Search Documents

By Joseph A. Norden, A. R. Willson

Much has been written concerning the hoisting equipment of various mines throughout the country, but most of the literature on the subject, if not all, describes equipment of extraordinary capacity and the small and medium-sized outfits have been neglected. The hoisting equipment here described is capable of handling about 1200 tons per day. It has not been worked to capacity, but 600 to 700 tons, together with men and materials, have been handled in two shifts without any great effort. Only one shaft extends from the surface to the lowest levels of the mine; see Fig. I. It is composed of three compartments, two of which are used for hoisting, and the third for water, air, and electrical lines, and for sinking purposes. In 1909, this shaft was raised from the 1000-ft. (304-m.) level to the 700-ft. There a station was cut and an air hoist installed. At this time all mining operations were confined to the upper levels of the mine; that is, above the 700-ft. The ore was dropped to this level through chutes from the 300-, 400-, 500-, and 600-ft. levels and then lowered in cars on cages to the 1000-ft., which is the main haulage adit, the mill and railroad bins being about 3000 ft. from the shaft. Orebodies were, in due time, developed on the 1000-ft. level and below. Therefore, in 1913,the shaft was sunk to the 1150-ft. level and thereafter sinking was continued at regular intervals until the 1500-ft. level was reached. But, hoisting from the 1300-ft. level was a difficult and slow process with the antiquated air hoist, so, in 1916, it was decided to raise the shaft to the surface and install a modern, electrically operated hoist. This was completed early in 1917 and the hoist was in operation in March of that year. The Hoist Figs. 2, 3, and 4 show various features of the hoist and electrical equipment. A Nordberg engine of the double-drum, geared type, designed for a 10-ton load, including rope, at a speed of 1500 ft. per min., is used. The drums are 7 ft. (2.1 m.) in diameter, with a 54-in. (137-em.)

Jan 1, 1922

By F. L. Ruhlman, Max C. Eastman

The termination of the recent war marked a major changc in the oil-field operations at. Elk Hills. Production was greatly curtailed, various experimental production projects were started, and a long-range exploratory program was inaugurated. Production was reduced from the wartime peak of 65,000 bbl per day to the current rate of about 8ooo bbl per day in order to maintain an available source of crude that can be drawn upon during a national emergency. The existing production rate is deemed essential to carry on the experimental production projects. These projects are being conducted to determine the best methods and means of maintaining the Reserve, shut in, but in a state of readiness to produce the maximum quantity of oil in the most expeditious manner. Projects now being investigated include: (I) protection of surface equipmerit; (2) repair of wells that might damage the reservoir; (3) gas injection; (4) water eneroachment, (5) gravity drainage. An exploratory program is being carried on to determine the amount of recoverable oil in the unexplored portions of the Reserve. The Shallow-zone exploration is practically completed, but owing to the complexity of the Stevens structure and the magnitude of the oil discoveries in the Stevens, exploration will probably continue in that zone for another two years. Introduction Summary of Development History of Elk Hills Although a detailed history of Elk Hills oil field is presented in various publications,1,2 a summary of its unusual development and production history will be given in order that the present problems may be more readily understood. The first development that resulted in commercial production was in the central part of the Reserve. The first well, Hay No. I, in sec 36-R,* was completed in January I9I9, and was soon followed by 37 Others in the same general area. It is commonly referred to as the Hay-Carman area and was shut in in stages from 1925 to 1934 as a result of litigation regarding the title to the land. Not being highly producti,, only about 6½ million barrels of oil with a large volume of gas was produced from the area prior to the shutdown. Shortly after the start of the development of the Hay-Carman area, drilling was started 6 miles east in sec 36-S' Tupman NO. I came in for 5000 bbl per day early in I92o, It was followed by more than 200 others in the next few years, and the field became one of major importance. Produc-

Jan 1, 1949

By Thomas Fraser, H. F. Yancey

BefoRe a new coal-washing plant is installed, or an existing washery is remodelled and improved, considerable experimental work on the coal to be washed should be done. A thorough examination of a coal, to determine whether it can be washed successfully and the best and most economical process for securing a satisfactory grade of washed coal, calls for mine sampling, analytical work, and experimental coal-washing tests which may require several months' time and the expenditure of considerable money. Having collected the required data on the characteristics of the coal and the results secured with different types of machines under different sets of conditions, it is necessary to decide what method of treatment should be used. The interpretation of the experimental data is no small part of the problem. The first step in the investigative work is to determine from the characteristics of the coal and the manner of occurrence of the impurities whether it is possible to separate the dirt from the coal and produce a washed coal of the required quality. Then it is necessary to select the type of washer or the system of treatment that will produce the greatest amount of washed coal of this grade for the least cost. To secure this information, tests may be made at laboratories operated by consulting or contracting engineers, at the testing plants of machinery manufacturers, at commercial washeries, or in an experimental washery erected on the site of the proposed plant. Such tests will produce figures in which every factor, including analyses of raw coal and washed coal and yield of washed coal, may vary widely. To reduce these results to a common basis for comparison is very difficult. A number of plans

Jan 1, 1923

By R. S. Rickard, H. D. Peterson, J. D. Miller, M. C. Fuerstenan

Pure chrysocolla is floated with chelating agents that form insoluble complexes with copper at ambient temperature. Complete flotation is obtained with potassium octyl hydroxamate as collector at pH 6. Flotation response is enhanced with increased temperature when low additions of hydroxamate are involved. A natural ore was floated with 0.4 1b per ton octyl hydroxamate at pH 6.5 and 58°C, and a flotation recovery of 76% was obtained with a concentrate grade of 31.6% copper. The flotation characteristics of the oxide copper minerals, malachite, azurite, and cuprite, have not presented the difficulty for concentration as have those of the copper silicate, chrysocolla. The copper carbonates and oxides respond reasonably well to flotation with conventional collectors, whereas chrysocolla will not respond to flotation with fatty acids or xanthates under normal flotation conditions. In this view then, other reagents will have to be devised to function as collectors for chrysocolla. The most obvious general class of reagents for this purpose would seem to be the organic copper chelating compounds. The utility of chelating agents as collectors in flotation systems has already been demonstrated. For example Vivian1 has floated cassiterite using ammonium nitrosophenylhydroxylamine. Holman2 studied the flotation of nickel oxide ores with dimethylglyoxime and also suggested the use of taurine on oxidized lead ores. A rather detailed study on the application of certain chelating agents to some flotation systems was presented by Gutzeit.3 This work indicated that the formation of surface insoluble chelates is probably responsible for flotation in many cases. The role that soluble chelating agents assume was also presented, that is with effective removal of polyvalent cations by complex formation, effective depression of quartz can be obtained. DeWitt and Batchelder4,5 have shown that oximes function well as collectors for chalcocite, malachite, azurite, and cuprite. Ludt and DeWitt6 have presented the possibility of using dyes, such as octyl malachite green, as a collector for chrysocolla. These works presented by DeWitt et al suggest that organic copper chelating compounds offer considerable promise for concentrating chrysocolla, but that the most suitable chelating reagent from the standpoint of cost and flotation response has yet to be determined. The present investigation was undertaken to study the flotation response of chrysocolla to selected copper chelating compounds, that is ethylenediamine, hexamethylenetetramine, potassium octyl hydroxamate, dimethylglyoxime, and benzoin a-oxime. EXPERIMENTAL MATERIALS Chrysocolla: The chrysocolla used in the microflo-tation studies was from New Mexico and analyzed 20.8% copper. X-ray diffraction showed this material to be amorphous in character, but some faint lines of quartz were noted. A natural oxide copper ore from Utah was also used in the investigation. This ore contains malachite principally but also contains chrysocolla and tenorite together with some sulfide copper. The sulfide copper comprises from 15 to 20% of the total copper content of the ore. Water: Conductivity water was used in the microflo-tation experiments, while Golden tap water was used in the experiments with the natural ore. Reagents: Reagent grade n-amyl alcohol was used as frother in the experiments with pure mineral, while MIBC was used in the work with the ore. Ethylenediamine, hexamethylenetetramine, dimethylglyoxime and benzoin a-oxime were reagent grade in quality. Pure potassium octyl hydroxamate was prepared as follows: 1) 1.0 mole of KOH in 140 cc of methanol was combined with 0.6 mole of hydroxyl-amine hydrochloride in 240 cc of methanol at 40°C (KC1 precipitates under these conditions and to effect complete removal of KCl, the system was cooled to 10°C and filtered), 2) a light methanol wash was then given the KC1 cake, 3) the filtrate was agitated at room temperature and 0.33 mole of the methyl ester of the organic acid (i.e. methyl octanoate) was

Jan 1, 1965

By J. A. Roberson, A. M. Adair

A plate of an 18 pct Ni (250 grade) maraging steel was homogenized, cold rolled, and re crystallized to produce a fine grained, homogeneous structure. The strength properties were then measured as a function of cold rollinq prior to aging. Material in the solution heat treated condition exhibited substantial strengthening from cold rolling. The strength increase and the rate of strengthening were proportional to the amount of cold work. This benefit, of up to 50 ksi, was retained during aging for 3 hr at 900°F. The structures produced by cold rolling and aging have been systematically investigated by transmission electron microscopy. The dislocation debris produced by cold work influences the morphology and distribution of precipitate particles as well as the rate of precipitation. An essential feature of the high nickel maraging steels is a martensitic transformation from y to a with the resultant production of a supersaturated bcc solid solution. This solution decomposes to form precipitates during a subsequent aging treatment.' The martensitic transformation in low carbon Fe-Ni alloys and the mechanical properties of these alloys have been studied extensively.2"4 The carbon content of maraging steel is kept very low to avoid the embrittling Fe3C phase and to minimize the acicular, or twinned martensite common to high carbon steels. Maraging steels are usually strengthened only by heat treatment, but cold working is also quite effective in this respect as was shown by Decker, Eash, and Goldman,' Kula and Hickey,5 and the present work. There is, however, some disagreement between the present results and those of Kula and Hickey. Because of their unusual combination of strength and toughness, the maraging steels are suitable for applications where reliability at high stress levels is required. The strength level is achieved through a combination of solid solution hardening,' substructure hardening,3 and precipitation hardening,6 of which the last is the most important. The toughness of these steels has been reported to be a result of the absence of brittle grain boundary precipitates,7 and the fact that the matrix is a massive martensite rather than a twinned martensite.' One of the problems in the manufacture and use of maraging steel has been its tendency to develop laminations in sheets parallel to the rolling plane.9,10 The present investigation was intended to determine whether or not an improvement in the microstructure of this steel by thermomechani-cal processing would improve mechanical properties. I) EXPERIMENTAL PROCEDURES A plate ($ in. thick) of an 18 pct Ni-250 grade maraging steel* was obtained and examined metallo- "Analysis, wt pct: 18.1 Ni, 8.6 Co, 5.3 Mo, 0.4 Ti, 0.07 Mn, 0.05 Cr, 0.05 Si, 0.002B. 0.1 Cu. 0.05 Al. 0.002 Zr, 0.016 C, balance Fe. graphically. It was found to contain laminations as described by previous investigators. This plate was then annealed at 2100°F for 24 hr to promote homogeneity. Pilot studies on small coupons had shown that this was the minimum temperature required to accomplish homogenization in 24 hr. Since this treatment caused excessive grain growth, the steel was subsequently cold rolled and recrystallized. Cold rolling was accomplished at about 5 pct per pass to a total reduction of 66 pct. The standard solution annealing treatment (1 hr at 1500°F) was adequate to cause recrystallization of the cold worked steel. The recrystallized steel was then cold rolled and ground by various amounts to produce specimens with uniform thickness and selected amounts of cold rolling. All tensile specimens with less than 90 pct cold reduction were ground to a thickness of 0.050 in. Specimens with 90 pct cold reduction were 0.025 in. thick. All

Jan 1, 1970

By T. S. Spice, H. L. Lovell, R. W. Utley

With the increased demand for lightweight aggregate, such materials have been manufactured from slags, clays, slates and, to a minor extent, the refuse of coal preparation processes. The latter source is attractive due to current interest in eliminating air pollution and improving the unsightly appearance of refuse piles as well as potential economic advantages, if a satisfactory product can be manufactured. A study has been made relating the composition of coal refuse, expressed as washability data, and the properties of a resulting aggregate. Three different refuse samples (bituminous and anthracite) were separated at gravities ranging from 1.4 to 2.96. Their chemical composition was determined. Aggregates were prepared in the laboratory from these fractions before and after ashing, at temperatures between 1250' and 1350°C. The specific gravity, expansion and compressive strength of the products were measured. The response of the refuse to bloat is related to the chemical composition of the ash to yield a glass of desirable viscosity and to the presence of a material which will form a gas in the appropriate temperature range. Carbonaceous material within the particle is a negative factor since after the glassy phase forms, oxygen cannot enter the mass thus inhibiting bloating. Precombustion of the refuse is also detrimental since it removes substances of gas-forming capability. The fraction sinking at 2.60 yielded an aggregate which had the lowest specific gravity, greatest expansion and greatest compressive strength. The purpose of this study was to seek and evaluate possible relationships between the composition of coal refuse, as expressed by washability data and factors affecting the production of a lightweight aggregate from refuse by bloating. The study was related to an intense program in the Department of Mineral Preparation at the Pennsyl- vania State University to study the possible elimination of air pollution from coal refuse fires by producing a low-carbon secondary refuse. Another phase of this program is considering possible uses of such a secondary refuse. Thus one study was to determine if a lightweight aggregate could be produced from this new refuse material and what effect the nature of refuse beneficiation would have upon the characteristics of a commercial aggregate product. It was hoped that these studies would assist in the design of a process and in establishing the optimum specific gravities at which refuse separation could be made in respect to the production of a low grade fuel from the float product and a raw material for lightweight aggregate from the sink product. The combination of producing a saleable low grade fuel and a lightweight aggregate, while at the same time minimizing disposal costs and air pollution contributions would be attractive to the coal mining industry. MECHANISM OF BLOATING Austin, Nunes and Sullivan,1 Ehlers,3 Everhart, 4 Riley7 and others, have stated that bloating is an expansion process in which two essential conditions exist simultaneously. These are: the development of a glassy phase over a relatively wide temperature range (the viscosity of this glassy phase plays a major role in the retention of the evolved gases) and the evolution of gases (SO2, SO3, CO2 and O2) from mineral decomposition in sufficient quantities at appropriate temperatures to form a cellular structure in the mass. Unless these two "type" reactions take place at the same time or in overlapping ranges of time and temperature, cellulation will not occur. In this event, the gas will escape or the melt will be too fluid to retain it; or if retained, the cells may be too large or the walls too thin for good strength in the product. The stages that occur when clays and shales are heated in an oxidizing atmosphere may be summarized as follows: 1) Drying and removal of free water to 200°C. 2) Absorbed water dehydration period (200 to 480°C). 3) Chemically-combined water dehydration period (480 to 700°C).

Jan 1, 1965

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 James T. Gow, Oscar E. Harder, Anton de S. Brasunas

This paper deals with the results obtained and the techniques employed in determining: 1. Liquidus and solidus temperatures of the HH and HT type heat-resistant alloys. 2. The relation of true temperatures (thermocouple) to apparent temperatures (optical pyrometer) for the molten HH and HT type alloys. This work was the outgrowth of needing to know something definite about these temperatures in order to study the effect of pouring temperatures, shake-out times, and heat-treatments on the properties of some heat-resistant alloys. The information developed is thought to be of interest to industry for guidance in foundry practices and the industrial heat applications of the alloys. This paper is based upon work that has been done for the Alloy Casting Institute at Battellc Memorial Institute. Liquidus and Solidus Temperatures of HH and HT Type Alloys Experimenta1 Procedures Temperature-time curves were obtained during the cooling and solidification of several HH and HT type alloys. The metal was melted in an Ajax high-frequency, induction furnace and poured into a core-sand mold for cooling. The temperature measurements were obtained by means of a platinum, platinum-rhodium thermocouple contained within a fused silica sheath, which was extended through the side wall of the dry core-sand mold to a point at about the center of the mold cavity, which was 3 in. in diameter by 5 in. deep. The cold junction of the thermocouple was kept in an ice bottle, and the temperature was recorded automatically on a direct-reading potentiometer recorder. Through a switching arrangement, the temperature recorded was checked occasionally with a portable Leeds and North-rup potentiometer indicator. Less than a 5°F. variation in the two instruments was always found. This is within the limit of accuracy of readings. The thermocouple used was checked for accuracy against a standard couple, both before use on this work and after completion of the work, by the instrument laboratory. Temperature-time Cooling Curves Fig. I is an enlarged reproduction of a part of one of the cooling curves over the temperature range of 2600' to 2200°F. for an alloy of the HT type with 0.45 per cent carbon, 16 per cent chromium, and 36 per cent nickel. This is a typically shaped cooling curve as obtained for the alloys. It will be noted that two temperatures "re shown on Fig. I for the liquidus temperature. It is thought likely that the higher temperature, A, at the point at

Jan 1, 1945

By L. S. Darken

The manufacture and treatment of metals comprises operations whose effectiveness depends in large measure upon diffusion phenomena. The significance of such phenomena has, for a few simple cases, long been understood, but in more complex cases it has not been adequately recognized. This paper presents a mathematical analysis of some of the cases in which diffusion leads to the formation of a new phase within the metal, and this analysis provides a basis for interpreting quantitatively many apparently diverse experimental observations recorded in the literature, some of which are outlined in the following pages. OBservt1ons Recorded in Literature An outstanding instance is the general type of phenomenon in which diffusion of an element into an alloy is followed by a phase change, which most frequently is a reaction (precipitation) of the diffusing element with one or more components of the alloy to form a compound. Thus in certain alloys heated in contact with oxygen not only the usual oxide scale appears but, in addition, beneath the surface, a well-defined zone in which free oxide particles are dispersed with notabIe uniformity. This zone, commonly referred to as the subscale or zone of internal oxidation, is remarkable in that the visible boundary between it and the unaltered metal is usually sharp. Such a subscale, which appears to have been recognized first in oxidized copper alloys, has been described and studied by C. S. Smith,l by Rhines,2 and by Rhines, Johnson and Anderson;3 but it appears also in other alloys when they are exposed to conditions that bring about the precipitation of some compound, not necessarily an oxide. Another clear case of this same kind is the diffusion of carbon into a steel containing an element such as molybdenum, chromium, vanadium or tungsten, whose carbide is less soluble in iron than iron carbide, the result being that some part of that element is precipitated as a dispersed carbide. A similar phenomenon may occur on the welding of two alloys each of which contains one of a pair of elements that react to form a compound or a new phase of any sort. But in addition to the precipitation of a compound formed from the diffusing element, other effects may result merely from the presence of the diffusing element. For instance, if homogeneous austenite, nearly saturated with respect to carbon, is heated in contact with silicon in the absence of oxygen, one would expect to find, behind the several iron-silicon phases analogous to a scale, a zone of austenite containing dispersed carbide, because of the fact that silicon decreases the solubility of carbon in iron. As still another example, oxidation of an alloy may result in diffusion toward the surface of one or more of the constituents; for instance, during oxidation of a steel containing an element such as "copper, which is less prone to oxidation than iron, there is a notable enrichment of copper just beneath the scale-metal interface. AS an example of a yet more

Jan 1, 1942

By Stephen Royce

[Stephen Royce.. I .......................... E. S. Moore.. 9 .......................... E. L. Bruce.. 10 ......................... J. E. Hawley.. 11 ........................ L. M. Scofield.. I 2] STEPHEN ROYCE, * Crystal Falls, Mich.Having examined and reported upon the Steep Rock Lake property and vicinity on several occasions for clients, and having kept in touch with the operation from 1938 on, the writer offers the following discussion of the valuable paper by Messrs. Roberts and Bartley, which was so kindly prepared in response to the writer's earnest request, made as a member of the Committee on Mining Geology. ORIGIN OF THE ORES The writer wishes to express complete agreement with the authors that the origin of these deposits was replacement action by hot ascending sulphide solutions of igneous origin, gaining access via the limestone and dolomite breccia channels, and acting mainly to replace the dolomite and limestone breccia with hematite and hydrated hematite. Some replacement in adjoining greenstones and tuffs by the same solutions may also have taken place, but the latter occurrences may represent replaced iron formations originally bedded in with the volcanics. Iron formation in the same relationship to volcanics is extensively developed on the western Menominee Range of Michigan, in the Middle Huronian belts; also appears in similar relations on the Vermilion Range; and is abundantly exposed in the same stratigraphic position no farther away than the southeast arm of Steep Rock Lake itself. Furthermore, the writer's classification of drill cores, made June 11, 1938, on the prop¬erty, contains the following, for hole 8, in the original strike of the A ore body: "332-342 ore; 342-350 much-leached sugary chert with thin ferruginous slate bands." This is a typical iron formation. The ore in this first group of holes (Numbers 6 to 13, incl.) is described in the writer's 1938 notes as follows: "The ore is hard, massive hematite with some limonite, and shot through with many small vugs." The breccia forms in the float ore and in the drill samples were also noted. As a field classification this does not check badly with Professor Schwartz's microscopic work. Notes supplied by Mr. Neeland, then geologist at the exploration, show: "Hole 23 has some iron formation, not logged yet." It appears therefore that there is some iron formation at the exploration, much iron formation in similar stratigraphic position in the vicinity, and iron formation similarly located as to volcanics on the Vermilion and western Menominee Ranges. Certain of the holes, notably Holes 25 and 27, showed some specular hematites resembling those known to be formed from cherty iron formation elsewhere. Hence the writer concludes that at least a small part of the Steep Rock ores results from replacement of iron formation. By far the greater part of it is, as the authors say, clearly a replaced limestone and dolomite breccia, with the breccia outlines still preserved in much of the ore, and all of the replacement is by hot ascending waters. The authors' citation of the neighboring Atikokan pyrite, pyrrhotite and magnetite

Jan 1, 1943

By E. E. Wahlstrom

The Harold D. Roberts tunnel, in Summit and Park Counties, Colorado, is a concrete-lined pressure tunnel finished to a circular cross section of 10.25 ft diam. The tunnel is 23.3 miles long and is designed to transport water for domestic use by the City and County of Denver from reservoir storage at Dillon to a tributary of the South Platte River at Grant. The tunnel passes beneath the Continental Divide at the crest of the Front Range and intersects rocks of diverse types and origins and geologic structures of extreme complexity. Geologic investigations preceded and were continued during construction and contributed materially to the safe and economical advance of the tunnel headings. Geological data obtained and evaluated during construction include plans and sections on a scale of 1 in. to 50 ft, rock temperatures at 2000-ft intervals, records of progress related to geology, records of ground-water conditions, summaries of the use of steel and timber supports, and summaries of grouting operations. Adverse conditions required use of steel and timber supports for 71.74% of the total length of the tunnel. Most supported sections reached a condition of equilibrium within a few hours to a few months after being excavated, and the concrete lining serves chiefly to provide an additional safety factor and to provide necessary hydraulic characteristics to the tunnel. Efforts to attain a maximum rate of advance of the tunnel headings necessitated use of more steel and timber supports than would have been required in a less accelerated tunneling operation. This report summarizes geologic investigations prior to and during construction of the Harold D. Roberts tunnel in Summit and Park Counties, Colorado (Fig. 1) and analyzes the tunneling operation as it is related to the various kinds of geologic conditions encountered in the tunnel. The tunnel is concrete lined to a circular cross section of 10.25 ft and is designed to transport 788 sec ft from reservoir storage at Dillon 23.3 miles along a dog-leg course passing under the Continental Divide to outlet and control works at Grant, Colo. The primary purpose of the tunnel is to augment the domestic water supply of the City and County of Denver, but plans include provisions for power generation at the outlet portal at a future date. Location of the tunnel was made after extensive preliminary geologic and engineering investigations by the Board of Water Commissioners of the City and County of Denver, and the U.S. Bureau of Reclamation. In 1946 the Board of Water Commissioners started tunnel driving on a limited scale from the east (outlet) portal and by 1955 had completed 9986 ft of tunnel. An urgent need for additional sources of water supply required acceleration of the tunneling operation, and, in 1955, Tipton and Kalmbach Inc. of Denver were assigned the responsibility of reviewing the project and preparing plans and specifications for the tunnel and appurtenant features. Additional geologic studies were made, and the geology of several possible alternate tunnel lines were compared before determining that the tunnel should be driven along the course previously adopted by the Board of Water Commissioners. On July 12, 1956, a contract for construction was awarded to Blue River Constructors Inc., a combine

Jan 1, 1962

By Earl Bardwell

THE structural relations existing between cuprous oxide and copper -were first systematically studied by Heyn1', who suggested that a study of the microstructure of refined copper might be substituted for the analytical methods in vogue for determining the oxygen content of copper coming from the refining furnace. Hofman, Green and Yerxa,2 in their paper A Laboratory Study of the Stages in the Refining of Copper, working along the lines suggested by Heyn, showed that a study of the microstructure could be made to yield a correct estimate of the oxygen content of the specimen of refined copper under examination. In order to determine the oxygen content of the specimen of refined copper under. examination a photomicrograph was first made. From the relative areas of copper and eutectic, the percentage of cuprous oxide and the corresponding percentage of oxygen were then readily calculated. It was found that the values for oxygen thus obtained corresponded closely with the analytical results. With the idea of utilizing this method of determining the oxygen content of each charge dipped from the refining furnaces at the Boston & Montana Reduction Plant, the physical testing laboratory was equipped with suitable apparatus for carrying out the work. It may be of interest to describe certain of the methods which have been worked out in this connection and give such data as have been obtained bearing on the effect of the oxygen content of refined copper upon its conductivity. Apparatus.-For the benefit of those who may be unfamiliar with metallographical apparatus and methods and may be interested to know what equipment is required, a few words with regard to the apparatus is in place at this point.

Jan 7, 1913

By C. D. King

The key to maximum production of ingots for the war effort is maximum production of pig iron. For any given furnace and ore, the most important single influence on blast-furnace production is the quality of the furnace coke. It is generally recognized that undoubtedly deterioration of Pittsburgh coking coals will be accelerated by the depletion of some of the better grades of coal; the extension of mine mechanization, with attendant degradation of coal quality; and the opening of new coal fields, some of which are known to be fairly high in sulphur. All of these factors, contributing as they do to higher ash and sulphur, and not infrequently to inferior quality of coke, made it desirable that an extensive study be undertaken to determine the effect of washing some coals that closely approximate the average chemical composition of coking coals to be expected in the future, by the subsidiaries of the United States Steel Corporation in the Pittsburgh district. It was deemed advisable to study the improvement obtained by washing such coals and to determine the need for supplementing the existing coal washery at Clairton to assure continued maximum production of pig iron from presently available furnace stacks. The existence of the cog! washer at Clairton made ii possible to conduct tests on coals used in raw form and on similar coals after washing. This study covers an investigation of the quality of furnace coke as affected by washing coking coals, and is confined to the use of IOO per cent high-volatile Pittsburgh seam coals from Fayette and Greene Counties in the manufacture of furnace coke. It covers the improvement in chemical and physical properties as well as the degree of uniformity in coke obtained by washing these coals, and includes the effect on blast-furnace production and practice. The test period involved is four months; the plants, Clairton by-product coke plant, and the Carrie No. I and No. 2 blast furnaces at Rankin, Pa., Carnegie-Illinois Steel Corporation. Clairton By-product Coke Plant and Coai, Washer The Clairton by-product coke plant carbonizes 30,800 net tons of coal daily in 22 batteries containing 1482 ovens. Three oven sizes are used, commonly known as wide, medium, and narrow types, recognizing the difference in oven width. These ovens also differ in height and length. The present coal washer has a daily capacity of approximately 14,000 tons, leaving I6,800 tons to be carbonized in raw form. Washed coal is used on all the narrow ovens and on three of the six batteries with ovens of medium width, raw coals being used pn all the remaining ovens. The customary arrangement is 7

Jan 1, 1944

By Kenneth M. Bravinder, Jonathan E. Koogle, Dean H. Sheldon

THE solution of subsurface geological problems requires an analysis of vertical and horizontal dimensional relationships. For many, the ability to visualize structures in three dimensions is not easily acquired, and for areas where the structural conditions are complicated by faulting the need arises for illustrative models. This has led to the many forms of peg models, in which the usual course of construction includes a peg for the well bore, and the suspension of strings from peg to peg to illustrate the formation correlation between wells, thus developing the usual folds and flexures. Such models proved unsatisfactory for the analysis of fault problems. Consideration was then given to the use of transparent materials. A model of the Dominguez oil field was constructed in 1935, which consisted of a series of glass panels from well to well on which the formations were shown in color. This model, though serving its purpose at the time, proved to be impractical from the standpoints of cost, breakage hazard, and extreme difficulty in maintenance. The cost of materials was estimated to be approximately $300, and the construction required the services of one man for a period of from 10 to 12 months. Maintenance after actual construction was an additional expense. Consideration was then given to the use of celluloid, and after some experimental work it was found that the advantages in its use were: (I) a short construction time, (2) ease of handling the material, (3) sections are easily reworked, (4) low cost. CONSTRUCTION It would be of little value to recount here the difficulties encountered in the building of the first celluloid model. The conclusions drawn, however, suggest the following procedure for those who contemplate the construction of such a model: I. Determination of the Scale.-Accurate geological interpretation requires the use of the same scale horizontally and vertically. It should be emphasized that the limited size of the celluloid sheets (20 by 48 in. maximum dimensions), the area to be studied, the maximum depths to be plotted, and area of possible future development are the controlling factors in adopting a scale to be used. 2. Construction of the Base.-A satisfactory base may be constructed by using two pieces of three-ply wood separated by strips of i by 2-in. material. This type of construction prevents warping of the base. The size of the base is determined by the dimensions of the area to be studied plus 5 in. on all sides. A map of the same scale as that adopted for the model, indicating the lines of section to be used, should be fastened to the base. 3. Construction of the Frame (Fig. I).-In the construction of the frame the use of 1/2-in. o.d. pipe welded in a box shape is suggested, with two horizontal rods on each of the four sides, top and bottom. The top rod should be at least I in. higher than the

Jan 1, 1941

By S. M. Moos, R. E. Dugger

In reviewing available vacuum filtration equipment in this paper, several types of filters have been included which are probably not familiar to operators in the mining industry. However, it was felt that these units might be of interest in the event that an application should arise for the filtration of a product in the future that does not exist today. In addition, exposure of our industry to equipment used in other industries might have a tendency to be thought provoking and perhaps lead to the development of equipment and/or applications in our industry which do not necessarily exist at this time. Most innovations in filtration equipment have been confined to the improvement or design change of components and accessories, particularly reflecting the availability of materials of construction that have been developed. Recent developments in the application of horizontal belt filters have been innovative. The application of this unit as a substitute for or an adjunct to thickeners in a CCD circuit can certainly be innovative and perhaps provide a basis for reduction in capital and operating costs in this area. For example, mill tails are generally pumped at the highest possible density that will adequately flow. Tailings ponds and dams with their attendant design requirements are based upon this estimated water volume. Availability of a much dryer, non-flowing filter cake opens up several new options for disposal. Considerable cost savings could result in the design and planning of disposal areas for containing dryer material. No attempt has been made in this paper to provide a manual for the operation of all of the many types of equipment that are available. Rather, the paper tries to present an overview. Fundamentals of Vacuum Filtration Filtration may be defined as the separation of solids from a liquid effected by passing the liquids through a porous medium. If the filtrate is induced to flow through the medium by hydrostatic head, we refer to it as gravity filtration. If we utilize super-atmospheric pressure upstream of the medium we call it pressure filtration and if sub-atmospheric pressure is applied downstream of the medium we call it vacuum filtration. Centrifugal filtration occurs when a centrifugal force is applied across the medium. Centrifuges will not be discussed in this paper. The process goal of filtration may be the production of dry solids where the cake is the valuable product, clarified liquid

Jan 1, 1980

By C. K. Stoddard, R. G. Knickerbocker, E. T. Leidigh, R. R. Lloyd, K. L. Mattingly

DURING July 1941, a study was initiated at the Boulder City Experiment Station of the Bureau of Mines on proposed methods for the production of magnesium metal. The major emphasis was placed upon development of a process that required minimum reliance upon the sale or utilization of by-products to meet economic competition from processes already in operation. It was felt that only through the development of such a process could the nation's vast resources of dolomite and low-grade magnesite that are contiguous to large blocks of firm power be properly utilized. During the course of exploratory tests on electrolytic and electrothermic processes, it was noted that fused magnesium chloride would dissolve appreciable quantities of magnesium oxide in the presence of small amounts of boric oxide. The use of such a solution in an electrolytic cell employing a graphite anode and a steel cathode caused the deposition of magnesium at the cathode and the evolution of carbon oxides at the anode with minimal amounts of chlorine The original results were sufficiently encouraging to justify further investigation on an enlarged scale, since the process seemed to present advantages over other proposed processes 1.2 for the direct use of magnesium oxide in a fused chloride electrolyte. When work on a larger scale was initiated it was found that the addition of boric oxide to the electrolyte resulted in dispersion of metal and constantly decreasing current efficiency. However, by this time solubility studies had shown that magnesium oxide was sufficiently soluble in magnesium chloride melts containing no boric oxide to effect the same anode reactions, and development work was continued. Over two years of development work have been carried out, and considerable work remains to be done. It is desired to emphasize the fact that many of the difficulties met require a long period of continuous operation before they become serious enough to affect operation of the cell. ACKNOWLEDGMENTS This paper is one of many reporting on various aspects of the Bureau of Mines' program initiated in August 1939 by passage of the Strategic Minerals Act the scope of which was greatly expanded by subsequent legislation. Some of these papers are published as war minerals reports. others as bulletins, technical papers, reports of investigations

Jan 1, 1945

By Neil H. Bostick

A specialized electronic data system processes data taken from normal gologic maps and stratigraphic tables to yield predictions of uranium resource location for an area such as much of the United States. Strata underlying specific quadrangle cells of the earth are evaluated for uranium favorability on the basis of selected criteria of lithology, structure, and stratigraphy but not specific past production in the area. Stored data are computer processed to yield volume of likely host rock beneath each cell to a given depth and at a given favorobility level. Results are plotted by machine directly as maps. A specialized and simplified procedure for predicting approximate uranium resources has been devised at Stanford Research Institute (SRI). The prediction is based on general geology beneath specific area cells of the earth in order to provide complete volume coverage; the strata underlying these areas are evaluated for uranium mineral favorability on the basis of a few selected geologic criteria in order to allow rapid data input. In order that the predictions be generally applicable, they are based on geological factors only, not past ore production of a given cell. Since the bulk of U.S. resources are in elastic host rocks, only such strata are considered in this work. This paper describes the structure and some details of the resource system and how it was tested in an area of the western United States.

Jan 1, 1971

By J. J. Mulligan

Introduction Coal is found in every major geographic region through-out Alaska and in two offshore areas. Since about 1852, more than 100 small mines have opened and closed. In 1981 and 1982 production was almost entirely from one mine in central Alaska. Clearly, the utilization of Alaska's abundant coal resources presents political, economic, and technical problems that are far from solution. Location and Physical Features Figure 1 illustrates the geographical and geological similarity between Alaska and the western US. Alaska's Arctic region generally resembles Wyoming and the Dakotas. The Brooks Range is an extension of the Rocky Mountains. The central region of Alaska between the Brooks Range and the Alaska Range bears a general resemblance to the basin and range country of the western US. The southcentral and southeastern regions have a similar resemblance to the Sierra Nevadas and the West Coast. The Arctic, the central, and the southcentral regions have major coal fields. The southeast region has scattered occurrences. The topographic, climatic, and economic characteristics of each are different. These differences have a large effect on practical mining methods. Description of Alaskan Coal Coal deposits known in Alaska are almost all of Cretaceous or Tertiary age, much younger than the Carboniferous coal deposits of eastern and midwestern states. Consequently, most Alaskan coals have a higher moisture content and a lower heat value. Most Alaskan coals also have a much lower sulfur content. Low volatile bituminous, and anthracite coals occur in Alaska, but most Alaskan coals range downward in rank from high volatile bituminous to subbituminous and lignite. Alaskan coals are similar to coals of the northern Great Plains and the Rocky Mountains. Arctic Region Northern Alaska Coal Fields Northern Alaska coal fields are by far the largest group of coal deposits in Alaska (Fig. 2). Coal-bearing rocks extend from the foothills of the Brooks Range about 160 km (100 miles) northward to within about 80 km (50 miles) of the Arctic coast. The principal known occurrences are between the Colville River and the western coast. However, coal has been found eastward almost to the Canning River and westward under the Arctic Ocean. The northern Alaska coal fields are in a series of Cretaceous rocks that suggest a deltaic depositional environment. The beds tend to occur in series of up to eight or nine beds totaling 6-9 m (20-30 ft) of coal. However, beds 6-12 m (20-40 ft) thick have been reported, and beds 1-3 m (3-10 ft) thick seem to be relatively common. Including thin seams, coal in some places makes up about 10% of the stratigraphic section. However, no area has been explored in sufficient detail to block out any large body of reserves or to make reliable estimates of the total resource. Observation points in drill holes and along river and stream banks are few and far between. Coals in northern Alaskan fields tend to range in rank from bituminous, where moderately folded and faulted in the foothills, to subbittuminous in the gently folded rocks under the coastal plain. Coals toward the eastern extremities of the coal fields may be of Tertiary age and lower in rank, but the same general correlation between degree of folding and increase in rank can be expected. Some higher rank coals in the foothills belt are known to have coking properties but most, and possibly all, require

Jan 1, 1984

By E. R. Borcherdt

Methods and equipment are subject to study in the Butte district at all times in order that advantage may be taken of any new developments that might serve to increase efficiency and lower costs. Detailed time studies of each operation involved in the entire mining process have been of invaluable assistance in analyzing the individual operation for possible improvement. In other words, the fact that a practice has been in vogue for many years does not exempt it from change provided an easier and cheaper method can be found. Exploration Ore occurs in the Butte district in such manner that structures must be drifted on, no matter how small or unimportant they appear when intersected by a crosscut. Parallel structures also add to the crosscut requirements. This prospecting work, which formerly was done entirely by drifting and crosscutting, is now being supplemented by diamond drilling, to aid in greatly decreasing the amount of waste. In long-shot prospecting, or in second or third class prospecting, where the chances of developing commercial ore are not very great, diamond drilling is of particular value. In prospecting where chances of encountering commercial ore are certain or where crosscuts can be used later for haulage or ventilation, drilling obviously is not applicable. Improvement in drilling machines and bits has contributed greatly to the present success of diamond drilling. For many Years square-set filled stoping was the principal mining method used in the Butte district. Later timbered and untimbered rill stopes were used, also some horizontal cut-and-fill stoping where both Ore and walls were particularly strong. Because of the amount of timber required for square-set stoping, and the relatively low production per stope unit, experiments with horizontal cut-and-fill stopes were begun in Ore bodies of which many had fairly weak walls. It was felt that increase in the speed of the stoping operation, followed closely by filling, to avoid weight being produced by permitting ore and walls to remain insufficiently supported over comparatively long periods, would maintain the walls and backs in safe condition without the use of timber and without ore dilution from wall slough. Cycles of drilling, scraping ore, and filling were devised, which have materially increased stope production and decreased timber consumption, so that both stope labor and supply costs were lowered appreciably without materially affecting grade control. In the larger mines of the district, where wide veins favor this stoping method, half of the total production is now being mined by horizontal cut-and-fill stopes (Fig. I). The increase of the filling rate demanded by the increased stoping rate was the greatest obstacle encountered in the accelerated stoping operations. Finally this problem was solved by replacing the time-honored method of transferring filling waste by small cars, which were hoisted or lowered in the auxiliary shaft, by a transfer

Jan 1, 1943

By E. F. La Vigne

On Aug. 10, 1884, the Retsof Mining Co. began the sinking of an 18 by 12-ft. shaft at Retsof, N. Y. Rock salt was reached in 1885, in September, at 1008 ft. below the surface. The first salt was shipped in the fall of 1885. A timber breaker was built in 1886. In July, 1895, there was a consolidation of the Retsof Mining Co., the Greigsville Salt & Mining Co., the Livonia Salt Co., and the Lehigh Salt & Mining Co. These consolidated salt companies are now known as the Retsof Mining Co., a subsidiary of the International Salt Go. Eventually all of the mines were closed except the mine of the Retsof Mining Co., at Retsof, Livingston County, New York. On July 6, 1921, about 1/2 mile south of the original shaft arid plant of the Retsof Mining Co., a new 9 by 28-ft. concrete-liried shaft was started. The shaft is 1063 ft. deep from the top of the shaft collar to the bottom of the salt bed. At the same time a reinforced-concrete breaker and other buildings were erccted. This new plant began operations on Dee. 15, 1923. It was designed for a capacity of 3000 tons in 8 hours. Geology The workable rock-salt bed is 1063 ft. below the surface and 326 ft. below sea level, and varies in thickness from 9 to 10 ft. The total thickness is greater than that but is contaminated with shale above and below the pure mineral. The characteristics of the formation are shown in Tables 1 and 2. The bed dips uniformly 0.87 per cent to the south. Prospecting As the salt deposil in New York State is known, very little prospecting has been necessary. Before the sinking of the Retsof No. 1 shaft, a well was drilled to prove the existence of salt. No. 1 shaft was sunk where the well was located. The several salt wells drilled in Livingston Courlty proved generally the existence of salt throughout the county. (See U. S. Geol. Survey Bull. 669: Salt Resources of the U. S.)

Jan 1, 1938