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By John A. Patterson

There are certain data handling techniques used by statisticans that can help one to calculate the size of an orebody and to develop some of the relationships between size, cutoff grade, average grade, mining cost, and profit. Knowing the relationships amongst these factors can help determine the most desirable operating conditions. A statistical technique that we will use is data grouping. This allows us to work with relatively few groups instead of many samples, If we were able to segregate the ore in an orebody into grade groups, add the amounts in each group, and plot the sums as a curve, we would get something like Figure 1. A smooth curve develops when the grade classes are small enough, otherwise a "stair step? histogram results. The area under the curve, being the sum of the tonnages in all of the grade classes, is proportional to the total number of tons in the orebody. If we take the data of a distribution such as Figure 1 and add the tonnages above a specific grade we have in effect a cutoff grade

Jan 1, 1958

By J. B. Fletcher

Miami Copper Company has used conveyor belts underground since 1947, At that time, the draw level for a portion of the orebody was only 25 feet above the haulage level To get the main haulage drifts as far away as practical from under the stoping area, 301, belt conveyors were used for lateral transfer, If the haulage drift was kept in good ground., a satisfactory draw of these stopes could be effected in that repairs to the conveyor drifts could be accomplished while drawing ore from the stopes. Some 7,400,000 tons of ore were mined with the 301d belt conveyor at a belt cost of approximately $0,004 per ton of ore. When the mining plan for a new section of the mine was laid. out in 1953, it was decided to use belts to collect the ore from the higher draw levels, and. move the ore laterally to centrally located concrete ore passes for transfer to the main haulage level (See Cut #1), The final plans used belt conveyors too collect from 3 stopes: on the 570 level, 2 stopes on the 720 level, and one stope on the 900 level. The six stopes to be mined in this manner would produce about 8 million tons of ore, and the sequence of mining was such that not more than four 400 feet conveyors would be in use at one time, Four conveyors using 36 inch, 4 ply belting were designed by the mine engineering office and fabricated at the mine shop, so that no piece was longer than 10 feet, and. the heaviest piece was the 25 horsepower motor, The head drive section, and the tail-pulley-takeup section were of bolted construction so they could be broken down for easing moving, Due to the size of the mine openings, the belt had to be transported into the mine in 400 feet rolls, requiring two splices for an installed belt. A torque arm speed reducer was used with a V -belt drive which made it possible to change the speed of the conveyor by changing the diameter of the V-belt pulley (See Cut #2).

Jan 1, 1958

By R. J. Grace

Carbon in the form of coke is used as a reactant in various chemical and metallurgical processes, The high cost of by-product coke along with its periodic lack of availability and high investment necessary for additional by-product recovery facilities indicate the need for a low-cost production unit based on a new procedure for coke manufacture, Traveling-grate coking-stokers were developed by Shawinigan Chemicals Ltd. to produce coke required for carbide manufacture, They were adapted to kilns so that the hot volatile gases could be used for lime burning, Their operation has been described in various patent disclosures (1, 2, 3), and their application to rotary lime burning kilns has been described by Andersen (4). Details of the process and plants have also been published by Kurt Baum (5,6).

Jan 1, 1958

By John C. Wangaard

A method of raising by use of a cage has been used with considerable success at the Cary Mine, Hurley, Wisconsin, on the Gogebic Range. The Cary Mine is owned by the Odanah Iron Company, Pickands, Mather and Company, Managers. This method is known to have been used previously at Tennessee Copper and probably in other districts. The Cary Mine was engaged in deepening the main shaft from the 34th to the 37th level. To accomplish this, and to facilitate development, a ventilation shaft 930 feet north of the main shaft was sunk from the 34th to the 37th level and crosscuts were driven on the 35th, 36th and 37th levels to the main shaft location, These levels are at 200 foot verticalintervals. All of this work was done in granite and the majority of the workings were not timbered. The next step was to get a raise through from the 37th to the 36th level, and from the 36th to the 35th level, which would later be stripped to full shaft size. Several methods of raising were considered including double-compartment cribbed and two single raises interconnected at regular intervals, but the cage method was chosen as it appeared to be a safe and efficient operation. The method in general consists in the use of a drill hole between levels through which to pass a hoisting rope to raise and lower a light cage.

Jan 1, 1958

By L. E. Mylting

Modern ash handling systems exist in considerable numbers and a great deal of design and operating experience has been accumulated, It is the endeavor of this paper to point to a number of frequently recurring problems, A general appreciation of these problems will be helpful in both design and operation of systems. The light fly ash and the heavy bottom ash present different problems, and each will therefore be treated separately. Fly ash is collected in the ducting from the furnace to the stack and beneath the mechanical and electrostatic dust collectors. The greatest amount of dust is collected beneath the mechanicals and electrostatics. Hoppers are generally provided as integral parts of the collector. Suitable valves and piping are connected to these hoppers for removal of the dust. The size and shape of these hoppers are important factors in determining frequency of handling and the capacity of the system.

Jan 1, 1958

By George K. Martin

This paper deals primarily with roof bolting in Kentucky coal mines from January 1, 1949, through December 31, 1954, a six year period. The Consolidation Coal Co? (Ky.), a division of the Pittsburgh Consolidation Coal Co., began systemmatic roof bolting in November 1948 in Mine #204, Elkhorn #3 Seam, at Jenkins, Kentucky, After the efficiency of roof bolting had been proven by the Consolidation Coal Co., this procedure was adopted by other companies throughout the state. In December 1954, approximately 46.4 percent of Kentucky's total production of coal from underground mines was obtained from roof-bolted areas. Statistics in this report will show the progress made in roof bolting during the six years, in addition to the roof bolting being done as of December 31, 1955. Significant progress has been made in the coal mines of Kentucky in reducing the roof-fall fatality, rates, during the past several years by adopting systemmatic roof bolting. The ground work by the Consolidation Coal Co, (Ky.) in cooperation with the Federal Bureau of Mines was based upon the previous experience of the St. Joseph Lead. Cc, in its Southeastern Missouri Mines and The Tennessee Coal and Iron Division of the United States Steel Corporation. The author gratefully acknowledges the cooperation and assistance given by officials of the various mining firms using roof bolts, the Kentucky Department of Mines and Minerals, and the Federal Bureau of Mines in furnishing photographs and information.

Jan 1, 1958

By E. O. Dearden

In the ten years since its discovery the Lac Tio ilmenite-hematite deposit has become one of the world's largest sources of titanium ore. The deposit lies 25 miles north Havre St, Pierre, a fishing village on the north shore of the St. Lawrence River some 400 miles northeast of Quebec City. In 1941, J, A, Retty made a reconnaissance survey of the region for the Quebec Department of Mines. His results, published in 1942 and 1944, described a number of small ilmenite occurrences on the shores of four lakes within the area. Several of these were staked by Montreal and Toronto interests, Two of the claim groups were later optioned by Kennco Explorations, (Canada) Ltd, and a third by the New Jersey Zinc Company, During a detailed exploration program in 1946, Kennco geologists discovered eight ilmenite deposits, one of which was Lac Tio, the largest known body of ilmenite-hematite ore in the world. Quebec Iron and Titanium Corporation owned two-thirds by Kennecott and one-third by The New Jersey Zinc Company as organized in 1948 to equip and operate the new mining property. A railroad 27 miles in length was built (completed in 1950) to transport the ore to the St. Lawrence River at Havre St. Pierre, An electric furnace treatment plant, yielding a high purity iron and slag containing 70 m 75 percent Ti02, to process the ore was built at Sorel, Quebec, 550 miles up the river from Havre St. Pierre.

Jan 1, 1958

By W. M. Aubrey

Cobbing, as used in this; paper, refers to coarse dry magnetics separation. It is applicable when, by its practice, plant cost can be reduced or the value of its products increased by the removal of a barren or lean reject or a high grade final product. This paper attempts to describe cobbing as it is now being practiced, and also the problems that are of concern to the operator and design engineer. In preparation, 16 plants were contacted, all of which were very cooperative with their time and data.

Jan 1, 1958

By D. A. Ellis

This paper review s several new applications of solvent extraction developed in conjunction with a research project carried out under the AEC Raw Materials Division by the Dow Western Division laboratories during the last two years. A process for recovery of uranium from ores by contacting the solid directly with organic solvent without a prior aqueous leach and a process for uranium recovery by contacting thick, unfiltered leach slurry with the solvent were studied in an attempt to simplify ad reduce the cost of ore treatment procedures. Processes to produce pure UF4 products at the concentrate mills thus eliminating the production and subsequent dissolution of the concentrate were investigated. Since many of the leach liquors contained vanadium as well as uranium, solvent extraction processes for by-product vanadium were also studied.

Jan 1, 1958

By P. J. M. van Heteren

The flotation system is usually distinguished from other physical chemical systems by the importance of interactions at the solid-liquid and solid-gas interfaces. In recent years, numerous quantitative studies have been made of the distribution of common collectors between the solid and aqueous solution phase. Although in flotation the main emphasis is placed on the adsorption process, chemical reactions will also take place in the bulk solution phase, These side reactions are often undesirable and will contribute towards the complex behavior of many a flotation system In sulfide flotation, the precipitation of insoluble base metal xanthates is a well-known phenomenon although the conditions under which the precipitation reaction will proceed and the magnitude of the solubility products are not known with accuracy. A knowledge of the solubility of these xanthate salts will not only be of value in flotation research, but the mill operator might be interested in knowing whether precipitation of metal xanthate can be controlled or completely prevented in the flotation circuit thereby reducing reagent consumption, For these various reasons the solubility of lead, copper, zinc, and silver ethyl xanthate was determined. The experimental techniques and results will be discussed in this paper.

Jan 1, 1958

By J. Ludden

By Sims Some two years ago we began investigating low coke button bituminous coal for use at our Bound Brook, New Jersey plant as a replacement for anthracite. Since August, 1957 the plant has been operating continously on this coal with the exception of a new 200,000#/hr, cyclone boiler. This study was motivated not from our desire to conduct combustion research, but for the sole purpose of reducing our steam costs, The indications were that the economics were greatly in favor of the use of the bituminous coal. The attempts to reduce the costs have developed some knowledge that has enabled this program to succeed reasonably well as will be discussed later by Mr. Ludden. The production of steam size anthracite is dependent upon either the reclamation of old culm banks or the breakage in the production of the domestic sizes. The anthracite industry has not as yet been able to cut, its mining costs to the point where the production of steam sizes has become a profitable business. This has resulted in two basic problems: (1) A constantly rising price in an attempt to have those sizes carry their share of the cost and, (2) Periodic shortages of supply when the demand is good but the market price prohibits the mining of anthracite to produce those sizes only, These conditions have made the long range forecast appear as though the price would go even higher and/or the supply at times be critically short.

Jan 1, 1958

By Kenneth E. Merklin

The heavy media process is now being used extensively throughout the world for the gravity concentration of a great number of different minerals. The process is a commercial adaptation of the heavy liquid procedure which has been used in laboratories for many years to separate mineral of different specific gravities. In the heavy media process, a suspension of finely divided particles in water is used to increase the apparent specific weight of the fluid to the desired separating point. The original heavy media process used sand, galena, pyrite and occasionally other materials as the solid constituent of the separating suspensions. As the process developed, the need for a more universal media meeting the following requirements was recognized: 1. High specific gravity. 2. Resistance to abrasion. 3. Resistance to corrosion, 4. Readily reduced to a suitable size consist for suspension in water. 5. Available in quantity at economical price levels. 6. Easy to recover by a simple method.

Jan 1, 1958

By Raymond Mancha

"Conventional" coal dryers really do not exist. What we really mean to say is that there are many and varied kinds of coal dryers -- and the "air handling problem" is really the only portion of it which has been approached for solution by conventional methods. It now appears that these conventional methods need supplementing by some new, relatively unconventional apparatus. In order to approach this matter with mutual understanding, here is a quick survey of the coal drying art. Figure #1 shows a McNally Pittsburgh Pulso Dryer, In the McNally Pulso Dryer, the coal is fed-onto the screen by a rotary feeder assuring even distribution of the coal. As the coal travels forward due to the action of the screen, it is subjected to a controlled pulsating flow of hot gases. T ese originate below the deck and are directed up through the moving bed of coal. Fines sifting through the deck are dried in the hopper below the screen as they fall through the hot gases to an evacuation screw. The coal remaining on the screen is dried to the desired degree when it reaches the discharge end. At this point, it is directed down into the evacuation screw to combine with the fines, or it may be discharged as a separate product. As the coal travel down the screen, it is subjected to intermittent upward currents of hot gases. These first distend the bed allowing the gases to come into intimate contact with the coal particles for rapid and efficient heat transfer. Then, as the flow of gas is cut off, the bed settles and the extreme fines sift through the screen preventing their being carried away into the exhaust system.

Jan 1, 1958

By Arvid E. Anderson

The paper describes the unique operating features of the Lebanon Concentrator Plant wherein both magnetic concentration and froth flotation are used. Chalcopyrite is selectively floated by the use of dual collectors, Aerofloats and secondary butyl xanthate. Both potassium ethyl xanthate and sodium isopropyl xanthate (dual collectors) are dissolved and emulsified with pine oil for the flotation of the pyrite. Sketches describing the plant operations are presented and plant operating metallurgical results obtained are also given. The optimum production of minerals by the froth flotation process is normally achieved [ ] smooth operating conditions, The process, because of its many variables, becomes an art as well as a science. Therefore, any phase of the operation which will aid the operator in his control of the numerous variables will help produce a smoother operating plant, One of the most important variables which the operator deals with is the addition of reagents, These are usually costly items and should be used discretely with minimum amount of waste, The danger is that ,this emphasis may reduce the reagent amount to a point below that required, or, on the other hand an excessive amount is added, Both conditions usually yield poor results. The drea of the metallurgist is to provide the operator with the tools enabling him to produce a good mineralized froth at all times. Normally the operator has no control concerning the type of reagents which he must use. This has usually been worked out for him in the laboratory and by plant-scale tests. He does have limited authority to control the amounts of the reagents at his disposal. This, I believe, is as it should be. However, the operator should be provided with the best combination of reagents for his particular job. The providing of this combination of reagents for his particular job, The providing of this combination is often difficult, yet it is very important in order to yield good results, With this in mind, the dual collector idea was introduced for the selective flotation of chalcopyrite, and also later for the flotation of pyrite at the Concentrator Plant of the Bethlehem Cornwall Corporation in Lebanon, Pennsylvania, Supplying several types of collectors has given the operator an additional effective tool, even though great variations are encountered in ore tonnage treated and in ore composition, This paper deals with the use of the dual collectors and with the addition of the reagents in the form of a pine oil emulsion, or more accurately, a pine oil suspension. It is hoped that other operators may find the ideas helpful. There does not seem, to me, to be definite scientific fact which would substantiate the necessity for the use of dual collectors. However, in actual operations, the metallurgical results at Lebanon became more uniform with the use of dual collectors, I look upon it as an "insurance policy" which greatly aids the operator during those periods when he is having trouble, One theory has been advanced, i.e. one collector has a greater tendency to float the coarser minerals, whereas the other one may float the finer ones. Consequently, a combination of the two collectors will result in optimum recoveries, There has been some evidence that this is true, but it is not conclusive, on the other hand, the use of a xanthate-pine oil suspension (or emulsion) should speed up the surface reaction of the collector with the sulfide minerals, result¬ing in more uniform flotation, This has definitely happened in our particular case. A more interesting unexplainable observation has also occurred; i.e., the xanthate solution does not appear to lose its strength as rapidly when it is in the form of an emulsion (or suspension), This was first observed when a 50 gallon drum of a 20% ethyl xanthate solution was taken to Wilmington, Delaware, for a laboratory test, using a homogenizer to produce the xanthate-pine oil emulsion. Upon arrival, the drum of xanthate had developed a considerable internal pressure, The relief of this pressure was accompanied by a strong odor, When the emulsion was returned to Lebanon for the plant test orders were issued to release the pressure prior to emptying the emulsion. Nopressure had developed within the drum and there was no unusual odor present, Since this time it has been observed that xanthate pine oil suspensions have greatly decreased the possible decomposition of the xanthate solutions. In fact, suspensions have been exposed for more than a month without apparently effecting their reactions in the flotation plant. On the other hand, we found that the regularly mixed concentrated xanthate solutions, after standing for several days, did not give the same results as the freshly made ones. This effect was greatly reduced by the use of xanthate pine oil suspension (or emulsion)

Jan 1, 1958

By Alex Stewart

Most iron bearing ores treated by the R=N Direct Reduction Process yield products lower in phosphorus than the initial ores. The sulfur is also removed and generally is in the range of 0.02 to 0.05 percent in the final R=N product, independent of the sulfur content of both the ore and carbon fuel used. The mechanism and thermodynamic justification of the principal reactions leading to the removal of both phosphorus and sulfur and the control of sulfur in the iron concentrate produced are discussed and supported by the results of laboratory experiments. In substantially all iron ores (containing up to 4% sulfur as sulfate) and solid carbon fuels (containing sulfur from 1 to 8%) tested, the rejection of sulfur has been in the order of 90-95%, with not more than 0.02 0.05% sulfur in the final metallic iron product. With respect to phosphorus-bearing iron ores ranging up to 0.70% P205, the rejection to tailings has been in the order of 70-95%, resulting in not more than 0.03 - 0.07% phosphorus in the final product.

Jan 1, 1958

By A. L. Slaughter

A mineral deposit has been discovered and explored sufficiently to establish the presence of ore in ?commercially marketable quantities," This is the point at which exploration ends and development begins, Of course, when the presence of commercially marketable quantities of ore are thought to be present the deposit had to be evaluated before this decision could be reached. If, on the other hand, it was thought that commercially marketable quantities of ore are not present, either more exploration has to be done or the project must be abandoned. It is this decision with which this paper is concerned, more particularly, the influencing factors behind this decision. The final question to be answered is: what profit can my company reasonably expect to make by developing this property and putting it into production? The answer must be expressed as a figure for final profit per ton, money which can be used to pay dividends or to build up cash reserves. Some acquisition cost has probably been paid and the exploration cost incurred at the time of decision, (abandonment or development). In addition to returning these acquisition and exploration costs, the deposit must pay development, mining, milling, capital expense of new plants and equipment, taxes, and various other related costs.

Jan 1, 1958

By G. P. Richards

RIP is the process of recovering uranium from an unclarified slime pulp which comes in contact with a moving bed of ion exchange resin, T e sand-slime separation of the acid-leach slurry is accomplished through the use of cyclones and spiral, akins type classifiers. Course sands are discarded to tailings and the combined slimes and leach liquor become the IX feed. Uranium in the IX feed is absorbed on the resin and after exhaustion the barren pulp is directed to tails, Following the uranium exhaustion cycle an eluant solution passes through the resin bed stripping the uranium from the beads, The uranium is precipitated from the clarified pregnant eluant, filtered, dried and packaged for shipment to the AEC, Normally, the ore received at the Mines Development plant is a loosely cemented sandstone containing small quantities of clay and bentonite, The ore is easily crushed and requires only a 1/3 load of rods in the rod mill to reduce the ore to grain size, Very little of the ore is recirculated by the 42" spiral classifier.

Jan 1, 1958

By E. J. Duggan

In the course of the grinding experiments conducted at Climax over the past twenty years one interesting phenomenon was discovered. This was that balls made of certain types of steel grind more ore than some others. The most significant variation was found in the results secured with a hard alloyed forged bail compared with the forged carbon ball supplied by most ball manufacturers These findings were confirmed in many repeated controlled tests. The value of these tests depends on the methods used, the facilities available and the care with which the investigations were conducted. These are described in the following. This testing has largely been done in eight 91 x 81 low discharge ball mills in closed circuit with 7811 duplex screw classifiers. The mills run at 20 RPM on a feed approximately minus 3/8 inch. The ore is ground through 28 - 35 mesh in one stage. Each mill and auxiliary classifier is in a separate, parallel grinding and flotation section The, ore to each: section is the same in every particular and is fed to each mill bin from a common conveyor tripper. The ore from the bin to the ball mill is weighed on a conveyor scale. The power input to each mill motor is determined with a watt hour meter which measures the power to the mill motor only. Both the scales and meters are checked periodically by standard methods and have been found to give consistent readings. Automatic samplers are employed on each section to obtain accurate measurement of the size reduction. A size analysis of the classifier overflow is made each shift and a composite size analysis is obtained from a weighted average of the various shift analyses. The mills are driven by 450 horsepower synchronous motors so that there is no speed variation and shell liner age is kept as nearly the same as possible in mills under test. The same set of operators is assigned to a mill under test as to the control mill. In all other particulars such as ball load, mill density, liner design, etc., the mills are operated in precisely the same manner. This is an ideal arrangement for plant testing. From observations of ball mill performance over the past twenty years, it is the opinion of the operating staff that variations of as little as one to two percent in mill capacities may be determined if sufficient operating hours are devoted, to a test. Relations between mill capacities, fineness of grinds and power consumptions have been established with a high degree of accuracy. In Table I are given the figures for one test. These show that while the grind with the hard ball was slightly coarser, the tons ground per hour was 5.4 percent greater with the soft ball, and 6.1 percent greater on the basis of minus 100 mesh produced. The power consumed per ton with the hard ball was 4.4 percent greater on a tons per hour basis. The soft ball was inferior in wearing qualities and steel consumption per ton of ore was 24 percent greater with it. In this test the mills were charged with the two different balls for a period of three months to allow the ball loads to come to equilibrium. In Table II are given the specifications of the two balls. The difference in grinding rates may be accounted for by the nature of the ball surfaces. The surfaces of the soft balls after being worn a while are rougher to the touch. Enlarged photographs of the worn-in surfaces of hard and soft balls are shown in Figure I. These show the presence of scratches or skid marks on the hard ball surface which indicates less nipping effect. The Climax ore pulp has a pH of 6.0 and is in a dispersed state. The viscosity is low compared with most ores especially those that are conditioned with lime. This would lower the tendency of the pulp to adhere to the balls with a lower grinding efficiency as a result. It was also thought that the relative rates of wear from surface to center might affect the make-up of an equilibrium ball charge. This probably has little effect on capacity as shown by the size analyses of the two ball charges in Table III. This hypothesis was subsequently investigated by the use of rationed charges. The rationed charge did not overcome the advantage in

Jan 1, 1958

By Clark B. Risler

Conveyors constitute a critical artery, essential to the successful operation of most ore processing plants. Technological progress is permitting the recovery of metal from low grade ores which previously could not be worked economically. Significant in this progress are improved material handling techniques highlighted by the frequent and successful application of conveyors. From the standpoint of the plant manager the successful conveyors are those he never hears about. They continue to economically transport the loads applied to them, requiring only routine maintenance attention. When such conveyors are subjected to closer scrutiny you will find that they have been carefully selected and engineered to perform the required function under the anticipated conditions. Among the primary elements in a conveyor drive are the electrical motor and the speed reducer. The selection of these elements can be aided by the systematic consideration of several factors. These factors which focus attention on the critical operating conditions, will be individually evaluated and their effect upon the selection of the electrical drive system and speed reducer analyzed.

Jan 1, 1958

By C. B. Kettering

In the early operations of the Inspiration Consolidated Copper Company, which extended from 1914 to 1926, the ore as mined contained principally sulphide copper minerals, manly chalcocite (Cu2S), This ore was milled in a 20,000-ton concentrator and the copper was recovered as a flotation concentrate which, was smelted at the nearby plant of the I ternational Smelting and Refining Company. In those days no attempt was made torecovercopper from any oxide minerals present. As mining operations progressed it soon became evident that the proportion of oxide minerals present would steadily increase, and. it was realized that some method of treatment would have to be developed whereby copper could' be recovered from both oxide and. sulphide minerals contained in the ore. A research program was instituted and after some years of experimental and, pilot plant work a new hydro-metallurgical process was developed whereby it became possible to extract copper from both the oxide and. sulphide minerals in a single leaching operation. The solvent adopted consisted of a mixture of sulphuric acid and ferric sulphate,' Following dissolution, the dissolved copper would be recovered by electrolysis. Also in the process of electrolyzing, spent solutions of sulphuric acid, H2SO4, and. ferric sulphate, Fe2(SOk)3, would be regeneratede and. somade available for continuous reuse. The process appeared to be a success and the decision was made to build a full-sized leaching plant for the future treatment of I spiration mixed ore. Construction was started and the leaching plant was completed in 1926, With the start of ferric-sulphate leaching the Inspiration Concentrator was shut down and, with the exception of a few short time runs has not operated since that time. The ferric sulphate leaching process served. Inspiration well for some 30 years and during that time won a name for itself as a classic example of hydro-metallurgical treatment of mixed sulphide oxide ores. Originally the plant was designed. to process some 7,500 tons per day of a mixed ore containing about 1,2% copper, of which 60% was present as oxide and. 40% as sulphide, Down through the years various improvements were made for the purpose of increasing plant capacity and as well to change the limiting 'factor of sulphide-oxide ratio, The final result was a leaching capacity of 9,600 tons per day when treating a 1% copper ore with a reversal of the sulphide-oxide ratio to 60% sulphide and 40% oxide.

Jan 1, 1958