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Vol. 54 will be sent about February 1 to members who have paid their dues for the current year. Members who do not receive copies within a reasonable time are asked to notify the Secretary.

Jan 2, 1917

By C. H. Hewitt, J. T. Morgan

The Fry cocurrent in situ combustion project was carried out in a 3.3-acre portion of a lenticular body of Robinson sandstone of Lower Pennsylvanian age. This particular sand body is about 12,000 ft long and 3,500 ft wide; it varies in thickness from 0 to 55 ft, trends northeast-southwest, and occurs at depths between 880 and 936 ft. The reservoir is made lip of three distinct types of sandstone, each having characteristic reservoir properties. (See Table I.) Textural gradients, sedimentary structures, erosional basal contact, absence of marine fossils, and geometry of the sand body all indicate that the sandstone is of fluvial origin and part of an extensive complex of river-deposited sedirnents. Geographically oriented cores reveal that the direction of depositing currents was to the southwest, parallel with the sand-body trend. Directions of maximum pernleability also parallel the northeast-southwest trend. The layered nature of the sandstone reservoir and its directional properties have controlled the areal distribution of produced gas, the ignition behavior, and the vertical profile and horizontal extent of the conbustion zone. INTRODUCTION This paper is the first of a series of three papers that describe Marathon Oil Co.'s Fry in situ combustion test. This first paper contains the results of a detailed study of the geology and reservoir characteristics of the Robinson sandstone reservoir in which the Fry test was conducted. The second and third parts of this series report on the field operations' and performance. The geologic work reiated to cores cut through the combustion zone, and to the effect of reservoir properties on the combustion process is included in the third paper.' GEOLOGIC STUDY OF THE FRY RESERVOIR The close spacing of continuous cores in the Fry combustion site provided an unusual opportunity to study the details of distribution of lithology and reservoir properties within a petroleum reservoir. The objectives of the study were: 1. To determine the distribution of rock types within the Robinson sandstone in the Fry area; 2. To determine the relationship of porosity and permeability to rock properties and establish the distribution and homogeneity of reservoir properties; 3. To determine the origin of the Robinson reservoirs so that attempts could be made to predict the size, shape, distribution, extent and continuity of Robinson reservoirs; and 4. To determine what effects, if any, the reservoir rocks might have on the combustion process, so that the process might be better evaluated. AVAILABLE MATERIALS Continuous cores were cut from 14 wells in the vicinity of the Fry site. Ten of the cores are from within the 3.3-acre pattern; six of the 10 were cut prior to combustion, and four were cut through the combustion zone. Cores from the Emma Fry No. 15, Wampler 0-2 and Fry B wells (Figs. 1, 2 and 9) were geographically oriented by the Eastman Oil Well Survey Co. Maps, electric

Jan 1, 1966

By R. F. Ashton, D. S. Thompson

In reporting results of precipitation hardening experiments, it is customary to include such conditions as solution heat-treatment temperature, specimen size, and quench medium as well as the aging time and temperature. One portion of thermal history that is often overlooked is the effect of the heating rate employed to achieve the aging temperature. It is the intent of this note to demonstrate the importance of the rate of heating to aging temperatures. This work was carried out on the aluminum alloy 7075 aged at 162oC, a temperature commonly used to obtain stress corrosion resistant tempers. A 2 in. thick plate of 7075 was quenched into 24°C water after solution heat-treatment for 2 hr at 470°C. Approximately 13 in. around the periphery of the plate was discarded to minimize quench rate variations, and the remainder was divided into 0.5 by 1.8 by 2.0 in. blocks which were held at room temperature, 24oC, for 4 days. This aging, or incubation period, at room temperature allowed a copious supply of GP-zones to form. Blocks were then heated at various rates to 162oC using the practices shown in Table I. Both single- and two-step aging practices were employed for each heating rate. The single step consisted of heating directly to temperature at the indicated rate while the two-stage practice included an interruption of heating for 7 hr at 107°C. In the case of the two-stage 100°C per min heating rate experiment, the specimens were returned to room temperature after 7 hr at 107°C where they were held for 16 hr be-fore finally heating at this rate to 162°C. This devia-tion in practice is not considered to be significant. Blocks were removed after various times at 162oC, and aging curves were determined from duplicate R. F. ASHTON and D. S. THOMPSON,Member AIME, are Senior Scientists, Metallurgical Research Division, Reynolds Metals Company, Richmond. Va. Manuscript submitted December 2. 1968. IMD Table I. Aging Practices Two-step Single-Step Heating Rate* Aging Aging Heating Method oC per Min A B Programmed Air Oven 0.23 * 0.01 C D Programmed Air Oven 0.93 ? 0.03 E F Air Oven at Temp -10 G H Oil Bath at Temp ˜100 *Heating rate was measured with a thermocouple at the center of a specimen. An average value between the starting temperature and 3°C below the aging temperature is reported for practices E-H. PRACTICES 70 — A B C E and G 50 - 2 4 6 810 20 40 60 HRS.AT 162°C Fig. 1—The influence of processing variables on 7075 aging curves. Fig. 2—7075 aged according to practice H (single-step aged, fast heating rate) for 3 hr. Magnification 100,000 times.

Jan 1, 1970

By W. E. Hess

AT the Vesta mines of Jones & Laughlin Steel Corp. on the Monongahela River, 35 miles south of Pittsburgh, JCM Joy continuous miners and 6-SC shuttle cars are used for pillar extraction in the Pittsburgh seam. The entire output of the mine goes into coke production for blast furnace operation. Because this coal, which has high metallurgical qualities, is close to the place where it is ultimately used and because supplies with these properties are being rapidly depleted, complete recovery is essential wherever possible. With the introduction of the continuous miner into the Vesta mines, therefore, it was necessary to plan a project which would facilitate development and provide for future total recovery. Use of continuous miners under varying conditions has necessitated changes with regard to roof action, method of development, method of pillar attack, and innovations in mining. In the following pages these changes will be discussed. Prior to the introduction of continuous miners the Pittsburgh seam at the Vesta mines averaged 72 in. of mineable coal, overlain with 12 to 16 in. of draw-slate, which weathers on exposure to air and is not firm enough to be held in place. Immediately overlying the drawslate is a wild or rooster coal, very irregular in thickness, varying from 12 to 30 in. Above the rooster coal the strata consists predominately of shales or lime shales and sandstone. Just before the introduction of continuous miners in the Vesta mines, however, a definite change took place in the character of the roof structure overlying the coal seam, see Fig. 1. Thickness of laminated roof immediately overlying the rooster coal varies with localities and introduces a problem in roof control. In fact, a set of eight main entries being developed in Vesta NO. 4 mine entered an area with so difficult a roof structure that management considered abandoning the section altogether. In this locality the roof consisted of thin alternate layers of roof coals and shales, with an occasional irregular seam of slick slate, the whole being laced with slips, and the coal itself being cut by clay veins and spars. In an effort to overcome some of the difficulties, roof bolting was introduced in 1949. This made it possible to recover coal from an area which otherwise would have been economically unmineable. There have been very few roof bolt failures. When conventional machines were replaced by continuous miners there were changes in the development of butt entries. By these changes it was hoped: 1—to speed development, 2—to provide as much solid coal as possible for the machine on retreat, 3—to avoid the necessity of crossing entries already driven, 4—to facilitate ventilation, 5—to provide a roadway for shuttle car haulage which would eliminate the necessity of running through doors or curtains, 6—to provide better roof control by increasing the distance between the center lines of entries, and 7—to provide more coal when the machine was in position to load.

Jan 1, 1956

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

In an earlier paper1 we discussed a simple, rapid method of taking samples of liquid steel and analyzing them for oxygen, which, though possibly not absolutely accurate (as is likewise true of all other methods), certainly yields results that are satisfactorily consistent. Accordingly, samples from hundreds of heats ill different shops were analyzed for oxygen, with results sufficient in number to be handled statistically and permit estimation of some general trends. Measurements were made in acid as well as in basic open hearths, throughout the refining period, to learn how stirring, blocking, adding ore or lime or sand, affects the oxygen content of the bath; we also followed the changes brought about by tapping, teeming, and deoxidizing. Some of these results, discussed in an earlier paper,2 bring out the dominant role of the carbon-oxygen reaction in regulating degree of oxidation both of metal and of slag, also the marked variability of concentration of dissolved oxygen [0] in excess of that which would be in equilibrium with carbon and with carbon monoxide at a partial pressure of one atmosphere. As a logical beginning on this complex problem, we take up in the present paper the variability of this excess oxygen, A[O], when other factors are as constant as they ever are in open-hearth operation; namely, toward the end of the refining period, when the bath may be regarded as in the comparatively steady state it approaches in absence of disturbing factors. This condition was chosen because by that time the significant reactions, except carbon oxidation, between slag and metal have come to substantial equilibrium; this facilitates interpretation of the complex situation and thus perinits more definite conclusions to be drawn. This excess concentration of oxygen, A[O], in the metal is needed, of course, to make the steelmaking process go at a reasonable rate; alternatively it may be regarded as a measure of the virtual pressure increment above one atmosphere, or of the degree of supersaturation, of carbon monoxide in the metal. According to our measurements, the excess concentration A[0] in liquid steel varies from about 0.009 to 0,035 per cent oxygen by weight. It is quite sensitive to change of conditions, being altered temporarily by operations such as stirring with green poles, additions of ore, lime, or sand; but as the disturbance works itself out, it resumes a level within the intermediate range characteristic of the steady state. The conditions most common in average furnace practice are covered by the following: i[O] Usual Per Cent Oxygen by Conditions Weight 1. Vigorous lime boil........... o.oog-0.01s 2. Steady state................ o.or;-0.025 3. Working of ore.............. o.023-0.035 Thus when the bath can be regarded as in a reasonably steady state—that is, when it is boiling normally, is at proper temperature with a well-shaped slag, and has been free from disturbing effects during the preceding half hour—A[0] returns from

Jan 1, 1948

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

In an earlier paper1 we discussed a simple, rapid method of taking samples of liquid steel and analyzing them for oxygen, which, though possibly not absolutely accurate (as is likewise true of all other methods), certainly yields results that are satisfactorily consistent. Accordingly, samples from hundreds of heats ill different shops were analyzed for oxygen, with results sufficient in number to be handled statistically and permit estimation of some general trends. Measurements were made in acid as well as in basic open hearths, throughout the refining period, to learn how stirring, blocking, adding ore or lime or sand, affects the oxygen content of the bath; we also followed the changes brought about by tapping, teeming, and deoxidizing. Some of these results, discussed in an earlier paper,2 bring out the dominant role of the carbon-oxygen reaction in regulating degree of oxidation both of metal and of slag, also the marked variability of concentration of dissolved oxygen [0] in excess of that which would be in equilibrium with carbon and with carbon monoxide at a partial pressure of one atmosphere. As a logical beginning on this complex problem, we take up in the present paper the variability of this excess oxygen, A[O], when other factors are as constant as they ever are in open-hearth operation; namely, toward the end of the refining period, when the bath may be regarded as in the comparatively steady state it approaches in absence of disturbing factors. This condition was chosen because by that time the significant reactions, except carbon oxidation, between slag and metal have come to substantial equilibrium; this facilitates interpretation of the complex situation and thus perinits more definite conclusions to be drawn. This excess concentration of oxygen, A[O], in the metal is needed, of course, to make the steelmaking process go at a reasonable rate; alternatively it may be regarded as a measure of the virtual pressure increment above one atmosphere, or of the degree of supersaturation, of carbon monoxide in the metal. According to our measurements, the excess concentration A[0] in liquid steel varies from about 0.009 to 0,035 per cent oxygen by weight. It is quite sensitive to change of conditions, being altered temporarily by operations such as stirring with green poles, additions of ore, lime, or sand; but as the disturbance works itself out, it resumes a level within the intermediate range characteristic of the steady state. The conditions most common in average furnace practice are covered by the following: i[O] Usual Per Cent Oxygen by Conditions Weight 1. Vigorous lime boil........... o.oog-0.01s 2. Steady state................ o.or;-0.025 3. Working of ore.............. o.023-0.035 Thus when the bath can be regarded as in a reasonably steady state—that is, when it is boiling normally, is at proper temperature with a well-shaped slag, and has been free from disturbing effects during the preceding half hour—A[0] returns from

Jan 1, 1948

By W. J. Babyak

The habit planes for hydride precipitation in large grains of alpha zirconium and Zircaloy-4 were determined in specimens containing 161 and 136 ppm hydrogen, respectively. In zirconium, the hydride platelets precipitated on various pyramidal planes inclined to the basal plane at angles ranging from 10 to 50 deg and 75 to 90 deg. Hydrides were most frequent at inclinations of about 20 deg and 90 deg. In Zircaloy-4, hydride plates were inclined to the basal plane at angles vanging from 5 to 25 deg with the predominant inclinatioiz equal to 12 deg; no platelets were found near the prism plane position. In both zircoinium and Zir-caloy-4, hydrides were principally, but not exclusively, parallel to (1011) type planes. When the Zircaloy-4 gmins were subjected to tension stresses of 12,100 psi and 19,400 psi during hydride precipitation, the hydride habit was unchanged except that precipitation favored those habit planes which were inore nearly perpendicular to the stress direction. The stress-orientatioiz effect increased with the magnitude of the stress. THE hydride habit plane in a zirconium has been determined by several investigators. Langeron and Lehr1 in 1956 found platelets parallel to (1010) slip planes. In 1962, Westlake and Fisher2 confirmed these findings and showed that hydride precipitation on twin planes, as reported by Kunz and Bibb,3 was probably due to incomplete removal of worked metal from specimen surfaces. Secondary habit planes, ( 1011) and (1015), were also found by Westlake and Fisher under particular charging conditions where a hydride layer may have existed at the hydrogenation temperature. Most recently, walters4 and westlake5 again observed the (1010) habit exclusively in zirconium. Louthan and hgermans have reported that hydrides precipitate on various ( 101l ) and (112l) planes positioned 31.5 to 90 deg to the basal plane in Zircaloy-2. The present investigation was concerned with the hydride habit in Zircaloy-4 in relation to that in zirconium and the possible change in habit when precipitation occurs under stress. The latter consideration was prompted by the stress-orientation phenomenon originally reported by Marshall and Louthan.7, 8 I) EXPERIMENTAL Materials and Specimens. Chemical analyses of all materials are given in Table I. The zirconium specimen was a 1 by 4 in. section cut from wrought crystal bar that was hot-rolled at 790°C to 0.300 in. and cold-rolled to a final thickness of 0.095 in. All surfaces were mechanically polished and then chemically polished with a solution of H2O, HNO3, and HF (10:10:l by volume) prior to annealing. Grains as long as 0.25 in. were grown by thermally cycling the specimen three times (in 10-5 mm Hg vacuum) between 825" and 1000°C at approximately 125°C per hr. The specimen was isothermally annealed at 825°C for 24, 40, and 120 hr after each cy-

Jan 1, 1968

By Ferit Gurses

EXTENSIVE coal and lignite deposits exist in Turkey. Bituminous coal is the nation's principal mineral resource; important not only as fuel for the industrial development of the country, but also as a means of obtaining foreign exchange. Of the electrical power generated during 1939, 82.8 per cent was derived from coal.1 It is anticipated that the lignite deposits will be one of the major sources of power for the electrification of the country. BITUMINOUS COAL The locations of the four principal bituminous-coal2 basins are indicated on Fig. 1. The Eregli-Zonguldak basin lies on the Black Sea northwest of Ankara; the Canakkale basin lies near the junction of the Dardannelles; the Trakya basin is in European Turkey northwest of the Bosphorus and the Erzurum basin lies far to the east. The Eregli-Zonguldak basin is the most important of the four. This coal basin was discovered in 1829 but did not arouse interest until 1848; up to 1865, its annual production had not exceeded 50,000 metric tons. In 1865, during the Sultanate, working of coal mines was started by the Navy department, and the annual production was 61,145 metric tons. In 1908, the Minister of Commerce and Agriculture was authorized to develop the whole coal basin. Production reached 264,397 metric tons in 1910, and during 1915-1922, preceding the revolution, the total production of the basin was 2,676,043, or an average of 334,000 metric tons annually. The period of national revolution that followed was marked by a steady increase in coal production in Turkey, although world production decreased, because of the crisis of 1929-1932, about 52 per cent. AREA AND GEOLOGY OF THE COAL BASIN The geology of the Eregli-Zonguldak area is not yet completely known. The basin begins at Gungazi, to the west of Eregli, and extends to Sogutozu, a distance of about 106 miles. The coal beds, however, appear only in places and the overlying Cretaceous strata hide them elsewhere. This area includes the most important centers, such as Kandilli, Alacaagzi, Teflenli, Kozlu, Gelik (see Fig. 2). The geological sequence of the district is correlated with the Cretaceous of the Balkans; which is, in turn, a prolongation of the Alpine zone. Geological surveys made by G. Ralli, R. Zeiller and P. Arni3 show that the Eregli-Zonguldak formations fall into three natural divisions. Series a, The Alacaagzi in the district of that name (Fig. 2), is of the Culm stage. The strata strike generally east-west and dip about 30° south. This series appears also around Kandilli and Camli, with a steeper angle of dip, but the deposits there lack regularity. The Alacaagzi stage appears also farther east, to the south of Kozlu and Gelik. The Alacaagzi series contains 15 seams of an average thickness of 2 ½ ft. of good coal. Series b, the Kozlu, best represented in Zonguldak, Kilimli, Catalagzi, and between

Jan 1, 1944

By J. H. East, E. R. Maize

Probably no other type of equipment is now being introduced into American underground mines about which less is known and about which there is more misinformation than the Diesel mine locomotive. Any equipment installed underground brings with it hazards that were not present before, and it is the purpose of this paper to show that Diesel-powered equipment can be used with safety in underground mining if certain feasible but very important precautions are taken. Diesel locomotives are not new in underground mining, since they have been used successfully in European mines for at least 19 years, but their use in American mines is exceedingly limited; so far as is known, only five Diesel mine locomotives are operating at present in United States mines. Diesel locomotives were permitted underground in Great Britain for the first time in 1939, and the rapid increase in the number in use in those mines testifies to their safety and efficiency. More than 1300 Diesel mine locomotives are in use in European mines, according to the latest statistics; these locomotives are employed in coal, iron, lead-zinc and mercury mines. SO far as is known, the Diesel mine locomotive has not yet caused an explosion or mine fire, and has not been the cause of any fatalities from toxic gases in the mines in which it has been used. Adjustments of Diesel Engines foR Underground Use A Diesel engine adjusted to work in an underground mine should not be confused with the same engine adjusted for use outside the mine. Adjustment of the air:fuel ratio for engines that are used on the surface is intended to give the maximum fuel economy, and little or no attention is given to the chemical composition of the exhaust gas produced. Many such engines are poorly adjusted and perhaps poorly maintained, with the result that the engine emits smoke in quantity as well as the disagreeable odor familiarly associated with Diesel engines. The air:fuel ratio in the case of Diesel engines operated on the surface is generally around 15:1; that is, 15 lb of air to burn I lb of fuel. Diesel engines adjusted to work underground usually have an air:fuel ratio of about 20:1, as at this ratio the minimum quantity of toxic gases is given off in the engine exhaust. Diesel engines operated underground must be maintained in better condition than similar engines used on the surface; otherwise, the smoke emitted in the exhaust becomes objectionable if not unendurable. Practically normal air is required for safe operation of Diesel engines underground or in confined places; failure to provide virtually normal air for Diesel engine operation is almost certain to result in trouble. Bureau of Mines engineers have determined that the air in which a Dicsel Operates hi confined places should contain less than % pct of carbon dioxide and at least 20 pct of

Jan 1, 1949

By A. D. Dorenfeld

Factorial design is a mathematical method of drawing valid conclusions from a series of tests made in a predetermined pattern. It is applied to flotation ore testing using, in this case, five variables, each varied twice, requiring a total of 32 tests. Representative conclusions are checked by repeat testing. THE factorial design method of testing is particu-J- larly applicable to mill-scale plant tests because 1—the conclusions take into consideration all variabilities of the plant, 2—there are no subjective guesses as to the conclusions, 3—the number of tests necessary to check large numbers of variables are at a minimum, and 4—interactions between variables can be discerned. It is equally applicable to laboratory testing because once the variables to be investigated have been decided upon, as well as the number of levels for each variable, a well-trained assistant can run the tests, since the test work follows a predetermined mathematical pattern, leaving the conclusions to be found by the person in charge. The conclusion takes into consideration all errors of the tests. To illustrate the use of factorial design, a sample of copper porphyry ore was tested, the copper content of the batches not being uniform. Five variables were investigated, each at two levels, requiring only 32 tests. After suitable arithmetical analysis, conclusions for copper recovery and concentrate grade are arrived at. Four cases of these conclusions are checked by repeat testing. Testing Methods Reproducibility of flotation tests varies widely, depending on the experimenter, the process, the equipment, and probably a host of unknown and perhaps interacting variables. If the reproducibility of tests is good, let us say to 0.2 pct recovery of a certain metal, and the minimum difference between tests upon the variation of the interested variables is, say 1 pct, then simple averaging and inspection is ample in interpreting results. However, often such a happy situation is not the case. This is particularly true in mill-scale operations where the ores are variable, and a host of minor operating fluctuations common to most milling, and familiar to all mill operators, is the normal state of affairs. Yet, mill scale tests are run, the results averaged, and conclusions drawn therefrom. The trustworthiness of such results should be measured by comparison to the experimental error—the measure of the degree of reproducibility of the results. This comparison is a test of significance. Thus, if daily results of recoveries average 80.0 pct, and a spread of 78.0 pct to 82.0 pct, is recorded, then, clearly, on changing mill operations, an 81 pct recovery cannot be construed as significant by using averages, since it is well within the previous test limits. By using rigorous statistical tests of significance, conclusions arrived at by inspection or simple averaging may be shown to be based on inadequate evidence or even wrong. In any event, a plant super-

Jan 1, 1952

By Andrew B. Crichton

THE oft repeated statements of the United States Geological Survey and the Bureau of Mines that the coal reserves in the United States are sufficient for 3000 yr have given us all a sense of security hard to disturb. If we discount that statement 66 pct, which to the uninformed seems drastic, we would still have 1000 yr supply, and why should mortal man worry about the difference between 1000 and 3000 yr supply of coal. That is exactly the attitude which many who should be better in- formed have taken toward the problem. A recent brochure "How Strong is America," by the New York Herald Tribune, states there is "no end to coal * * The truth is staggering * * U. S. has enough coal to last 4000 years." What the statement actually means is that the calculator got as high as 4000 yr and decided it was pointless to go on. The President's ERP Committee on Natural Resources and others in high places continue to assure the country of our superabundance of coal. In the past few days the “Wall Street Journal" carried an item from Director Boyd of the Bureau of Mines saying that within a hundred years we would need to triple our annual coal production of 6oo,ooo,ooo tons, making it 1,800,000,000 tons annually, and at that rate we had coal reserves sufficient for 1000 yr. In my humble opinion, we cannot now or ever produce 1,800,000,000 tons annually, be- cause we do not have the reserves in the producing districts to do it. Past reckoning in geology may be, as in the Psalms, a thousand years are as yesterday, but the future must be considered more realistically, and "a thousand years" is really a "heck" of a long time. In most developed areas, mining engineers can count the coal reserves in terms of a hundred years or less, and in some important areas now producing large annual tonnages in much under fifty years. They think when they hear of thousands of years coal reserves, reference is made to distant fields and "the fields are always green far away." The estimates of our coal reserves upon which we are relying were made in 1907 under the direction of Marius R. Campbell, Director of the United States Geological Survey. It is apparent this work was not well done. It was necessarily based on rather meager data because of the little development and prospecting as of that date. There were mistakes in the correlation of seams, and they assumed the continuity of thick seams throughout large coal bearing areas where coal often did not exist. The overstatement of coal bearing areas in the Campbell figures are common. The 1907 survey includes all coals over 14 in. in thickness to a depth of 3000 ft. The total coal reserves according to Campbell were 3157.2 billion tons. The production and mining losses that year would amount to -570 million tons and

Jan 1, 1948

By Andrew B. Crichton

The oft repeated statements of the United States Geological Survey and the Bureau of Mines that the coal reserves in the United States are sufficient for 3000 yr have given us all a sense of security hard to disturb. If we discount that statement 66 pct, which to the uninformed seems drastic, we would still have 1000 yr supply, and why should mortal man worry about the difference between 1000 and 3000 yr supply of coal. That is exactly the attitude which many who should be better informed have taken toward the problem. A recent brochure "How Strong is America," by the New York Herald Tribune, states there is "no end to coal * * The truth is staggering * * U. S. has enough coal to last 4000 years." What the statement actually means is that the calculator got as high as 4000 yr and decided it was pointless to go on. The President's ERP Committee on Natural Resources and others in high places continue to assure the country of our superabundance of coal. In the past few days the "Wall Street Journal" carried an item from Director Boyd of the Bureau of Mines saying that within a hundred years we would need to triple our annual coal production of 600,000,000 tons, making it 1,800,000,000 tons annually, and at that rate we had coal reserves sufficient for 1000 yr. In my humble opinion, we cannot now or ever produce 1,800,000,000 tons annually, because we do not have the reserves in the producing districts to do it. Past reckoning in geology may be, as in the Psalms, a thousand years are as yesterday, but the future must be considered more realistically, and " a thousand years" is really a "heck" of a long time. In most developed areas, mining engineers can count the coal reserves in terms of a hundred years or less, and in some important areas now producing large annual tonnages in much under fifty years. They think when they hear of thousands of years coal reserves, reference is made to distant fields and "the fields are always green far away." The estimates of our coal reserves upon which we are relying were made in 1907 under the direction of Marius R. Campbell, Director of the United States Geological Survey. It is apparent this work was not well done. It was necessarily based on rather meager data because of the little development and prospecting as of that date. There were mistakes in the correlation of seams, and they assumed the continuity of thick seams throughout large coal bearing areas where coal often did not exist. The overstatement of coal bearing areas in the Campbell figures are common. The 1907 survey includes all coals over 14 in. in thickness to a depth of 3000 it. The total coal reserves according to Campbell were 3157.2 billion tons. The production and mining losses that year would amount to 570 million tons and

Jan 1, 1949

By E. V. Potter, H. C. Lukens

In general, during all electrowinning processes, large volumes of gas are liberated at the cathodes of the electrolytic cells. Most of this gas escapes from the electrolyte, but much of it may be absorbed or entrapped in the metal being plated. In many cases, this is not objectionable, but since it is liberated when the metal is heated, it is a potential source of trouble where the metal has to be heated or melted. Since this gas is principally hydrogen, in escaping from a melt it may cause minor explosions and scattering of the molten metal, while in other cases it has been reported to make the metal rise in the molds when cast. Therefore such metals would be more generally useful if all or most of the gas were removed before they were used. In an earlier investigation1 the gas content of electrolytic manganese was determined and methods found for removing most of it without detrimentally affecting the properties of the metal. In this paper results of a similar investigation on electrolytic chromium are presented. Apparatus The apparatus used in these investigations is essentially the same as that used in the manganese work1 and is shown in Fig I. It consists of sample chamber A, vacuum or pressure gauge B, volume gauge C, vacuum pump D, hydrogen container E, auxiliary gas chamber F, and furnace G. The pressure in the system is measured by the mercury manometer B. Changes in gas volume are measured by volume gauge C which is essentially another mercury manometer in which the volume of the mercury column is large and can be varied by raising or lowering the well H. In this apparatus the tube had a volume of 100 cc for a 60 cm length. Larger changes in gas volume can be accommodated by auxiliary containers attached to the system at F. The sample is heated by furnace G, the temperature of which is regulated within 5°C by an automatic controller. The only change in the apparatus was the sample container which was made smaller because of the chromium metal's containing more gas than the manganese metal and a smaller sample was used. Method The measurements were made as follows: The complete system when cold was evacuated and refilled with H2 until all the air was removed. It was then refilled with H2 to the desired pressure, as indicated on manometer B. The temperature was then raised to the desired value and as gas was either liberated from or absorbed by the sample, the height of well H was changed to alter the volume of the system so as to keep the pressure constant. In determining rates of liberation of gas, volume readings were taken at convenient tirne intervals depending on the rapidity with which the gas was liberated. In determining total volumes, the specimen was held at constant temperature and pressure until

Jan 1, 1949

By E. O. Kirkendall

THE investigation reported herein was undertaken to obtain data to supplement those presented in an earlier paper,' particularly for longer times at a higher temperature. This information may further develop our knowledge of the mechanism of diffusion in substitutional solid solutions. Diffusion of solute atoms interstitially dissolved is not difficult to visualize; it can take place without much disturbance of the solvent atoms. The mechanism of diffusion of solute atoms substitutionally dissolved is not so easily understood. Many theories have been advanced. Rosenhain,2.3 Frenkel,4 Smekal,5 Langmuir,3,6 Steigman, Shockley and Nix,7 and Johnson8 all proposed theories to explain diffusion. Langmuir's proposed mechanism consisted of a simple cyclic interchange of atoms involving an interchange in lattice positions of four or more atoms. This mechanism, accepted by the author at the time of his previous publication,' and apparently quite generally accepted, required a net exchange in positions of both solute and solvent atoms in equal numbers. In other words, the rate of diffusion of the solvent and solute atoms would be the same. That this cyclic interchange in lattice positions of solute and solvent atoms in equal numbers is the only true mechanism of diffusion is denied by the evidence presented in this paper. EXPERIMENTAL PROCEDURE This modification of similar procedure1 will be outlined briefly. The method consisted of the following steps: I. A pure Muntz metal (60.6 per cent Cu, 0.1 per cent total impurities) that was almost all beta brass at 780°C. was chosen. 2. The brass was turned between centers in a lathe to a diameter of 0.5965 inches. 3. After the brass was cleaned, it was electroplated with pure copper to a diameter of about one inch. 4. The sample was heated slowly to 400°C. under reduced pressure and held at that temperature for 4 hr. to remove occluded gases. It was then cooled to room temperature. 5. The sample was heated to 780°C. ±3°C. and held at that temperature for different lengths of time; then quenched. Samples were cut off each time after quenching. Metallographic samples 34 in. thick were removed at I hr., 6 hr., 24 hr., 96 hr. and 701 hr. After 96 and 701 hr. samples of about I in. were removed for X-ray analysis. 6. The metallographic samples were polished on the cross section and photo-micrographs were taken. 7. The I-in. samples were polished and etched on both ends. With the aid of a microscope each was carefully mounted in a four-jaw chuck in a lathe, so that they could be accurately bored for an arbor. 8. Turnings were taken for each X-ray sample. After each set of turnings, the sample was heated in an oil bath for about 3a hr. at 300°C. It was then etched in 25

Jan 1, 1942

By Richard M. Foose

Most of the roofing granules produced in Pennsylvania are made by two companies at three plants. The Advance Industrial Supply Co. has three quarries and a mill at Gladhill Station, in southern Adams County. The Funkhouser Company is actively engaged in mining, quarrying and milling at Charmian, about one mile west of Gladhill Station, and at Delta, in southern York County. These three plants produced approximately 100,000 short tons of granules in 1943. The production of rooling granules in 1937 for the whole country was only about 500,000 tons, generally indicating the large growth of the industry since that time. The fundamentals of producing roofing granules were discussed comprehensively by G. W. Josephson in a paper before the Industrial Minerals Division in October 1943 (TP 1725, this volume page 460). The purpose of this paper is to describe the actual methods of mining the rock and milling it to make roofing granules, as practiced by the three Pennsylvania plants. Mining and Milling Funkhouser Company (Delta) The Funkhouser Company, near Delta, in southern York County, Pennsylvania, operates three joined quarries and a mine in the Peach Bottom slate. The joined quarries extend 1200 ft. in a northeast-southwest direction along the strike of the slate cleavage and are about 200 ft. .wide. The working face is about 45 ft. high. The quarrying operation is rather simple: Ingersoll-Rand Jackhamers, I-in. and 1 1/4-in., with hollow drill steel, are used to undercut the quarry face with 10-ft. holes. The face is "blown" and allowed to slump on the quarry floor. Loaded by a Marion electric shovel into trucks, the shattered rock is taken to the primary crusher 2500 ft. away. In addition to the normal quarrying operations, a drift into the northeast wall of the middle quarry was begun about 1935, and has advanced about 2000 ft. since then. Rooms have been driven at angles to the drift along choice beds of slate. The drifts and rooms have a roof more than 35 ft. high and the main drift and rooms are all more than 40 ft. wide. There is no cribbing at any place in the mine. Several of the rooms were driven upward on a 5' slope and in some of them there was stop-ing. Ingersoll-Rand Jackhamers are used, and only wet drilling is practiced. When visited in January 1940, an electric shovel was operating at the head of the drift, scooping up the slate blasted out of the working face. This, too, is trucked to the primary crusher. Some of the better quality slate is saved for making slate shingles. The primary crusher, which is a No. 30 McCulley gyratory, receives the slate from the trucks and crushes it to about a-in. size (Fig. I). This product travels by a 30-in. belt conveyor to Niagara vibrating screens with 2-in. holes. The minus a-in. product moves to a storage silo by bucket

Jan 1, 1948

By Richard M. Foose

Most of the roofing granules produced in Pennsylvania are made by two companies at three plants. The Advance Industrial Supply Co. has three quarries and a mill at Gladhill Station, in southern Adams County. The Funkhouser Company is actively engaged in mining, quarrying and milling at Charmian, about one mile west of Gladhill Station, and at Delta, in southern York County. These three plants produced approximately 100,000 short tons of granules in 1943. The production of rooling granules in 1937 for the whole country was only about 500,000 tons, generally indicating the large growth of the industry since that time. The fundamentals of producing roofing granules were discussed comprehensively by G. W. Josephson in a paper before the Industrial Minerals Division in October 1943 (TP 1725, this volume page 460). The purpose of this paper is to describe the actual methods of mining the rock and milling it to make roofing granules, as practiced by the three Pennsylvania plants. Mining and Milling Funkhouser Company (Delta) The Funkhouser Company, near Delta, in southern York County, Pennsylvania, operates three joined quarries and a mine in the Peach Bottom slate. The joined quarries extend 1200 ft. in a northeast-southwest direction along the strike of the slate cleavage and are about 200 ft. .wide. The working face is about 45 ft. high. The quarrying operation is rather simple: Ingersoll-Rand Jackhamers, I-in. and 1 1/4-in., with hollow drill steel, are used to undercut the quarry face with 10-ft. holes. The face is "blown" and allowed to slump on the quarry floor. Loaded by a Marion electric shovel into trucks, the shattered rock is taken to the primary crusher 2500 ft. away. In addition to the normal quarrying operations, a drift into the northeast wall of the middle quarry was begun about 1935, and has advanced about 2000 ft. since then. Rooms have been driven at angles to the drift along choice beds of slate. The drifts and rooms have a roof more than 35 ft. high and the main drift and rooms are all more than 40 ft. wide. There is no cribbing at any place in the mine. Several of the rooms were driven upward on a 5' slope and in some of them there was stop-ing. Ingersoll-Rand Jackhamers are used, and only wet drilling is practiced. When visited in January 1940, an electric shovel was operating at the head of the drift, scooping up the slate blasted out of the working face. This, too, is trucked to the primary crusher. Some of the better quality slate is saved for making slate shingles. The primary crusher, which is a No. 30 McCulley gyratory, receives the slate from the trucks and crushes it to about a-in. size (Fig. I). This product travels by a 30-in. belt conveyor to Niagara vibrating screens with 2-in. holes. The minus a-in. product moves to a storage silo by bucket

Jan 1, 1948

By H. O. Hofman, H. L. Norton

Zinc smelters in the central western States have established a very high standard of purity for blende-concentrates, viz., zinc 60, iron less than 3, and lead less than 1 per cent. The very low percentage of iron is made possible by the blende containing less than 0.5 per cent. of isomorphous iron sulphide and by the sparing occurrence of pyritic material in the large blende-deposits of southwestern Missouri and southeastern Kansas. The blende in these districts occurs in sub-carboniferous limestone, and is associated with chert, residual clay, cal-cite, pyrite or marcasite and bitumen. The different minerals are coarse and therefore readily separated by mechanical washing from the blende, with the exception of pyrite and marcasite, as blende has a specific gravity of from 3.9 to 4.1; marcasite, from 4.85 to 4.90; and pyrite, from 4.95 to 5.10. The common procedure of dressing the ores' is to crush in a breaker, size in a trommel with 0.5-in. holes, pass the over-size through roughing-rolls and then jig the ore, now reduced to 0.5 in. maximum size, in roughing multiple-sieve jigs. The products of these are waste-tailings, side-discharge coarse concentrates, side-discharge coarse middlings and bottom-discharge fine middlings. The coarse concentrates are sent to the market. The coarse middlings are crushed in middlings-rolls and brought back into the system by feeding into the main trommel. The fine middlings are cleaned upon finishing-jigs. Any pyrite or marcasite present follows the blende and increases the percent- +1MASSACHUSETTS INSTITUTE OF TECHNOLOGY, BOSTON, MASS.

Jan 1, 1905

By W. F. Holbrook, C. E. Wood, E. P. Barrett

THE high-frequency laboratory induction furnace with a rotating liquid crucible enables research workers to conduct certain investigations heretofore very difficult or impossible to realize because vessels are not obtainable that are capable of resisting the physical conditions or the chemical actions to which they are subjected. Applications for patents covering commercial uses of centrifugal liquid crucibles were filed in the United States in 1919,1 in France in 1925, in the United States and Germany in 1926,2 and in France in 1927.3 In 1928, Schuette and Maier,4 of the Bureau of Mines, reported the construction of a whirling table beneath a high-frequency induction furnace. SHAPE AND SIZE OF CENTRIFUGAL LIQUID CRUCIBLES When a vessel containing a liquid is rotated, the inner surface of the liquid takes the shape of a paraboloid under the simultaneous action of centrifugal force and gravity. The internal dimensions of the paraboloid rotating liquid crucible are governed by the diameter of the container and .its speed of rotation. Melted Wood's metal was poured into cylinders 2 1/2 in. in diameter, which were rotated at 180, 300, and 468 r.p.m. When the metal was solid, a slurry of plaster of Paris was added, and rotation continued until the plaster had set. Fig. 1, a photograph of sections of the cylinders, shows the effect of speed of rotation of the containers upon the shape of the inner surface of the rotating liquid crucibles. TYPES OF FURNACES Three types of high-frequency induction laboratory furnaces using liquid-metal crucibles were designed and constructed by the Blast

Jan 1, 1938

By E. W. Wagy

The year 1927 witnessed numerous developments of significance in California. The State's shut-in production increased from an average of 58,000 bbl. daily in January to a maximum of 93,000 bbl. daily in August, and a total of about 25,000,000 bbl. was withheld from the market during the year. This not only lessened the effect of large production at Hunt-ington Beach, Seal Beach and Alamitos Heights, but also was a most effective conservation move. Since 1922, it is estimated that approximately 100,000,000 bbl. of oil that would otherwise have been produced have been held in the ground. I. Outstanding Field Developments of 19271 1. Long Beach Deep Sand. 2. Rincon (Seac1iff)Discovery. 3. Alamitos Heights Discovery. 4. Ventura Avenue Extension. 5. Round Mountain Discovery. 6. Mt. Poso Extension. 7. Potrero (Cypress) Discovery. 8. Buttonwillow Gas Development. 9. Goleta Discovery. 10. Union Avenue, Edison and Fruitvale Areas. (See Fig. 1.) II. Crude Production The most important factor in new crude production in 1927 was the Seal Reach field with its adjacent Alamitos Heights townlot area. Production increased from about 10,000 bbl. daily in January, to a peak of 66,000 bbl. in June. Huntington Beach finished the year 1926 with a production of 102,000 bbl. daily, which declined to 60,000 bbl. by the end of 1927. Due to production in these three areas and the tempering effect of shutting in numerous wells, the California production curve for 1927 was held at a fairly even level. Fig. 2, shows the State's production in 1927 and 1926 as compared with 1923, which was the year of excessive

Jan 1, 1928

By G. T. Horne, P. J. Wray, S. Kinoshita

CONSIDERABLE evidence exists for a strengthening mechanism operative at intermediate temperatures in bee metals containing interstitial solute atoms. The strengthening is indicated by a maximum in the ultimate tensile strength1 as well as the fatigue

Jan 1, 1965