Table of Contents
The recovery of copper from low-grade oxidized and mixed oxide-sulfide copper ores by acid leaching is standard practice in the Southwest. Some ores, however, are unamenable to this method of treatment owing to excessive acid consumption by calcite or other acid-consuming gangue minerals. Other ores present a handling problem because of the presence of clay or other clime forming gangue minerals. The Tucson Metallurgy Research Laboratory has investigated for the past several years the applicability of the copper segregation process to the treatment of different types and grades of oxidized and partly oxidized copper ores.
This paper briefly outlines the history of the segregation process, discusses the principal reactions involved, and summarizes the results of the bench and pilot-plant research, with the objective of delineating the process conditions for treating different types of copper ore. A more detailed discussion of the process and the results of research completed to date are given in a recent Bureau of Mines report.
The copper segregation process, which was discovered in 1923, comprises heating copper ore with a reducing agent and sodium chloride to about 750° C., followed by flotation of the metallic copper produced. By 1931 the process was sufficiently developed so that two plants were constructed in Africa for treating oxidized ores. One plant having a daily capacity of 50 tons, was constructed in Southern Rhodesia by the Minerals Separation Co. At this plant minus-10-mesh ore was heated and segregated in a seven-hearth, mechanically rabbled furnace. The first five hearths were direct-fired to heat the ore to a reaction temperature of about 700° C. Salt and coal were added to the hot ore and segregation was carried out on the sixth and seventh hearths, which were heated by the exothermic nature of the reaction and by indirect heating with waste gas from the upper hearths. The segregated calcine subsequently was cooled in a rotary cooler, screened, and floated to recover the copper. The small plant operated for 4 months in 1931. About 3,500 tons of ore assaying 5 percent Cu was processed, and recovery of copper was 87 percent.
The other plant was built in 1931 at Katanga in the Belgian Congo by the Union Miniere du Haut Katanga. The plant had a rated capacity of about 350 tons of ore per day. A direct-fired rotary kiln was used to preheat the ore before segregation in a second rotary-reactor kiln. The exothermic heat of reaction maintained the temperature of the reactor kiln between 700° and 750° C. The calcine was cooled in a rotary cooler prior to recovery of the segregated copper by flotation. Operating results in the plant, which had a life of about 4 months, are not available. The low price of copper that prevailed during the depression was a factor in shutting down the two plants but mechanical difficulties also may have contributed to the cessation of activities.
Segregation of copper from oxidized ore minerals probably involves a number of complex reactions. These include (1) the decomposition of sodium chloride, (2) the volatilization of copper chloride and (3) reduction of the copper chloride to metallic copper. The principal chemical reactions presumed to occur are as follows:
Decomposition of sodium chloride. Hydrochloric acid is produced by interaction of sodium chloride with hydrous clay minerals, such as kaolinite and montmorillonite, according to the following reaction:
4NaCl + Al2O3·2SiO2·2H2O=4HCl + Na4Al2O3·2SiO3………………………………………..(1)
in the presence of traces of water vapor released in the furnace by the slow dehydration of the clay or other hydrous minerals, the reaction is rapid at temperatures of 600° C. or higher. Moisture is obligatory for the reaction to take place.
Volatilization of copper chloride. The hydrochloric acid produced by reaction 1 attacks the copper oxide minerals to form volatile cupric and cuprous chlorides. The cupric chloride decomposes into chlorine and cuprous, chloride at temperatures well below 500° C. When treating a chrysocolla ore the reaction between the hydrochloric acid and chrysocolla is as follows:
2(CuSiO3·2H2O) + 4HCl = Cu2Cl2 + 6H2O + Cl2 + 2SiO2……………………………….(2)
At the temperature of the reaction cuprous chloride has a vapor pressure of about 20 mm. and diffuses from the mineral grains. The chlorine liberated in reaction 2 also attacks the copper minerals to form more cuprous chloride as shown in reaction 3.
2(CuSiO3·2H2O) + Cl2 = Cu2Cl2 + 4H2O + 2SiO2 + O2………………………………………..(3)
Formation of metallic copper. The cuprous chloride vapor released by reactions 2 and 3 is reduced directly to metallic copper upon coming in contact with the hot carbon particles. Although a number of side reactions occur, the principal reaction taking place is as follows:
2Cu2Cl2 + C + 2H2O = 4Cu + 4HCl + CO2………………………………………..(4)
More detailed discussions of the reactions involved in segregation are presented in earlier reports.
During segregation, a negligible quantity of metallic copper may be reduced in place as micron-size particles, probably by the small quantity of carbon monoxide and hydrogen that is formed by interaction of the hot carbon and steam. Such metallic copper is generally unrecoverable because it cannot be liberated and froth floated.
A common side reaction that occurs during segregation is the absorption of cuprous chloride by the gangue or by the siliceous skeletons formed from decomposition of the chrysocolla. Enough cuprous chloride is often absorbed by the gangue particles to render them floatable with xanthate after fine grinding.
The Scope and Results of the Investigation
The initial phase of the investigation was confined to bench-scale research designed to evaluate the segregation method as applied to different grades and types of copper ore. A total of 23 domestic and foreign copper ores that contained 0.8 to 5 percent Cu was tested. The samples were either wholly or partly oxidized ores containing different proportions of chrysocolla, malachite, azurite, cuprite, brochantite, melaconite, chalcopyrite, chalcocite, bornite, and covellite associated with various siliceous, calcareous, and ferruginous gangue minerals.
The laboratory research was sufficiently promising to warrant further evaluation of the segregation-process by continuous tests in a small pilot plant having a capacity of 65 to 70 pounds of ore per hour. The ore, salt, and coke or coal were premixed for segregation in an indirect-fired rotary kiln. The rotary kiln was 6 inches in diameter by 14 feet long and consisted of a type-316 stainless steel reactor tube, 6 feet long, enclosed in a natural-gas-fired furnace. A 2-foot length of mild steel tube was attached to the reactor tube at the feed end and a 6-foot length of mild steel tube was bolted to the discharge end for indirectly cooling the segregated calcine by an external water spray. The feed end of the kiln was sealed by a choke feed screw and the discharge end by a ring seal and vented hood open at the bottom. The lower part of the vent hood was immersed below the pulp level in a drag classifier, allowing the partly cooled calcine to drop directly into the pool without coming into contact with the atmosphere. The quenched calcines were wet-ground to pass 100-mesh in a ball mill in closed circuit with the classifier. The ground pulp, after conditioning with amyl xahthate, was floated in a 5-cell rougher flotation machine. The rougher, froth was cleaned once. The contact time in the roughing circuit was 28 minutes.
Several grades and types of oxidized ore were successfully processed in the test plant and enough operating and engineering data were obtained to permit making tentative estimates of the direct costs of operating a commercial-size plant.
The results of the laboratory and pilot plant tests demonstrated that different types and grades of ore could be effectively segregated and the metallic copper recovered by flotation. The completely oxidized ores were processed to recover 73 to 96 percent of the copper in rougher flotation concentrates assaying 10 to 62 percent Cu.. The mixed oxide-sulfide ores responded equally well to treatment. From 85 to 92 percent of the copper was recovered in concentrates assaying 12 to 27 percent Cu.
The tests made on the oxidized ores showed that those containing 1 percent Cu or less gave low copper recovery in low grade concentrates. The relatively poor results obtained on such ores can be attributed to the absorption of copper-by gangue. Part of the copper so absorbed was lost in the flotation tailing and the remainder activated some of the gangue to xanthate flotation. The copper, in low-grade calcareous-type ores proved somewhat difficult to segregate. Apparently the calcite in the ore reacts with some of the salt, hydrochloric acid, and chlorine to form calcium chloride, which is not an effective copper chloridizing agent. Tests using calcium chloride instead of sodium chloride confirmed this hypothesis.
The quantity and ratios of salt and coke required for good segregation of the copper differed for each ore. From 1 to 1.5 percent salt and 0.5 to 1 percent coke gave the best results. Less salt gave incomplete segregation and resulted in a slightly lower copper recovery. Siliceous ores usually required less salt, and the ratio of salt to coke was not as critical as for calcareous or hematitic ores.
The free segregated metallic copper floated rapidly and completely, and the flotation tailings rarely contained any visible flake copper. Microscopic examination of tailings from the better tests failed to disclose any unreacted copper oxide minerals. The copper lost in the flotation tailings from these ores can be attributed principally to copper chloride absorbed by the gangue or by the siliceous skeletons formed from decomposition of chrysocolla. An insignificant quantity of copper occurred in the tailings as micron-size metallic inclusions in the gangue. These particles probably were copper that failed to migrate and was reduced in place.
Some gold and silver in both the oxidized and mixed ores was segregated and floated with the copper. The gold and silver recoveries ranged from 60 to 80 percent from ores assaying 0.2 to. 8.2 ounces per ton of total gold and silver. Efforts to identify free gold or silver in the concentrates were not successful. Presumably the precious metals were carried in the segregated copper. Segregation of the silver or gold from ores devoid of copper so far has proved unsuccessful. It appears that copper must be present to serve as a carrier for precious metals.
Factors Influencing Segregation and Flotation of Copper
Time and Temperature
The time and temperature employed to reduce and segregate the copper were critical process variables. Good segregation was achieved on typical siliceous ores at 600° to 850° C. Calcareous ores required a temperature of about 750° C for best segregation. Although moderately good segregation was obtained on a few calcareous samples at 700° C., erratic results were obtained on other ores, indicating that segregation at this temperature was sensitive to slight variations in the processing conditions. Oxidized ores with a ferruginous gangue were almost as sensitive to temperature variation as strictly calcareous ores.
The time required for good segregation varied with the roasting temperature. Heating a siliceous ore for 15 minutes at 700° C. gave nearly as good segregation as was obtained with a 60-minute contact time. When calcareous ore was treated, heating for at least 60 minutes was necessary for effective segregation at 700° C., whereas 15 minutes sufficed at 800° C. These and other tests demonstrated that siliceous or high-alumina ores generally are less sensitive to time and temperature than those containing appreciable calcite, limestone, dolomite, hermatite, or magnetite.
The free-moisture content of the feed had a pronounced effect on segregation, particularly at low furnacing temperatures. At 750° C. or higher the addition of moisture to ores, which usually contained about 5 percent combined water, had little influence on segregation of charges treated batchwise, but at 700° C. about 1.5 to 3 percent of added water was desirable for good segregation.
The influence of moisture on segregation was more pronounced for charges segregated continuously than for those treated in batches. Continuous segregation of dry charges or feeds low in free moisture gave inferior results even at a reaction temperature of 750° C. The divergent results obtained on dry or low-moisture charges in the batch and continuous tests may be attributed to a difference in composition of the reaction gas in the furnaces or to a catalytic action of moisture on segregation. In general, a contact time of 40 minutes was desirable when dry charges were segregated, whereas 20 minutes generally sufficed for moist charges.
Coke Particle Size
Segregation was noticeably affected by the upper limiting size of the coke used, particularly on calcareous ore. Coke crushed to pass 20- or 48-mesh gave good segregation, but coarser or finer coke gave poorer segregation. In tests with minus-100-mesh coke the conditions in the retorts probably were too strongly reducing, and some copper was reduced in place within the ore particles. Coke coarser than 20-mesh, on the other hand, failed to provide enough nuclei for rapid and efficient reduction of the volatilized copper chloride, thereby causing incomplete segregation. Apparently the copper chloride must be reduced simultaneously with its migration from the copper mineral to release the chlorine or hydrogen chloride for additional copper volatilization.
Type of Reducing Agent
Tests were made to compare the merits of different kinds of reducing agents. Samples of petroleum coke, two bituminous coals, mesquite charcoal and a char made from bituminous coal were compared. The results obtained on several ores indicated that satisfactory segregation could be obtained with any of these reductants. Based upon the quantity of coal used and the fixed carbon content, the two coals were more effective as reducing agents than coke, char, or charcoal.
Size of the Furnace Feed
A number of tests were made to define the effect of ore feed size on the segregation of copper. In general, segregation improved with fineness of the feed. In typical batch tests on a calcareous ore, for example, 75 percent of the copper was recovered when using a 3-mesh feed, as compared to 86 percent from a 65-mesh feed. Crushing the ore much finer than 10-mesh did not significantly improve segregation as reflected in the amount of copper recovered by flotation. The data obtained in the study of the influence of feed size also indicated that a relationship exists between the size of the ore feed and the retention time at reaction temperature. The coarser ore feeds required longer retention times than fine ore feeds.
Flotation Feed Size
Under ideal segregation conditions all of the copper migrates from the copper minerals and reports in the calcine as friable nodules about the size and shape of kidney beans and consisting of free flakes of finely divided metallic copper with some coke and gangue particles. Several methods for recovering copper from the calcines were investigated: One was to screen the calcine on a 6-mesh sieve to remove the nodules, followed by flotation of the screen undersize to recover the remaining copper; a second was direct flotation of the minus-10-mesh calcines to recover the flake copper; and a third procedure was to grind the calcine to different limiting sieve sizes before flotation.
A high recovery of copper was obtained only by grinding the segregated product to pass 100-mesh for flotation. Recoveries of copper were lower when coarser feeds were floated. Removing the copper nodules by screening, followed by flotation of the undersize after grinding to minus-100-mesh, gave about the same recovery as fine grinding and flotation of the entire charge. The copper nodules removed by screening assayed about 30 percent Cu and accounted for about 50 percent of the copper in the ore.
A small continuous two-stage segregation plant consisting of a rotary kiln for preheating oxidized and mixed oxide-sulfide ores to about 750° C. and a second fire-brick-lined rotary reactor kiln for segregation is now being tested at Tucson. The hot ore is transferred from the pre-heat kiln to the reactor by means of a screw feeder enclosed in an insulated transfer box. Salt and coke for segregation are introduced into the reactor through a second screw feeder. The exothermic heat of reaction accompanying the volatilization and reduction of the copper minerals, with a small amount of heat supplied by an auxiliary burner, maintains the temperature of the charge in the reactor at the segregation temperature.
Preliminary tests in the plant have been successful. The two-stage procedure appears to have considerable merit for treating mixed oxide-sulfide ores containing pyrite.
The sulfides are oxidized during the preheating step, which facilitates segregation of the copper and production of a high-grade flotation concentrate substantially free of pyrite.
Tests of an exploratory nature also are in progress at the Salt Lake City Metallurgy Research Center to investigate the feasibility of segregating ore in a fluidized-bed reactor. Tests in an indirectly heated 4-inch unit, using an inert gas for fluidization, have been encouraging, but much research remains to be done to fully evaluate the method. Results of segregation tests in the continuous two-stage plant and fluidized bed reactor will be presented in a later publication.
Estimated Operating Costs
Based upon the data obtained from the laboratory and pilot plant tests, the direct operating cost of a 1,000-ton-per-day furnacing and flotation plant for treating oxidized or mixed oxide-sulfide copper ores would be about $3 per ton when coke was used as the reducing agent or $2.85 when coal was used. Heat required for furnacing is estimated to be 2,500,000 B.t.u. per ton of ore. At 27 cents per thousand cubic feet of natural gas the cost would be about 70 cents per ton. The estimated cost of furnace maintenance is 40 cents per ton; power and labor are estimated at 15 cents, and salt at 12 cents. The coke or coal costs are estimated at 25 and 10 cents, respectively. Crushing, grinding, and flotation are estimated at $1.35 per ton.