90% for copper, 77% for lead and 58% for iron