Search Documents

By R. R. Angel

Drilling rate is a parameter that should be considered in determining the volume requirements for air and gas drilling. The use of past methods which ignore the effects of the solids content upon the pressure and velocity of the annulus flow stream can result in undercalculation of the required volurne by as much as 50 per cent. A vertical-flow equation is presented for determining volume requirements. This equation includes the effect of the solids that are transported up the annulus in the flow stream by incorporating the drilling rate as one of the parameters. The effect of down-hole temperature on required circulation rates is also ana-lvzed. A simple approximate method of determining volume requirements is presented. This method is more accurate than the methods used in the past. Hole cleaning difficulties are analyzed for a recent air drilling job where past methods indicated that - excess air was being used. Sample curves of calculated boitom-hole pressures are presented for air and gas drilling in several hole sizes. INTRODUCTION In certain areas the use of air or natural gas as a circulating medium for drilling oil and gas wells is becoming a common practice. Large increases in penetration rate and bit life are achieved through the use of these media in preference to mud or water. Drilling rates as high as 90 ft/hr have been obtained in shales. The importance of maintaining adequate circulation rate is generally recognized; however, much disagreement exists among drilling operators as to what constitutes "adequate" circulation rate. In quarry drilling, where annular velocities can be accurately determined, an annular velocity of 3,000 ft/min of standard air is required for best results in rocks having approximately the same density as those commonly penetrated in drilling oil or gas wells'. Although this standard air velocity has proven satisfactory for quarry drilling, some oil and gas well drilling operators believe that an equivalent annular velocity of more than 4,000 ft/min is required; others believe that as little as 2,000 ft/min . is sufficient. Much of this disagreement results from determining the required circulation rates with methods which fail to incorporate the drilled solids in an equation which is applicable to vertical flow. Hughes Tool Co. Bulletins No. 23' and 23-A3 present data for determining circulation rates based on the Weymouth formula. These data do not include the drilling rate as a parameter and, therefore, neglect the effect of the solids being transported up the annulus. In spite of this apparent defect and the fact that the Weymouth formula is not applicable to vertical flow, the Hughes data have well served the drilling industry in many areas and are important and timely contributions to the science of air and gas drilling. The Hughes data purposely omit a correction for increasing down-hole temperature. At slow drilling rates this effectively compensates for the use of a formula which is not valid for vertical flow; however, volumes determined by the Hughes method are not sufficient to support rapid drilling rates at moderate and great depths. For example, Phillips Petroleum Co.'s Cauthorn "D" No. l in the Vinegarone field for Val Verde County, Tex., was air drilled from 1,500 to 9,300 ft using a compressor delivering 1,400 cu ft/min. The 8 3/4-in hole was drilled with 5-in. drill pipe, and drilling rates as high as 90 ft/hr were obtained between 7,000 and 9,300 ft. No water or caving hole was encountered. At 7,728 ft it was necessary to wash-out 60 ft of cuttings to reach bottom after a trip to change bits. At 8,130 ft a twist-off occurred and the drill col-lars were stuck in drill cuttings. These difficulties indicate that the 1,400 cu ft/min of air was not sufficient to keep the hole clean. Hughes data indicates that 1,180 cu ft/min is sufficient to produce an annular ve-

Jan 1, 1958

By P. E. Busby, C. Wells, M. E. Warga

Fundamental data on the rate of diffusion of boron in austenite and solubility of boron in the a and y phases of iron and steel have been obtained from deboronizing experiments and provide partial explanations for some of the phenomena observed in boron steels. The rate of diffusion of boron is about the same as carbon in austenite. The solubility of boron in austenite at normal heat-treating temperatures is less than 0.001 pct. A partial tentative Fe-B phase diagram in the important low boron concentration ranges and an equation representing the diffusion of boron in austenite are presented. A LTHOUGH a search of the literature revealed ALTHOUGHthat boron diffusion had not been studied quantitatively and systematically prior to 1948, a few qualitative observations which were made by following the diffusion of boron into iron packed in ferro-boron1-3 suggest that boron obeys normal diffusion laws and that the rate of diffusion increases with increasing temperature. The claim of Cornelius and Bollenrath1 that boron does not appear to influence the rate of carbon diffusion has recently been substantiated by Wells, Batz, and Mehl.4 Calculations based on data from the paper by Campbell and Fayv indicated that the diffusion coefficient (D) for boron in iron at 900°C is approximately 3x10 sq cm per sec; a value of 2x10-7 sq cm per sec at 1038°C has been reported by Digges et al in connection with decarburizing experiments on a commercial boron steel containing 0.43 pct C. During the period 1948 to 1950 several diffusivity constant (D) values for the diffusion of boron in the y phase of iron and steel and a number of solubility values for boron in both a and y phases were reported. It was concluded tentatively as a result of Metals Research Laboratory studies that D values for boron in y and a phases are about the same as for carbon in the comparable phasesa and that carbon up to 0.4 pct did not affect the rate of diffusion of boron in y iron above 1000°C within the limits of experimental error. Whether saturation of the y or a phases with carbon would significantly affect rates of diffusion of boron through them was then and still remains in doubt. The best estimate of Q (activation energy) was reported to be about 25,000 cal, somewhat lower than that for carbon in y iron. The solubility of boron in y iron at 1000°C was thought to be higher than 0.004 pct B but not as high as a more recent analysis of available data shows it to be. The solubility of boron in a iron at 700°C was reported to be 0.0004 pct B, but it would not have surprised the authors if the true solubility value is actually lower than this. Breaks in diffusion curves which were not understood when first obtained prior to 1950 have now been recognized as indicating solubility limits, and these are included in the present paper. Experimental Procedure Two methods were employed in an effort to determine diffusion coefficients (D) for boron in austenite. Early experiments involved the use of welded couples prepared in accordance with the procedure of Wells and Mehl,' but when the results of these experiments proved to be unsatisfactory for calculating D values, studies were continued by means of deboronizing experiments. In both cases, boron analyses following the diffusion anneal were carried out spectrographically and all D values were computed using the Grube solution of Fick's law. The precision of the spectrographic method used proved to be about equal to that reported by Corliss and Scribner" —average deviation from the mean about 5 pct of the amount present when the boron content is 0.003 pct. Concentration-penetration curves obtained from spectrographic analyses of welded couples generally indicated that practically no boron was transported across the weld interface during the diffusion anneal. Furthermore the concentration-penetration

Jan 1, 1954

it may be clearly seen that one of the most rapidly increasing items in mining operations is the cost of the timber alone, to say nothing of the cost of installation. Consequently, any practical means that may be used to lengthen the life of mine timber will greatly tend to decrease mining costs and result in important economies in operation. The four principal agents by which mine timbers are destroyed are, decay, insect attack, fire and mechanical abrasion. Decay and insect attack are credited with more than 50 per cent of the total amount of timber destroyed in the mines. Numerous methods may be used to conserve mine timbers, such as better selection, preparation, storage and segregation. It frequently happens that good timber goes into places where poor timber would do as well, and timber of inferior quality is sent into places where only the best grade should be used. However, the most effective method of conservation is by treatment with some standard preservative to prevent decay. Some of. the benefits to be noted from the use of preserved timbers are; the reduction in operating costs by decrease in timber purchases; cuts down interference with mining operations due to timber failures on main haulageways; prevents excessive falling of roof in entries by maintaining timber in sound condition; conserves timber supplies and makes possible the utilization of inferior woods; reduces the danger of mine accidents; improves mine ventilation; lessens fire hazard. Our experience thus far with preservatives, while not of a sufficient period of time to establish just what length of life we may obtain from timber so treated, results obtained during the past four years indicate that the intelligent use of preservatives will effect sufficient economies in mine timber costs to justify the expenditures made for this purpose. MR. BAKER (Concluding): Now, if there are any questions you gentlemen would like to ask about our particular operations, I will attempt to answer them. J. H. EMRICK (Sullivan Machinery Company, Denver, Colo.) ; I was wondering how much penetration you get. MR. BAKER: We get between one-sixteenth and one-eighth of an inch; it is a dipping process. With pressure treatment you get-a greater absorption; that is, better than one-sixteenth of an inch. As long as there is no mechanical abrasion, it seems to serve the purpose. MR. EMRICK: Does that apply to the condition of the timber before it is treated? If it is perfectly dry, will it absorb more treatment than if it is partly green? MR. BAKER: Certainly; one of the things we do is to have it thoroughly seasoned before it is put into the preservative. We put some green timber into the mine once, and the fungus went after it right away. MR. EMRICK: What was the difference in the depth reached by the preservative? MR. BAKER: It hardly penetrated at all when it was green. It has to be sound and well seasoned. MR. EMRICK: How long was the timber dipped or treated? MR. BAKER: According to the seasoning. If it was well seasoned, it was left in the hot bath for an hour; if not so well seasoned, it was left for about a couple of hours. C. M. BILLINGTON (General Superintendent, Rocky Mountain Fuel Company, Lafayette, Colo.): Have you found it necessary to treat timber that has been charred? MR. BAKER: No. However, I suppose it has the sane effect; in lots of places they char timber for the purpose of treatment. MR. BILLINGTON: Have you ever tried lime and salt? MR. BAKER: No, this is the only thing we have tried, the creosote and the Aczol; the Aczol is a foreign product. WM. LITTLEJOHN: Which I is preferable, the Aczol or the creosote? MR. BAKER: I couldn't say as to that. MR. LITTLEJOHN: Have you had any tests as to the value of either? MR. BAKER: We have had some of the Aczol timbers in service for a little over four years; we have had some creosote timbers in service

Jan 1, 1926

By G. M. Schwartz, G. A. Thiel

THE first record of the quarrying of dimension stone in Minnesota dates back to 1820 when limestone was quarried locally for part of old Fort Snelling. Limestone quarries were operated at Stillwater, Mankato, and Winona as early as 1854. Granite was quarried first at St. Cloud in 1868, and within a few years thousands of tons were shipped to widespread points. Rough dimension stone for large buildings furnished the first important market, but beginning in 1886 paving blocks were in demand. The largest shipment was in 1888, when 1925 cars were shipped from the St. Cloud area. Quartzite was quarried first at New Ulm in 1859 and somewhat later at Pipestone and elsewhere in southwestern Minnesota: The productive dolomite quarries at Kasota were opened first in 1868 and have continued as large producers of a variety of stone to the present time. At present, the industry is controlled by relatively few operators, and for that reason detailed figures on dimension stone are not released for publication. A general idea may be obtained from the data in the Minerals Yearbook for 1948. The figures for total stone produced in Minnesota are 1,804,000 tons valued at $5,090,652. Probably the largest item in the latter figure is received from dimension stone. A better idea of the situation, in relation to the country as a whole may be gained by using the data for 1930 when more companies were operating in Minnesota, and complete figures were published. In that year Minnesota produced granite valued at $2,668,119 and ranked third among the states in value. Minnesota's production of granite was almost exclusively for dimension stone. In the same year Minnesota produced 300,000 tons of limestone (dolomite) valued at $840,860, and this likewise was mainly dimension stone. In finished limestone Minnesota ranked second among the states in 1930. Sandstone and minor amounts of quartzite are the only other dimension stones that have been produced in Minnesota, but the quarries are now inactive. The commercial stones of Minnesota have been described in two reports by Bowled and by Thiel and Dutton.2 The early history of quarrying in Minnesota and extensive notes on the various rocks are given by N. H. Winchell.3 Small limestone and dolomite quarries were numerous throughout the area of Paleozoic rocks in southeastern Minnesota. Early production was largely dimension stone. With the increased use of Portland cement, most of these ceased production, and today only those at Kasota and Winona remain in operation. In recent years many quarries have reopened and new ones started, but these are devoted to the production of crushed rock and agricultural lime. As the application of modern quarrying and finishing methods increased, small companies in the granite business have dropped out, and the remaining companies have modernized their plants, purchased old quarries, and opened up new ones, thus furnishing a wide variety of granites suitable for most of the customary uses. It is the purpose of this review to present notes on' the geology and operations of each of the quarries now operating within the state. Granites and Related Igneous Rocks The term granite as used in this report includes granites, gneisses, diorites, gabbros, and other igneous rocks. The granites of greatest economic importance are found in three widely separated regions, see Fig. 1. 1-Central Minnesota in the region of the city of St. Cloud, 2-the upper Minnesota River valley region, 3-the northeastern portion of the state, commonly referred to as the Arrowhead region. The St. Cloud Region: The rocks of the St. Cloud region are mainly granites and related rock types such as monzonites and quartz diorites. The stones may be grouped into three major types, namely, pink granite, red granite and gray granite. Most of the pink granite occurs in the area to the southwest of St. Cloud. The rock is best described as stone with large pink crystals set in a finer grained black and white background. The minerals of the matrix occur in remarkably uniform sizes, and the pink crystals are sufficiently uniform in. their dis-

Jan 1, 1952

By Harlan W. Nelson, James B. Purdy

A study of data obtained during laboratory tests to determine the suitability of bituminous coals for use in residential underfeed stokers of the clinkering type has led to the following general conclusions: 1. Performance data obtained during the test furnish useful and practical information upon which to judge the performance of stoker coals. 2. Conclusions derived from the results of evaluation tests are in general agreement with use experience of the coals in the field. The method is of practical use in this regard. 3. Few direct relationships were found between data furnished by the evaluation test and determinations by standard laboratory tests. This is particularly true with regard to coke formation and clinker formation, two of the most important factors in determining the degree of satisfaction to be obtained in applications of the small underfeed stoker. It has not been the intention in this paper to prescribe this evaluation test as the best method for the evaluation of coals for use in residential stokers. That there are other test methods having similar objectives serves to point out the necessity for full-scale laboratory tests. At the present time, a standard test procedure for the evaluation of coals for use in residential stokers is not available. For the past several years a joint committee, appointed from the membership of the Residential Stoker Committee of Bitulninous Coal Research, Inc., and from the Engineer- ing and Research Committee of The Stoker Manufacturers Association, has been working to develop a standard procedure for testing and evaluating bituminous stoker coals. This procedure has been completed to the extent that it has been drawn up in tentative form. It is now planned to install the equipment in several laboratories and to run check tests, using portions of identical lots of coal, to determine the degree of reproducibility offered by the method. Completion of this standard test procedure should greatly facilitate the evaluation of coals for use in residential stokers. Acknowledgments The cooperation of coal operators sponsoring the coal evaluation tests, in granting permission to use data and results from their tests, is gratefully acknowledged. Acknowledgment is also made to Carroll F. Hardy, Chief Engineer, Appalachian Coals, Inc., for his interest and timely suggestions on the preparation of this paper. Special thanks are due to Ralph Sherman, Assistant Director, Battelle Memorial Institute, who was largely responsible for development of the evaluation procedure, for helpful suggestions and advice. References I. H. A. Sherman: The Evaluation of Coal for Use in Domestic Stokers. Univ. of Ill., Eng. Experiment Station, Circular No. 39 (1939). 2. . Q. Shotts: The Relation of Free- swelling Indexes to Other Characteristics of Some Alabama Domestic Stoker Coals. TP 2314, Coal Tech., (Feb. 1948); Trans. AIME (1948) 177, 502. 3. . J. Helfinstine and C. C. Boley: Correlation of Domestic Stoker Combustion with Laboratory Tests and Types of Fuels 11. Combustion Tests and Preparation Studies of Representative Illinois Coals. Ill. State Geol. Survey, Re1. of Zntlestigations No. 120 (19467 ) p. 35. DISCUSSION E. R. Kaiser*—AS a former fuel engineer assigned to residential stokers at Battelle, the writer has read with considerable interest the paper by Messrs. I'urdy and Nelson. It is gratifying to note that earlier test techniques have been developed and that criteria found important in the years 1935 to 1938 are still important in judging the performance of stoker coals for residential heating. The authors have plotted a number of the primary coal data singly against performance data of the stoker in an effort to establish relationships. Unfortunately, wide scattering of points usually resulted, which did not illustrate what experience generally indicates to be the case. For example, Fig 3 indicates almost no trend in the percentage of released ash converted to clinker with change in ash-softening temperature. Other factors, such as fuel bed temperatures and zones of heat release must have influenced the results. In the present state of our knowledge, we cannot explain all of the reasons why suitable stoker coals perform satisfactorily, nor why one satisfactory coal may be better than another. It would therefore

Jan 1, 1950

By G. B. Thomas

Several methods of analyzing pressure build-up data in wells have been presented by various authors. This paper reviews the theory and method of D. R. Horner and presents example calculations performed on data obtained by testing several different types of wells. These calculations include, (1) graphical estimation of final static pressure, (2) determination of the productive capacity of the pay away from the well bore and. (3) the degree to which the formation adjacent to the well bore has been damaged by completion or other causes. The methods of testing and precautions which should be taken to assure the best data possible are discussed. Limitations and reliability of calculated results are also treated. INTRODUCTION Pressure testing of wells is generally limited to the determination of producing and static mean formation pressures. The so-called "static" pressure determination.. along with PVT, electric log and production data, enable the reservoir engineer to determine, within reasonable limits, the drive mechanism of the reservoir and in some cases. the amount of edge water encroachment. Producing pressure tests enable calculation of productivity indices and allow the engineer to plan the systematic production of a pool for optimum conservation of suhsurface energy. The radial flow formula advanced by Muskat' has been based on the assumption of incompressible radial fluid flow. It has been known that reservoir fluids do not behave in an ideally incompressible manner. For example, incompressible flow theory indicates a simple logarithmic relationship between the difference of the instantaneous and static well pressures when plotted against time. 'The latter stages of this type of plot of pressure build-up data generally show a marked devia-, tion from the earlier straight line trend. which deviation may be shown to be due to the compressible flow of fluids toward the well bore. Shut-in times of 24, 48. 72. or at most 96 hours are currently in wide use for determining so-called "static" reservoir pressure. Due to the continuation of compressible flow of fluids into the well bore long after this arbitrarily taken shut-in time, the determination of static pressures has almost invariably resulted in lower than equilibrium values. Materials balance calculations made early in the life of a reservoir often result in a calculated reserve which later observations prove to be too low. Failure to obtain reliable "static" reservoir pressures within the prescribed 24 or 48-hour build-up period has undoubtedly been a major factor in obtaining these low estimates. Comparison of theoretical and actual productivity performance has indicated that formation damage and not being able to attain true static pressures have been partially responsible for the observed discrepancies. Research into the theory of compressible flow behavior has resulted in methods of applying these theories to the testing of wells. Application of the developed theories to pressure build-up performance indicates that in many cases the following can be estimated. 1. The static reservoir pressure. 2. The in-place effective formation permeability away from the well bore. 3. The degree or extent by which the formation has been damaged adjacent to the well bore, either through completion methods or subsequent damage due to fluid entry. REVIEW OF THEORY Horner2 has shown that the pressure build-up within a "point source" well is approximated by the following formula: Equation (1) is the approximate "point source" solution to the radial compressible flow equation advanced by Muskat. The solution assumes the following: 1. A point source well is producing at constant rate from the center of an infinite reservoir with a constant pressure at its external boundary. 2. The fluid flowing is present in one phase only. 3. The compressibility and absolute viscosity of the fluid remain essentially constant over the range of temperature and pressure variation encountered. 4. The well i.; shut-in at the sand face arid there is 110 after production into the well bore. 5. The formation permeability is homogeneous in the direction of flow. From Equation (1) it can be seen that if an ideal well were shut-in while producing from a reservoir under the conditions assumed, the pressure build-up would be a logarithmic func-

Jan 1, 1953

By A. Romero-Juarez

This paper is concerned with the effect of propping-agent concentration on flow capacity of a fracture in the case in which there is embedment of the propping agent. Previous published studies have shown definitely that there is a relationship between fracture flow capacity and prop ping-agent concentration, and it has been shown that the theoretical results are confirmed by laboratory experiments. The problem of directly finding sand concentration for maximum flow capacity of a sand-propped fracture is solved, and formulas and charts are given to obtain this concentration under various conditions of effective overburden pressure, medium sand-grain diameter and rock properties. It is shown that, for the same effective overburden pressure and the same rock characteristics, optimum sand concentration in pounds per gallon does not depend on medium sand diameter. For conditions met in bydraulic fracturing operations, it is found that sand concentration in grains per square inch for maximum flow capacity varies within a wide range of values; this indicates the convenience of using data of fracturing pressure and rock characteristics for calculating sand concentration in order to achieve the best results in fracture treatments. INTRODUCTION It has been pointed out in the published literature1 that one important factor controls the success of a hydraulic fracturing operation — the propping of the fracture. The main effect of the propping agent is to hold the fracture open by means of the reaction forces that oppose the pressure due to the overburden. It is assumed in this paper that the propping agent is sand, that the grains are spherical and uniform in size, and that this sand is distributed in a monolayer in the fracture. The existence of a sand concentration value for maximum fracture flow capacity has been recognized since the publication of one of the first studies on the subject.2 In the reference cited, flow capacity was assumed to be proportional to the cube of the fracture volume per square inch not filled by sand. The obtained experimental curves relating to flow capacity and sand concentration had the same appearance as the curves obtained with the afore-mentioned assumption. Recent work3 has shown that fracture permeability, in the case of a monolayer, can be calculated accurately by means of a modified Kozeny-Cacman equation; thus, it can be shown that fracture pqmeability is a function of free volume in the fracture and of the rock and sand wet surfaces. The flow capacity of a fracture is dependent on the permeability of the fracture and on the fracture width. If all excessive quantity of sand is present in the fracture, its width may be large but its permeability i:; low due to the greater flow resistance; too few grains of sand per square inch are present in the fracture, sand embedment in the rock is greater and the width becomes small enough to appreciably decrease the rate of flow. To determine optimum conditions of flow, therefore, it is necessary to know the dependence of fracture width on density or concentration of propping agent. This relation is given in a study4 in which the approach to embedment is similar to the one made in a penetration hardness test of metal. With the equacions presented by Darin and Huitt,3 the number of sand grains per square inch that results in maxirnum fracture capacity can be determined graphically by plotting fracture capacity vs sand concentration. Since the relationship between fracture capacity and sand concentration may be considered known, one is led to formulate the analytical problem of directly determining sand concentration for maximum fracture flow capacity, in order to study the effect of the various quantities that come into the process. The solution of this problem appears to be interesting both from the viewpoint of fluid mechanics and from the practical viewpoint of oil production in hydraulic fracturing treatments. EQUATION FOR FRACTURE FLOW CAPACITY It can be shown3 that the fracture flow capacity of a monolayer is given by the following expression,

By Mark A. Krall, Robert L. Geddes, James C. Frost

Introduction The Conda Partnership's Dry Valley phosphate mine is a thinly bedded, multiple seam open-pit mining operation where selective mining techniques are used to recover phosphatic shales. The mining methods used are truck/shovel and scraper/dozer operations. Ore is shipped 32 km (20 miles) by rail to a beneficiation facility. The ore is upgraded by washing and calcining. The mine and beneficiation complexes are owned by the Conda partnership. It is a joint venture between Beker Industries Corp., of Greenwich, CT, and Western Co-Operative Fertilizers (US) Inc., of Alberta, Canada. The Partnership operates as a separate entity of the two partners. The Dry Valley mine is located 48 km (30 miles) northeast of Soda Springs in Caribou County in southeastern Idaho. The mine is situated on the Caribou National Forest. Mining operations take place between 2 and 2.4 km (6400 and 7900 ft) in elevation. It is accessible partly by 32 km (20 miles) of paved roads and 16 km (10 miles) of dirt roads. The winters are long and severe, and the summers are short and mild. This article describes the history, geology, exploration, mining, and reclamation that makes this mine Idaho's largest producing mine and the western US' leading phosphate producer. History and production In the mid-1950s, Western Fertilizers of Salt Lake City, UT, drove an exploratory drift in Maybe Canyon. A large bulk sample of phosphatic shales was analyzed for phosphate content and processing characteristics. No large scale mining or processing operations were undertaken. In the late 1950s, the Dry Valley property was sold to Central Farmers of Chicago, IL. No major operations took place. In 1964, Central Farmers sold the property to El Paso Products Co. of Odessa, TX. El Paso Products supervised the mining operations of Wells Cargo Mining Co. from 1965 through 1967. During this time, El Paso Products built a beneficiation facility and a fertilizer complex in Conda. A 32-km (20-mile) railroad was also constructed from the mine to this facility. From 1968 through 1972, the mine was shut down due to a depressed fertilizer market. In 1972, El Paso products sold its ore reserves, beneficiation plant, and fertilizer complex to Beker Industries Corp. In 1979, Beker Industries sold 50% of its ore reserves and 50% of its beneficiation plant to Western Co-Operative Fertilizers (US) Inc., of Alberta Canada, forming the Conda Partnership. It has operated the mine and beneficiation plant since January 1979. From the mid-1950s to the mid-1960s, no substantial production took place. From 1965 to 1967, El Paso Products stripped 3 Mm3 (4 million cu yd) and mined 2.3 Mt (2.5 million st). From 1972 through 1983, 50 Mm3 (66 million cu yd) were stripped and 18 Mt (20 mil¬lion st) were mined. Geology The Wells Formation forms high ridges and hillsides in the Dry Valley area. It is best exposed along the west face of Dry Ridge. It forms the imposing wall on the east side of Dry Valley. The formation is divided into two members. The lower member, about 213 m (700 ft) thick, is dominantly thin to medium-bedded limestone and silty limestone. It contains nodules and stringers of chert and minor sandstone. The upper member is composed principally of thick-bedded to massive cross-bedded, light-gray to orange-yellow, fine grained sandstone. There is some interbedded brown to light-gray limestone. This member varies from 369 to 457 m (1300 to 1500 ft). Recent investigations indicate that the upper Wells is of Permian age. Under some conditions, the Wells may be water-bearing. Otherwise, it has no apparent economic significance. Grandeur Member (Park City Formation) Overlying the Wells Formation is a distinctive light-gray to white dolomitic fossiliferous limestone. This unit has been identified by the US Geological Survey (USGS) as the Grandeur Tongue Member of the Park City Formation. This member is sometimes absent due to its contact with the Meade Peak Member of the Phosphoria Formation. It is easily detectable by its color, hardness, and fetid odor. Phosphoria Formation The Phosphoria Formation of Permian age was named from Phosphoria Gulch, Bear Lake County. The formation has been studied extensively and developed for its economically valuable phosphate reserves.

Jan 11, 1985

By L. D. Jaffe

From the limited experimental data in the literature, preliminary values were derived for the thermal diffusivity of titanium alloys and for the quenching severity of various mediums used in heat treating them. The thermal diffusivity, under quenching conditions, was found to be approximately 0.006 sq in. per sec. The quenching severity H for brine is approximately 8, for water 4, for oil 0.9, for air 0.08, and for argon 0.02 in. -' The quenching severity of the Jominy-Boegehold end quench water jet against titanium alloys appeared effectively infinite. Using these results, graphs were prepared showing ideal round sizes and Jominy positions having cooled conditions equivalent to those at the center and surface of titanium alloy rounds, plates, and sheets quenched in various media. THE importance of proper quenching of steel parts has long been recognized, and much quantitative work has been done on the subject. Recently, it has become evident that for titanium alloys, too, the rate of cooling from high temperature is important in obtaining high strength and hardness. Maximum hardness and strength may not be attained in the as-quenched condition but rather after a subsequent tempering or aging treatment. Nevertheless, if cooling from temperatures near the ß transus is too slow, for the hardenability of the alloy used, a soft a-ß structure will be produced which will not harden on subsequent aging.'-" The quenching of parts of practical interest can usually be approximated by simple shapes, such as cylinders (rounds) and plates (or sheets) of various thickness and combinations of these. A standard that has been used extensively for comparison purposes is the center of an ideal round: a cylinder quenched in a medium that instantly cools its surface to room temperature." For steel, considerable practical use has been made of the Jominy-Boegehold end quench hardenability specimen,".' and its value for titanium alloys is becoming evident. It therefore appeared worthwhile to attempt comparison of cooling conditions in titanium alloy rounds and plates of various sizes quenched in various media with cooling conditions in ideal round and Jominy quenches. These comparisons are based upon measurements reported in the literature. Quenching Severity of Jominy Water Jet and of Brine As one starting point, there were available the data of Phillips and Tobin8 for positions on a Jominy specimen having the same hardness as the centers of round bars of various size quenched in 5 pct NaOH brine, shown in Table I. The Jominy speci- men is a round bar, normally 1 in. diam and 4 in. long, quenched by a water jet at one end and air cooled on the other surfaces.' After quenching, hardness is measured vs distance from the quenched end on longitudinal flats ground 0.015 to 0.10 in. deep. The data of Table I were based on a subsized specimen of 3/4 in. diam. Since the positions tested are close to the jet quenched end, cooling there may be taken as due solely to the jet, and air cooling can be neglected. These positions are then equivalent to various locations near one surface of an 8 in. plate jet quenched on both sides. Assuming that the thermal diffusivity is constant and that Newton's law of cooling holds, temperature vs time curves at corresponding positions in Jominy specimen and round bar then may be calculated for various assumed quenching severities of the water jet and of the brine. Cooling curves are generally taken to be metallurgically equivalent when they show the same time to reach a temperature halfway between the initial (heat treating) temperature and the final (quenching medium) temperature.5 On this basis, using Russell's tables,9 upplemented by heat flow tables derived in similar fashion by the author for positions near the surface of severely quenched plates, it was found that the experimentally observed equivalence, Table I, was best matched if the quenching severity H of the Jominy water jet was taken as infinite (H = co). This means that the jet quenched surface should be considered as instantly cooled to the temperature of the water

Jan 1, 1956

THE Secretary of Commerce is to be urged to as-sume leadership in assembling data as to produc-tion and consumption of minerals and metals, and studies of their importance in trade, as a result of the joint conference of the Institute and the Mining and Metallurgical Society of America, held on Tuesday afternoon, Feb. 15. With no set program and no especial effort to promote interest 120 members of the two soci-eties assembled and for two hours discussed in the most informal and frank manner the whole situation. George Otis Smith acted as chairman, Clinton H. Crane, who was to have served in that capacity, having been called abroad unexpectedly. Major Taylor for the Army and Capt. R. E. Bakenhus for, the Navy paid high tribute to the value of the reports already made by the committees of the two societies and pointed out the need of addi-tional data and of constant review of the figures, as a measure of national defense. J. E. Spurr, R. C. Allen, W. O. Hotchkiss, Fred Hellmann, W. Spencer Hutchin-son, P. N. Moore, A. C. Lane, E. E. Ellis, and others discussed various phases of the subject, reviewed what had so far been done, what was now in progress, what had been planned for the future, and the necessary limitations of the work. A. W. G. Wilson related the experience of the Canadian Government in this field, emphasized the importance of studies of the rate of consumption, and brought up the special case of the waste of the sulfur in sulfide ores with correspondingly increased demand on raw sulfur supplies. Mr. Lane pointed out that conservation, to be soundly based, must recognize the necessity of the present value of the ex-pected future returns exceeding the cost of the conserva-tion. Mr. Ellis called attention to the remarkably rapid change in technology and other factors that made predic-tions for more than a very short period extremely risky. Mr. Hotchkiss noted the large part of modern industry that is fundamentally dependent on minerals. Studies made by him indicated that even in a furniture manu-facturing center, such as Grand Rapids ids with its large consumption of wood, the metal and mineral industries constituted 60 per cent of the industrial life. For Michi-gan as a whole 55 per cent of the capital used in all in-dustries was employed in the mining and the metal in-dustries. This, he said, was little recognized because the buyer of alloys and metals was so far from the pro-ducer of raw materials, .but the ultimate supply of the latter was evidently of fundamental importance to the whole people. Mr. Spurr held that the more knowledge as to these matters the better. Mr. Hellmann felt that in the especial case of copper, wastage was less impor-tant than markets and economical routing of metal to them. Mr. Hutchinson classified minerals into three groups: (a) those, such as gold, that are used with very little loss; (b) those, such as coal, as are entirely consumed in use; and (c) the intermediate group. The problem, he thought, was necessarily different for the minerals of each group. Mr. Manning proposed definite-ly that since it was manifestly a great national problem, the Department of Commerce be asked to assume leader-ship here and to devote the necessary funds and staff to keeping up such a perpetual inventory of mineral resources as would best serve the public needs, both for war and peace. Both George Otis Smith and Scott Turner agreed that this was probably already within the authority of the Department but Mr. Turner told of the difficulty in obtaining sufficient funds for the work. Recognizing that the job was large enough to require the help of both public and private interests and the desire of the two societies to be helpful to the maximum extent, the Chair was asked to appoint a committee from the meeting to present to the Council of the Society and the Board of Directors of the Insti-tute a definite plan for further work. P. N. Moore, Van H. Manning, J. E. Spurr, R. C. Allen, R. H. Richards and W. S. Hutchinson were appointed and drew up the following plan for, consideration

Jan 3, 1927

By J. B. Cheatham, J. G. Yarbrough

Although adequate removal of cuttings from beneath a drill bit is important for efficient drilling operations, very little basic data are available relative to the fundamentals of chip removal by hydraulic jets. A discussion is presented in this paper of an experimental investigation of the jetting action of hydraulic jets in removing loose particles from the bottom of a cylindrical hole. Conditions for which the jet is no longer capable of removing chips from the bottom of the hole are determined. This situation represents equilibrium between the chip removal force and chip Bolddown forces such as gravity and pressure. In most of the tests loose particles were jetted with water or a water-glycerine mixture to determine the dependence of chip removal on hole size, jet size, height of jet off bottom of hole, flow rate, particle density and fluid viscosity. ,A test with a pressurized mud system indicated that relatively low pressures can completely overcome the removal action of a hydraulic jet. Although the system studied herein is not directly applicable to a rotary drill bit, the work with such simplified systems can provide a better understanding of the chip removal action of jets, and with logical extensions it may provide a reasonable basis for the best use of fluid jets in drilling. INTRODUCTION The primary deterrent to maximum drilling rates is the inability of the drilling system to remove rock cuttings efficiently enough to prevent interference with the drilling action.1-2 The objective of chip removal studies is to permit predicting and controlling removal forces under downhole drilling conditions. Conditions at the bottom of a hole during rotary drilling are exceedingly complex and are not likely to be described in a quantitative way by investigations in terms of the total drilling action until a better understanding is developed of the simplified components of the problem.3 The present study is concerned with the elementary condition of removal of chips by a single central jet. Even this relatively simple model provides mathematical difficulties because of the turbulent nature of the flow from the jet and because of the shape of the bottom of the hole beneath the jet. Theoretical and experimental studies have been made of turbulent jets impinging normally on an infinite body and deductions based on analytical solutions to simplified problems can give some insight into the problem of cutting removal by a jet. However, because of the present lack of understanding of the behavior of the interaction between the fluid jet and the chips being removed, an experimental approach was chosen for the present study. Methods have been developed for maximizing hydraulic horsepower, impact force and jet velocity; but whether maximizing these parameters maximizes chip removal with present drilling bits has not been demonstrated. Simplifying the problem of chip removal may make it possible to develop some understanding of the manner in which the jet velocity is dissipated. Better understanding of a simple case should materially assist in extending analysis to more complicated cases. Thus, we are not concerned in the present study with the rock fracturing process itself but only with the removal of the debris from the bottom of the hole. A problem which is quite similar to the chip removal problem is the suspension of solids in stirred vessels. This problem has been studied by the chemical industry and correlations have been obtained by dimensional analysis which permit the design of mixing vats7 An approach similar to that used in the mixing vat problem is used in the analysis of the jetting data in the present paper. EXPERIMENTAL PROCEDURE The test equipment arrangement shown sche-matically in Fig. 1 allows the jetting action to remove particles until an equilibrium height is attained for each combination of hole size, jet size and flow rate.*** Equilibrium conditions require that the removal force is unable to remove additional particles. This balance between holddown and removal forces implies a relationship between the two forces which is constant for the particular system. When the holddown forces are constant,

Jan 1, 1965

By G. Britsianes, T. L. Joseph

THE reduction of a lump of iron ore is a complicated sequence of up to three reactions proceeding simultaneously in a gas-solid system. As the ore moves down the blast furnace into zones of higher temperature and higher reducing power, it is successively reduced through the three oxides of iron into metallic iron. The reduction process involves much more than chemical problems. Physical factors add to the complexity of the overall process. Under optimum conditions, reduction of the ore is completed at a level about one-half way down the blast furnace stock column. At this point, the ore undergoing reduction has attained a temperature of about 1000°C and has been in the furnace for about 6 hr. On the practical side, the behavior of the ore during smelting has been of great interest to operators. Unsatisfactory blast furnace operation on burdens containing magnetite ore or badly slagged sinter has often been attributed to poor re-ducibility. The question of reducibility has also been raised in formulating quality standards for agglomerates such as nodules, briquettes, and pellets. In the present investigation, the solid phases formed during reduction were studied as a step toward a better understanding of the overall process. Equilibrium Studies The iron-oxygen equilibrium diagram shown in Fig. 1 reveal; a number of facts pertinent to the gaseous reduction of iron ores. This diagram is from the work of Darken and Gurryl, 2 and represents a correlation of the best available data. Four solid phases may exist during the complete reduction of hematite to metallic iron. These are hematite (Fe2O3), magnetite (Fe3O4), wustite (FeO), and iron (Fe). The wustite phase is a solid solution which is not stable below 570°C. At this temperature the solid solution undergoes a eutectoid-type of decomposition into the phases, magnetite and iron. Thus above 570°C, the diagram dictates that a hematitic ore should pass through a four-phase sequence on reduction to metallic iron. Below 570°C, only hematite, magnetite, and iron should appear. Information on the iron-oxygen system has been derived largely from CO and H2 reduction equilibria. The Fe-C-0 relationships have been studied extensively by R. Schenck and his coworkers and well summarized by H. Schenck.3 More recent studies have been made by Darken and Gurry.1, 2 Data from these sources have been combined and plotted in Fig. 2. With respect to the Fe-H-O system, the works of Emmett and Schultz4 seem the most reliable, and these data have also been included in Fig. 2. Certain physical properties of the solid phases of the iron-oxygen system are summarized in Table I. The crystallographic information is of special interest as much of the present work has been concerned with the X-ray analysis of the products of reduction. Reduction with Hydrogen The reduction of ore with hydrogen is the net result of two or more gas-solid reactions. Above 570°C, the reaction sequence may be represented by stoichiometric stages as follows: 3Fe203+ H2e2Fe,O, + H20 [1] 2Fe3O, + 2H2 6FeOw + 2H2O [2] 6FeOw + 6H3 ^ 6Fe + 6H=O [3] Fe2O3 + 3H2 ^ 2Fe + 3H,O. [4] These reduction reactions follow the general form: A (solid) + B (gas) e C (solid) + D (gas). This type of gas-solid reaction has been investigated by Langmuirl' who has shown that such reactions can occur only at the boundary between the two solid phases. Furthermore, a nucleus of the second phase must initiate the reaction. Once such an interface exists, the reaction proceeds through a layer of the solid reaction product (C). The specific mechanisms involved will depend a great deal on the properties and condition of this particular layer. A number of heterogeneous reactions such as the dehydration of single crystals of copper penta-hydrate and the calcination of limestone follow this type of process. It should be noted that the inter-facial type of reaction also occurs even in dense polycrystalline material which simulates a mono-crystalline behavior.

Jan 1, 1954

By Robert F. Reeves, Stephens A. Avary

Introduction People, systems, methods, and organization make up maintenance productivity. Methods promote efficiency in operations, systems ensure consistency and continuity, and organization maintains direction and control. The common denominator in these techniques, though, is people. Management systems can be technically correct, but fail to work because management systems do not work by themselves - people make them work. Therefore, the foundation for mine maintenance productivity must begin with the people involved. Traditionally, the emphasis placed on mine maintenance, as opposed to production, is not proportional to the impact that maintenance has on a mining operation. The performance and productivity of mine maintenance substantially affects the operating costs, revenues, and profits of the mine. The mining industry has recently experienced a period of economic depression. While it has created financial problems in all segments of the industry, it has caused the industry to focus on the need for improved productivity. Mining companies that have made investments in productivity during this period will have a competitive edge during the recovery, and mine maintenance may be the best place to put that investment. Future lost production and maintenance cost problems can be avoided if mine management has the foresight to seize the opportunities that now exist. Background Technological improvements over the last 30 years have been dramatic in the industry. Before these changes, mining equipment was not very sophisticated. So maintenance needs were not very sophisticated. Maintenance people were not required to have any formal training in mechanical or electrical skills. At many mines there was no formal maintenance organization. The mechanic was a part of the section crew and received direction from the face boss. Mine maintenance was not a function, it was a task. Technological change in mining came about in response to an expanding economy and greater demand for mined products. The past 20 years brought about a change in the kind of individual needed to mine ore and coal, and an even greater change in the person needed to maintain equipment. Equipment went from mechanical and pneumatic to complicated electrical and hydraulic. Maintenance progressed from being a task to being a function. Types of Mechanics Mine maintenance workers can be grouped into three categories based on experience: mechanics with more than 20 years, five- to 10-years, and less then five years. These groupings have influenced mine maintenance productivity. The 20-plus mechanic has been the foundation of the maintenance force. He had no formal training and his skills were acquired on the job. As new equipment was introduced, he learned to repair it by trial and error. Not all of what he learned was correct. Personal initiative and conscientiousness are characteristics of this group. They have experienced the bad times and the good, and appreciate the opportunity to work. The five- to 10-year group worked with the older, more experienced mechanics to acquire maintenance skills. These mechanics came in during industry boom times, and many progressed quickly by being in the right place at the right time. Many in this group have difficulty with diagnostic trouble-shooting, and depend on changing out parts to identify problems. This group is generally more mobile and less apt to have a strong work ethic. Mechanics with about five years experience have vocational training, some formal education or basic skills training. Like the five- to 10-year group, these miners are mobile and tend to move readily to satisfy their changing needs. These two younger groups are less apt to feel any loyalty to the employer or obligation to be productive. Their attitudes and abilities reflect the influence of their general age groups, and has a definite impact on mining productivity. Past Requirements In the 1960s, the basic requirements for a maintenance supervisor was that he be a top mechanic, get along well with the men, and be a hustler and improviser. He learned the details of new equipment as repairs were needed. Oftentimes, this individual would be the only person on the property with the ability to read prints, use diagnostic testing equipment, and troubleshoot using an analytical approach. This approach to maintenance is sufficient when the equipment design is reasonably uncomplicated. Management Changes As technology progressed, organized labor gained strength, government regulation increased, and the job of supervising and managing people had to change. The environment has changed and

Jan 11, 1983

By R. G. Bowman

DISCUSSIONS of processes for the manufacture of white lead generally open with the statement that white lead is the oldest chemical pigment known to man. This fact is of more than historical interest; in the light of the present extensive use of white lead, ii indicates that the .compound possesses characteristics that make it unique among white pigments, and superior to all ;in its particular field. A greater variety of processes have been proposed for manufacturing white lead than for any other one chemical compound. Only a few of these processes have come into commercial use, however, and a large part of the present production here and abroad is made by the process used in ancient times. The outstanding disadvantages of the old process are that it is slow and laborious, affords no control of the product, and yields at best an impure material lacking in uniformity. Improved processes, sometimes classed all together as "quick processes," have had as their main object the shortening of the time required. Most of them accomplish this and some have yielded a superior product as well, but only two or three are now in use commercially. For centuries the manufacture and mixing of paints has been treated as an art, and the practice has been much influenced by tradition and precedent. There is a reluctance to abandon old materials for this purpose, and as few new uses for white lead have been developed, the adoption of the quick processes has been slow. The electrolytic production of white lead is not-new, a large number of processes and forms of apparatus have been proposed and patented within the past 25 years, all directed toward the, production of white lead, wholly or in part, by electrolysis. One or two of these processes were tried commercially but proved unsuccessful.

Jan 9, 1925

By Warnie Flint

According to statistics compiled by the National Safety Council, US Bureau of Mines, American Medical Association, and other agencies, accidental injuries cause more deaths than all the leading diseases. No one is immune to an accident. Probably the best way to describe an accident is to say that it is an unexpected happening. The automobile driver is not expecting a terrifying crash, the electrician does not expect to be electrocuted, nor is the coal miner expecting the roof to fall on him. All personal injuries will not result in death. There is a multitude of victims who are living proof of what accidents can mean to an individual-victims who are blind, paralytic, missing arms, legs, etc. In most instances the earning power of these individuals is gone, and they are utterly dependent upon family or welfare. For many years there seemed to be an attitude that prevailed in industrial America that accidents were inevitable, and that so many fatalities and injuries should be expected. This, of course, was utter nonsense. Since the birth of such organizations as The National Safety Council, American Society of Safety Engineers, and. the US Bureau of Mines, much progress has been made in reducing accidents. The enactment of compensation laws greatly contributed to the progress that has been made. Does this mean that all accidents can be prevented? Unfortunately, it does not. Human psychology indicates that men will work unsafely or perform unsafe acts for the following reasons : 1) Lack of knowledge or skill. 2) Conflicting motivations, such as: (a) to be more comfortable, (b) to save time or effort, (c) to be accepted by a group. 3) Physical or mental condition. Nearly everyone desires his own safety, but not all individuals are willing to do what they know to do to insure their own safety.

Jan 1, 1981

By G. A. Kosick, M. Freberg, L. A. Kuehn

"Highly variable and higher than forecast lead miff feed grades were problems experienced at the Polaris Concentrator. Limited flotation capacity and poor operating flexibility prevented the plant operators from overcoming these difficulties, and resulted in grades and recoveries which were lower than necessary.Two column cells were integrated into the lead circuit in an effort to give the operators the tools they required to achieve the best grades and recoveries possible, regardless of the miff feed grade. Installation of the column cells increased lead concentrate grade from 76% to 79.3% Pb, and increased lead recovery from 9/% to 91.3% Pb.This paper reviews the design, development and operation of column flotation cells in the lead circuit at the Polaris Mine.IntroductionThe Polaris Concentrator was constructed in 1980 on a barge in Trois Rivieres, Quebec, and towed to its pre sent location on Little Cornwallis Island, Northwest Territories, about 1450 km from the North Pole. The plant presently treat s about 2800 mtpd of new feed with average feed grades of 3010 Pb, 13% Zn and 5% Fe. Current production averages 36 000 tonnes of lead concentrate and 220 000 tonnes of zinc concentrate annually.The barge design and min e location dictated that the concentrator .be compact with very little unused floor space. As a result, flexibility was limited in terms of expanding the plant.Mill feed grades were highly variable and above forecast and the lead flotation circuit was found to be under capacity. This caused continued overloads in the lead cleaning circuit, resulting in low concentrate grades and recoveries. It was evident that an increase in lead flotation capacity and flexibility was necessary. The existing floor space did not allow for the installation of additional conventional flotation cells, but one alternative was the installation of column cells."

Jan 1, 1988

By Walter C. Labys, William A. Vogely, David O. Wood

INTRODUCTION The past decade has witnessed a literal explosion in formal modeling of energy systems and markets. This rapid development was stimulated primarily by policy issues arising from the energy price shocks of 1973- 1974 and 1979- 1980. The suddenness of this interest is illustrated by comparing the current and earlier editions of Mineral Economics. The third edition, published in 1976, includes an addendum to the section of projection and forecasting methods because, Since the foregoing chapter was written, major events in petroleum have caused an explosion in formal modeling efforts in the energy field.a In contrast, this fourth edition includes the present chapter on energy modeling, and a separate chapter on nonfuel minerals modeling. The considerable interest in formal energy models is based on the complexity of the system, in particular the interaction between technical and engineering data and designs, and economic behavior of energy producers and users. Energy is a vital component in the economic and social well-being of nations, and must increasingly be explicitly considered in developing economic growth and welfare policies. Energy models can contribute to policy development and analysis, both by improved understanding of the energy system, and of the interaction between energy markets and the economy, and as an explicit means of analyzing and comparing the implications of alternative policy actions. At their best, formal models can clarify and illuminate policy options. In this chapter we will survey the objectives, methods, and potential contributions and limitations of energy system modeling. Section 2 considers the objectives, scope, and approaches to energy system and market modeling. Methodological approaches are then discussed more formally in section 3. Our approach is to intro- duce each model type and method via relatively simple generic models, illustrating their use by a selective review of the applications literature. Section 4 shifts the focus from modeling methods to procedures for model evaluation and improving model credibility especially in policy research and analysis. We conclude with some brief remarks on directions and opportunities for future modeling research. ENERGY MODELS: OBJECTIVES, SCOPE AND APPROACHES The concept of a model usually evokes an image of a complex, computerized system of mathematical or econometric equations providing detailed information concerning the operation of the process being modeled. In fact. models may be simple or complex, formal or mental, depending upon the purposes for which the model is intended. Simple judgmental models may be most appropriate when monitoring the overall performance of a process. When more detailed information is required and/or when the model

Jan 1, 1985

By H. W. Rogers

The first steam shovels used in this country were built by the Otis Company, of Boston, about 50 years ago, but as they were of very crude construction and rather unsuccessful only a few were built. For possibly 10 years prior to 1884 successful steam shovels were made and used on certain classes of work, but it was not until that time that they were manufactured in quantities and began to play an important part in all classes of excavation. From that time up to the present day there have been gradual but continuous improvements on the original shovel, not only in the mechanical construction, but also in the design of boilers and engines which are best adapted to this class of service. In proposing a change from steam to electric operation we have to deal with a steam equipment which has not only proved its worth but has probably reached its highest stage of development and efficiency. That the electric shovel is a possibility cannot be denied, as at present from 12 to 18 shovels are in operation in this country. These shovels may be divided into three classes: the friction electric, which is operated by a single constant-speed motor with friction clutches; the three or four motor direct-current equipment, and the three or four motor alternating-current equipment. It is the second and third classes that I wish to deal with, as the first class does not compare favorably with the steam shovel so far as speed is concerned, although it may be operated as cheaply. There is probably no other class of machinery that presents a duty cycle as severe as that of the shovel, which is very short, varying from 7 to 12 sec. on the hoist, from 7 to 12 sec. on the thrust, and from 10 to 18 sec. on the swing, making a complete cycle in from 17 to 30 sec., and the motor to meet these requirements must have a sufficiently low armature inertia to permit of rapid acceleration and quick reversals under small power. It should also be a motor of rugged design, as it must be subjected to severe overloads and shocks and frequent reversals. This is especially true of the hoist motors and, to a lesser degree, of the swing motor; the thrust motor being practically stalled during the digging operation, although it may revolve or overhaul, according to conditions, and is operated at full speed only after the hoisting operation is completed.

Jan 1, 1915

By R. J. Arsenault

The yield drop that occurs in tantalum, Cu-AZ. and Ag-Al was investigated as a function of strain rate and at several temperatures. From the strain-rate dependence of the yield drop an activation volume was calculated and found to agree with that obtained from differential strain-mte tests. The temperature dependence of the yield drop can he explained in terms of the parameter temperature divided by activation volume. The yield drop is attributed to the rapid multiplication of dislocations during their motion. THE yield-point or yield-drop (upper yield stress minus the lower yield stress) phenomenon has been the subject of numerous investigations1-9 in both bcc and fcc metals. Also, there is no dearth of theoretical explanations10-20 to account for the yield-point phenomenon. The present explanations can be divided into two categories: the concept of dislocation locking and subsequent unlocking and the concept of multiplication of dislocations due to their motion. The Vottrell10,11mechanism of locking and unlocking, as applied to bcc metals, is the best known of the three proposed locking mechanisms. The two additional mechanisms, short-range order (SRO)7,16,17 and Suzuki-type locking,14,15 have been employed to account for the existence of yield points in fcc metals. The experimental evidence supporting the locking of dislocations, especially in bcc metals, is well-established. The dependence of the yield point on interstitial concentrations,' the kinetics of strain aging,2 and the observations by electron-transmission microscopy18 are some of these experimental results. The number of experimental observations supporting dislocation locking in fcc metals is not as extensive as in bcc metals. The unlocking of dislocations, which actually is supposed to account for the yield drop, is by no means well-established by experimental results for either bcc or fcc metals. The similar temperature dependence6 of the yielding and plastic flow is one of the experimental factors that cannot be explained in terms of the unlocking of dislocations. The mechanism based on the rapid multiplication of dislocations has been proposed by several investigators. 18-20 Johnston and Gilman19 have proposed a detailed model of yielding, which was later employed by Hahn21 to predict quantitatively the yield drop in iron. This mechanism was discussed in a previous paper22 and several limitations were pointed out. In the same paper it was shown that an activation volume could be obtained from the strain-rate dependence of the yield drop and that a correlation existed between the activation volumes obtained from the yield-drop tests and from differen-

Jan 1, 1964

By W. A. Alexander, L. M. Pidgeon

Metallic magnesium and similar meta!s near the top of the electromotive series have been commercially produced by the electrolysis of a suitable molten salt. Despite the success of electrolysis, sufficient inherent objections exist to justify the search for a direct reduction method. Such a method would widen the choice of raw materials, relax the rigid electrolytic requirements of raw-material purity, and, by obviating direct current, over simplicity of plant equipment. This paper describes the pilot-plant development of one such method, which has subsequently formed the basis of operation of six commercial plants-—one in Canada and five in the United States. Unlike aluminum, magnesium is volatile at relatively low temperatures. This physical property, which produces its greater inflammability, also offers possibilities of direct reduction that are not available with aluminum. In the basic reaction MgO + X = XO + Mg if X is nonvolatile, the reaction may be forced to the right at high temperatures by the evolution of magnesium vapor, despite the large negative value of the heat of reaction which is almost bound to follow the use of any available commercial reducing agent. The cheapest X, of course, is carbon, but this notable advantage is modified by the production of a volatile oxide CO, so that both Mg and CO are evolved simultaneously. Elaborate devices are required to shock-cool the equilibrium mixture in order to prevent back reaction. At best, a pyrophoric powder is produced, requiring redistillation to provide marketable metal. The practical development of this process, is due to F. Hansgirg, and his methods have recently been given a trial on a grand scale in the plant at Permanente, California, and smaller plants were previously built in Austria, England and Japan. A second class of reducing agents is available, those which produce nonvolatile oxides. When XO and X are relatively nonvolatile, magnesium is the only volatile member of the system. It may be evolved in a simple distillation step and condensed without the formation of powder. Two basic requirements must bc fulfilled in order that such a reaction may proceed: I. The reducing agent when heated to a reasonable temperature in presence of MgO must cause the production of an appreciable equilibrium vapor pressure of magnesium. This magnesium must be removed continuously. z. A supply of heat must be maintained. In order to fulfill the first requirement, operations must be conducted in vacuo or in a stream of Hz. Owing to its greater efficacy and relative safety, vacuum has been employed in the work to be described. The choice of a suitable reducing agent formed the subject of an extensive series

Jan 1, 1944