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By E. F. Ketterer, D. B. Collyer, A. B. Wilder

AISI constructional alloy steels are a series of steels in which a numbering system signifies the chemical composition limits. SAE (Society of Automotive Engineers) and AISI (American Iron and Steel Institute) use basically the same numbering system. The steels contain, in part, alloy additions of chromium, molybdenum, nickel, and vanadium. They are used widely in the automotive industry but have not been adopted for high temperature service. The purpose of this investigation was to study changes in microstructure observed in typical AISI constructional alloy steels in the welded and un-welded conditions after exposure for 34,000 hr at 900" and 1050" and, in several instances, 1200°F. The tensile and creep rupture properties after exposure for 10,000 hr were also evaluated. The investigation is a continuation of the work previously discussed. Little was known concerning the existence of graphite in medium or low carbon steels before 1943. It is important to note that continued exposure at elevated temperature is required in order to develop graphite in steels. Interest in the presence of graphite in steel, and also in changes in the properties of steels after long service at elevated temperatures, developed after 1943 when a failure of a welded joint of high temperature piping was associated with the graphitization of a relatively low carbon content steel with 0.50 pct Mo,8,0 Numerous investigations have beeen made in recent years studying the relationship of this type of failure to changes in properties. Attempts have been made to evaluate the performance of a number of steels, particularly in the heat-affected zone of welds. It was the unfortunate failure of a weld-heated zone in high temperature piping in the steam-power industry which focused attention upon possible failures, and upon the need for improved steels for high temperature piping service. Investigation by the National Tube Div., U. S. Steel Corp., Lorain Works, has included a wide variety of steels in an attempt to learn the effects of numerous factors and elements upon the long time performance of steels, although many of the compositions studied will never be used at elevated temperatures.1,4-7 Included were the well established AISI steel series, which offered opportunity to test moderately high carbon content steels as well as observe the effect of nickel, chromium, molybdenum, etc., upon the properties of these well known steels. The properties of the AISI steels were also studied for comparison with the previously reported properties of the Cr-Mo steels, now used by the steam-power industry.5,7 Materials and Testing Procedure The AISI alloy steels in Table I were exposed for 10,000 and 34,000 hr at 900°, 1050°, and 1200°F.

Jan 1, 1958

By G. W. P. Rengstorff, H. B. Goodwin

A method is described for purifying iron in batches of 150 Ib or more. Oxygen, carbon, nitrogen, and sulphur are removed from flakes of electrolytic iron by treatment in wet and then dry hydrogen. A special consumable-electrode arc furnace is used to remove hydrogen and to melt the flakes into ingots. A LTHOUGH iron and steel have been made and used for hundreds of years, there are still wide gaps in the knowledge of these metals. One reason for this is that the true properties of the base metal, iron, are not well known, simply because really pure iron has never been readily available for research purposes. Commercial iron, and even the "pure" irons which are readily available, are really complex alloys of iron with oxygen, nitrogen, hydrogen, carbon, sulphur, silicon, and other elements. With such materials, it has not been possible to prepare steels and other iron-base alloys in which all elements are present in exactly controlled amounts. It is an almost impossible task to determine the separate effects of each element or the intereffects of several elements unless alloys can be prepared containing only the desired elements. The ability to make controlled alloys of low impurity content is of far more than academic interest. It will assist in the study of such practical problems as the brittle failure of ship hulls or high pressure gas pipelines, determination of the factors which affect the deep-drawing properties of steels, the improvement of magnetic and electrical properties, the phenomenon of aging, and many others. In 1949, the American Iron and Steel Institute asked Battelle Memorial Institute to prepare high purity iron on a larger scale than had been attempted up to that time. The work was directed toward the preparation of iron as low as possible in carbon, oxygen, nitrogen, hydrogen, and sulphur because these are the contaminating elements of major interest in many research investigations on iron-base alloys. The initial goal was to prepare iron with not more than a few thousandths of one percent of each of these impurities. The purpose was to set up a "bank" of high purity iron on which researchers could draw when a supply was needed and which could be replenished from time to time as the stock became depleted. Of several possible purification methods, heating in hydrogen was selected as the most promising for the removal of both carbon and oxygen from solid iron. Metallic impurities are not removed by hydrogen. Electrolytic iron was chosen as a starting material because it is the most readily available form of iron low in metallic impurities. Decarburization and desulphurization were obtained by annealing in wet hydrogen at relatively low temperatures. The optimum temperature, as found by previous investigators,' is about 1400°F. This treatment also removes nitrogen. Deoxidation is favored by dry hydrogen and higher temperature. There is evidence from previous work that an anneal at 2200°F in dry hydrogen will deoxidize the iron, reduce the nitrogen content, and reduce the sulphur content. These facts suggested the use of a two-stage process: treatment in wet hydrogen followed by treatment in dry hydrogen. To carry out the purification process in a reasonable time, it was necessary that the iron be in thin pieces. Flakes of electrolytic iron were used. Because most research requires that the iron be in massive form, it was nccessary to melt the purified iron to consolidate it into large pieces. For this work, it was decided to prepare ingots weighing at least 15 lb by using a cold hearth arc-melting process in which the iron was melted in an inert atmosphere of argon or helium. This method of melting was selected because the water-cooled copper crucible is not wet by the iron and does not contaminate it as a refractory crucible would. Because the crucible is also the ingot mold, contamination from mold materials was also eliminated. Contamination from electrodes was avoided by making them from purified electrolytic iron. The melting procedure finally developed was an adaptation of the Kellogg electric ingot process.' Two batches of purified iron weighing about 100 Ib each in ingot form were prepared with analyses as shown in Table I. The remainder of this paper describes the procedures in more detail. In general, the process con-

Jan 1, 1956

By L. L. Davis

Current research by the US Bureau of Mines to develop technology for improving efficiency and productivity in metal/nonmetal mines is described. The research program addresses a wide variety of problems associated with each of the major categories of mining: underground mining, surface mining, m situ mining, and premining investigations. An overview of the Bureau's research activities is given with emphasis on underground mining. In particular, two innovative and promising concepts currently under development are described: a unique shaft sinking system that uses an impact hammer for rock fragmentation, and a low-profile portable crusher for underground hard-rock applications.

Jan 1, 1982

Mr. A. Bement, Chicago, Ill. (communication to the Secretary*) :—Mr. Campbell, in proposing his method of sampling coal-seams, has rendered an important service in insisting on the presentation of an adequate history of the sample; and it is to be hoped that his suggestion will meet with further recommendation. In my opinion it is fully as important that a complete description of the sample and the conditions surrounding its taking be given, as the results of the analysis itself; and if such should become standard practice, coal-analysis would assume greater value. As it is, the published data are very confusing, and only to an extremely small extent are they really useful. The method suggested by Mr. Campbell, however, cannot have a general application; for example, in the Eastern Interior coal-basin, particularly at the present time, the output consists largely in grades other than mine-run, which in Illinois, for the year 1905, was only about 24 per cent. of the total product; and inasmuch as Mr. Campbell's method of sampling cannot give values for any other grade than mine-run, my contention that, at best, it can have but a limited application, is apparent. To illustrate the argument, there are two coal-seams, with the characteristics of which I am particularly familiar; each seam, as it lies in the ground, contains almost exactly 10 per cent. of ash; the lump coal from one of these seams ranges from 10 to 11 per cent. of ash, but from the other seam it does not average more than 6 per cent., although, as before mentioned, the average percentage of ash in the full seam is the same for each case. A careful and proper sampling of coal-seams, to be followed by analyeis, is to be commended in connection with all investigation; in fact, it should precede any careful study. A serious error, however, may be made in attempting to use such results to set values on the product as shipped or received. The only

Jan 1, 1906

By James C. Fisher

Traditional uranium exploration techniques were utilized and comparatively evaluated in the Front Range, Colorado uranium province. Intense surficial leaching and thick colluvial cover render traditional airborne and ground radioactivity reconnaissance prospecting useless. Radon gas emanometry and sulfur gas emanometry were slightly more effective but the long and short term variations In soil gas emanation rates make it extremely difficult to gain useful information from these systems. The technique judged most useful was Track Etch radon prospecting. The excellent definition of the technique, as well as its repeatability, indicate that it is an ideal prospecting tool for narrow vein-type deposits similar to those of the Front Range, Colorado.

Jan 1, 1976

By Wilson D. Michell

The Magma copper vein trends east-west, dips 70" south, and cuts through a 6000-ft thickness of limestones, quartzites,. shale, diabase, and schist. The vein is itself a fault with a horizontal offset of 500 ft on the formation contacts in the two walls. There are two principal faults of large displacement and numerous smaller steep and flat faults that complicate the prediction of ore occurrence. A study of the geology, its systematic recording to facitliate ready reference, and graphic methods of interpretation of mineralization have aided the exploration, development, 2nd mining of the ore. In making predictions of the distribution of mineable ore bodies and their shape in critical areas, studies are made of "mineralization contours" and "mineralization trends," which are developed on a longitudinal section of the vein from variations in the concentration of the combined base metal content as shown by assays of available openings in the mine. Introduction This paper describes an example of geological work applied directly to mining and exploration problems in a deep vein deposit, and it is presented as being of interest especially to those engaged in working more or less similar veins. In 1944 to 1945 the writer spent a year as assistant to J. K. Gustafson in a thorough geological examination of the Magma mine. Much of the writer's subsequent independent work described in the present paper had its inception in ideas originated by Gustafson during that examination. Objectives of Geological Work at Magma Mine The objectives of geological work at Magma, as at operating vein mines of a similar nature, can be summarized as follows: 1. Direct assistance to the mining department by means of predictions as to the size, shape, grade, and location of ore bodies; by solution of fault problems and irregularities in the veins; by prepara-?ion of detailed ore-reserve estimates in specific areas; by keeping a complete record of the geology and assay results in the mine; and by guiding development work. In short, geology can aid in the realization of the prime purpose of the mining department, which is that of producing the greatest amount of ore in the most economical and expeditious manner possible. 2. Guidance for exploration, especially diamond drilling, in undeveloped areas of the mine. 3. Contributions to the understanding of the general geology of the deposit and its regional setting, including ore genesis, structure, and mineral relationships. Magma Ore Deposit The east-west trending Magma vein, dipping about 70" south, cuts through a 6000-ft thickness of concordant limestones, quartzites, shale, diabase, and schist, named in descending order. These

Jan 1, 1949

By Wilson D. Michell

The Magma copper vein trends east-west, dips 70" south, and cuts through a 6000-ft thickness of limestones, quartzites,. shale, diabase, and schist. The vein is itself a fault with a horizontal offset of 500 ft on the formation contacts in the two walls. There are two principal faults of large displacement and numerous smaller steep and flat faults that complicate the prediction of ore occurrence. A study of the geology, its systematic recording to facitliate ready reference, and graphic methods of interpretation of mineralization have aided the exploration, development, 2nd mining of the ore. In making predictions of the distribution of mineable ore bodies and their shape in critical areas, studies are made of "mineralization contours" and "mineralization trends," which are developed on a longitudinal section of the vein from variations in the concentration of the combined base metal content as shown by assays of available openings in the mine. Introduction This paper describes an example of geological work applied directly to mining and exploration problems in a deep vein deposit, and it is presented as being of interest especially to those engaged in working more or less similar veins. In 1944 to 1945 the writer spent a year as assistant to J. K. Gustafson in a thorough geological examination of the Magma mine. Much of the writer's subsequent independent work described in the present paper had its inception in ideas originated by Gustafson during that examination. Objectives of Geological Work at Magma Mine The objectives of geological work at Magma, as at operating vein mines of a similar nature, can be summarized as follows: 1. Direct assistance to the mining department by means of predictions as to the size, shape, grade, and location of ore bodies; by solution of fault problems and irregularities in the veins; by prepara-?ion of detailed ore-reserve estimates in specific areas; by keeping a complete record of the geology and assay results in the mine; and by guiding development work. In short, geology can aid in the realization of the prime purpose of the mining department, which is that of producing the greatest amount of ore in the most economical and expeditious manner possible. 2. Guidance for exploration, especially diamond drilling, in undeveloped areas of the mine. 3. Contributions to the understanding of the general geology of the deposit and its regional setting, including ore genesis, structure, and mineral relationships. Magma Ore Deposit The east-west trending Magma vein, dipping about 70" south, cuts through a 6000-ft thickness of concordant limestones, quartzites, shale, diabase, and schist, named in descending order. These

Jan 1, 1949

By W. E. Showalter

This paper discusses some of the results of combstion-drive tests which were made in a test cell using a sand bed 10 in. in diameter x 10-ft long. The test method is illustrated and described. The relationship between the API gravity of the in situ oil and the amount of air required for combustion drive is discussed in detail. Other things constant, the air requirement for combustion drive increases as the APl gravity of the in situ oil decreases. If the test results apply to actual reservoirs, the lowest-priced oils may cost the most to recover by this method. Information is shown which indicates that the effect of pressure on the amount of hydrocarbon burned is not large. A method of predicting air requirements from the API gravity of the in situ oil is presented. INTRODUCTION Combustion drive is the term used to identify the process of interstitial or in situ burning as an oil recovery method. Part of the in situ oil is burned to generate the energy needed to produce the remainder of the oil. Combustion drive as an oil recovery mechanism remains an economic uncertainty in spite of all the work that has been done by the industry in both laboratory and field?-6 This paper will show some of the results of tests which were made in a test cell for the purpose of studying the nature of the combustion-drive process. It will present data which indicate that the API gravity of the in situ oil is a significant indicator of the amount of air required to drive a burning front through oil sand. Air requirement varies inversely as the API gravity of the in situ oil. EXPERIMENTAL The tests were performed in a cell which utilized a cylindrical sand section 10 in. in diameter x 10-ft long. The thin-walled metal pipe which held the sand was wound with twenty external electrical resistance heaters which, by means of an automatic controller, maintained adjacent sections of the wall of the pipe at temperatures equal to the temperatures of the contained sand. Each heater covered 6 linear in. of the pipe. By this means lateral heat loss from the sand section was minimized, thereby causing the sand section to simulate more closely a horizontal increment of a combustion-drive reservoir. Fig. 1 shows a schematic diagram of the test assembly. Thermocouples to measure the temperature in the sand were located every 6 in. along the length of the sand section. The pipe containing the sand was enclosed in a cell designed for an operating pressure of 500 psig. The inlet air pressure was controlled at the inlet, and the gas flow rate was controlled and measured at the outlet of the cell. The oil sand used for the tests was prepared by mixing first water and then oil with the noncon-solidated sand using a closed mixer similar to a cement mixer. Table 1 shows a screen analysis of the sand. Ninety percent of the sand was 100 mesh or finer. This sand was a mixture of 80 per cent No. 120 Nevada White Sand and 20 per cent Tennessee Hi-Fusion Moulding Sand No. 3. The Nevada sand was a clean silica sand. The Tennessee sand con-

By P. I. Brooker

The semi-variogram, the basic tool of geostatistical estimation procedures, must itself be estimated from the available data. From a numerically simulated two-dimensional deposit, with known variability expressed through its semi-variogram, the standard arithmetic mean estimator is used to calculate experimental semi-variograms. These are determined on a series of data sets obtained from the deposit with various sampling grids that cover different areas of the deposit and have a mite range in the number of available data. The orders of discrepancies between estimated and theoretical semi-variograms are determined in this idealized situation; and this provides a basis against which to assess experimental semi-variograms made on field data.

Jan 1, 1984

By John J. Sense

The Mission unit of the American Smelting and Refining Co., located 20 miles south of Tucson, Ariz., is a high-volume, 15,000 tpd operation which utilizes a relatively light equipment fleet. As such, management is very conscious of the mechanical and operational availability of the equipment in use. One of the key items in attaining this efficiency is a complete and reliable record of mechanical and operational machine availability which is an invaluable guide for: 1) Estimating future equipment and personnel requirements of an operating property. 2) Planning an expansion program of an operating property. 3) Estimating equipment and personnel requirements of a new property. 4) Selection of replacement equipment. 5) Comparison of separate operating properties.

Jan 3, 1964

By George W. Mitchell

THE property of Eureka Corp. Ltd. is located in the approximate geographic center of Nevada, 2 miles from Eureka, the county seat. The great sources of power, the Colorado, Snake, and Salmon Rivers and the rivers of northern California, are 300 to 500 miles distant, and no lines serve areas closer than 150 miles. Fuel for diesel and steam generation is available in Utah, 300 to 400 miles to the east. Eureka's railhead is 80 miles north where two trunk lines cross the county. A spur line serves Ely, 77 miles east. Good highways connect Eureka to the railheads. Activity in the Eureka mining district began in the early 1870's. The oxidized high grade lead-sil-ver-gold ore terminated against the footwall of the Ruby Hill fault, and in 1890 the main operations ceased. In 1938 Eureka Corp. Ltd. discovered ore in the hanging wall of the fault by diamond drilling. The history of Eureka in the late 1800's indicates that there was some water at 600 to 800 ft in the old workings, probably accumulations above the water table which did not seriously interfere with mining operations. Both the Locan and Richmond shafts were sunk to a level below the table, but apparently the only serious difficulty with water occurred in the Locan. The steam pump used when the last work was done on the 1200 level in 1923, many years after exhaustion of the main orebodies, is still installed on the Locan 900 level. The capacity was about 500 gpm, lifting 750 ft to the 100 level, which connected with the surface. In addition to this. bailers were used to keep the 1200 level free of water. It is said that pumping in 1923 lowered the water in the Holly shaft, about a mile and a half away, but this seems doubtful. The pumping was of short duration because no ore was found. When work at the new Fad shaft was started in 1941 Eureka Corp. Ltd. engineers were fully aware of the probability of encountering water in large volume. Their primary exploration and development had to be carried on at the 2250 level. The first water was encountered at 300 ft. This was undoubtedly surface drainage in the bedding of the Pogonip limestone and was less than 100 gpm. The fractured, loose Hamburg dolomite at the water table was not well cemented, and relatively little water, 300 gpm, percolated through it with difficulty. At 1350 ft well-cemented dolomite containing some open fractures was encountered. These fractures produced the first water of consequence, 750 gpm. At 1700 ft the volume was 1000 gpm increasing to the maximum during shaft sinking, 1600 gpm, at the 2000 level. Secret Canyon shale, a dry formation, was entered at 2100 ft, where a concrete water ring was placed to catch all of the water. The volume decreased rapidly to a constant flow of 1200 gpm. Below 2100 ft the shaft and stations remained in the shale and water was not a problem. Several faults of moderate displacement, including the reverse Martin fault, had been intersected during the traversing of 1000 ft of wet Hamburg, but no undue quantities of water were encountered. Observations in the diamond drill holes in the ore zone area showed a rapid lowering of the water table. The shaft was flooded when it left the dry shale and entered the water-bearing Eldorado dolomite on the 2250 level, crossing a fissure which paralleled the Martin fault. High pressure water doubled the volume then being pumped. Pipe failure through a water door bulkhead was a contributing factor. Immediately following this flooding in March 1948 preparations were made to recover the shaft as rapidly as possible by increasing power and pump capacities as needed. Measurements before flooding indicated the water could be lowered at a fast rate. However, the water table did not recede as rapidly as expected and volumes required to lower the water in the shaft were higher. Obviously the size of the main water channel on the 2250 level was increasing because of erosion, allowing greater volumes to enter the workings and draining beyond the cone originally being drained during shaft sinking. Eroded material was being deposited in the shaft below the 2250 level in serious proportions. In December 1948 a second flooding of the Fad shaft was allowed for the purpose of reassessing existing conditions and studying alternate methods of attack. The detailed geology of the Eureka mining district, see Fig. 1, has been described during the past 75 years by many geologists.' Only the general features and those which seem to affect the drainage problem will be discussed. The old ore zone, mined between 1870 and 1890, is located in a wedge-shaped block of Eldorado dolomite between the footwall of the Ruby Hill fault and the underlying Prospect Mountain quartzite, see Fig. 2. Production of high grade oxidized lead ore containing high values in gold and silver has been variously estimated at $50 to $90 million. The tonnage mined was probably close to 1,500,000, nearly all of which was found above the water table. The new ore, discovered by diamond drilling in the hanging wall of the Ruby Hill fault, is a flat-

Jan 1, 1954

The following list contains the names of members of whose deaths notices have been printed in MINING AND METALLURGY during the year ending April 1, 1939. Biographical sketches published in MINING AND METALLURGY are indicated in the last two columns.

Jan 1, 1943

By R. E. Zinkham, R. J. Goodwin

A mathematical study has been made of the amount of support a cement sheath could provide to casing cemented into the earth. Several assumptions were required to make the analysis, but only two of them are limiting: (I) the pipe must be completely surrounded with cement, and (2) any mud filter cake between the cement and formation has the same physical properties as either the cement or formation. The calculations showed that little support would be provided to the pipe before an unsupported cement sheath failed in tension; however, when the cement is confined between the pipe and wellbore and is loaded in compression, the pipe could receive a considerable amount of support. In fact, the theoretical results indicate the lower grades and larger sizes of pipe could have their working pressures doubled when reasonable compressive loads were applied to a surrounding cement sheath. These data are shown in six charts. Other down-hole conditions such as setting the cement under pressure, increased temperature and cement confinement all tend to increase the potential usefulness of the sheath. Because of size limitations, a laboratory program to verify the most important results of this mathematical study would be very difficult. However, small-scale field tests would be practicable. This paper shows that, if a solid cement sheath can be obtained in the field by either primary cementing or by repair after detection of flaws by surveys such as the new cement-bond logs, the use of this approach to reducing pipe costs merits further consideration. INTRODUCTION A modification in casing design practices is proposed which may either reduce the amount and grade of steel required to contain a specified internal pressure or permit the working pressure to be increased for a specified weight and grade of pipe. One of the more important considerations in casing design is its resistance to collapse; however, Bowers' and, more recently, O'Brien and Goins' have shown many casing programs are unnecessarily conservative in this respect, and they have indicated how savings can be made by designing for more realistic down-hole conditions. Earlier, Saye and Richardson howed that pipe costs could be reduced by considering the cement sheath as a part of the casing string when collapse resistance was being calculated. More recently, Rogers4 has raised the question as to whether a cement sheath might be considered in the design for burst resistance of the cemented casing. Calculations have been made for the increased burst resistance a cement sheath would provide for casing in a wellbore, and the results show that a sizable amount of support could be obtained in some instances. These data are presented in addition to a discussion of several other factors that are considered to affect the burst strength of pipe supported by cement. Two types of support are treated: Case I for tensile loading of the unconfined cement sheath, and Case for compressive loading of the confined cement sheath. ANALYTICAL TREATMENT AND RESULTS CASE I—TENSILE STRESSES IN AN UNCONFINED CEMENT SHEATH Conditions like this would most likely occur in a greatly enlarged portion of the hole where the cement was not in immediate contact with either the formation or a thin and hard mud cake. The mathematical analysis for this condition, as shown in the Appendix, rests on the following concepts. Pressure inside a unit length of pipe causes: (1) a tensile or tangential stress to be exerted over the longitudinal cross-sectional areas of the pipe and cement; and (2) an equal amount of strain in both the pipe and cement that is uniformly distributed over the wall thickness of each. This analysis was then used to make several calculations for a cement sheath around 51/2-in. OD pipe. The results are illustrated in Fig. 1, which shows that a tensile stress of 500 psi is imposed on a 5-in. thick sheath when the casing contains a pressure of only 1,450 psi. It also shows that a 10-in. thick sheath would be stressed to 500 psi in tension when the pipe contained a pressure of only 2,350 psi. Alternatively, if the stress analysis is made by means of the Lame thick-wall cylinder theory, the inner fibers of the 10-in. thick sheath will be stressed to 500 psi in tension when the pressure in the pipe is only 990 psi. This, of course, reveals that an unconfined sheath is of little support to the pipe in burst; however, an entirely different result is obtained when the cement is confined between the pipe and formation.

By W. F. Schwedes, S. S. Rumsey

In order to realize the enconomics which would result in operating the mine pumps at the Chapin mine with electricity generated by water power, the Oliver Iron Mining Co. recently installed a hydro-electric plant at the upper Quinnesec Falls on the Menominee river, and constructed a transmission line to the mine.

Jan 1, 1915

By W. L. Phillips

Nickel alloys containing approximately 0.5, 2.0, and 6.0 at. pct of Os, Pd, Ru, and Rh were Prepared by vacuum melting. Tension tests were carried out at 25°, 500°, 800°, and 1000°C; stress-rupture tests were carried out at 650° and 800°C. The room- and elevated-temperature tensile strength, as well as the stress-rupture life, increased for the same atomic percent in the order palladium, rhodium, ruthenium, osmium. The ductility decreased with increasing concentration or temperature for a given concentration. The results are discussed in terms of solid-solution theory. BOTH solid-solution strengthening and weakening have been observed in alloys. The simple rules which have been suggested to explain alloy strengthening are subject to several exceptions and deviations based on atom size and valency effects. The purpose of the present investigation was to extend the knowledge of the effect of solid-solution alloying on the mechanical properties of several Ni-Pt group metal alloys at room and elevated temperatures. The platinum group metals, osmium, ruthenium, rhodium, and palladium were chosen as the solute elements because of their different valencies and atom sizes. Nickel was chosen as the base material because of the large amount of previous literature available on nickel alloys. EXPERIMENTAL PROCEDURE Grade A Carbonyl nickel powder having the following composition: C, 0.07 pct; Fe, 0.013 pct; Mn, 0.0005 pct; Si, 0.016 pct; Mg, 0.0003 pct; S, 0.003 pct, was purchased from the International Nickel Co. The platinum-group metal powders were purchased from A. D. Mackay Co. The powders were weighed, cone-blended for 4 hr, and vacuum-melted into 1-in.-diam billets, approximately 2 in. long. The ingots were homogenized for 5 hr at 1800°F before being extruded into 1/4-in. rods at 2100°F. The alloy compositions are shown in Table I. Typical impurity values were not significantly different than that of the as-received nickel. All alloys were single phase. The lattice parameter of each alloy after the heat treatment described below is also included in Table I. Each alloy was annealed at 2200°F for 1 hr and air-quenched to comparable grain sizes (ASTM G.S. No. 4-5). The rods were then machined into test samples 0.160-in. diam and 0.750-in. gage length. Tension testing was carried out in an In-stron tension-testing machine at a strain rate of 0.02 in. per in. per min. An extensometer was used to determine the room-temperature yield stress. Stress-rupture testing was carried out in an Arc Weld stress-rupture unit. All testing was performed in air. EXPERIMENTAL RESULTS A) Room-Temperature Mechanical Properties. The 0.2 pct yield stress, ultimate tensile strength, and elongation for all alloys tested at room temperature are plotted as a function of atomic percent solute element in Fig. 1. The ultimate tensile strength for the same atomic percent increases in the order palladium, rhodium, ruthenium, osmium. The yield strength at low concentrations, -0.5 at. pct, increases in the order osmium, palladium, ruthenium, rhodium. At higher concentrations, 3.0 and 10.0 pct, the yield strength increases in the same order as the tensile strength. Both yield strength and tensile strength increase rapidly ini-

Jan 1, 1964

By William L. Langdon

PRELIMINARY COMMENTS Recent years have been difficult for the mining industry, as reported in the most recently published Minerals Yearbook (Volume III), the estimated value of world crude mineral production has been generally declining since an historical peak in 1979 (16).* The costs of mineral extraction have, however, been rising rapidly. Individual project costs are, in general, exceeding the growth rate of mining companies, while operating and technical risks continue to increase (16). Risks resulting from currency flotation and political actions in the host country are also on the increase. Labor difficulties (such as those experienced by Phelps Dodge) and the Reagan Administration's recent rejection of a proposal by the International Trade Commission for higher tariffs and/or quotas on copper imports have also contributed to the mining industry's difficulties. The net result of these factors has been financial difficulties for mining companies with respect to paying for their current financial obligations as well as trouble finding new financing. It may be that one of the main factors in determining the future of the mining industry will be centered upon capital structure, i.e. how its assets are financed. Historical Perspective on Capital Structure In The Mining Industry As discussed by Isreal Borenstein in (4) and Thomas Navin in (7), mining firms typically have used a financial model based upon high financial leverage, i.e. large amounts of debt relative to equity. This has been especially true for American firms developing orebodies discovered overseas. Problems sometimes occurred for those who were forced to commit themselves to completion guarantees in order to obtain bank loans. Mines under these circumstances must be brought into operation regardless of market conditions. Politicians in overseas countries also may use the high profits required by these high risk ventures to argue for nationalization of mines. Coupled with the frequent use of high leverage, mining companies traditionally have paid out a higher percentage of earnings in the form of dividends than manufacturing firms. This is significant, in that such funds paid out in dividends are not available for other uses. Hence, a greater reliance upon debt and/or additional stock issues than would ordinarily be the case for manufacturing firms. Thomas Navin has stated in (7), that the theory behind the high dividend payout was that, "most mining investors thought of themselves as investing in a wasting asset, one that would gradually deplete over time until it ultimately became valueless." Investors, therefore, required a receipt, in the form of dividends, of a percentage of current earnings and a return on original capital. This view of reality, apparently, has not changed in some quarters, even though modern mine management ordinarily replaces depleted mines with new ones. It may be that this practice of high dividend payout coupled with relatively high financial leverage should be reexamined. Current capital structure theory could prove interesting and useful for those concerned with current conditions. Capital structure patterns and related practices in this industry are not, in many cases, congruent with those thought by some to be theoretically optimal. Prelude to Capital Structure Discussion With a few notable exceptions (such as the Guggenheims and Anaconda's chief operating,. officer in 1971) mining companies' chief executives and boards of directors have not been persons with financial backgrounds. This, one could conjecture, might have accounted for the relatively nonsophisticated nature of capital structures of mining companies. In order to appreciate the possibilities with respect to alternative influence on a firm's performance, one must be acquainted with the basics of financial theory. This paper is an attempt to familiarize the reader with (1) capital structure in general, (2) the usual components of this structure, (3) the general theory behind the use of financial leverage, and

Jan 1, 1985

By R. H. Jamison

A relatively new haulage system is described. Employed by the Delmant Fuel Co.. the "Full Dimension" system provides an uninterrupted flow of coal from a loader or continuous miner at the face to the main line transportation system. This system is said to provide a higher percentage of recovery as well as additional safety and production. Delmont Fuel Co. is employing a comparatively new system of transportation known as a Full Dimension system. Cne of these systems has been in operation for a year at the company's 10-B Mine as a part of a conventional section. A second was installed at the No. 10 mine in late 1960 to handle the production of a Colmol in a pillar section. SYSTEM COMPONENTS A Full Dimension system is a haulage system that provides an uninterrupted flow of coal from a loader or continuous miner at the face to the main line transportation system. The equipment required for this system consists of a series of interconnected chain conveyors that are mobile and articulated. They will retract or extend a sufficient distance for the development of a five-entry system; or, in the Colmol pillar section, it provides reach of 210 ft in all directions from the section belt. The components of this system are: l)One 160-ft chain line placed in tandem with the belt conveyor. It has a self-propelled drive, is 20 in. wide and 9 in. deep. Moving this conveyor requires the assistance of a loading machine or cutting machine. 2) One 40-ft piggyback that discharges along the entire length of the 160 ft chain conveyor. 3) A mobile bridge carrier, which is a self-propelled conveyor with four wheel steer and four wheel drive, twenty-eight feet long, it delivers coal to the receiving end of the piggyback. Axles steer individually making possible almost lateral movement. 4) Another 40-ft piggyback, duplicate of item 2 that delivers coal along the entire length of item 3 (mobile bridge carrier). 5) A second mobile bridge carrier, similar to the first, which deliver coal to the piggyback (item 4). 6) A third 40-ft piggyback, duplicate of items 2 and 4. This pig is attached to the loading machine and delivers its coal along the length of the second mobile bridge conveyor. Since the original preparation of this paper, the Delmot Fuel Co. has been able to eliminate the 160-ft chain conveyor. This was accomplished by connecting the outby piggyback directly to a loading machine with an extended boom. The loading machine loads directly onto the belt. This change has resulted in a substantial reduction in moving time and greatly increased flexability. A single trailing cable powers the entire string of equipment. It is attached to the side of the equipment in such a way as to keep it off the ground and afford maximum protection. The tramming rate of this equipment is 90 fpm. The conveyor capacity in a conventional section at Delmont's mines is 7.5 tpm and in the Colmol section is 5.5 tpm. This regulation is a simple function of conveyor speed. To visualize operation of this equipment, it would be well for me to touch briefly on local conditions in the Upper Freeport seam in which we mine. (Also, see the photographs of some of the equipment in use.) DELMONT'S TOPOGRAPHY The Delmont Fuel Co. operates two mines in this seam in Westmoreland County, Pa. The No. 10 mine, which was opened in about 1912, is now almost worked out. Depending on economics in the industry, it has a life of two to four years on a declining production basis. A year ago a new drift mine was opened which is called No. 10-B. It is about two miles from the cleaning plant and is connected thereto by an overland belt conveyor. The new mine is being developed at a rate calculated to take up the slack as the old mine plays out. The Upper Freeport seam averages 4.2 ft in thickness in the area of the Delmont mines. It carries 4 in. of boney coal at the top of the seam and a middle man of from 2 to 4 in. We mine just above a 1-in. slate parting which has 4 to 6 in. of highly laminated coal beneath it. This material normally makes a very firm bottom. The roof varies from dark shale to sand rock and 36-in. bolts are placed on 4-ft centers for roof support. All working places are driven 20 ft wide on development and 25 ft wide on retreat. Selection of mobile chain conveyor equipment when it became available, was a very natural move for Delmont Fuel to make, because chain conveyors and piggybacks had been in use at the company's mines for about 12 years. Grades in the new mine

Jan 1, 1961

By P. Bolsaitis, L. Skolnick

A model of metallic solutions is postulated which explains the energy of formation of alloys on the basis of changes in electron density around solute and solvent atoms and changes in pairwise interactions between atoms. The model is based on the Wigner-Seitz cell model of metals, with the electronegativities of the metals as the parameter for the electron-attracting power of the constituent atoms. In contrast to models based on the description of cohesive energy in terms of pairwise interactions only, which require an empirical, adjustable correction parameter to the Berthe-lot relations in order to predict negative energies of formation, the electron cell model accounts for both negative and positive energies on the basis of parameters of the pure components. Calculated heats of mixing based on published parameters of the pure metals are found to compare favorably to experimental values, despite the absence of any adjustable parameters in the model. THE formulation of an "exact" quantum mechanical model for the energy of cohesion of metals reduces ultimately to the many-body problem and for a complete solution would require the knowledge of wave functions and energy levels of all the particles in the system. The case of alloys is further complicated by the fact that the potential fields around the constituent atoms are different and the matching of boundary conditions in terms of the resulting wave functions is still an unresolved mathematical problem. It appears that in spite of the numerous investigations that have led to an elucidation of many properties of the metallic state, the formulation of the Hume-Rothery rules in 1931 marks the lone accomplishment toward simple, qualitative prediction of the energy of formation of alloys. More recently Friedel'' and Arafa have utilized the modern concepts of the electronic structure in metals in combination with the Hume-Rothery rules to arrive at quantitative estimates of the energy of formation of alloys with low solute contents. The relative success of the Hume-Rothery rules and Friedel's and Arafa's studies encouraged broaching the subject of formulation of a model based on realistic physical parameters and of conceptual and mathematical simplicity that could be applied to alloys regardless of concentration. On the basis of the free-electron theory, the cohesive energy of pure metals can be expressed as a function of "valence" (number of free electrons per atom), lattice parameter (or density), and the ioniza-tion energy. In addition to the electronic contributions to cohesion, one has to consider the pairwise exchange interaction energy between neighboring ion cores and the contribution from van der Waals forces. The energy changes upon alloying are viewed as resulting primarily from a redistribution of electron densities around the ions (a simplification of the more basic concept of a change of the wave functions of the electrons in the presence of a perturbing potential field). In addition there is a change in the pairwise interactions between neighboring ions which is usually the only factor taken into account in commonly used adaptations of the quasichemical theory. The electronegativities of the metals forming the alloy are used as the measure of the electron-attracting power of the ions in the alloy and the difference in electronegativities as the driving force for changes in electron energy on alloying. The interplay of pairwise interactions and electronic forces results in positive or negative energies of formation of alloys. Results obtained for systems to which the free electron model is best applicable are in good agreement with experimental values. 1) COHESIVE ENERGY OF METALS IN TERMS OF THE ELECTRON CELL MODEL The simplest physical model of a metal is that of a lattice of positive charges immersed in a sea of uniform, negative charge distribution represented by the

Jan 1, 1969

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 Paul A. Taylor

MUCH information concerning pegmatites which was thought to be true a few years ago has been proved false, and what is now actually known about some pegmatites is not true of many others. The erratic and seemingly unpredictable structure and variable composition of this class of mineral deposits has been widely emphasized. Even parts of the same pegmatite body exhibit marked differences in texture, mineral composition, width, and attitude. Constructive geological thinking in respect to pegmatites now aims to establish general laws rather than to stress the confusing diversity of features having no special economic significance. Substantial progress has been made in classifying different types of these deposits according to general features, internal structure, mineralogy, and origin. In some cases it has even been possible to block out tonnage reserves in advance of mining. It is still easy, however, to make highly erroneous predictions after a preliminary examination of a pegmatite prospect. Pegmatites are important to the economic well being of the country and to its military security. They furnish virtually all the feldspar, strategic mica, beryl, columbium, tantalum, and caesium utilized in the United States, as well as sundry other minerals and significant amounts of lithium and rare earth minerals and gems. With the exception of vermiculite, occasional ilmenite-rutile, and perhaps soda-lime feldspar and garnet deposits, basic pegmatites are of little economic importance. Consequently in this paper, as in common parlance, the term pegmatite generally relates to coarsegrained acidic rocks or what is aptly called giant granite. Available data indicating the size and importance of the production and trade in specified pegmatite minerals are summarized in Table I. Geological Features Much of the latest thinking on the economic geology of pegmatites is now available in a 115-page monograph' by a group of experts who participated with geologists of the Federal Geological Survey in the widespread wartime investigations. Doubtless the most significant feature of the monograph is indicated by the title, The Internal Structure of Pegmatites, but it also contains a vast amount of other new information and includes the assimilated concepts of many earlier writers, whose works are given in a comprehensive list of references. The shape, size, attitude, and continuity of many pegmatite bodies is controlled by the structure of the older rocks in which they occur. If the older rocks are easily penetrated, e.g., biotite schist, most of the pegmatites in a given district will be found outside the parent granite mass as exterior pegmatites. Marginal pegmatites are more prevalent if the older rocks are massive, unsheared, and sparingly jointed. Networks of pegmatites are abundant in highly-jointed rocks. In strongly foliated schists the bodies are usually lenticular, whereas in highly-folded areas they assume tear drop, pipe or pod-like, bow-shaped, or sinuous forms. Jahns2 recognizes five major shape classes: l—dikes, sills, pipes, and elongate pods; 2— dikes, sills, pipes, and pods with bends, protuberances, or other irregularities; 3-—trough-or scoop-shaped bodies with or without complicating branches; 4—bodies with the form of an inverted trough or scoop; and 5—other bodies, including combinations of the above and miscellaneous shapes. Many pegmatite deposits are scarcely big enough to be recognizable as such. Most of them, in fact, are small tabular deposits less than 4 in. wide and usually without economic concentrations of minerals. On the other hand, some pegmatites are more than a mile long and over 500 ft wide. The ratios of length to breadth range from 1 : 1 to 1 : 100. Although the vertical dimension bears no invariable relationship to strike length, tabular deposits or large lenses are often symmetrical enough to show nearly as much continuity down dip as in their horizontal extension, and some pipes or pods are amazingly persistent in the vertical plane. Small pegmatites often string along a fairly definite trend line; in a given district major bodies may lie roughly parallel, and where only a few of them do not, the erratically disposed bodies generally differ in composition from those conforming to the regular pattern. This does not apply, however, in all districts. Characteristically, pegmatite veins pinch and swell or split into branches. When they pinch out entirely it is often possible to find a new body by prospecting the extension of the strike or dip, but the chances of finding a hidden deposit are ordinarily too uncertain to justify much subsurface prospecting. Diamond drilling may yield valuable information as to the continuity of known deposits whose upper portions are well-exposed. Some deposits, in fact, can be proved up for hundreds of feet by surface trenching and then intersected by drill holes at various depths like any other vein-like deposit. Others twist and branch, apparently defying all efforts to explore them short of actual mining.

Jan 1, 1954