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By J. R. McWilliams

Several of the current concepts of brittle fracture involve consideration of the existence of defects or flaws. Griffith 1 observed that the tensile strength of brittle materials was several orders of magnitude less than might be expected from the calculated strength of the molecular bonds. He was able to show that the stress concentration at the tip of an elliptical crack is sufficient to raise the local stress level many fold. He therefore proposed that real materials do contain such stress risers in the form of minute cracks and that this accounts for the rather large difference between theoretical and actual strengths of brittle materials. The degree of stress magnification depends to a great extent upon the orientation of the crack with respect to the general stress system, and to the general geometry of the crack.

Jan 1, 1967

By L. D. Jaffe, R. K. Pitler

IT would clearly be advantageous to produce molten titanium, suitable for alloying and casting, directly from the relatively inexpensive tetra-chloride, without using a metallic reducing agent. Accordingly, a preliminary investigation has been made of the production of titanium by hydrogen reduction of the tetrachloride in an electric arc furnace. For the reaction: TiCl4(g) + 2H2(g) ? Ti(s) + 4HCl(g), Lockhart and his colleagues' have calculated equilibrium constants at 800" to 1500°K, based on tables of Brewer.' Extrapolating these data on the assumption that the heat of reaction varies linearly with temperature, and using Brewer's value for the heat of fusion, there is obtained for the reaction: TiCl4(g) + 2H2(g) ? Ti(l) + 4HCl(g) an equilibrium constant at 2100°K (100" above the melting point) Kp = (PHCl)4/pTlcl4 (Ph2)2 = 0.167 atm. If this is correct, starting with hydrogen saturated with TiC14 at 25°C, where the vapor pressure of the latter3 is 12.6 mm, and maintaining the overall pressure at 1 atm, 99 1/4 pct of the TiC14 would be reduced to Ti at equilibrium. An arc furnace designed for the melting of titanium and lined with water-cooled copper was used for the preliminary experimental check. Its inside diameter was 2 7/16 in. and height 6 in. The water-cooled electrode was tipped with 3/4 in. diam tung-sten. In the one run carried out, the furnace was charged with a starting batch of 293.35 g of scrap titanium,' analyzing 99.24 pct Ti. After the system was twice pumped out and flushed with argon to remove air, the arc was struck in argon and kept at 400 amp dc, electrode negative, for 1 min to melt the starting batch. Commercial tank hydrogen, dried by passing over silica gel and bubbled through technical grade TiC14 at room temperature, was then admitted to the furnace. The inlet tube directed this gas downward toward the arc and melt; the gas outlet was at the top of the furnace. The quantity of gas passed through was not measured accurately but is estimated to be about 50 liters per min at a few mm of mercury above atmospheric pressure. The arc was held in the H2-TiC14 mixture at 400 amp with little difficulty. After 10 min the arc burned through the copper lining above the melt, admitting water and ending the run. (This had happened frequently in straight melting runs under argon, and is not thought due to the atmosphere used.) In the furnace were found an ingot, several small pieces, and a little metallic powder, evidently formed when water entered the hot furnace. Some titanium also clung to the electrode. The portions were weighed and separately analyzed for titanium. Their total weight was 321.76 g of 94.82 pct overall purity. Much of the impurity consisted of tungsten broken off when the titanium was removed from the electrode. No detailed impurity analysis was made because the starting materials were impure and flooding the furnace introduced further contamination. Thus, the metal recovered contained 305.08 g Ti, as compared to 291.12 g in the starting batch. Titanium could not have been carried over from previous melting runs in the furnace, as the interior was sandblasted and the electrode ground clean before the run. Evidently at least 13.96 g of Ti were produced from TiCl4. Since this would require complete reduction of 430 liters of TiCl, gas at 12.6 mm of Hg, the efficiency of the process must have been rather high. Experimental work is continuing. No reason is seen why the process could not be scaled up to the size of a 30-ton steel-melting arc furnace. In a large unit there would be a large molten pool from which liquid metal could be withdrawn for alloying and casting. The problems of electrodes and power supply are similar to those arising in arc furnaces for melting titanium. Unreacted TiC14 and the HC1 in the exhaust gases could be recovered; unreacted H2 could be dried and re-used. The melt would be saturated with hydrogen, whose solubility in liquid titanium is unknown. Small amounts of hydrogen seem to have little effect on the properties of titanium.5 If the amount introduced was so large as to impair properties or cause casting difficulties, it could be reduced by holding the melt under an inert gas or by pumping on it, prior to casting. Acknowledgment The authors would like to thank I. Preble and M. Goldman for their valuable aid in the experimental work and the thermodynamic calculations, respectively. References 1R. J. Lockhart, J. J. Ward, M. J. Hussey, and J. W. Clegg: In unpublished report, "U. S. Air Force Project Rand. Titanium and Titanium Base Alloys." Battelle Memorial Institute (1949) 37-44. 2 L. Brewer et al: Chemistry and Metallurgy of Miscellaneous Materials—Thermodynamics. Ed. by L. L. Quill. pp 13-39, 60-192. (1950). McGraw-Hill Book Co. 3 K. Arii: Scientific Reports. Tohoku Imp. Univ. (1933) 22, 182. 4 R. S. Dean, J. R. Long, F. S. Wartman, and E. L. Anderson: Transactions AIME (1946) 166, 369-381; Metals Technology (Feb. 1946) TP 1961E. * R. I. Jaffee and I. E. Campbell: Transactions AIME (1949) 185, 646-654; Journal of Metals (Sept. 1949) TP 26813.

Jan 1, 1951

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air-pressure drop p in. water gauge across the airshaft. † The airshaft pressure drop is calculated by the formula: R X l X O X Q2 P= 5-2 X A3 wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: w= 0.746Q X P 6345 With a combined over-all unit efficiency of the fan and drive of E per cent, the air-shaft power requirements become W = 100w There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C¢ per kw-hr. the annual airshaft power cost MI dollars per year becomes: M1 = 87.60W X C which, expressed by the component parts, becomes: M1 V 5.2X4' / XQ Z 0.746 X 100 X 87.60 X C V 60 X E ) The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available molletary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air-shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1947

By Raymond Mancha

TO determine the optimum inside dimension of an airshaft, it is necessary to strike the proper balance between the cost of power for air friction and turbulence losses within the airshaft, on the one hand, and the cost of sinking the airshaft, on the other. An airshaft larger than necessary is extravagant because the increased amortization rate exceeds the reduction in power cost. Conversely, an airshaft smaller than is needed wastes more power than can be justified by the corresponding reduction in amortization rate. The airshaft power is a function of the air volume Q cu. ft. per min. and the air-pressure drop p in. water gauge across the airshaft. † The airshaft pressure drop is calculated by the formula: R X l X O X Q2 P= 5-2 X A3 wherein R is the air-friction coefficient for the portion of the airshaft 1 ft. deep, 0 ft. in perimeter and A sq. ft. in area, when passing an air volume of Q cu. ft. per minute. The airpower w kw. dissipated within the airshaft is computed by the formula: w= 0.746Q X P 6345 With a combined over-all unit efficiency of the fan and drive of E per cent, the air-shaft power requirements become W = 100w There are 8760 hours per year, so with energy at I cent per kilowatt-hour the annual cost per kilowatt for continuous operation is $87.60; therefore, with energy at C¢ per kw-hr. the annual airshaft power cost MI dollars per year becomes: M1 = 87.60W X C which, expressed by the component parts, becomes: M1 V 5.2X4' / XQ Z 0.746 X 100 X 87.60 X C V 60 X E ) The amortization rate a dollars per year per dollar invested includes both depreciation and interest on the investment. These in turn involve the expected useful life of the airshaft and available molletary interest rate. For the purpose of this discussion the cost of airshaft sinking is assumed to comprise a fixed move-in charge Y dollars, which covers the cost of transporting, erecting and removing the shaft-sinking equipment and is independent of the air-shaft size. Also, it is assumed that the cost of shaft lining, if any, is fixed at $X per cu. yd. of lining material and that excava-

Jan 1, 1947

By Ernst F. Eurich

(Cleveland Meeting, October, 1912.) ALEXANDER PARKES patented in England in 1851-52-53 a process for desilvering lead by means of zinc, making use of the greater affinity of silver for zinc than for lead, discovered by Karsten in 1842. The efforts in England to develop it commercially were not particularly successful, nor was the attempted introduction into the United States. This last was not surprising, as it was about the beginning of our Civil War, and at that period there was very little lead-bullion produced in the united States and there was no field for the process here. There was at that time on Staten Island an establishment that pattisonized lead, probably from Rossie, N. Y., and Chester, Pa., and possibly some Spanish lead, but the production of lead-bullion in the Far West had not yet begun. In 1864, Edward Balbach, of Newark, N. J., patented in this country a process for desilvering lead by means of zinc, and in 1867 Edward Balbach, Jr., patented a " movable black-lead retort with a neck, placed in a furnace," for the purpose of distilling the alloy of zinc, lead, and the precious metals. A brief comparison of the Parkes and the Balbach processes is interesting. The Parkes process as practiced in England consisted in stirring into silver-bearing lead, melted in a kettle, from 1 to 2 per cent. of zinc, cooling, skimming the alloy of lead, zinc, and silver that rose to the surface, liquating it in an iron retort, distilling the residual rich alloy in fire-clay retorts, thus regaining the zinc, and refining the desilvered lead at a low heat in a shallow cast-iron pan that formed the hearth of a reverberatory furnace. If the lead was too impure, it was first softened * Presented also at the Joint Meeting of the Institute and Section IIIa, Metal¬lurgy and Mining, of the Eighth International Congress of Applied Chemistry, New York, September, 1919, and here published by mutual agreement.

Dec 1, 1912

By D. B. RUSHJIORE

(New York -Meeting, February, 1912.) TORCHES were used by the early Romans for mine-lighting, and these were followed by open lamps or earthen jars filled with tallow or oil, and later by candles. In early coal-mining, explosive gases seldom occurred, and, if they were encountered, the danger of explosion was materially diminished by the absence of any ventilating-system and the consequent insufficient circulation of air to form with the methane an explosive mixture. As the art advanced, the danger of fire-damp became more apparent, and methods for testing the air were resorted to as early as in the fifteenth century. The method of testing consisted in lowering a clog in a, basket down the shaft. As soon as he encountered fire-damp he would commence to howl, after which he was withdrawn, and a leafy bush was fastened to the end of the rope and run rapidly up and clown the shaft. This so disturbed the accumulation of gas as to cause it to rise out of the shaft. Later this odd method of testing was improved upon through the discovery that a light inserted in a body of fire-dam p would produce a blue flame. Candles were used for this purpose, and, when fire-damp was thus discovered, brushing or some other primitive method of fanning was resorted to for driving out the gas. In some instances the gas was burned away, regardless of the danger thereby incurred. A man, wearing a well-moistened coat and a mask with glass spectacles, descended in the mine a few hours before each shift. With a long stick, to which was fastened a burning torch or candle, lie would crawl on the floor with the stick so raised that the light would sweep the roof and ignite the gas which it encountered. In this position the flame of the explosion would pass over him, after which he would retreat from the gases resulting from the combustion. The occurrence of serious accidents resulted in

Jul 1, 1912

By Dennis J. McCarthy

TAX TREATMENT OF EXPLORATION AND DEVELOPMENT EXPENSES Exploration Expenditures General Principles. Exploration expenditures are expenses incurred within the United States or the Outer Continental Shelf for the purpose of ascertaining the existence, location, extent or quality of any deposit of ore or other mineral. Section 617 of the Internal Revenue Code permits a taxpayer to elect to deduct these expenditures currently. In default of election, exploration expenditures must be capitalized and included in the basis of the mineral property for purposes of determining cost depletion. Entitlement to Deduction. In order to qualify for current deduction, exploration expenditures must be incurred prior to the commencement of the development stage, i.e., when mineral deposits are shown to exist in sufficient quantity and quality to reasonably justify commercial exploitation. Generally, expenditures for geological and geophysical investigations, reconnaissance, surveying, testpitting, trenching, drilling, driving of exploration tunnels and adits and similar types of work constitute exploration expenditures if incurred prior to the commencement of development. However, expenditures to further delineate the extent and location of an existing commercially marketable deposit are not exploration expenditures. Exploration expenditures may be incurred from within a producing mine to ascertain the existence of what appears to be a different ore deposit. Expenditures for property which is subject to an allowance for depreciation are not deductible under §617. However, depreciation properly allowable on assets used in exploration is an exploration expenditure. Election to Deduct Exploration Expenditures. The election to deduct exploration expenditures is made by deducting these expenditures in the return for the first taxable year for which such treatment is desired. The election applies to all properties of the taxpayer. Once the election is made all exploration expenditures, in the year of election and thereafter, must be deducted, unless the election is revoked with the prior consent of the Commissioner. The Tax Equity and Fiscal Responsibility Act of 1982 (TEFRA) made substantial changes in the treatment of exploration expenditures. For exploration expenses incurred after 1982 and before 1984 by a corporation, the amount currently deductible is limited to 85% of the expenditure. The remaining 15% must be capitalized and recovered over a five-year period. Although exploration costs capitalized under this provision are not part of the basis of the property for cost depletion purposes, investment credit may be claimed with respect to these expenditures. Further changes have been added by the Tax Reform Act of 1984 with respect to expenditures made after 1984. The Act reduces the amount of otherwise currently deductible costs from 85% to 80%. Additionally, it limits the availability of the investment credit for capitalized mineral exploration costs to deposits located in the United States. TEFRA also allows individuals a special method of recovery under which exploration expenditures may, at the election of the taxpayer, be amortized ratably over a 10-year period beginning with the year in which the costs are paid or incurred. If taxpayer does not use this method, the excess of the exploration expenditures actually deducted over the amount that would have been deducted if 10 year amortization had been elected is an item of tax preference subject to the alternative minimum tax.

Jan 1, 1985

Jan 1, 1882

By William P. Blake

This camp is just now startled by the alarming suggestion that unless there shall speedily be an appreciation in the value of silver, and a decrease in the cost of salt, it will be prudent business policy to close the Alice mines and mills until the conditions of production are more favorable. Such an event would be a great blow to the prosperity of Butte. It would throw some hundreds of men out of employment, and it might lead to the adoption of the same policy by other companies, which, like the Alice, have not been able to pay dividends to their shareholders for a long time. The question arises, and becomes emphatic: If, under the present adverse conditions, the mines and mills cannot be made to pay a proper profit, why should they continue to be worked ? The owners mentally ask: Shall we go on exhausting our mines, pouring out a stream of silver for the maintenance of the community, and the support of the rail ways, or shall we husband our reserves, withdraw the miners from our slopes, and wait for more favorable conditions ? This is the impending question, and all concerned dread to meet and answer it. Let us briefly review the situation and forecast, if possible, what the solution of' the problem will be. It must be understood that the discussion does not include, or refer to, the immense copper-producing interests of the camp, which may or may not be upon a very much more satisfactory basis. It will be confined chiefly to the operations of the Alice company, sta tistics of which are at hand, without involving the other silver com panies in the conclusions beyond those inevitable from well-known facts, patent to all. There are at present in Butte and in its vicinity in the principal mills an aggregate of 290 stamps dropping upon silver-ores: Alice,...........80 stamps. Moulton,. .........40 " Lexington,...50 " Silver Bow,.........30 " Blue Bird,..........90 " The Alice mills at present, with their 80 stamps, devour, in round numbers, from 100 to 120 tons a day. The actual amount

Jan 1, 1888

By G. T. Jackson

The Alaska Gastineau Mining Co.'s mine is located at Perseverance, about 4 mi. east of Juheau, Alaska. Its property consists of a group of claims, the lode system traversing these claims for a distance of 11,000 ft. (3352 m.). It is shown in Fig. 1. The ore deposits occur in what is known as the Juneau gold belt. This gold belt can be traced for about 50 mi. northwest of Juneau and 40 mi. southeast. It consists of a single band, several hundred feet wide, in which stringers and veins of quartz, carrying gold, occur in a slate formation near its contact with some altered volcanic rock known as greenstone. The Perseverance orebody consists of stringer lodes, having a strike northwest and southeast, dipping about 60° to the northeast. These

Jan 1, 1920

By A. E. Dwight, A. F. Berndt

ThE published U-Ru equilibrium diagram indicates the existence of six intermetallic compounds in this system.' The structure of only two of the compounds, u2Ru2,3 and URU3,4 have been identified. This communication reports on the crystal structure and density of Uau3, which is formed by a peritectoid transformation at 1025° ± 5°C: U3Ru4 + U3Ru5 = U2Ru3. An alloy with the stoichiometric formula U2Ru3 was prepared by arc melting electrolytic uranium and ruthenium of 99.9 pct purity on a water-cooled copper hearth under an argon atmosphere. After homo-genization at 1000°C for 3 days, a metallographic examination revealed that the peritectoid transformation was complete and less than 5 pct of a second phase was present. A powdered sample, obtained by crushing the brittle compound, was annealed for 1 hr at 1000°C and air-cooled. Debye-Scherrer patterns were obtained with filtered CrKa, FeKa, CoKa, or CuKa radiation. The X-ray patterns of U2Ru3 weFe indexed as an fcc structure with a. = 12.895 0.00lA. The unit-cell constant and its standard deviation were calculated by a computer program described by Mueller, Heaton, and Miller.5 The density, measured by an immersion method, was 14.45 g cm-3. A calculated density of 14.48 g cm-3 was obtained on the assumption that twenty-four formula units (120 atoms) comprise a unit cell. The space group is either oh7 Fd3m, or the less symmetric, T4h, Fd3, since reflections indexed as hk0 with h + k + 4n are absent. All values of h2 + k2 + 12 permitted by both space groups are listed in Table I together with dc, do, and I. values. The calculated d spacings were obtained by the computer program of Mueller, Meyer, and simonsen.6 The observed d spacings above 0.77A were obtained from X-ray" patterns prepared with CuKa radiation, above 0.90A with CoKa radiaaion, above 0.98A with FeKa radiation, and above 1.16A with CrKa radiation. An attempt was made to identify the crystal structure of U2Ru3 by comparison with published data on other A2B3 compounds. A recent compilation7 of intermetallic compounds lists thirty-two structure types with the A2B3 stoichiometry. Only two of the structure types are fcc and they are Sb2O3(r), senarmontite, with 16 Sb2O3 per unit cell, and In2Te3 for which several structures have been proposed.8 However, the two cases cited do not have twenty-four formula units in the unit cell. We therefore, cannot identify UPRU3 with any of the other known A2B3 compounds. Ternary alloys were made in which 5 at. pct Os, Rh, or A1 were substituted for 5 pct of the Ru in U2Ru3. The osmium substitution increased the unit cell constant to 12.899 * .001A, which is to be expected since the atomic volume of osmium is slightly larger than

Jan 1, 1969

By Jennings S. Cox

(Glen Summit Meeting, June, 1911.) THE following notes, prepared in 1908, as the result of a personal examination and extensive explorations under my direction in 1906, have been revised and greatly augmented after two subsequent visits and further explorations in 191.0. The hard iron-ores of the south coast of Oriente Province, in the island of Cuba, have been known for many years, in fact one mine has been in operation for more than 20 years ; but it is only within the past four years that the large mantle-deposits of brown ores of the north coast have attracted serious attention. A. C. Spencer 1 has described these ores in a paper, entitled Three Deposits of Iron Ore in Cuba, which refers to the Mayari, Moa, and Cubitas deposits. C. M. Weld,2 in a scholarly paper on The Residual Brown Iron-Ores of Cuba, has discussed the character and possible genesis of these deposits with particular reference to the Moa field, calling attention also to the Taco and Navas fields in addition to those described by Spencer. Descriptive articles with reference to the Mayari field, and a few words about the Moa deposit, have also appeared.3 It is the purpose of the present paper to give some account of the exploration of the Moa deposit and to call attention to its commercial importance. LOCATION. The Moa district is situated on the north coast of the . Province of Oriente, which is the most easterly province of the island of Cuba. It lies on the northern or seaward slope of 1 Bulletin No. 340, U. S. Geological Survey, pp. 318 to 329 (1908). 2 Trans., xl., 299 to 312 (1910). 3 Iron Age, vol. lxxx., No. 7, pp. 421 to 426 (Aug. 15, 1907), and vol. lxxxi., No. 15, pp. 1149 to 1157 (April 9, 1908).

Mar 1, 1911

By Harrison Schmitt

The definitions of the terms commonly used in connection with metamorphism at elevated temperature near igneous contacts are re-examined. As the study of tactite zones associated with igneous contacts progresses, more and more emphasis is put on the localization of the ore by structural forms. The generally early age of the silicates and late age of the sulphides is well established. Postsilicate structural deformation is likely to localize ore bodies. The silicates are often replaced by the ore minerals. we may not be far from proving that the elements involved in the formation of silicate aureoles are merely normal "volcanic" emanations, that silica, lime, magnesia, and alumina, and probably soda and potash, are likely to undergo local transfers only. The work of Newhouse suggests the agents of the formation of sulphides in deposits of various classifications are the volcanic emanations. Introduction The purpose of this paper is to define the terms commonly used in connection with the subject, to briefly review and discuss the broad features of the aureoles, to indicate some of the features lately recognized: to discuss the genesis, and to suggest how the associated ore deposits can best be studied and developed. When developing a subject in mining geology one is soon conscious of the vast amount of pertinent literature and that most relevant ideas, no matter how obscure, have been considered. One can only hope to make some small contribution in point of view, a few confirmatory observations, or, perhaps, suggest a small new idea indebted to the most recent data. It is nearly impossible, also, to give specific credit, and adequate credit, to all the contributors to a given subject. This study, furthermore, was not intended to cover the field exhaustively, nor is the bibliography complete. Definition OF Terms There is much confusion in the use of the broad terms concerned. Needless to say, metamorphism means change of form; metasomatism, change of substance. The term contact metamorphism, however, has been widely misused possibly by a majority of economic geologists and engineers. As originally adopted by European investigators it meant changes effected largely by heat (thermal effects) in the rocks invaded by plutonic intrusives, but did not include those local changes usually at the immediate contact brought about by the addition of substance, i.e., by metasomatism. The term has become clouded. In Europe it has

Jan 1, 1949

By Harrison Schmitt

The definitions of the terms commonly used in connection with metamorphism at elevated temperature near igneous contacts are re-examined. As the study of tactite zones associated with igneous contacts progresses, more and more emphasis is put on the localization of the ore by structural forms. The generally early age of the silicates and late age of the sulphides is well established. Postsilicate structural deformation is likely to localize ore bodies. The silicates are often replaced by the ore minerals. we may not be far from proving that the elements involved in the formation of silicate aureoles are merely normal "volcanic" emanations, that silica, lime, magnesia, and alumina, and probably soda and potash, are likely to undergo local transfers only. The work of Newhouse suggests the agents of the formation of sulphides in deposits of various classifications are the volcanic emanations. Introduction The purpose of this paper is to define the terms commonly used in connection with the subject, to briefly review and discuss the broad features of the aureoles, to indicate some of the features lately recognized: to discuss the genesis, and to suggest how the associated ore deposits can best be studied and developed. When developing a subject in mining geology one is soon conscious of the vast amount of pertinent literature and that most relevant ideas, no matter how obscure, have been considered. One can only hope to make some small contribution in point of view, a few confirmatory observations, or, perhaps, suggest a small new idea indebted to the most recent data. It is nearly impossible, also, to give specific credit, and adequate credit, to all the contributors to a given subject. This study, furthermore, was not intended to cover the field exhaustively, nor is the bibliography complete. Definition OF Terms There is much confusion in the use of the broad terms concerned. Needless to say, metamorphism means change of form; metasomatism, change of substance. The term contact metamorphism, however, has been widely misused possibly by a majority of economic geologists and engineers. As originally adopted by European investigators it meant changes effected largely by heat (thermal effects) in the rocks invaded by plutonic intrusives, but did not include those local changes usually at the immediate contact brought about by the addition of substance, i.e., by metasomatism. The term has become clouded. In Europe it has

Jan 1, 1949

By William L. Batt

FROM the time the Japs overran the Far East, the United Nations faced a serious military problem in the critical shortage of many raw materials desperately needed to prose¬cute the war on two fronts. For many months this critical shortage threatened substantially to limit the gigantic effort to get shooting equipment into the hands of the fighting men on the far-flung battle lines. Now this crisis has been met. Solvency in raw materials has been achieved. Simply stated, this job of producing solvency has been done by "cutting cloth to fit the garment." This does not mean that we have all the raw materials we need. Unfortunately, a few raw materials are still short, thereby limiting the total production of certain war goods, or making these goods less effective. Where the supply of raw materials is known to be limited, however, substitutions have been made or production schedules have been geared to the amounts available, and an over-all balance has been achieved between various parts of the war-production program.

Jan 1, 1943

By H. Evans, L. A. Hunt

In the pellet-producing industry it is sometimes claimed that "The pellets are made in the concentrator," or that "The key to good pellets is in the balling operation." However, one claim is generally agreed on: If you can put in, on, and through the indurating furnace a good green ball of Proper size and adequate strength to withstand the handling, load, dehydration, and thermal shock until it is properly heat-hardened, you will have the pellet product you desire-that is, a "competent" green ball. The strength requirements that enable a competent green ball to withstand the environment of the indurating furnace are usually less restrictive than those for transporting the green ball from its source of production to the furnace. Experience has shown that different types of material may have different balling characteristics which may impose limitations on the quality of the green ball that can easily be produced under normal conditions. Therefore, if a specific material will produce a green ball that meets the physical requirements for competency in the furnace, but not for handling or transportation, various practices must be instituted to overcome this deficiency. In other words, if a normal handling system is too rigorous for the green balls that can be produced under standard conditions, then "handling with care" should be tried.

Jan 5, 1967

By J. J. Rutledge

I. DESCRIPTION OF THE CLINTON ORES AND ASSOCIATED POCKS. THE Clinton rocks in Stone Valley comprise (1) thick layers of deep-red shale, (2) layers of reddish-gray shale interspersed with beds of sandstone and thin beds of extremely fossiliferous limestone, and (3) yellowish-gray shales alternating with thin layers of olive-colored shales. These layers of shale are all very thin and are at no point massive.1 A bed of gray-white sandstone varying in thickness from 12 to 30 ft. occurs in the tipper portions of the Clinton shales. This is called the ore sandstone, a name given to it by Rogers and adopted also by the Second Pennsylvania Geological Survey. It is the surest guide to the presence of the iron-ore bed, which lies below its lower edges at distances varying from 10 to 20 ft. This ore sandstone, in the valley, lies generally at very low inclinations, but often is locally subjected to considerable changes in dip. At its outcrop this ore sandstone weathers very easily and yields a rather sparse crop of stone, Composed of dirty, iron-stained boulders. Upon the roads these boulders are quite well rounded, and bear fragments of crinoid stems and other fossil forms. The ore sandstone exhibits at many points evidences of former movements in this very great thickness of the shales, and as it is the only stratum in this series competent to transmit and record stresses due to structural movements, its evidence is of considerable importance in determining the occurrence of movements in times past.

Nov 1, 1908

(Members are urged 'to, send in for this column any notes of interest concerning themselves or, their fellow-members.) Members and guests who registered at Institute headquarters during the period Sept. 10 to Oct. 10, 1915: George H. Blakemore, Sydney, N. S. W. W. B. A. Dingwall, San Antonio, Tex.. H. Haege, Sydney, N. S. W. R. H. Jeffrey, London, England. C. A. Filteau, New York, N. Y. H. S. Emlaw, Spring Lake, Mich. Myron R. Walker, Cerro de Pasco, Peru. Y. T. Woo,' Tientsin, China. William R. Grunow, Morenci, Ariz. A. L. DeDeken, LouVain, Belgium. E. DeGolyer, Norman, Okla. K. P. Swensen, Tokyo, Japan. William W. Elmer, Tuttletown, Cal. Charles Enzian, Wilkes-Barre, Pa. Huntington Adams, Iquique, Chile. C. F. Schnepp, Cobalt, Ont., Canada. W. Ray Cox, for several years Mineral Inspector of the General Land Office at Portland, Ore., has been appointed Professor of Mining and Metallurgy in the School of Mines, Oregon Agricultural College, Corvallis, Ore.

Jan 11, 1915

CARL E. SWARTZ, Chairman D. K. CRAMPTON, Past-Chairman CYRIL S. SMITH, Vice-Chairman ARTHUR PHILLIPS, Vice-Chairman FRANK T. SISCO, Secretary 29 West 39th Street, New York, N. Y. H. A. MALONEY, Treasurer Executive Committee 1943 1944 ERIC R. JETTE WILLIAM C. ELLIS LOUIS W. KEMPF ALAN MORRIS GEORGE SACHS KENT R. VAN HORN

Jan 1, 1942

NEW MEMBER, The following list comprises the names of those persons who became members during the period Mar. 10 to Apr. 10, 1915: Members ADAMS, Louis WINFIELD, Supt., Rolling Mills, Lehigh Plant, Bethlehem Steel Co., So. Bethlehem, Pa.. ALLEN, ARTHUR WATTS, Mine Mgr Rodeo, San Juan, Argentina., BALL, EDWIN MARCOTTEL, Mine Supt., Ishkooda. Mines, Care Tenn. Coal, Iron & R. R. Co., H.F.D. 1, Birmingham, Ala. BOLN, CHARLES A., Oil salesman 135 S. 2d St., Philadelphia., Pa. BOYD, DAVID F., Naval Officer Navy Department, Washington, D. C. CAPERTON, GEORGE HENRY Charleston, W. Va. COGGESHALL, GEORGE W., Chem. Engr., Institute of Industrial Research, Washington D C

Jan 5, 1915