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By Michael Hoch, Seong Kwan Rhee

The ternary system Zr-Cr-0 was investigated at 1200°, 1500°, and 1700°C. The isotherms at these temperatures were determined by metallographic and X-ray diffraction analysis of carefully selected alloys. AS a part of a series of investigations1-4 of the ternary and quaternary phase diagrams on refractory metals and their compounds, the ternary system Zr-Cr-O was investigated at 1200°, 1500°, and 1'700°C. The binary diagrams Zr-Cr, Zr-O, and Cr-0 as reported by Hansen5 and Levin et a1.' and the existence of a ternary coompound ZrsCrsO (q-carbide structure a. = 11.96A) provided a good base for this investigation of the Zr-Cr-O phase diagram. MATERIALS The materials used and their suppliers were: Zirconium powder, 99.4 pct pure, Charles Hardy, New York; Zirconium dioxide, Fairmount Chemical Co., Newark, N.J.; Chromium powder, 99.85 pct pure, Fairmount Chemical Co., Newark, N.J.; Chromium sesquioxide, 99.86 pct pure, J. T. Baker Chemical Co., Pillipsburg, N.J. Though the materials used were not "hyperpure", the impurities present do not affect the results (lattice parameters, phase boundaries) within the experimental accuracy. EQUIPMENT AND PROCEDURE The detailed procedures and equipment used for this investigation were similar to those described elsewherel' "ith the exception that in this study an induction-heated argon-atmosphere furnace was used. The latter furnace was used to prevent the vaporization of chromium metal in samples in the higher temperature range (above 1200°C). The tungsten crucible was induct ion-heated in a vertical quartz tube filled with argon gas. The quartz tube was 1-1/2 in. in diameter and 10 in. long; it was connected at the top and bottom to water-cooled copper plates. The argon gas of a purity 99.995 pct was further purified by being passed through zirconium metal chips which were maintained at 800°C in a silica tube 1 in. in diameter and 30 in. in length. The furnace was flushed with argon gas for an hour before starting heating; during heating, a slight (3 cm oil) overpressure of argon was maintained in the system. Power was supplied by a 20-kw Ther-Monic Induction Generator. The samples rested on a tungsten-wire basket; the wire basket was connected to a tungsten wire which passed through the tungsten crucible lid, and entered the system on top through a Wilson-type seal. By pulling on the wire, the sample was removed from the crucible and dropped to the water-cooled copper base of the furnace. Because of the presence of argon gas, the cooling of the sample was even more rapid than in the vacuum furnace.' EXPERIMENTAL RESULTS AND DISCUSSION 1) Isothermal Section at 1200°C. The compositions investigated have been numbered within the phase diagram, Fig. 1, and all the pertinent information about the samples can be found in Table I. This table lists the composition of the samples, the phases present in the room-temperature diffraction patterns, the phases detectable in the microstruc-ture, and calculated lattice parameters of the a-zirconium phase. The shapes of the various phase

Jan 1, 1964

By B. M. Larsen, T. E. Brower

INTRODUCTION AND SUMMARY BY "thermochemical factors" we refer to those variables which affect the net heat which must be put into the bath in order to make a heat of steel from any given set of charge materials. This is admittedly only one phase of the production rate problem. The whole problem includes also consideration of rate of charging, flame radiation intensity, rate of optimum fuel input obtainable without undue damage to the furnace structure and any variations in conditions such as stirring, slag volume or viscosity which affect the rate at which heat can flow down and away from those exposed surfaces where the heat is first received from the flame. Because a given type of charge in a shop is often considered as being nearly constant in both physical and chemical characteristics from one heat to another, and because such a state of affairs may be thought of as resulting in a reasonably constant net heat requirement, this requirement is usually neglected as an invariable, inevitable matter, about which nothing can be done. With normally variable charge, and variable percentages of limestone and cold pig, however, the net heat requirement can be subject to very appreciable change; it may also be affected by such things as the currently popular procedure of blowing oxygen into the bath, To bring the general problem of production rate into better perspective, therefore, it seems worthwhile to try to evaluate the net bath heat requirement and the amount by which it may change through variation in different factors. Because our calculated values of net heat are admittedly uncertain, we have tried to check them in part by statistical data on actual production rates over rather large groups of commercial heats. The results of calculation of net heat requirements in a wide range of actual commercial practices may first be summarized briefly. We find a wide range between the extremes of different practices, as would be expected, with duplex practice (essentially blown metal plus hot metal) having essentially a zero net heat requirement, an all-cold charge practice somewhat over 1 million Btu per ton of steel and a high hot metal practice about 0.75 million Btu per ton. Over the commercial range of scrap plus hot metal practice, increasing the percentage of hot metal from 40 to 78 pct, with other factors nearly constant, decreased the net heat by only some 5-10 pct, The net heat is increased by about 1.2 pct for each I pct of the metallic charge present as cold pig. Calculated values of oxygen supplied by air varied from about 25 up to 60 lb per ton of steel, and in a 60 pct hot metal practice, an increase from 32 to 52 lb oxygen from air decreased the calculated net heat from 0.88 to 0.74 million Btu per ton. Use of pure cold oxygen to substitute for about one half of the total charge ore (about 28 lb oxygen replacing 145 lb ore per ton of steel) decreased the

Jan 1, 1948

By Richard H. Jahns

Gem materials, comprising those minerals and closely allied natural substances used for personal adornment, for the fashioning of ornamental objects, or for other decorative purposes, have been valued in human commerce for thousands of years. Amber, jet, natural glasses, and at least 15 mineral species were sought and used by primitive man, and most of the materials in common use today were known well before the time of Christ. Their variety is remarkably great, and the traceable relationships among their respective sources, modes of occurrence, properties, and histories of production, preparation, and valuation are extremely complex. Their uses have been affected by numerous shifts in fashion, and from time to time by superstition, by decrees of church and state, by personal interpretations of religious significance, and by various presumptions concerning medicinal value and influence on human behavior (Ball, 1935, 1950; Kunz, 1913; Pogue, 1915; Whitlock, 1946). Thanks to their durability, portability, and relative scarcity, some gem stones have been employed as media of exchange, and the investment value of the more precious stones has long been recognized and exploited, especially in Asiatic and European countries. Demands for gem materials have stimulated trade among various peoples for centuries, and they also were responsible for much early geographic exploration. During more recent times they have prompted numerous geological studies and a considerable amount of research on the growth and properties of crystals. Apart from decorative uses, some naturally occurring gem materials also are employed directly in numerous fields of industry. Substances that are manufactured for gem or industrial uses are termed synthetic gem materials; most of these correspond in composition and crystal structure to certain minerals found in nature. Imitation materials are those whose use is based upon their resemblance to substances of different composition and greater intrinsic value. Gem stone is a designation commonly restricted to those materials suitable for personal adornment, and the complementary term ornamental stone is applied to the remaining materials not used industrially. This distinction is arbitrary and somewhat inconsistent, owing to numerous geographic differences and historical shifts in end uses. Further, there is a complete gradation between the category of ornamental stones and that of dimension stone. Substances that lose their identity in use are excluded from the gem category; thus a transparent crystal of beryl would not be classified as a gem stone if it were processed as a source of beryllium metal. The adjectives precious and semiprecious have been employed as a means for distinguishing different gem materials on the basis of recognized value. Diamond, emerald, ruby, and sapphire consistently have been regarded as precious stones, and in cut form they commonly are referred to as the noble gems. Opal, pearl, turquoise, and chrysoberyl also have been classed as precious at various times. In recent years, however, it has been increasingly recognized that no satisfactory distinction can be based upon assessed value or any group of specific properties, and now all gem stones normally are referred to as precious. Indeed, the tendency is to extend this term to include many of the so-called ornamental stones, and the designation semiprecious has been essentially discontinued. Unlike most mineral commodities of economic value, gem materials are little altered in processing for end uses. Some stones are used in essentially their natural form, and most are reshaped and polished without significant changes in their fundamental properties. Thus the preparation of gem and ornamental stones is intended chiefly to emphasize or enhance

Jan 1, 1983

By Erskine Ramsay

The following description of the Alabama coal-,fields is taken (with slight omissions) from the report " On the Warrior CoalField, by Henry McCalley, A.M., C. and M. E., Assistant State Geologist:" The coal lands of Alabama are of the great coal basin of the Ohio, or of the great Appalachian coal-field, and are estimated at 8660 square miles. They formed one connected field prior to the great Appalachian revolution, when there was pushed up through them long, narrow, anticlinal ridges which have been denuded into the present well-known anticlinal valleys, and thus this great field has been divided up into three more or less distinct parts. These parts were named by Prof. Tuomey, in 1849, the Warrior, the Cahaba and the Coosa coal-fields, respectively, from the names of the rivers which drain them. They are of very unequal size, the Warrior, with an area of 7810 square miles, being nearly ten times as large as the other two combined, and about two-thirds as large as the coal-area of Great Britain. This Warrior field is the most northwestern of the three coal-fields of Alabama, and comprises all of the coal-measnres in Alabama, northwest of the Alabama Great Southern railroad, or all of those drained by the Warrior and Tennessee rivers. The greater part of this field, almost all of it away from the present railroad lines, has never received more than a surface-examination. Still, we have in the Warrior coal-field one of the richest coal-basins of this or any other country. As a whole, the Warrior coal-field is a broad, shallow, tray-shaped depression, sloping towards the southwest, and with its southwest end covered by a newer formation and its southeast side, especially, complicated by folds and fractures. As is the case with all of our Coal Measures. it is composed of a series of sandstones. conglomerates, shales, slates and coal-seams, with some little limestone at several horizontal positions, but contains no thick bed of this last rock like the Measures of Pennsylvania and other States. The thickest conglomerates are near the bottom and top of the series, and hence these coals might be said to belong mainly to the

Jan 1, 1891

By T. Petrov

This study was performed to analyze the behavior of the blowing curtain face ventilation system modified with a specially designed regulator, called Wing Regulator. The purpose of the regulator is to extend the airflow penetration depth toward the face and allow effective face ventilation at curtain setback up to 13.7m (45ft) from the face. A small scale model of the Wing Regulator was successfully tested in the Mine Ventilation Laboratory at the University of Kentucky. Then a prototype of the proposed regulator was tested in full-scale in mine environment. Series of three dimensional simulations were performed to support this research using validated CFD code. This paper presents results of the performed studies and discuses the advantages of using the Wing Regulator. Presented information relates to the US Provisional Patent Application No 61/818,112.

Feb 23, 2014

By R. M. Hardgrove

Pulverized coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the dificulties incident to the burning of coal. In a. cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverizeti in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. Process of Burning Cement Clinkers The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output, This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,' in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. Wet vs. Dry Plants The study presented here covers 88 plantsl of which 41 are wet and 47 dry. Of

Jan 1, 1942

By R. M. Hardgrove

Pulverized coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the dificulties incident to the burning of coal. In a. cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverizeti in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. Process of Burning Cement Clinkers The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output, This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,' in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. Wet vs. Dry Plants The study presented here covers 88 plantsl of which 41 are wet and 47 dry. Of

Jan 1, 1942

By G. H. Lair, B. J. Langford, S. H. Rockwood

Volumetric reservoir analysis is dependent upon a firm relationship between porosity, connate water, and net pay. Capillary presswe data on core samples interrelate these three factors. It is shown that other reservoir problems may be resolved: (I) whether water production is indigenous or extraneous to the oil producing interval, (2) an oil-water contact is being approached, and (3) whether the oil-water contact is remotely situated. INTRODUCTION Volumetric analysis of carbonate reservoirs challenges the ingenuity of the reservoir engineer due to an absence of data on the minimum oil-bearing and oil-producing pores of a given reservoir rock. Many studies with this as the ultimate objective are available on the physical characteristics and fluid-flow behaviors of the carbonate reservoir. Some of these publications are: Hassler, Brunner and Deahl's1 study of the role of capillarity in oil production; Jan Law's' and A. C. Bulnes3 statistical treatment of core analysis data; Chalkley's method4,5 or estimating specific surface areas and porosity; Walter Rose's9 report on porosity, reserves, and recovery; Archie's78 tex-tural classification of carbonate rocks; Stewart and Spurlock's9 large core analysis; Purcell'sl0 mercury injection method in studying capillary pressure phenomena; Stewart, Craig, and Morse's" model multiple-phase flow test in investigating the relative permeability effect; and Fatt's12 illumi-nating network model of porous me-dia. The present study shows that the static reservoir parameters of porosity, connate water, and net productive thickness may be interrelated by utilizing a statistical arrangement of mercury capillary pressure data in addition to routine core analysis. This report summarizes the techniques that are followed to obtain these relations which are applicable to carbonate and sandstone reservoirs that do not contain clays. Capillary pressure statistical studies may be extended to include clay-bearing rocks by utilizing water capillary pressure data. POROSITY It is necessary that the laboratory determine the porosity value that includes the pores that are filled with hydrocarbons and water in the reservoir. This is the reservoir's total porosity. When true total porosity values are not available, correlations between porosity and capillary pressure characteristics lose their significance since they cannot be related through oil-base mud core data to the porosity and connate water values found in reservoirs. Methods used to obtain reliable porosity data will not be discussed; however, care should be exercised in selecting samples for measurement. In homogeneously porous formations, small core plugs sampled frequently will furnish adequate information. Where heterogeneity, particularly vugginess, is present, large-core porosity data may be required; in such reservoirs a suite of capillary pressure samples may be selected from small plugs. CAPILLARY PRESSURE Capillary pressure data provide a method for relating two carbonate reservoir parameters, porosity and connate water saturation, which in turn permit delineation of productive rock types necessary for selecting the third parameter, productive thickness. Specific applications of capillary pressure data to petroleum reservoirs were suggested by Archie8 and Pur-cell."' Their work demonstrated that in preferentially water-wet rock, porc

Jan 1, 1958

By Nicholas J. Grant, Italo S. Servi

Extensive data of minimum creep rates and rupture times for high purity and commercial aluminum confirm the existence of a transition range from the low temperature-type to the high temperature-type behavior. The data are analyzed in line with the suggested theories of deformation of metals. CONSIDERABLY more work has been done on the rupture and creep behavior of commercial alloys than on the less complex, high purity metals. As a result, certain fundamental behaviors are overlooked. One of the lesser known variables affecting the high temperature behavior of metals is the impurity content (or alloy content). Impurities are known to affect strongly the recrystallization temperature, electrical resistivity, and other properties of metals, but the effect on creep and other high temperature mechanical properties has been determined for very few materials. In view of these facts aluminum appears to be a fairly ideal metal, since it is available in a wide range of purity contents from 99.995+ down through the 2s (99.0 + Al) and 3s (1.2 Mn) grades in small steps. It is highly resistant to oxidation and other surface instabilities, leaving recrystallization and grain growth as virtually the only instabilities. Control of grain size is also well standardized. Limited data on the creep of aluminum have been obtained by Dushman et a1.l for high purity aluminum; by Sherby2 on 2s aluminum of variable grain size; and by Dorn and Tietz³ on cold-worked 3s aluminum. Experimental Procedure The apparatus consisted of a constant stress, creep testing unit similar to the type described by Hop-kin.' The specimens were held at temperature for 30 min before loading. The loading time was less than 2 min. Factors affecting the accuracy of the data were: 1-—the temperature gradient and control, ± 3°F; 2-the stress determinations, ± 1.5 pct; and 3—elongation measurements (1 in. gage), ? 0.001 in. Three grades of aluminum were tested: 1— high purity (99.995 pct Al), 2—2s (99.3 pct Al), and 3—3s (98.2 pct Al). The spectrographic analyses of these materials are reported in Table I. The specimens were annealed then electropolished in a perchloric-acetic mixture before testing. Data on heat treatments and grain size obtained appear in Table 11. The entire creep curve was determined for almost all of the tests. A few tests were interrupted after the beginning of the third stage of creep, and in these cases the rupture time was estimated. Experimental Results Constant stress creep-rupture testing gives a much longer, more accurately measurable second stage of creep than constant load testing. This behavior is in agreement with the results obtained by Andrade for lead (see Fig. 1). If the test is run to completion, the third stage of creep appears as soon as stress intensification occurs. In the case of specimens which fail in a ductile manner, the stress intensification begins when the sample starts "necking". This differs from the "high temperature brittle" specimens in which the stress intensification is caused by intercrystalline cracking. These specimens

Jan 1, 1952

By H. Kalsi

"IntroductionRoyal Oak Mines Inc. is currently operating the Parnour No.1 mine located 10.5 miles east of Tinunins, adjacent to Highway No. 101. The operation produces gold ore at a rate of 3300 tpd. A variety of mining methods including bulk open stopes with ITH drilling, narrow vein panel, longhole and open pits contribute to produce 1.2 rnillion ore tons annually. The milling process requires water at an average rate of 1 rnillion gallons per day (700 U.S. gpm). The Pamour No.1 mine has historically had two major problems with regard to water control. The flfSt is the underground pumping system's inability to handle excess inflows of water generated from spring thaws. When the thaw takes place over a short period of time, large inflows of water are created which quickly migrate underground and sometimes result in the loss of the lower loading pocket. During the spring runoff the pumps at Parnour No.1 work continuously for 24 hours at a rate of 1200 U.S. gpm. The system was further taxed in 1991 when the Hoyle mine was connected to the Parnour No. 1 mine and whatever excess capacity the pumping system had for the spring runoff was taken up handling the inflow of mine water generated by the Hoyle.The second problem facing the mine was the lack of water for the mill during the severe cold winter months and dry summer months. At those periods inflows of water into the mine decline and the underground water normally pumped to the mill is severely decreased. In times of dry spells mill personnel increase the water diversion from the Three Nations Lake in order to meet water demands and maintain tonnage throughout. In essence, the mine was going through an annual ""feast or famine"" situation and a solution was needed.A number of options were considered but were narrowed down to the two most practical. The first was to modify the existing pumping system and increase its capacity in order to handle the excessive runoff during spring thaw. This involved a heavy capital expenditure and would have caused additional operating costs (i.e. increased hydro). The option chosen by the management was to divert the spring runoff into an unused section of the mine called The West End. A bulkhead placed at the 2400 level would provide an ultimate underground water storage potential of 150 million gallons. This would also cut down the load on the existing pumping system, save electric consumption in pumping and provide water storage which could be used for milling during the two dry spells. Pumping stored water from the mine could be scheduled for off peak hours to take advantage of lower power rates. To meet these objectives an underground bulkhead capable of withstanding high water pressure up to 2000 ft head (866 psi) was constructed in the spring of 1992. The bulkhead site was at the 2400 level at the Pamour No.1 mine. A major consideration was the requirement to construct a safe bulkhead and to minimize water leakage from the bulkhead.Figure 1 shows the bulkhead location."

Jan 1, 1994

By Savita Schlesinger, Mike Hanna, Nathan Scalla, James O’Shaughnessy, Steve Miller

Loudoun Water’s Quarry A (Milestone Reservoir) project involves conversion of a rock quarry to a billion-gallon raw water reservoir. The project includes a new 40 MGD pump station, deep intake shaft, and three drill and blast intake tunnels for accessing the reservoir. Due to concerns regarding constructability and construction risks, the project design was paused while a Construction Manager At-Risk (CMAR) contractor was procured. The CMAR contractor was engaged collaboratively with the design team to provide risk and constructability input for refining the design. Aspects of the original design, design modifications, and benefits of the CMAR collaboration will be presented.

Jun 13, 2023

By Campbell Brown G F

THE Mount Morgan mine has, from the time of its historical discovery, been the subject of investigation and speculation by geologists and mining engineers. The first examination was made by Dr. R. L. Jack, in 1884. He came to the conclusion, from theevidence then available, that nothing but a thermal spring in the open air-otherwise a geyser-could account for the formation of the ore body. This report attracted world-wide attention at the time. Since then the mine has been examined by most of the leading geologists of Australia, and by many from elsewhere. In all cases the examination has only extended over a period' of a few days or weeks, and in the light of present knowledge of the deposit we know that sufficient data could not be obtained in such a short time to warrant the statement of definite conclusions. It is not to be wondered at, therefore, that a considerable difference of opinion exists as to the genesis of the deposit. That the company considered more detailed investigation necessary is shown by the fact that some thirteen years ago it employed Mr. Charles Walker as geological surveyor on the .mine. At that time also the data available was insufficient to give satisfactory results. It is, however, a matter for comment that none of the geological reports on the property took notice of the increase of copper values in the ore with depth, though the present workings and diamond-drilling...

Jan 1, 1911

By John B. Campbell, George G. Binder, Loren H. Jenks

Conventional methods of producing crude oil generally leave in the depleted reservoir 1 to 3 bbl of oil for every barrel brought to the surface. This paper describes a field test which explores the possibility of using gas-driven liquid propane to improve oil recovery efficiency. Laboratory theoretical and experimental studies leading to the field test indicated that low residual oil values could be obtained using limited amounts of propane driven by gas. The field test involved (I) repressuring the reservoir to the vapor pressure of liquid propane, (2) injecting the liquid propane, and (3) injecting dry natural gas to drive the propane and oil to the producing wells. Considerable preliminary work, including new drilling, work-overs, and a pressure survey was necessary. During the test, oil production rates have averaged about nine times the final primary rates. The field average GOR is still below the pre-test value. Oil recovery to date is 112,000 STB. About one-third of the 143,-000 bbl of propane injected has been produced to date. Results of the C-2 block test are encouraging and larger, more definitive field projects ore indicated. INTRODUCTION The cost of finding new crude oil reservoirs continues to increase. Today the oil industry is drilling more and deeper wells and making fewer big discoveries. Nevertheless, the demand for oil and oil products increases each year. It is natural that oil companies, faced with this situation, concentrate on recovering more of the oil existing in our known crude oil reservoirs. There is room for much improvement. For every barrel of oil brought to the surface by present day methods of producing crude oil, generally 1 to 3 bbl are left behind in the depleted reservoirs. The "unrecoverable" oil in petroleum reservoirs is largely "trapped" by capillarity. Since flushing with a solvent is exceedingly effective in combatting oil loss due to capillarity, much effort in this direction has been made in recent years1,2,3,4,5. Liquefied petroleum gas (LPG) has been suggested as a relatively inexpensive. abundant, easily recoverable solvent for crude. Some operators have observed striking responses in oil recovery at production wells offsetting input wells in LPG storage reservoirs6,7,8. In spite of the low unit cost of LPG, however, it is uneconomic in general to inject a volume of LPG into a reservoir to remove a like volume of oil. A modification of solvent flushing in which a limited quantity of liquid propane is gas driven through a reservoir appears capable of recovering much more than 1 bbl of oil per barrel of propane injected. Also, most of the propane injected should be recoverable for sale or re-use. Since such a process could well have widespread and profitable application, The Carter Oil Co. is making an extensive evaluation of the technique's possibilities. Laboratory development of the gas-driven propane recovery technique was sufficiently encouraging to justify testing in the field. Since 1952, the Production and Research Departments of The Carter Oil Co. have cooperated in a field test in the C-2 Block leases of the

Jan 1, 1958

By K. J. Gill, P. Chiotti, J. T. Mason

Thermal, metallographic, and vapor pressure data were obtained to establish the pkase boundaries and the standard free energy, enthalpy, and entropy of formation for the compounds in the Y-Zn system. Three coinpounds with stoichiometric formulas of YZn, YZn2, and Y2Zn17 melt congruently at 1105", 1080°, and 890°C, respectively. Four compounds with stoiclziometric formulas of YZn3, YZn4, YZn5, and YZn,, undergo perztectic reactions at 905", 895", 870º, and 685ºC, respectively. Three eutec-tics exisl in this system with the .following eutectic temperatures and zinc contents in wtpct: 875ºC, 23.2 Zn; 1015ºC, 51 Zn; 865ºC, 82 Zn. The YZn, pkase undergoes an allotropic transformation. In the two phase YZn2 -YZn alloys the trans.formation gives a weak thermal arrest at 750°C, whereas in the two phase YZn2-YZn3 alloys no thermal arrest is observed and the transformation occurs over a temperature range below 750°C. At 500°C the free mzergies of formation per lnole vavy from —18,090 for YZn to —53,430 fov YZr11 and corresponding enthalpies vary from -24,050 to -92,080. The free energies and enthalpies per g atom as a function of composition show a maximum for the YZn2 phase; the 500°C values are -9580 and -13,180, vespectively. 1 HE only information found in the literature on Y-Zn alloys was the observation reported by Carlson, Schmidt. and speddingl that Y-20 wt pct Zn forms a low melting alloy. The alloy was produced by the bomb-reduction of YF3 and ZnF2 with calcium in an investigation of methods for producing yttrium metal. The solubility of yttrium in zinc has been determined by P. F. woerner2 and reported by Chiotti, Woerner, and Parry.3 In the temperature range 495" to 685°C the solubility may be represented by the relation In these equations N represents atom fraction of yttrium and T is the temperature in degrees Kelvin. The purpose of the present investigation was to establish the phase diagram for the Y-Zn system and to determine the standard free energy, enthalpy, and entropy of formation for the compounds formed. MATERIALS AND EXPERIMENTAL PROCEDURES The metals used in the preparation of alloys were Bunker Hill slab zinc, 99.99 pct pure, and Ames Laboratory yttrium sponge. Arc-melted yttrium buttons contained the following impurities in parts per million: C-129, N-12, 0-307, Fe-209, Ni-126, Mg-13, Ca < 10, F-105, and Ti < 50. Some of the alloys containing 70 wt pct or more of Zn were prepared from yttrium containing 5000 ppm Ti as a major impurity. Tantalum containers were found to be suitable for all alloys studied and were used throughout the investigation. The pure metals, total weight about 30 g, were sealed in 1 in. diam tantalum crucibles by welding on preformed tantalum covers. A 1/8 in. diam tantalum tube was welded in the base of each crucible for use as a thermocouple well. Welding was done with a heli-arc in a glove box which was initially evacuated and filled with argon. The sealed crucibles were enclosed in stainless steel jackets and heated in an oscillating furnace at temperatures up to 1150°C. Homogeneous liquid alloys were obtained within a half hr at these temperatures except for alloys containing less than 20 pct zinc. The latter alloys were held at 1000º to 1100°C for 2 to 3 days in order to obtain equilibrium. After the initial equilibrations the tantalum crucibles containing the alloys were removed from the steel containers and used directly for differential thermal analyses. Further annealing heat treatments for alloys in which peritectic reactions were involved were carried out in the thermal analyses furnace. After thermal analyses the tantalum crucibles were opened and the alloys sectioned and polished for metallographic examination. In the following discussion alloys referred to as "quenched" were obtained by quenching the sealed stainless steel jacket containing the tantalum crucible and alloy in water. The differential thermal analyses apparatus used was a modified version of the one described in an earlier paper., The graphite crucible was replaced by an inconel crucible, the nickel standard and sampie container were separated by a 1/8 in. MgO plate, no getter was used, and provisions were made to

Jan 1, 1963

By Gordon M. Pugh, David G. Rasmussen

INTRODUCTION Variations of cut-and-fill stoping are the most com¬monly used methods for mining narrow, nearly vertical veins in the United States. These methods offer the best results for maximum recovery with minimum dilution of ore from veins that are irregular along either the strike or dip. The vein is developed along the strike length in 61 to 122-m (200 to 400-ft) blocks or stopes and vertical lifts of 48 to 61 m (160 to 200 ft). The level development is either on the vein or parallel to the vein with crosscuts intersecting the vein at right angles. Chutes, manways, and haulageways are established in the center of the stopes at the lower level of the ore block. Mining then proceeds by blasting down horizon¬tal cuts of the vein in both directions from the chutes. As the blasting continues to the outer limits of the stope, the broken ore is slushed into chutes. When a complete horizontal cut has been made, the service raise, including chutes, is raised to a height of a few feet from the back and the stope is filled with either sand fill or waste rock. Then the mining cycle is re¬peated until the stope reaches the level above. LEVEL DEVELOPMENT PRACTICES FOROVERHAND CUT-AND-FILL STOPING The two basic methods used are: the development of the haulage drift within the vein, or development of the haulage drift paralleling the vein in the foot or hang¬ing wall with crosscuts intersecting the vein perpendicu¬larly at 61 to 122-m (200 to 400-ft) intervals. The il¬lustrated methods of development show some of the variations which can be used to develop a stope block. Development for cut-and-fill stoping does not require the more elaborate schemes outlined for shrinkage or sublevel long-hole stoping. Many factors must be considered before deciding on a particular development plan and usually the nature of the vein or host rock will dictate which specific method to use. If the vein is relatively straight, then develop¬ment on the vein is feasible. Excessive ground pressures will necessitate the leaving of a sill pillar (Fig. la) and highly fractured or "heavy" ground will require timber¬ing (Fig. 1b). Development on the vein has the advan¬tage of reducing capital outlay and waste handling prob¬lems since the mined rock has value. But removal of the sill pillar is costly and time-consuming. Also, where no sill pillar is left, excessive timbering and subsequent level repair is a drawback. Development in either the footwall or hanging wall is the more common procedure and offers greater flexi¬bility in overall development of consecutive stoping blocks by partially disassociating the level development from stoping (Fig. 2). Chutes, manways, and service compartments com¬prise the next stage of development and are established in a variety of ways. They provide access to the stope, ventilation ducts, and muck removal. Several different development systems are in common use and average costs for each have been developed. Raise borers are used to establish ventilation and ac¬cess from level to level. They are normally reamed with 1.5 or 1.8-m (5 or 6-ft) cutterhead and may either be left bald or lined with a 1.2-m (4-ft) diam corrugated liner plate, depending upon ground conditions. The use of an Alimak raise climber is another effec¬tive method of driving a raise from level to level. The size usually is 2 m (7 ft) square. This will accommodate square-set timber which is later installed. This method of raising is not common, although it is a valuable tech¬nique especially when level spacing is several hundred feet.

Jan 1, 1982

ASSETS Working Fund Assets Cash-unappropriated $62,41294 Members&apos; 1951 dues receivable 14,336 98 Accounts receivable Advertising, publications and miscellaneous $43,22055 Less-Reserve for delinquent accounts 12,775 76 $30,44479 Petroleum Branch-Special Projects Fund¬ Advances re Petroleum Conservation Volume 5,37085 35,81564

Jan 1, 1952

By T. J. Ballard

The Rickard mine in York County, S. C., named for the original owner, Alexander Rickard, was operated about 50 years as. No records are available of the quantity of corundum shipped. Pratt states, "Many tons were picked up from the surface and shipped." The mine was visited on August 6, 1943, by the author, accompanied by Gilbert W. Withers of Atlanta, Ga. Investigation of the property was begun by the Bureau of Mines after some initial work by the Mi-Cor Corp., a company organized by Gilbert P. Withers of Atlanta to explore the property. In conducting the Bureau program, 754 linear feet of trenches were dug, 110 feet of old drifts cleaned out and examined, and 38 feet of shaft retimbered.

Jan 1, 1948

By W. G. Sinclair

The wartime momentum of exploration continued throughout the year 1945 despite the end of hostilities in mid-August. The table below illustrates drilling activity in the various categories: Completed wells. . . . . . . . . . . . . . . . . . . . . . . . 401 Wildcats.. . . . . . . . . . . . . . . . . . . . . . . . . . . . . . .104 Wildcat discoveries, oil.. . . . . . . . . . . . . . . . . . . 2 Wildcat discoveries, gas-condensate.. . . . . . . 4 Development wells.. . . . . . . . . . . . . . . . . . . . . .297

Jan 1, 1946

By Regis M. N. Pelloux, N. J. Grant

Five or six alloys each in the systems Ni-C.v, Ni-Mo, and NL-W, spaced to cover the single phase areas as well as a part of the adjacent two-phase field, were prepared as uacuum-melted alloys. Tensile tests at room temperature and stress rupture tests at 1200° and 1500°F we.ve run for time periods to give fracture in 0.1 to 500 hr. Observations mere made of solid-solution and second-phase strengthening or uleakening, coupled with studies of ductility and the role of structure on the noted behavior patterns. THE development of stable creep-resistant alloys while broad and systematic is also highly empirical. A number of simple rules have been suggestedL-" but these are based on scant data. Whereas rules for low-temperature alloy strengthening can be applied with some success in more simple systems, the extension to behavior at high temperatures (greater than about 0.4 of the absolute melting temperature) is very poorly understood and has encountered important exceptions and deviations in terms of atom size or valency effects. In particular, the occurrence of both solid-solution weakening4,5 and second-phase weakening2 at high temperatures is of interest. MATERIALS AND PROCEDURE Nickel was chosen as the base material because of its important role in current high-temperature alloy applications, because of its freedom from a phase transformation, and because of wide solubilities for many elements of interest. The selected alloys were single- and two-phase structures in the Ni-Cr,6 Ni-Mo,7 and Ni-w 8 systems. The alloys were vacuum melted and cast, utilizing high-purity raw materials, in the form of 15 lb ingots. They were readily forged into 1/2-in. bar stock (except the two highest chromium alloys). The alloy compositions are shown in Table I. Typical impurity values are: Carbon: Max 0.03 pct (except for 30 and 35 pct Cr alloys which had 0.07) Nitrogen: Max 0.005 pct (except for two of the Cr alloys which had 0.05) Sulfur: Max 0.004 pct Silicon: Max 0.05 pct Each alloy was solution treated (air quenched) to produce comparable grain size. The two-phase alloys were then overaged in an effort to produce a stable structure for testing at 1200° and 1500°F, The results are summarized in Table 11. Following the heat treatments outlined in Table 11, the bars were machined to a gage section of 1.25-in. long by 0.250-in. diam. In addition to room-temperature tensile tests, stress-rupture tests were run at 1200" and 1500°F for time periods from 0.1 to about 500 hr. RESULTS Room-Temperature Mechanical Properties—The ultimate tensile strength, total elongation, and reduction of area values for all the alloys are plotted vs atomic percent of the solute element in Fig. 1. It can be seen that for the same amount of solute element the ultimate tensile strength increases in the order chromium, molybdenum, tungsten. Elongation increases (reduction of area decreases) with increasing amounts of solute for all the solid solution alloys, the values ranging from 50 to 70 pct. The ultimate tensile strength values for the two-

Jan 1, 1961

By R. M. Hardgrove

PULVERIZED coal was first used for firing cement kilns about 45 years ago, with such success that it has continued in general use. Based on cost, pulverized coal is usually the most economical fuel. In some localities natural gas or oil is cheaper, but 85.5 per cent of all cement kilns at present are fired with pulverized coal. The pulverized-coal flame is well suited to kiln conditions. Its soft, luminous nature gives it a high radiating property and the length of the flame can be readily varied to suit conditions in a kiln of any length. Ash and slag cause most of the difficulties incident to the burning of coal. In a cement kiln, however, the ash, being similar in composition to the raw mix and being deposited uniformly in the mix and clinker, occasions no difficulty as long as it remains constant in quality and quantity. Until 6 years ago almost all cement kilns using pulverized coal employed the bin system. The coal was dried and pulverized in a separate plant and delivered into a bin at the kiln. Thus the bin system requires driers, storage-type pulverizers, conveying systems, individual kiln bins, pulverized-coal feeders and primary air fans. The direct-fired system requires only a pulverizer and a blower. Fig. I shows typical plants. The first installations of direct-fired pulverizers were of the impact type but fineness of pulverization was both inadequate and variable. About 1927 a few direct-fired installations employing tube-type pulverizers were made, which were much more satisfactory. As more efficient pulverizers were developed from 1934 on, equipment for direct firing was installed at a large number of plants, until now 196 of about 900 kilns in use in the United States are direct-fired. Most of these replaced antiquated bin systems. PROCESS OF BURNING CEMENT CLINKERS The process of burning cement clinkers is fundamentally one of heat transfer. In boilers and other heat-exchange apparatus the area of the surface receiving heat from the gases is the best unit for comparing output. This seems to be equally true for kilns. The inside kiln surface is the heat-receiving surface and its area is used as the unit of kiln size in the following comparisons. Heat input is based on British thermal units in the fuel. E. G. Bailey,1 in his work on boiler-furnace heat absorption, has used the unit "B.t.u. available." This is the practical calorific power of the fuel, plus the heat in the air used for combustion. In other words, it is the heat available for useful work in the furnace or kiln. The unit of kiln output is barrels per hour. The composition of the raw mix, the percentage of water in the slurry of wet plants and hardness of burning all affect the output. These conditions, however, are hard to evaluate and probably have less effect on the output than irregularities of raw feed and clinker discharge and inexpert control of combustion. WET VS. DRY PLANTS The study presented here covers 88 plants, of which 41 are wet and 47 dry. Of

Jan 1, 1941