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By Louis D. Huntoon

(New Haven Meeting, February, 1909.) THE Hammond Mining and Metallurgical Laboratory is the gift of Prof. John Hays Hammond to the Sheffield Scientific School of Yale University. Professor Hammond was graduated from this school in the class of 1876, and has always shown great interest in its welfare and progress. In 1903, he offered to build and equip for it a mining and metallurgical laboratory. This offer was gratefully accepted by the trustees; and in October of the same year the architect, W. Gedney Beatty, and the writer were employed to draw up plans for the building, which is the largest of its kind in this country. The Treasurer's report shows the total gift of Professor Hammond, including land, building, and equipment, to have been $128,741.53. This does not include $5,000 given by Professor Hammond to defray incidental expenses. The detailed cost of the machinery and installation, included in the above sum, is as follows: Machinery, $9,846.00 Pipe and fittings, 925.05 Hangers, shafting, pulleys, belts, etc., . 357.83 Steel-work, tanks, launders, etc., . 272.19 Screens for sizing ore, 411.37 Lumber, 1,896.36 General Supplies Hardware, . $135.96 Tools, 263.66 Paints, . 04.63 Electric and incidental, . 439.57 - 1,043.82 Machine-shop material, 259.44 Labor, 722.29 - 981.73

Mar 1, 1909

By O. J. C. Runnals

The intermetallic compounds NpAl2, NpAl3, and NpAl have been prepared, and examined by X-ray diffraction methods. The compounds are isostructural with the corresponding U-Al compounds. NpAl3 is face-centered cubic with a = 7.785A and has the MgCu2 structure. NpAl is simple cubic with a = 4.2624 and has the AuCU3 structure. NpAl4 is body-centered orthorhombic with a = 4.428, b = 6.268, and c = 13.71A. The published description for the atomic positions in UAl, applies also to NpAl4 THE U-Al system has been examined by Gordon and Kaufmann.1 They identified the following intermetallic compounds: UAI2, UAl3, and a third, tentatively identified as UAl5. A structure analysis of the latter compound by Borie2 proved that the formula UAl4 was more consistent with the X-ray data. The crystal structures of UAl2 and UA have been reported by Rundle and Wilson.3 UAl2 was found to be isostructural with MgCu. The atomic positions in the simple cubic UA1, cell were found similar to those in the AuCu3 structure. Three analogous compounds exist in the binary system Np-Al. X-ray data for the compounds NpAl2, NpAl3, and NpAl4 are reported below. The neptunium used in the investigation was supplied through the courtesy of the Argonne National Laboratory as neptunium (V) nitrate. It was stated to have a radioactive content of 99.25 pct Np237. The impurity limits had been established by a spectrographic analysis given in Table I. Any titanium, boron, or bismuth impurity should be effectively removed during the fluorination of NpO2 to NpF4, and sodium, lithium, magnesium or strontium during the alloying procedure. The oxalate ignition technique, described by Westrum for the preparation of finely-divided reactive PuO2, was used.The neptunium (V) nitrate solution, containing 100 mg Np, was evaporated to dryness on oxalic acid crystals. The resulting oxalate was calcined to NpO, by heating in platinum at 650°C for 2 hr. An equimolar mixture of anhydrous hydrogen fluoride and purified hydrogen was passed over the NpO2 for 2 hr at 500°C. According to Fried and Davidson" the above fluorination conditions should produce the purple-black fluoride, NpF3 However, the light-green product showed only NpF4 lines on an X-ray diffraction pattern and contained only an estimated 5 pct of finely dispersed black NpF3 particles on microscopic examination. NpAl3: An intimate mixture of 50 mg of neptunium fluoride and aluminum powder (99.99 pct pure, —28 to $60 mesh) was heated to 900°C in a BeO crucible in vacuum by a tungsten coil for 2 hr, to allow the following reaction to proceed: NPF4 + 7Al + NpAl, + 4AlF. [1] The formation and sublimation of aluminum mono-fluoride in this system was verified from an X-ray diffraction analysis of the sublimate. The diffraction pattern was similar to one obtained in the aluminum reduction of UF4 to UF3,6 showing both Al and AlF3 lines. The resulting alloy was sintered at 1100°C for 1 hr. A powdered specimen of the crystalline product was sealed in a thin-walled quartz capillary for X-ray examination. The photograph was taken on a 14.32 cm General Electric powder diffraction camera, using filtered copper radiation. The films were calibrated in terms of a NaCl standard pattern. The wavelengths used in the calculations were 1.5418A for Cu K and 1.5405A for Cu K The powder diffraction pattern showed that NpAl is simple cubic with a = 4.262 % 0.005A, and is isostructural with UA13 where a = 4.27A1 or a =

Jan 1, 1954

By Messervey, J. P.

NOVA SCOTIA is sharing in the rapid advance of the mining industry that is one of the remark- able features of Canada's recent progress. The production of coal and gypsum has increased rapidly, and there has also been a marked development in the mining of gold, salt, zinc, copper and other minerals. Coal, the basis of the principal mining industry in the Province, was discovered by the early explorers on the shore of Cape Breton Island, and used by the garrison at Halifax as early as 1749. The first government record shows total annual sales amounting to 1668 long tons in the year 1785, while last year production was 6,643,000 long tons. The principal mines are discussed in a separate paper by Alexander L. Hay1.

Jan 1, 1928

I. C. WHITE, Morgantown, W. Va. (written discussion *).-The writer's conclusions with reference to the possibility of finding commercial deposits of petroleum in Brazil were published in 1908 in the Final Report of the Coal Commission of Brazil.1 Since this volume is not generally accessible, and the author has nothing to acid thereto, the following quotation found on pages 243-7 may prove of interest: The Possibility of Petroleum Deposits in Brazil "The writer has often been asked to express an opinion as to the possibility of finding petroleum in commercial quantity in Brazil. "The question is one that does not admit of an adequate answer except for the regions visited by the writer, which as already stated, include only, a portion of that great territory, and only the southern portion of Brazil; namely, the states of Rio Grande do Sul, Santa Catharina, Paraná, and a few rapid excursions across São Paulo, and Minas Geraes, as well as the brief visit to Marahú in Bahia. "That the sedimentary beds of the Brazilian Permian once held some petroleum is attested by the evidence of the Iraty black shale which gives off the characteristic odor of this substance from São Paulo to Rio Grande do Sul.

Jan 1, 1918

By AIME AIME

THE report of the united. States Steel Corporation for the' year 1928 gives the, following ,figures of raw materials and' finished products . Raw materials, tons :

Jan 1, 1929

By Roy E. Dickerson

CUBA differs considerably from the other Greater Antilles in many geologic fundamentals. Cuba is geosynclinals; whereas Jamaica, Hispaniola. (Haiti), and Puerto Rico are geoanticlinal. (Scliuchert, Chas., Historical Geology of the Antillian. Caribbean Region, p. 13, 1935.) Rocks of lowermost Cretaceous age exist in large volume in Cuba, where- as they are generally absent in the other islands. Since many of the surface indications of petroleum are intimately associated with lowermost Cretaceous rocks, the Vinales limestone,, this difference is most important. The Vinales limestone in many places in Pillar del Rio and Santa Clara provinces reeks with light tars when broken with a hammer. The late David White and his associates on the U. S: Geological Survey studied

Jan 1, 1937

By Wayne A. Greenwalt

A method that determines venture participation when the decision maker's level of risk aversion varies according to venture profitability and risk investment is presented. The decision maker's fundamental aversion to risk is established by the selection of a desired working interest for a venture that anticipates only a discounted payout.

Jan 1, 1982

By Thomas F. McLoughlin, David K. Hylbert

A new method of satellite imagery analysis has been developed during underground mine roof research. Landsat positive film transparencies were projected onto topographic base maps to scale and lineaments enhanced with a Ronchi grating (method A). Thirty-five millimeter slides of video-enhanced scenes were also projected onto base maps (method B). Both methods permit direct plotting of lineaments to scale and direct correlation with surface and subsurface data wing transparent overlays.

Jan 1, 1981

By C. S. Simons

Manufacturers of filtering, thickening or clarifying equipment, have been active in developing machines to meet increasingly rigid operator requirements, and the innovations resulting from this effort are moving into commercial applications. In general, the results of these applications are reported to be good. Little has been reported on progress in theory. The growing insistence on the part of pollution control agencies that waste water discharged into streams contain less and less suspended solids is being felt by a broad cross section of industry. A new clarifier, based on the often rediscovered fact that suspended solids settle more rapidly in inclined channels, is being marketed by Neptune MicroFLOC of Corvallis, Ore. This unit employs modules of inclined tubes, mounted in the upper zone of the clarifier, which carry the water flow. These were developed for light solids, such as those encountered in domestic water treatment and in paper mill waste water. In these applications, flow rates up to about three times those permissible in conventional clarifiers are reported to give comparable overflow turbidity. By combining these units with mixed media filters, high clarity may be obtained with minimum capital. Use of such units in the metallurgical industry, particularly where water reuse or effluent quality is important, should grow.

Jan 1, 1970

By L. Hartsell Cash

INTRODUCTION Created in 1944 to help rebuild those economies, principally in Europe, which were seriously damaged or destroyed during the Second World War, the World Bank--or to use its correct name, the International Bank for Reconstruction and Development-began its operational phase by making a loan to the Government of France to accomplish two tasks: (i) to import coal for space heating purposes; and (ii) to buy new machinery and equipment to help rehabilitate its coal mines. Once the thrust of reconstruction in postwar Europe was largely accomplished, the World Bank Group (for today the World Bank has three separate lending windows) turned to mobilizing funds to improve living standards in developing countries, which remains its primary goal today. Over this period of time, the World Bank Group has continued to assist the world mining industry in various ways and, since 1955 the World Bank Group has financed mining or mineral processing projects with loans/credits totalling $1.5 billion. The World Bank Group consists of two institutions--the World Bank itself and the International Finance Corporation (IFC), which was founded as a separate institution in 1956 and three separate lending windows--the World Bank, the International Development Association (IDA), and IFC. The World Bank loans are provided at commercially-related interest rates but have grace periods usually of up to 3 to 5 years and repayment periods ranging from around 15 to 20 years. Since the World Bank raises its operating funds on the international capital markets in varying currencies and in varying interest rates, it relends a basket of currencies at floating interest rates which are fixed every six months to reflect the Bank's own borrowing cost. At the present time, the interest rate is just below 10% per year and in addition, a one-time front-end fee of 1/4 of 1% is charged. Also, a commitment fee of 3/4% per year is charged on the undisbursed amount of the loan itself. For loans to industrial and mining projects, it is normal to charge an additional guarantee fee of 1 to 1.5% in addition to the annual interest rate. This guarantee fee is payable to the government of the country in which the project is carried out since, by its charter, the World Bank is required to have the guarantee of the host government for each loan made in a specific country. By far the greatest source of lending funds within the World Bank Group is that of the World Bank itself and for the fiscal year ending June 30, 1984, the World Bank approved loans totalling $12 billion. The second lending window within the World Bank Group is IDA which performs the same functions as the Bank itself and has the same staff. Nevertheless, its loans (called credits) are made on much softer terms and are made to the poorest nations only. IDA funds are subscribed by the wealthier countries, usually as grants, and these are relent as credits to those countries with the lowest per capita income. IDA credits are made for 50 year periods, including 10 years of grace, to the government itself. The credits are interest free, but there is a service charge of 3/4% per year on the amount withdrawn and outstanding to defray the administrative costs of IDA. Credits approved during fiscal 1984 totalled $3.5 billion. The third lending window is IFC. While IFC is a separate corporation and operates quite separately from the World Bank and has its own staff and regulations, it works very closely with the World Bank on its overall approach to the mining sector and to a particular country. IFC's mission is essentially to supply venture capital and to stimulate development of the local private sector in capital markets as well as to promote the international flow of private capital to developing countries. Unlike the World Bank and IDA, the IFC does not lend directly to governments

Jan 1, 1985

By T. P. Harrington, P. G. Doctor, K. A. Prisbrey

Crushing and grinding ore consumes so much energy that it represents a major component of total processing costs. While it has been estimated that only about 0.6% of the total energy expended in grinding is actually used to create new surfaces, accurate measurement of this quantity is difficult. If energy expended in crushing and grinding could be more directly measured, alternatives for increasing efficiency of the processes could be evaluated more effectively. Since the energy intensity associated with rock fracture is related to fracture sonority, it should be possible to use this sound (acoustic emission) to measure the actual energy expended. However, it is difficult to apply this approach because the mill environment is so noisy. To make the approach practical, a method for differentiating acoustic emissions caused by rock fracture from those caused by extraneous noises must be found. Pattern recognition analysis may provide such a method. This paper presents the results of pattern recognition analysis to make two types of differentiations: acoustic emissions caused by crushing two types of minerals (quartz and chrysocolla) and those caused by two extraneous noise sources, and acoustic emissions produced by each mineral.

Jan 1, 1982

ALABAMA Aldrich.-Lloyd, T. W. Ashland.-Downs, F. G. Geary, E. S. Auburn.-Brown, R. L. Bessemer.-Abbott, C. E. Ball, T. L. Ferguson, V. Salmon, H. S. Stuart. Q. W. Birmingham.-Adams, J. H. Aldrich, T. H. Aldrich, T. H., Jr. Allen, A. W. Armstrong, W. D. Ball, E. M. Bonnyman, J. Bush, M. W. Crawford, G. G. Davis, C. B. Ellis. E. E.

Jan 1, 1923

By Joseph F. Shirley

A froth flotation plant to recover byproduct molybdenite must be designed specifically for the selected separation process and the type of ore being treated. However, some general rules concerning byproduct molybdenite plant design can be suggested that would be valid for most processes. The term byproduct molybdenite in this paper is limited to the recovery of molybdenite from copper porphyry ores, which essentially amounts to selectively separating molybdenite from copper and iron sulfide concentrates. In addition, most of the discussions would apply only to the almost universally used approach of floating the molybdenite and depressing the copper and iron sulfide minerals. Thickener It is considered essential that the feed to a byproduct molybdenite plant be thickened prior to being treated for molybdenite recovery for the following reasons: 1) Thickening allows the rougher and any preconditioning stages to be operated consistently at the optimum percent solids. 2) Generally, the size of equipment used for pretreatment and the rougher flotation is considerably smaller and less expensive than if the copper flotation concentrates were treated directly. 3) The removal of as much water as practical by thickening removes flotation reagents added in the copper flotation circuit, especially frother which can be quite troublesome in a byproduct molybdenite plant. Although thickening of the molybdenite plant feed has long been recognized as necessary to optimum metallurigcal results, very often this area has been designed poorly. The thickener should be sized so that it can produce a consistent thickened product containing 60 to 65% solids without the use of flocculants. Some plants are only able to thicken to 45 to 50% solids consistently while other plants let the percent solids vary from 45 to 60% over a 24-hr period. While the former is more desirable than the latter, in many cases better metallurgical results could be obtained if the percent solids could be held consistently close to 60%. Probably the simplest and most satisfactory system of handling the thickener underflow is to spigot it into a sump, using a sphincter type valve for the orifice, controlled by a density meter. Another satisfactory approach is to use an ODS type pump to meter the underflow. There are other methods of handling thickener underflow which work satisfactorily as long as the amount of copper concentrates entering the thickener is relatively uniform. However, they often fail to operate satisfactorily when there are large fluctuations in the thickener feed rate due to fluctuations in ore tonnage and/or copper grade. If the percent solids in the plant feed could be held close to 65%. There is the possibility the molybdenite plant tailing would not have to be thickened prior to the copper filter plant. This means that the molybdenite rougher float would have to be conducted at 45 to 50% solids with all the cleaner tailings returned to the rougher for dilution. Admittedly, in some situations this approach would be impossible, but if capital or plant area is limited, it should be considered.

Jan 1, 1981

By K. R. Price, M. Gauna

A high seam geologic condition exists at Jim Walter Resources No. 4 mine that requires immediate rib support after a continuous miner cuts a place. High seam mining heights are typically 2.4 to 2.6 m (8 to 9 ft) and a thick rock layer is present in the upper portion of the coal rib side. The 610-m (2000 ft) mining depth at No. 4 mine induces horizontal rib line movement after excavating an entry. Horizontal rib movement greater than 305 mm (12 in.) have been noted. When lateral movement occurs in the high seam areas, the thick rock layer, and the coal above the rock layer, become fractured and tend to roll into the entry. If not adequately supported, large rib rolls develop in the place being bolted as well as in previously mined areas that endanger men and equipment. To control this, rib support must be installed in the roof bolting operation as the bolting crew works toward the face. A rib control system has been developed for this high seam mining condition using a mechanical anchor bolt placed at an angle in the rib and installed with a modified Fletcher DDM-15 bolting machine. More than 120 km (75 miles) of entry have been supported with angle rib bolts. These angle bolts constitute 43% to 60% of the total roof and rib bolt footage in the high seam portions of No. 4 mine. The angle rib bolting system has proven safer, more effective, and more efficient than the earlier support method using bolts installed horizontally into the rib. No rib roll accidents have occurred since the angle drill bolting machines were placed in service in 1981. By securing the broken rib material with an angle bolt anchored above the coal pillar, a more effective means of rib control is achieved. Angle rib bolting does not interfere with continuous miner productivity. In addition, bolting crews are now reduced from three to two people. Total drilling supply costs are about 60% the previous cost, when compared to the previous support method using horizontal rib bolts:

Jan 1, 1985

US 4,132,635 -In the froth flotation beneficiation of low- grade siliceous iron oxide ore, an aqueous pulp of undeslimed ore is conditioned first with a reagent formed by incorporating a water-soluble polyvalent metal salt into an aqueous solution of sodium silicate and then with oleic acid or other anionic collector and optionally pine 011 or other frother. The conditioned pulp is subjected to flotation to concentrate Iron oxide values in the froth D. C. Yang, assigned to Michigan Technological University Jan 2, 1979 7 pp (209-1 66) Filed 1-1 3-77 (759,094) Drawing below.) [ ] US 4,133,038-In the froth flotation beneficiation of copper sulfide ore or other ores, an aqueous suspension of a mineral sample of uniform grain size and composition is conditioned with a strongly adsorbable collector in an amount which is a fraction of the amount sufficient to create a monomolecular layer on the mineral grains, and the conditioning continued until the entire amount of collector is adsorbed The conditioned pulp is subjected to flotation, the floated particles separated at timed intervals, and the flotation rate coefficient determined. By varying the degree of adsorption in repeated treatments, the dependence of the flotation rate coefficient on the degree of adsorption may be determined A Nieml Jan 2, 1979 10 pp (364-502) Filed 12-7-77 (858,353) Priority Finland, 5-26-75 (75/1,525) Same: British 1,533,723, dated Nov 29, 1978 Same Canadian 1,065,066, dated Oct 23, 1979 US 4,133,746System and method for maintaining substantially steady state conditions in the froth flotation beneficiation of copper ore, molybdenum ore, or other metal ores The mass flow of flotation concentrate from the cell is monitored continuously, and the rate of admission of aeration air into the pulp in the cell is varied inversely with respect to variations in the mass flow control of rougher flotation cells is accomplished with a single valve G W Dopson, assigned to Magma Copper Co Jan 9, 1979 6 pp (209-1) Filed 11- 23-76 (744,306

Jan 1, 1980

By F. M. Morris

A large reserve of thick coal in southwest Virginia was developed by Clinch-field Coal Co. in 1957-1958 to produce a nominal rate of 1500 tph raw coal. Operation features coal cleaning in transit. Refuse removed averages 35 pct. Evolution of plant from initial conception to completion is discussed, selection of means applied is explained, and performance to date us design expectation is described. The long-range plan calls for ultimate handling capacity of 40,000 tpd raw coal with anticipated clean coal capacity of 26,000 tpd on a three-shift basis. From a materials handling viewpoint, the Moss No. 3 operation is principally of interest as an ensemble. Generally speaking, it uses time-tested equipment and ideas but some of these are employed on a scale that may be new in the industry. At present about 20,000 tpd of raw coal are being handled. This is expected to increase to 40,000 tpd as soon as business conditions justify it. The purpose of this paper is to discuss the evolution of the plant as to materials handling practices, to describe briefly the equipment and methods used, and to comment on performance in certain areas. The subject divides itself naturally into four phases: 1) operations at the mines, 2) transportation from mines to plant, 3) raw coal handling into the plant, and 4) transportation of refuse away from the plant. OPERATIONS AT THE MINES The coal reserves for Moss 3 are in southwest Virginia where Dickenson, Russell, and Buchanan Counties come together. This area contains about 15 square miles and over 100,000,000 tons of coal. Here the No. 4 (Tiller) and No. 5 (Jawbone) seams of the Norton Formation lie so closely together that for practical purposes, they constitute one seam of coal. This seam, which is called the Thick Tiller, varies from 10 to 18 ft in thickness and underlies Sandy Ridge, a mountain cresting between 2400 and 3300 ft above sea-level. At 18-ft seam height (which is considered to be the maximum practical mining height) the parting between seams will be about 3 1/2 ft thick and each bench of coal will be about 7 ft thick. Depending on the amount of impurities in the seams Drover. total reject in this height coal may approximate 50 pet by wt. It will average about 35 pet for the property as a whole. The first move in developing this resource was made in 1953 when contour maps of several square miles around Duty, Va., including all the known outcrop, were made by photogrammetry. At this time, it was felt the prospective operation would be served by the Clinchfield Railroad and a photogrammetric route survey was made by this railroad from Haysi to Duty. Study of the resulting maps indicated only one site—adversely owned—which might accommodate the size washing plant to be erected. Water resources of a dependable nature seemed nonexistent. In 1954 bulk washability tests were made on the Tiller Bench at the Moss No. 1 preparation plant. The tests indicated this portion of the seam, mined separately, would wash to 4 pet ash with good recovery. Also in 1954, development of Moss No. 2, south of Sandy Ridge, was begun. This mine is in the Tiller Seam where it is about 100 ft below the Jawbone Seam. The reasons for developing a mine in the normal Tiller Seam before tackling the Thick Tiller seemed compelling: Railroad service could be established quickly, communications were better (though not good!), more was known of the seam (it had been mined in the years 1911 to 19241, and there was no essential property to be acquired. After some legal skirmishing, the Norfolk & Western Railroad was granted the right to serve the new mine. Three decades earlier the old mine had been served by the Clinchfield Railroad. The event which triggered active development of Moss 3 was the Appalachian Power Co.'s decision in 1956 to build a 450,000-kva power plant on Clinch River at Carbo. This solved the problem of marketing the steam coal which inevitably must be a product of a mine in the Thick Tiller. Management promptly decided to build the preparation plant at Carbo where an excellent site was owned; where railroad service existed; where telephone service could be obtained; and where roads, bridges, and water supply were tolerable.

Jan 1, 1961

By C. J. Osborn

In oral discussion at the Columbus midyear meeting, September 26, 1949, these pertinent questions were asked: Was a satisfactory separation of metal and matte obtained? The matte was quite fluid and in most cases the separation was satisfactory. Were the mattes chlorinated after aging? The mattes were stored in sealed glass mason jars, and chlorina-tion experiments made after standing for several weeks gave satisfactory results. Was ferric chloride volatilized? Most of the iron content of the matte was converted to ferric chloride. However, because of the low chlorination temperatures used (150" to 250°C) the major part of the ferric chloride remained with the residue. Undoubtedly, some ferrous chloride was also formed, because of the reducing conditions within the reaction vessel, but no effort was made to separate it from the ferric chloride. Could the reaction be stopped before making metallic iron? No attempt was ever made to stop the reaction before making metallic iron. Generally, 20 to 60 pct of the iron in the charge was obtained as metallic iron which could be readily separated from the matte. In certain cases, metallic iron was dispersed through the matte in the form of globules, which could be removed by magnetic separation. Could other sulphides, i.e., calcium sulphide or sodium sulphide be used in place of iron sulphide? No other sulphides were investigated. It was felt that titanates of calcium and sodium would be formed should these sulphides be used to replace the pyrite. What were the sulphur volatilization losses during matte smelting? In the course of this laboratory investigation, a rather wide variety of variables was introduced from highly oxidizing to reducing conditions and various ratios of pyrite, rutile, and coke. These and other factors all have a definite bearing on the sulphur losses. In general, the higher the sulphur in the charge the higher were the sulphur losses and the percentage of titanium in the matte decreased. With sulphur additions more nearly the theoretical amount, the sulphur losses averaged approximately 10 to 15 pct. With a higher percentage of sulphur on the charges, the losses were in the order of 15 to 30 pct.

Jan 1, 1951

By David White

The submission of carefully prepared estimates of the oil reserves of the United States calls for no apology or explanation. In this country, petroleum is a rapidly wasting asset and an occasional appraisal of the amount remaining in the ground is a simple business procedure to safeguard the general welfare and the prosperity of the republic. The rumored danger of failure, in the near future, of the foreign supply on which we depend in constantly increasing amounts makes the immcdiatc review of our available reserves the more imperative. The Director of the U. S. Geological Survey, in March, 1921, requested the American Association of Petroleum Geologists to cooperate in the preparation of new estimates of the oil remaining in the graund. With the members of this cooperative committee, state geologists, regional geologists, oonsulting specialists, and geologists and engineers attached to many companies worked both independently and through subcommittees. The reports for districts, counties, fields, etc., after review by local or state committees, were discussed by the joint cooperative committee in conferences that, in some instances, were attended by the original compilers of estimates for the states or smaller areas. Finally the data were reviewed and the figures revised by the joint cooperative committee, so that the totals given in Table 1 represent the opinion of the committee, which was composed of F. W. DeWolf, W. E. Wrather, Roswell H. Johnsont Wallace A. Pratt, Alexander W. McCoy, Carl H. Beal, C. T. Lupton, G. C. Matson, K. C. Heald, W. T. Thom, Jr., A. E. Fath, Kirtley F. Mather, R. C. Moore, and David White, chairman. Oil geologists and engineers understand well the speculative nature and the chances of error that must attend estimates of the petroleum in the ground when they cover not merely prospective territory, but possible territory. Neither the character nor the magnitude of the factors of uncertainty is to be disguised. Further, the inevitable differences of opinion between geologists of equal rank and experience must also bc

Jan 1, 1923

By David White

The submission of carefully prepared estimates of the oil reserves of the United States calls for no apology or explanation. In this country, petroleum is a rapidly wasting asset and an occasional appraisal of the amount remaining in the ground is a simple business procedure to safeguard the general welfare and the prosperity of the republic. The rumored danger of failure, in the near future, of the foreign supply on which we depend in constantly increasing amounts makes the immcdiatc review of our available reserves the more imperative. The Director of the U. S. Geological Survey, in March, 1921, requested the American Association of Petroleum Geologists to cooperate in the preparation of new estimates of the oil remaining in the graund. With the members of this cooperative committee, state geologists, regional geologists, oonsulting specialists, and geologists and engineers attached to many companies worked both independently and through subcommittees. The reports for districts, counties, fields, etc., after review by local or state committees, were discussed by the joint cooperative committee in conferences that, in some instances, were attended by the original compilers of estimates for the states or smaller areas. Finally the data were reviewed and the figures revised by the joint cooperative committee, so that the totals given in Table 1 represent the opinion of the committee, which was composed of F. W. DeWolf, W. E. Wrather, Roswell H. Johnsont Wallace A. Pratt, Alexander W. McCoy, Carl H. Beal, C. T. Lupton, G. C. Matson, K. C. Heald, W. T. Thom, Jr., A. E. Fath, Kirtley F. Mather, R. C. Moore, and David White, chairman. Oil geologists and engineers understand well the speculative nature and the chances of error that must attend estimates of the petroleum in the ground when they cover not merely prospective territory, but possible territory. Neither the character nor the magnitude of the factors of uncertainty is to be disguised. Further, the inevitable differences of opinion between geologists of equal rank and experience must also bc

Jan 1, 1923

By Robert S. Baker

It has been demonstrated in the laboratory by A. J. Phillips1 that a transformation or conversion from beta directly to alpha may take place in a brass of 61 to 62.5 per cent copper content. The completion of the transformiation is dependent on a rapid rate of cooling from the all-beta field through the alpha + beta range to the alpha field. Phillips states that in a slowly cooled alloy containing 62 per cent copper and 38 per cent zinc the copper content of the beta is progressively decreased from 62 to 56 per cent by precipitation of alpha. Rapid cooling tends to prevent the separation of alpha with a copper content of more than 62 per cent by reducing to a minimum the time required for passing through the alpha + beta range. Upon crossing beyond the (alpha + beta) (alpha) boundary to the all-alpha region below, the material is transformed from beta to alpha. It, is essential that there be no change in composition. Phillips obtained his results by quenching a 62:38 copper-zinc alloy from 850" C. in iced brine at -8" C. No reference to the occurrence of this transformation in a manufacturing process has been found in the literature. An investigation of the microstructures of a number of hot forgings made from leaded brass rods yielded one forging in which this striking conversion from beta to alpha had occurred. The purpose of this paper is to describe the transformation and certain conditions under which it may occur during the hot pressing of brass. The conversion was noted in a shape forged from a leaded brass rod, the composition of which was: Cu, 60.26 per cent; Zn, 37.86; Pb, 1.88. In this study the lead content was disregarded and the alloy was considered to be equivalent to an alloy of 61.40 per cent copper and 35.60 per cent zinc (9812 parts = 100 parts). The possibility that the lead might obstruct or accelerate the precipitation of alpha was not studied. The lead appears as black globules in the photomicrographs and has a negligible solubility in the alloy.

Jan 1, 1932