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By D. B. Hobbs, H. O. Burrows, T. D. Stay

ALUMINUM which has lost its original identity as to source may be considered as secondary. This would include scrap originating in the fabrication of aluminum, which is not consumed at the plant of fabrication, and all material resulting from fabrication, which has lost its original identity regardless of whether it be consumed at the plant of fabrication or not. In general, secondary aluminum may be divided into two classes, commercially pure and alloyed. The chief impurities found in commercially pure aluminum are copper, iron, titanium and silicon; while in aluminum alloys, copper, iron, titanium, silicon, zinc, manganese, magnesium, nickel, or tin, may be present in varying proportions. Occasionally lead and antimony are found in secondary aluminum alloys, especially if the alloy results from remelting vender's automotive scrap. Such alloyed elements as are found in the scrap cannot be removed on remelting. In this, the remelting of secondary aluminum differs from the smelting of secondary copper or most other secondary metals. The majority of secondary materials reach the consumer in the form of remelt ingot or remelt No. 12 ingot. This ingot varies greatly in quality, depending upon the source of the material, and the manner in which it is refined and blended. The principal differences between primary and secondary ingot are: 1. The composition of secondary ingot may be more or less indefinite. 2. Secondary ingot may contain more dross and oxide. Secondary ingot may, however, be used successfully in a great many products. RELATION OF PRIMARY PRODUCTION TO SECONDARY PRODUCTION Table 1, compiled from data published by The American Metal Market, gives an approximate idea of the relation of primary to secondary production. Information obtained from other sources leads the authors to believe that the figures for secondary production are somewhat low.

Jan 1, 1928

By F. C. Newton

RefineRs of lead by the Parkes process have always been solicitous of recovering the zinc used in the desilverization, and justly so, as the loss in zinc constitutes one of the heavy costs in this method of refining. Part of this loss is due to the absorption of zinc by lead, and in the present state of knowledge is not recoverable as metallic zinc. The remainder of the zinc, which finds its way into the precious-metal-bearing crust, is recoverable and naturally engages the operators' most exacting attention. Careful elimination of the impurities in base bullion has enabled the refiner to produce a zinc crust better adapted to retorting, which reflects favorably in the zinc recovery;- and to produce a richer crust, thereby lowering the zineking cost, as the increased efficiency of the zinc results in less of it being in circulation. It is obvious that if the ratio of concentration be increased, a lessening of the amount of crust formed will ensue, throwing less work on the retorts and diminishing the amount of retort metal sent to the cupels, thus making an actual money saving in all these operations. In the effort to obtain a rich crust, the metallurgist has increased the heat to a considerable degree, upon the assumption that at the higher temperature the zinc-dver alloy could be squeezed free from the adhering lead. A further possibility for improvement suggested itself in the freezing-point curve of Prof. H. C. Carpenter, upon zinc and silver. Study of this curve disclosed the existence of a definite chemical compound, Zn3Ag2, with a freezing point of 665" C., a promising compound for practical application. In the experiment undertaken it was endeavored to force this compound into the crust, hoping thereby to obtain a higher concentration than formerly. It was recognized that in bringing a large kettle of lead to such a heat, more fuel would be used, the life of the kettle would perhaps be shorter, and other disadvantages would be experienced. AS a commercial operation, the project would have to stand on its own feet,

Jan 1, 1915

By H. Van F. Furman

TO those engaged in the purchase and sale of gold- and silverores, mattes and bullion, the frequent differences in the gold- and silver-contents, as reported by different assayers, is a source of much annoyance and expense. The desirability of uniform methods of assay has long been recognized by certain refiners who specify in their bullion-contracts the method to be adopted in assaying the bullion. One large producer of base bullion employs at each refining-works with whom it has a bullion contract an assayer whose duty it is to see that the bullion is sampled and assayed according to contract-specifications. For the assay of gold- and silver-bullion, standard methods are now generally adopted; the reeult being that differences seldom occur. While we cannot hope for anything near perfection from fireassay methods for ores, mattes and base bullion, as carried out cornmercially, such methods may be made to yield results which answer all commercial requirements, and they will probably be retained for obvious reasons. The object of the present paper is to point out some of the sources of error and to show the importance of more uniform methods of assay and close attention to the details of the work. The sources of error in the assay of ores and mattes are: improper sampling; losses of the precious metals during fusion, scorification and cupellation; imperfect elimination of the base metals on the cupel and imperfect extraction of the silver during the parting in nitric acid. It is not the object of the present paper to enter upon the subject of sampling; but a few remarks may be of value. There has been considerable discussion recently in Colorado upon the frequent differences which occur in the assays of Cripple Creek gold-ores, and these differences have been attributed to various causes. The writer is informed by two of our most prominent smelters that since they have adopted 100-mesh screens (the general practice is 80-mesh) the

Jan 1, 1895

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Jan 1, 1938

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Jan 1, 1938

By A. A. Collins

OF widespread interest to all lead metallurgists is a dross smelting process that will consistently give mattes and speiss of low lead and high copper contents. It is a problem that has intrigued operators through the years. In the past few years, with increased amounts of copper and overcrowded metallurgical furnaces, it has been almost imperative to evolve some method. Experience at El Paso has shown that high copper-to-lead ratios can be consistently attained by use of the new soda process. † Older Methods Three older methods, i.e., 1. Dross plus siliceous ore and limestone. 2. Dross plus one fourth blast-furnace speiss and siliceous ore, 3. Dross plus sulphides and siliceous ore and scrap iron, do not give consistent results and the products are higher in lead and lower in copper than those produced with the soda process. Any radical changes in the blast-furnace drosses will throw any of these older methods into a tailspin. For almost ten years, El Paso used the third method, with varying results. It was noted that as the arsenic and anti- mony contents of the dross increased, the speiss and matte steadily increased in lead and decreased in copper. A dross containing four parts of copper to one part of arsenic was considered to be ideal. At times this balance of copper to arsenic was almost impossible to obtain, especially when handling copper Cottrell dusts, which vary so greatly in copper and arsenic. Custom smelters like El Paso, handling a great variety of dusts, concentrates and ores, are apt to have anywhere from I to 15 pct copper on charge, with a variable in arsenic of from 0.5 to 2 pct. With such conditions, and drosses varying from 12 to 40 pct copper, an operator using the older dross reverberatory practices would find it just about impossible to control the reverberatory operations and to produce a good product low in lead and high in copper. These conditions prevailed at El Paso from the year 1940 to the present. In May 1942 it was decided to experiment with the soda process. The Soda Process After considerable work the soda method was finally evolved to the present state. It has never failed to produce a good product even under the most adverse conditions. It might be well to remember that many men of thought in the smelting industry have at one time or other sugge~ted the removal of all copper, arsenic and antimony from the lead blast-furnace charge as an aid to better metallurgy.

Jan 1, 1949

By Thomas Egleston

The uncertainty of finding the exact equivalents fortechnical expressions in different languages has led me to think that a glossary of furnace-terms would be useful to members of the profession. I was led, some years ago, to prepare a small one, which has been found so very useful, that I have arranged one for presentation to the Institute, for four different 'types of furnaces, in the hope that it may prove of use to those who are in the habit of consulting foreign technical journals, I was further induced to make the attempt in the hope that it might tend to make the terminology less complicated than it now is, and also to make it more uniform over larger districts. The idea of referring to a drawing which expresses exactly what is meant, resulted from the necessity of presenting such subjects to students, and will, it is hoped, prove much more useful than the mere dictionary arrangement in alphabetical order. The drawing of the blast-furnace was prepared for me by Mr. J. P. Witherow, of Pitsburgh ; the drawing of the open-hearth furnace by Mr. S. T. Wellman, of Cleveland, and the reverberatory furnace by Mr. John Fritz, of Bethlehem. The bloomary-furnace is taken from my own paper* on Bloomaries, but, in order to have a complete furnace, I have introduced into the drawing of those used at Saranac the guard-plate used at Ironville. The interest which has been taken in the preparation of this paper by those gentlemen who have contributed the drawings, has been supplemented by gentlemen in Europe who have taken the trouble to read the proof. In France it has been read by Mr. Lodin, of the School of Mines, and by Mr. Chaper, of Paris, and in Germany by Dr. Wedding, of Berlin, and Mr. Luhrman, of Osnabruck. The preparation of the paper has been much more difficult than was anticipated, and its value has been very much enhanced by Dr. Raymond's kind assistance and cooperation; and to him, as well as to all the other gentlemen, Iam very much indebted. It was at first proposed to add to the paper the names of all the tools used in these various furnaces. It was found, however, that this would make such

Jan 1, 1888

By C. R. Boyd

Without attempting to do more than give a preliminary or skeleton report upon the geology and minerals of Southwestern Virginia at this time, I am led to hope that the great commercial importance of the sixty miles of cross-section here crudely treated of

By C. V. O. Hughes

DIATOMACEOUS earth is probably the most widely used, and certainly the least well publicized, of the important non-metallics of the United States. The very multiplicity of the names by which this whitish earthy sediment is described has served to confuse its identity, and to dissipate a realization of its full importance.

Jan 3, 1953

By Eugene Guccione

For at least 10 years, various industry people had warned about the alarming decline in the nation's oil and gas reserves. Then, in October 1973, the Organization of Petroleum Exporting Countries launched an oil embargo and sharply increased crude oil prices. Suddenly, everybody discovered the "energy crisis." In November 1973, former President Nixon advocated a Project Independence program so that ". . . by 1980 we shall be able to meet America's energy needs from America's own energy resources."

Jan 5, 1975

By Frank R. Milliken

Before the rather sudden economic downturn in recent months, shortages of materials were painfully felt throughout the US. The current recession has provided a temporary relief here and there-but energy fuels and a number of staples such as sugar are still in short supply. This is an unsolved and fundamental problem that will haunt us again as the economy turns up. There is an awareness that the nation faces a widening supply gap in many of its most basic materials-from asbestos to tungsten-as evidenced by recent studies and government reports, and by the Secretary of the Interior's annual warnings about the mounting seriousness of the problem.

Jan 3, 1975

JOHN B. HASTINGS, Los Angeles, Cal. (written discussion *).-This, paper is so admirable for its brevity that one hesitates to enter into discussion. The cycle of eruptives makes one wonder, "Where's the basalt?" and the youthfulness of the Idaho flows suggests they may yet come. Veins and Dikes.-Veins and dikes seem to be closely aligned in the minds of some authors, while to me they appear quite distinct, the first representing quiet, long-continued action, typified by metasomatic replacement, far removed from volcanic action; the latter, sudden and cyclonic, emanating from the heart of vulcanism. The rock openings occupied by both types were formed by kindred forces, but in the one they were closed for indefinite periods before invasion by the solfataric solutions, and in the other their formation and occupation were simultaneous. Open Fissures.-I do not believe in the existence of open spaces of discission that may be occupied by veins. Occasionally they occur; I found one with. a machine drill 350 ft. below surface, fortunately in an adit; the water spouted for 12 ft. in a 6-in stream for nearly 24 hr. It was only a flattish fissure, 6 in. wide, crossing the perpendicular vein, and indicating how tabular veins of ore might be formed. The practical absence of open fissures should be a joy to miners. We all know the misgivings with which we feel about for an old winze having only 30 ft. of water in it, in an unsurveyed and inaccessible part of a mine. It is said that veins are sometimes formed in open spaces left by Protuberances of the walls, which meet after having been displaced; but I believe that in such circumstances there would be sufficient attrition and brecciated material to fill the fissure, though perhaps loosely. Parts of veins may possibly have been formed in originally open spaces, but if so, the spaces were made by solution shortly preceding the deposition of the veins. That total solution of rock adjacent to a fissure may occur under favorable conditions is proved by the existence of large bodies of sulphide which contain only a few per cent. of silica; while on the same level, and in other levels above and below, all being within the same ore shoot, assays may show as much as 50 to 60 per cent. of silica, although the gold, silver and copper assays may be almost identical in all the

Jan 1, 1918

By Francisco Cadena

The construction of Norris Dam, built by the Tennessee Valley Authority on the Clinch River, a tributary of the Tennessee River, involved the production of coarse and fine aggregate for approximately 1,000,000 cu. yd. of concrete. A hammer-mill plant for the manufacture of sand from quarried rock was indicated as the most economical method of providing this material. The abrasive character of the feed as compared to the average hammer-mill feed made it necessary to control operations rather closely, and this resulted in the accumulation of many data hearing on production and maintenance. This paper will deal only with three specific sets of data, accompanied by such general information as may be necessary to explain the conditions under which they were obtained and followed by discussions on the relation of silica content to hammer wear, comparative costs of slugger and stirrup hammers, and comparative costs of tool-steel and manganese-steel grate bars. Coarse Aggregate Aggregate for the concrete poured at Norris Dam came from a dolomite quarry within 2000 ft. of the west abutment of the dam. Geologically this rock is known as a Knox dolomite. A typical chemical analysis is given in Table 1. Table 1.—Typical Chemical Analysis Per Cent H2O, 110° F............................................. 0.02 lgnition loss.............................................. 43.56 MgO.................................................... 19.19 CaO ................................................... 29.07 Fe²O³.................................................... 0.39 Al2O³ .................................................. 1.94 SiO2..................................................... 5.75 SO2...................................................... Tr. Rock was carried by 12-yd, dump trucks to a 42-in. gyratory crusher set at 6 1/2in. The product of the primary crusher was taken by a belt conveyor to a double.-deck vibrating screen followed by a 5 1/2-ft. cone

Jan 1, 1938

By Francisco Cadena

The construction of Norris Dam, built by the Tennessee Valley Authority on the Clinch River, a tributary of the Tennessee River, involved the production of coarse and fine aggregate for approximately 1,000,000 cu. yd. of concrete. A hammer-mill plant for the manufacture of sand from quarried rock was indicated as the most economical method of providing this material. The abrasive character of the feed as compared to the average hammer-mill feed made it necessary to control operations rather closely, and this resulted in the accumulation of many data hearing on production and maintenance. This paper will deal only with three specific sets of data, accompanied by such general information as may be necessary to explain the conditions under which they were obtained and followed by discussions on the relation of silica content to hammer wear, comparative costs of slugger and stirrup hammers, and comparative costs of tool-steel and manganese-steel grate bars. Coarse Aggregate Aggregate for the concrete poured at Norris Dam came from a dolomite quarry within 2000 ft. of the west abutment of the dam. Geologically this rock is known as a Knox dolomite. A typical chemical analysis is given in Table 1. Table 1.—Typical Chemical Analysis Per Cent H2O, 110° F............................................. 0.02 lgnition loss.............................................. 43.56 MgO.................................................... 19.19 CaO ................................................... 29.07 Fe²O³.................................................... 0.39 Al2O³ .................................................. 1.94 SiO2..................................................... 5.75 SO2...................................................... Tr. Rock was carried by 12-yd, dump trucks to a 42-in. gyratory crusher set at 6 1/2in. The product of the primary crusher was taken by a belt conveyor to a double.-deck vibrating screen followed by a 5 1/2-ft. cone

Jan 1, 1938

By Kenneth J. Kurry

12.2-1. Introduction. Successful management constantly strives to improve its methods and equipment in order to produce a better product at reduced cost. It is not enough that managers be cost conscious and industrious in these days of rapid, dramatic technological development. They must keep abreast of these developments-new or improved methods, machines, and equipment. Most of the machines and equipment have been carefully designed and perfected by reliable companies. However, improvements and progress often evolve from a definite need and are most successful in a particular operation. What may be the "answer" to the requirements of one company may be totally useless or inefficient to those of another. The manager of today must continuously evaluate the myriad proposals for change that result from modern technology-in terms of the particular needs and objectives of his company-to determine correctly which will best satisfy those needs and reach those objectives. Each new method or piece of equipment must receive careful study to make certain it is compatible with the entire operation. It may solve a problem in one department only to create a new problem in another. There are many pitfalls to be avoided in obtaining the necessary facts, and often these problems are unrelated to the actual use of the equipment or method being studied. All facets of the operation being considered must be thoroughly studied to determine if the method is compatible to the entire operation. For example, in considering a possible recommendation that a certain method and type of equipment be used to transport ore, some of the factors that should receive close scrutiny are: blasting procedure, maintenance, turning radius, percent grades, weather extremes, dumping problems, tire problems, length of haul, operator acceptance, or mismatched equipment. Failure to consider every factor-large and small, present and future-in evaluating a proposed change can mean a large investment in a new method or machine that will have to be continued for years at higher-than-antici-

Jan 1, 1968

By John Vanderwilt

THE value of molybdenum as a ferroalloy has been known for many years. For a long time, however, the metal was used only on a small scale in the iron industries, because the known source of supply appeared inadequate and the cost to the consumer was not stabilized. The world's production of molybdenum prior to 1914 was only a few thousand pounds per year and even during the war period annual production was only about 800,000 lb. Production during 1929 and 1930, however, was over 4,000,000 lb. per year and that of 1931 promises to be of the same order of magnitude. This remarkable increase in production is due primarily to the development of a large molybdenite deposit at Climax, Lake County, Colo., which has accounted for 80 per cent of the world's production of molybdenum during the past two years. Based on conservative esti-mates, the deposit at Climax has the capacity to maintain its 1929 and 1930 rate of production for 100 years or more. The production of molybdenum ores has been established on a new basis-in recent years as compared with the period during and before the world war. During the earlier period, production came from small deposits and the cost to the consumer fluctuated. At present, production comes from a few large deposits and the price is not only lower but also fairly constant. In recent years the price of molybdenum has consistently dropped as consumption increased. The present price is $0.95 per pound of molybdenum in ferromolybdenum and $0.80 for molybdenum in calcium molybdate. Before and during most of the war, Australia, Norway, and a few other foreign countries were the chief producers, while at present the United States is producing over 90 per cent of the world's production. With molybdenum available at a moderate cost, the steel industry has been adopting more and more the use of molybdenum as an alloy addition. A relatively small demand for molybdenum exists in the chemical and electrical industries, but the recent large increase in con-sumption has in large part been by the iron and steel industries, which

Jan 1, 1932

By John P. Harmon

This paper was written with two objects in view: (1) To describe in detail what has been done toward the ventilation of the main unit of the Portovelo mines and the results; (2) to give information that may be useful to anyone who may have to make a preliminary mine ventilation survey, with recommendations for the improvement of the existing system and finally the maintenance of the improved system. General Description The mining operations of the South American Development Co. at Portovelo, Province of El oro, Ecuador, are described here only briefly, as a full discussion does not fall within the scope of this paper. The orebodies are of the steeply dipping, gold quartz vein type, averaging about 2.0 meters wide. As shown in Fig. 1, the ore chutes are relatively far apart, making considerable development work necessary before connecting raises can be driven between the levels. Faulting' on a large scale and a parallel vein system have made long dead-end drifts and many crosscuts necessary in the exploration of veins. The ventilation of these dead ends is described under Auxiliary Ventilation. The method of mining2 is the filled rill stope method, similar to the method that is largely employed at the Butte mines. The main shaft of the mine is in the Amarillo Valley, on either side of which mountains rise abruptly to a height of several hundred feet above the collar of the shaft. Entrance is made to the mine: (1) Through the American shaft, which extends vertically from the surface to a depth of 1100 ft., of which only the first 700 ft. are in active use at the present time; (2) levels are driven from the American shaft at intervals of 30 m. (100 ft.), of which only the third, fifth and seventh levels need be considered in the ventilation problem; (3) adits at intervals of 30 m. (100 ft.), with a few exceptions, above the collar of the American shaft.

Jan 1, 1931

By W. A. Tiller, S. O’Hara

Grain refinement in stirred melts has previously been shown to arise from dendrite segmentation. The present work discusses experiments capable of distinguishing between remelting and mechanical effects which result from fluid motion. Nucleation of still un-dercooled Sn-1 pct Pb melts with a single-crystal seed shows that the grain size of the resulting ingots decreases as the undercooling at nucleation increases up to about 1Z0C, then increases again for larger undercoolings. Pure melts exhibit little refinement as a result of nucleation at any undercooling up to 12°C. The presence of fluid motion gives rise to grain multiplication in pure melts, but the magnitude of this effect is partly obscured by solid-state pain growth. The results indicate that two grain-multiplication pvocesses can operate, one which involves dendrite remelt ing, and another associated with fluid motion. Small ingot experiments indicated that an equiaxed transition can be caused by fluid motion. THE presence of forced convection during recales-cence of undercooled Sn-1.3 pct Pb melts has been shown to give rise to extensive grain refinement during solidification of small ingots of this alloy.''2 In Ref. 2, this refining was shown to result entirely from dendrite fragmentation and not from any nucleation mechanism. The purpose of the present work is to establish what processes lead to dendrite segmentation. Experiments were therefore carried out to determine the resulting structure of both pure tin and Sn-1 pct Pb melts, nucleated at various undercoolings by single-crystal seeds in the absence of forced or natural convection. By comparing the information obtained from these experiments with the results of Refs. 1 and 2, the effects of fluid motion on the ingot structure, and the details of the segmentation mechanisms, should become clear. The results of these experiments should also be helpful in understanding the solidification structures of larger ingots since natural convection, which is known to play an important role in determining ingot structure13 differs only in degree from the fluid-motion effects studied here. I) EXPERIMENTAL PROCEDURE The apparatus shown in Fig. 1 was constructed to permit controlled nucleation of stirred or unstirred supercooled melts using a seed oriented in the tin dendrite growth direction. Because the melts, in all cases, were capable of larger supercoolings than the one used in each experiment, the structure observed always resulted from a single nucleation event at the surface. The sequential procedure was as follows: Tin was melted under purified hydrogen and the oxide removed at a temperature of 450°C . Electromagnetic stirring was then used to clear the surface completely and during this time the guide tube was lowered underneath the melt surface. A pointed Czochralski-like crystal was then propagated from a specially prepared tin dendrite seed4 inside the guide tube while the temperature of the bulk melt was held just above the melting point, ensuring a perfectly clean seed which was epitaxial with the original dendrite. The dendrite and epitaxial seed, still in the guide tube, were withdrawn from the melt which, in turn, was undercooled to 2C° below the temperature at which the nucleation experiment was to be performed. The melt was held there for several minutes, both in stirred and in still-melt experiments. The temperature was next raised to the required value and the melt nucleated by the carrot seed inside the guide tube, the heater power being shut off just after nucleation. In experiments where quenching after nucleation was required, a high-density graphite crucible was used, but otherwise the procedure was as described above. The temperature gradient in the melt before nucleation was kept below 1C° per cm by carrying the experiment out just below the center of the heater shown in Fig. 1.

Jan 1, 1968

By Nigel B. Pearson

INTRODUCTION Chingola Division of Nchanga Consolidated Copper Mines Limited is situated in the Copper- belt of Zambia, approximately 30 kilometres south of the border with Zaire. It is the country's leading producer of copper and its principal foreign exchange earner, and has been, in its time, the largest single copper producer in the British Commonwealth and the second largest in the world. There are two 'primary' copper sources at the Division, namely Underground producing approximately 100 000 tonnes of contained copper annually, and the Open Pits, producing approximately 200 000 tonnes annually. Following the commissioning of the Tailings Leach Plant in 1974, treating final tailings, metallurgical recovery increased from approximately 70% to in excess of 80%. Extensive deposits of cobalt are present and following the rise in the market price in 1979 and 1980, plans were made to extract this separately. Cobalt production commenced in 1980 from Underground and will increase substantially when the Open Pit cobalt reserves become available in approximately 2 years time. This paper covers the development of the production method of the principal Underground Orebody, known as the Lower Orebody, and the control systems applied to it. HISTORY The Lower Orebody, formerly known as the Nchanga West Orebody, was discovered by extensive Diamond Drilling operations in the late 1920's. Development of the Shaft complex and the orebody commenced and continued until 1931 when flooding from development on the 480 foot horizon necessitated closure. The mine remained closed until 1937 and production commenced from the Lower Orebody in 1939. Various expansion schemes of the Metallurgical Plants and the Underground workings were completed over the years with the result that Underground production increased steadily to in excess of 150 000 tonnes copper annually in the 1960's. In recent years production has levelled out at slightly less than 100 000 tonnes annually. This decrease is as a result of producing from a lower grade reserve, available within a more limited production front, and the application of production constraints in the light of experience gained. GEOLOGY Within the Nchanga Mining Licence Area of Chingola Division, there are a total of 14 known copper orebodies in the basal part of a thick succession of sediments of Pre-Cambrian age. The sediments rest on a Basement Complex of granites, gneisses and schists. Three superimposed stratiform orebodies, known locally as the Lower, Intermediate and Upper Orebodies, occur over a vertical interval of approximately 150m. Current annual copper production is approximately 300 000 tonnes. The ore originates from four different mining operations, one Underground mine and three Open Pits. Approximately 9 million tonnes of ore grading between 3 and 3.5% copper are milled annually. The two largest sources of ore, the Under- ground mine and the Nchanga Open Pit are on the gently dipping south limb of the Nchanga Syncline (Figure 1). In the Nchanga Open Pit the major source of ore is the Upper Orebody with lesser tonnages from the Intermediate and Lower Orebodies. In the Underground mine the Lower Orebody is the only orebody being mined, underground mining of the Upper Orebody being suspended for economic reasons in 1978. The typical stratigraphic column is illust- rated in the legend on Figure 1. The footwall rocks comprise either Nchanga Red Granite or the Arkose formation, both of which are hard competent rocks, offering generally good mining conditions for footwall development. Copper mineralisation of the Lower Orebody extends from the upper part of the Arkose to the top of the Lower Banded Shales. Overlying the Arkose is the Transition zone (0-18111 in thickness). The Transition zone comprise a lower member, Transitional Arkose, consisting of incoherent kaolinised sandstone, and an upper member, Transitional Lower Banded Shales consisting of normally soft bleached and silicified shale. The soft nature of the Transitional rocks is

Jan 1, 1981

By L. E. Snow, L. F. Pett

AT the Utah Copper Pit of Kennecott Copper Corp. a versatile mobile drill unit has been developed and tested. Through increased drilling speed and additional available drilling time, an improvement of 233 pct in drill performance has been realized. Operators using drill power can quickly and easily convert the unit from a vertical to a toe hole drill. Drilling is done with the machine in line with the unit or at right angles to it. Table I covers the 1950 drill record at the mine, showing the types of machines used and the footage obtained. [ ] Development of the new units was prompted by a desire for solution of the following problems: 1-The ever-expanding operations within the Utah Pit, requiring some 65 miles of transmission pipe lines radiating from a central compressed air plant, have now taxed this compressor plant beyond capacity. While 100 lb per sq in. pressure is maintained at compressors, transmission losses from innumerable small leaks result in low terminal pressure at the drills. Fig. 1 illustrates that while terminal pressure varies according to size of air hose and fittings used, the average drill is operating on air pressures within the 60 to 70-lb range. The compressor plant is operated primarily for air rock drills. All drilling is done on day shift, but it is necessary to run two of the three 3720 cu ft capacity compressors during night shift in order to prevent drastic loss of pressure through leaks. On day shift all three compressor units are operated at capacity, so that there is no reserve safeguard available for extensive repairs to any compressor. The calculated volume of compressed air required for operation of 25 drills is but 33 pet of three day-shift compressors. While compressed air is being used to a minor extent for purposes other than drilling, it is felt that satisfactory substitute facilities could be provided in the event that air lines were no longer available. 2-Drill performance has increased beyond the capacity of present 3 ½ -in. drifter machines. 3-The installation, maintenance, and moving of air lines in routine mining is a growing problem. The 2-in. level lines are placed outside electrification towers and must be moved to new location upon completion of each shovel cut. In addition to these moves it is often necessary to dismantle level lines so that excessive shovel spillage, or material covering track following a blast, can be pushed in the clear with bulldozers. Where level room is limited, it is often necessary to remove air lines in giving bulldozers room to prepare grades for wagon drills. In spite of precautions taken, many level lines become buried, twisted, and constricted, with excessive air loss resulting. Excessive back break in blasting often leaves several sections of air line suspended in mid-air. Occasionally air lines come in contact with defective electric transmission towers, so that holes are burned through pipe. 4-Rail-handling of drilling equipment such as wagon drills from one location to another requires considerable locomotive crane crew time. Any new unit developed should travel under its own power. 5-Present practice requires accessory material including hose, water pressure tanks, water barrels, drill steel, and bits to be transported by means of push cars when drill locations are changed on any level. 6-In order to supply numerous drill locations with material, many stock piles of new drill rods, drill bits, and other essentials must be left at strategic points, usually in the vicinity of powder magazines. This practice makes for poor housekeeping and contributes to the loss of supplies. Planning the Unit It was felt that a mobile drill unit which would rectify the foregoing problems should meet certain demands: 1-It must furnish compressed air to operate a 4-in. rock drill at a pressure of 90 to 100 lb.

Jan 1, 1952