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Discussion of the paper of E. J. Hall and C. W. Drury, presented at the Butte meeting, August, 1913, and printed in Bulletin No. 78, June, 1913, pp. 1059 to 1066. A. M. SMOOT,* New York, N. Y. (communication to the Secretary +) :-The nails method is given in many text books on assaying as a standard method for gold or for gold and silver in ores. It appears to be taught in the laboratories of technical and scientific schools as a standard method, and to some extent it is used by assayers in actual practice, although the class of ores to which it may be applied is limited to those free from reducible metallic compounds, the metals of which would enter the lead button. The experiments of Messrs. Hall and Drury show the inaccuracy of this rnethod for silver, in heavy sulphides, but not so clearly the short-comings with regard to gold. It seems certain that low silver is caused by the solubility of silver in iron-alkali sulphide as compared with its solubility in molten lead; that is, silver is distributed between iron-alkali sulphide and metal in the charge in proportion to solubility at the particular temperature of the fusion. Lead is a fixed quantity in these assay charges, but iron-alkali sulphide increases with increased sulphur, so the more sulphur in the charge, the lower the silver results.

Jan 11, 1913

By J. J. Calaman, D. H. Fleming

DURING 1950 the jet-piercing process was used commercially in the piercing of primary blast-holes in magnetic taconite at the preliminary taconite plant of .the Erie Mining Co., Aurora, Minn. The Erie Mining Co. purchased a jet-piercing machine and associated equipment in 1949, and the machine has been in operation since its installation in the summer of that year. An overall view of the pit with the JPM-1 jet-piercing machine in operation is shown in Fig. 1. During three months of 1949 and continuously since February 1950, production piercing took place under all pit conditions. By the end of October 1950, a total of 15,000 linear ft of hole had been pierced in an open pit handling an average of 1200 to 1800 tons of crude a day in one 8-hr shift. While it is recognized that this output does not approach the millions of tons a year planned for the developing Minnesota taconite projects, the machine as a unit has been operating under conditions probably more adverse than those that eventually will be encountered in the large operations. This pilot operation has permitted evaluation of the piercing problem in the full-scale taconite plant. This paper describes jet-piercing process funda- mentals, important features of the latest jet-piercing burners and machine, actual piercing results obtained, and new developments that give promise of materially increasing overall process efficiency. To produce the blasthole, jet-piercing utilizes thermal energy, as contrasted with the application of mechanical energy in churn drills and jackham-mers. The heart of the process may be termed a tailored flame produced by a rocket-type burner, thermodynamically similar in every respect to that of the giant rockets being developed for military purposes. The heat release in the combustion chamber of the burner is about 50x10' Btu per cu ft per hr as compared with 5x10' for most modern steam generators.* Combustion of liquid petroleum fuels (kerosine, diesel oil, or No. 2 fuel oil) with gaseous oxygen in the burner chamber at pressures of 100 to 200 psi results in temperatures of about 4300°F. Expansion of the hot combustion gases through divergent nozzles produces high-temperature supersonic jets having measured velocities of about 6000 fps. Impinging the flame jets on rock causes a thin surface layer to expand and break away from the base as a result of the thermally induced stresses. This spalling action is aided further in chert and quartzite by the 0.82 pct volume change resulting from the 6 to p quartz inversion at 1066"F.t The dynamic action of the jets whisks the spalled rock

Jan 1, 1952

This brief resume of the events transpiring in Mexico, culled from the daily New York newspapers since the last Bulletin went to press, does not show any degree of improvement in the situation. U. S. WITHDRAWS ARMY FROM MEXICO The American invasion of Mexico lasted less than 24 hours. The troops which crossed the Rio Grande June 15 to put a stop to the firing that was causing death and injury in El Paso were back in their billets on the American side of the border the next night. They had broken up the battle between Villa rebels and the Carranza garrison and driven the rebel forces headlong out of Juarez and into the hills. The Carranza Consul, Garcia, and Gen. Francisco Gonzales are reported by American officials to be pessimistic, if not indignant, at the swift reprisals made by the Americans against the stray shots which were falling in El Paso during the Villa attack on Juarez. In announcing his action in ordering Americans into Mexico Brigadier General Erwin in a formal statement said he had done-so to prevent the indiscriminate firing into El Paso by rebel soldiers, endangering the lives of Americans. "But on no account were they to undertake an invasion into Mexico," continued the statement. Americans with interests in the vicinity of Chihuahua are alarmed lest Villa, smarting under the defeat administered his troops by the American forces, should make a campaign of destruction. General Candido Aguilar, President Carranza's confidential ambassador to the United States, issued to the press, on June 17, a formal statement that "the government and people of Mexico consider as a violation of Mexican sovereignty the crossing of United States soldiers into Mexican territory" and expressing the hope ?that the situation created by the latest occurrences in Juarez will be satisfactorily adjusted between the two countries." Discussion of Mexican affairs was evoked in both houses of Congress by the crossing of the Rio Grande by American troops.

Jan 8, 1919

By J. M. W. Mackenzie

The interactions between oil drops and mineral surfaces have been examined for the cetyltrimethylammonium bromide (CTAB)-quartz and sodium dodecyl sulfate (SDS)-hematite systems. The results have been interpreted using a electrical double layer interaction model, a "sweeping-penetration" model based on steric interactions and a free energy of adhesion model. The adsorption of long-chain ions at the mineral interface was also studied and found to follow a Fuerstenau-Modi model of combined Coulombic and chain-chain interaction forces. The basis of the emulsion flotation process and the feature which makes it differ from conventional flotation is: encounter, followed by adhesion, between oil drops and mineral particles which are suspended in an aqueous environment. The manner in which the adhesion between these two phases contributes to the flotation process may be through the formation of min-eral-oil drop-mineral linkages, mineral-oil drop-air bubble linkages, or a combination of both. Possible flotation mechanisms derived from these adhesion processes are shown schematically in Fig. 2 while Fig. 1 shows an example of the first mechanism for quartz particles which have been floated using a cetyltrimethyl-ammonium bromide (CTAB)-Nujol emulsion as collector. In Fig. 2 the three phases, mineral, oil, and air, for convenience have been shown as having approximately equal size. This would not always be the case in an actual emulsion flotation operation. For the flotation of sand-sized particles of approximately 0.5 mm diam, the mineral particles would be at least ten times the size of the largest emulsion drops present in the system, while in the flotation of —10u slime particles, the mineral phase may be of the same order of size as the oil phase. The air bubbles in a flotation cell are normally about 1 to 2 mm in diam and therefore, in general, would be larger than the oil drops. Although the relative size of the three phases does not alter the basic adhesion processes between oil drops, mineral particles, and air bubbles, which depend on the physical and chemical properties of the interfaces involved rather than on the extent of these interfaces, it does alter the degree of deformation of the oil drops and air bubbles which will occur on adhesion. For mechanisms 1 and 2 to take place, the oil-water and mineral-water interfaces must be partially replaced by a mineral-oil interface while in mechanism 2 there is the additional requirement that the oil-water and air-water interfaces must be replaced by an oil-air interface. In view of the importance of the oil drop-mineral particle encounter and subsequent adhesion stages to both of the emulsion flotation mechanisms discussed, a study has been made of some physicochemical factors which control these processes for Nujol-cetytri-methylammonium bromide (CTAB)-quartz and Nujol-sodium dodecyl sulfate (SDS) -hematite systems. Experimental Electrokinetic Measurements.—The zeta potential of mineral particles was measured using a streaming potential technique similar to that described by Fuerste-nau.1 The streaming potential was measured using a multirange electrometer while the plug conductivity was measured using a 50 cycle sec1 bridge. Measurements were made on 28 x 65 mesh particles. Contact Angle Measurements.—A captive bubble technique was used to measure the contact angles for air bubbles and oil drops. The angles were measured on fracture faces of unmounted minerals which proved easier to clean than polished mounted specimens and should more closely represent a flotation system. Reagents: These have been described in a previous publication.0 Minerals: Rock crystal quartz from the New England district, N.S.W., was used for contact angle and streaming potential measurements. The quartz was hand

Jan 1, 1971

THERE are two general classes of problems in connection with electric furnaces. First, those relating to the best utilization of the electrical power inside the furnace; second, those connected with the bringing of the electrical power to the point where it is to be utilized. Referring particularly to the latter class of problems, many calculations have been made, covering a wide variety of furnaces, to check the electrical capacity and power factor found from test. The following analysis, made for the purpose of determining the maximum capacity and approximate performance of a new furnace, to operate at 160 volts on a 60-cycle circuit, may be of interest. The first characteristic necessary to determine is the inductance, since this is the factor that limits the amount of power which can be put into the furnace. Inductance is calculated from the dimensions of the furnace, according to formulas of the U. S. Bureau of Standards,1 as

Jan 10, 1918

By W. A. Anderson, R. F. Mehl

RECRYSTALLIZATION of cold-worked metals has long been known to proceed by a process of nucleation and growth.1 When a cold-worked metal is heated to a temperature at which recrystallization will ensue, nuclei of new grains, essentially strain-free, appear and grow. As the process proceeds, nuclei continue to appear and to grow, until the cold-worked matrix has been entirely consumed and recrystallization has been completed (Fig. I). Conducted isothermally, a measurement of the fraction of the matrix recrystallized plotted against time, yields an isothermal recrystallization curve, showing the time course of the recrystallization process (Fig. 2).2 Evidently, then, the rate of recrystallization is determined by the rate at which nuclei form and the rate at which they grow. The rate of nucleation N is the number of nuclei that form in unit time in unit volume of the unrecrystallized matrix, and the rate of growth G is the increase in radius of the recrystallized grain per unit time. In recent years attempts have been made to determine N and G as these vary with the well-recognized variables affecting recrystallization: time, temperature, degree of deformation, etc.' Such studies yield information that is more fundamental than that conveyed by the conventional recrystallization diagram,' since the rate of recrystallization is thus analyzed into its two component rates.2 Methods are now available by which N and G can be measured, and analytical expressions have been developed for the interrelationships among N, G, the isothermal recrystallization curve, and the final grain size, which greatly facilitate analysis.2-4 The application of these methods has been demonstrated recently for the recrystallization of silicon ferrite.2 Data on N and G in dependence upon the recognized recrystallization variables are, however, very fragmentary;* values of N and G have been determined for some metals, but the effect of temperature, which would yield the fundamentally important activation energies, has been too little studied. Thus far there has been no detailed study of the recrystallization of any one metal in terms of N and G over appreciable ranges of the variables time, temperature, degree of deformation, and original grain size. It is the purpose of the present paper to provide such a study, employing aluminum.

Jan 1, 1945

By A. G. Metcalfe

Observations at temperature are used to investigate the phase changes in alloys containing more than 50 pct Co and above 1000°C. The nonsuppressible transformations in cobalt above 1120°C and in the intermetallic compound Co2Cr3 are studied. The liquidus and solidus surfaces are determined. Insufficient data is obtained to complete the constitution diagram. A REVIEW of some of the earlier work on cobalt' and on the binary systems with chromium2-4 and with molybdenum,5 showed that one likely source of differences might be attributed to the presence of phases at high temperatures which could not be retained by quenching. Consequently, the decision was made that only methods of examination at temperature should be used in this work. The correctness of this decision was confirmed during the course of the work by the investigation of the martensitic nature of the transformation in cobalt- and by the decomposition of the high temperature form of the phase Co2Cr3 on quenching.' The use of high temperature methods, including dilatometry and thermal etching led to the postula-tion of the existence of a high temperature form of cobalt with a probable hexagonal close-packed structure, transforming from cubic cobalt over a range of temperature from 1119° to 1145°C.8 One of the important pieces of evidence for this transformation was that it was observed in alloys with chromium and molybdenum which were nonmagnetic at this temperature so that confusion with magnetostriction effects was impossible. These results were not reported in this earlier paper, so that the data given here supply further evidence for the existence of a second allotropic transformation in cobalt. High temperature observations are necessarily much slower in producing results than the examination of quenched specimens at room temperature by conventional means such as X-ray diffraction or microexamination. The investigation has only covered a limited part of the field, as a result, extending from the liquidus down to about 1050°C and from the cobalt corner up to 50 pct alloying elements. The differences between the various reported diagrams for the constitution of the Co-Mo and Co-Cr systems have already been mentioned. The best available diagram in the first system is due to Sykes and Graff' which is shown in Fig. 1. Fig. 2 shows the latest Co-Cr diagram due to Elsea, Westerman, and Manning.' A comparison reveals that these are incompatible because of the disagreement with regard to the number of allotropes of cobalt. Elsea, Westerman and Manning took their liquidus and solidus curves from Wever and Haschimoto.2 The temperature horizontal in the two-phase field beneath the eutectic was placed 35" to 39°C higher than in earlier work2,4 with a new interpretation of the nature of this transformation. This interpretation also required a peritectic horizontal at about 1470°C, which has been confirmed here for an alloy with 53 pct Cr. No investigation had been made of the ternary system until after this work had been terminated, when a section at 1200°C was published." In this work a ternary compound was found by the X-ray diffraction examination of quenched alloys. This compound contained less than 50 pct Co, and was thus outside the field investigated here. However, some reactions were found dilatometrically which could not be fitted on the diagram and are now believed

Jan 1, 1954

By D. E. Light, J. Y. Somnay

The use of alkyl acid phosphates and their alkali salts as collectors for the uranium minerals brannerite and uranothorite was investigated. In particular a detailed flotation study was carried out using isooctyl acid phosphate as a promoter. Tests were conducted on an Elliot Lake ore analyzing 0.11% U 3 O 8, chiefly as brannerite, and approximately 9% pyrite. Pyrite was floated with conventional reagents; the flotation of brannerite from the sulfide tails was studied to determine the optimum pH and pulp density to employ with isooctyl phosphate. 4 uranium recovery of over 90% was possible with a selectivity index of 8.95 when these tails were floated at 17% solids and a pH of 1.7. The effect of various modifying agents was also investigated. Aluminum sulfate, lactic acid and sodium silicate improved the selectivity of collection yielding selectivity indices of 9.5 to 11.5. Ferric chloride was found to be a depressant for brannerite. A Bancroft ore which analyzed 0.13% U308, and contained uranothorite was also amenable to flotation using isooctyl acid phosphate. The composite uranium concentrate obtained assayed 0.41% U308 at a recovery of 95.3% and a ratio of concentration of 3.24. Previous investigations1,2 on the flotation of brannerite, pitchblende, uraninite and/or uranothorite from various Canadian ores showed that fatty acids, including oleic acid and tall oils, petroleum sulfo-nates and alkyl acid phosphates, were suitable collectors for these uranium minerals. Additional work carried out by the Mines Branch in Ottawa on brannerite ores from the Elliot Lake area3 used the tall oil Acintol FA-1, or FA-2, as a collector. A recovery of 92% of the uranium in a concentrate which contained about 55% of the weight and assayed 0.2% U3O8 was obtained from a feed of 0.1% U308 in this latter instance. Eigeles et a1.4 carried out flotation on pitchblende ores and found that the most promising results were obtained with alkyl acid phosphates and their alkali salts as collectors. Within the homologous series of alkyl phosphates, the best flotation characteristics were displayed by isooctyl phosphates. Further, isooctyl phosphate derivatives were much less sensitive to hardness forming salts than oleic acid, enabling all flotation tests to be made with tap water. This paper describes tests employing alkyl acid phosphates for the flotation of ores containing brannerite and uranothorite. The flotation work described herein was conducted in conjunction with the development of a leaching process for Elliot Lake, Ontario, brannerite ores. In this process pyrite in the ore was concentrated by flotation and roasted, providing sulfur dioxide which was used to produce sulfuric acid in situ in a liquid-solid extraction process. A 40% conversion to sulfuric acid was indicated at a leaching temperature of 80°C; this resulted in the extraction of 93% of the total uranium after a contact period of 6 hr. A rotary kiln was used as a leaching vessel. Preconcentration by flotation was studied in conjunction with the above process to reduce the bulk of solids for leaching and, consequently, capital expenditure. The materials and methods used in this study will be described initially. Then the results will be presented in two sections, the first dealing with the brannerite ore, and the second with the uranothorite. Finally an economic analysis of the results on the brannerite material will be made. MATERIALS AND METHODS The two ore samples examined were from uranium-producing mines. One was a brannerite ore, a uraniferous quartz pebble conglomerate from the Elliot Lake area, which contained 0.11% U3O8, 4.3% S and 4.5% total Fe. Mineralogical studies of similar samples indicated the presence of quartz, microcline, sericite, pyrite and minor quantities of rutile, pyrrho-tite and zircon. The other ore was from a biotite-rich pegmatite in the Bancroft area; feldspar, quartz and biotite formed the major gangue components. Uranothorite was the principal uranium mineral; the head assay was 0.13% U 3 O 8. Preparation of Feed: The ores were crushed to —10

Jan 1, 1963

By Mark Sheppard

DURING the winter of 1937, the writer visited a West Virginia stone, quarry at which the overburden is stripped hydraulically. The quarry is in a bed of limestone, about 200 ft. thick, which outcrops on a hillside above a small river; the quarry face is 175 ft. high. The top of the stone is cut in all directions by erosion channels, which are filled with a tough, red clay. Some of the larger channels are 30 ft. deep and 8 ft. wide, The thickness of the clay overburden ranges from a few inches to 10 ft., the average being estimated as less than 3 feet. The topography of the district is rugged, and the hillside above the quarry face is not suited to the use of the usual mechanical methods of moving dirt. Until hydraulicking was started 10 years ago, the over-burden was dug by hand labor, dumped to the quarry floor and trans-ported to the dump in quarry cars, Cleaning deep channels was a slow and expensive process. When hydraulicking was started, the top of the quarry face was over 200 ft. above the level of the river. The pipe line was carried in a semi-circle around the top of the quarry about 200 ft. from the face and about 50 ft, above it. As there is no adequate supply of water above the quarry, the water must be pumped from the river: As quarrying pro-gressed the pipe line was moved until the discharge points are now over 300 ft. above the river. Water is pumped from the river by a four-stage centrifugal pump having a 5-in. intake and a 4-in. discharge. It is direct-connected to a 200-hp,, 2300-volt, 1760-r.p.m. induction motor. At 1760 r.p.m. the rated capacity of the pump is 750 gal. per minute under a head of 600 ft. The pump is installed in a house of frame construction on the river bank, about 10 ft, above normal water level, which also houses the pump for the mill-water supply. Above the pumps are hoisting hays into which the motors are raised whenever the river reaches flood stage.

Jan 1, 1938

By Samuel S. Wyer

Bethlehem Meeting, February, 1906 THE following abbreviations have been used in the text:¬ Am. MF. and Iron World. American Manufacturer and Iron World, Pittsburg, Pa. Can. Min. Rev. Canadian Mining Review, Ottawa, Can. Cass. Mag. Cassier's Magazine, New York, N. Y. Col. Eng. Colliery Engineer, London, Eng. Col. Guard. Colliery Guardian, London, Eng. Eng. and Min. Jour. Engineering and Mining Journal, New York, N. Y. Eng. News. Engineering News, New York, N. Y. Engr. Engineering, London, Eng. J. I. and S. I. Journal of the Iron and Steel Institute, Lon¬don, Eng. Jour. W Soc. of Engrs. Journal Western Society of Engineers, Chicago, Ill. Lond. Eng. The Engineer, London, Eng. Mines and Min. Mines and Minerals, Scranton, Pa. North of Eng. I. M. E. North of England Institute of Mining Engineers, Newcastle-on-Tyne, England. Oester Zeitsch f. Berg u. Hüt. Oesterrcichische Zeitschrift für Berg. und Hüttenwescn, Vienna, Austria. Pract. Engr. Practical Engineer, London, Eng. Proc. I. C. E. Proceedings Institution of Civil Engineers, London, Eng. Trans. A. I. M. E. Transactions American Institute of Mining Engineers, New York, N. Y. Trans. A. S. M. E. Transactions American Society Mechanical Engineers, New York, N Y Trans. Fed. I. M E. Transactions Federated Institute Mining Engineers, Newcastle-on-Tyne, England. Trans. I M E Transactions Institute Mining Engineers, Newcastle-on-Tyne, England.

Mar 1, 1906

By Paul A. Taylor

MUCH information concerning pegmatites which was thought to be true a few years ago has been proved false, and what is now actually known about some pegmatites is not true of many others. The erratic and seemingly unpredictable structure and variable composition of this class of mineral deposits has been widely emphasized. Even parts of the same pegmatite body exhibit marked differences in texture, mineral composition, width, and attitude. Constructive geological thinking in respect to pegmatites now aims to establish general laws rather than to stress the confusing diversity of features having no special economic significance. Substantial progress has been made in classifying different types of these deposits according to general features, internal structure, mineralogy, and origin. In some cases it has even been possible to block out tonnage reserves in advance of mining. It is still easy, however, to make highly erroneous predictions after a preliminary examination of a pegmatite prospect. Pegmatites are important to the economic well being of the country and to its military security. They furnish virtually all the feldspar, strategic mica, beryl, columbium, tantalum, and caesium utilized in the United States, as well as sundry other minerals and significant amounts of lithium and rare earth minerals and gems. With the exception of vermiculite, occasional ilmenite-rutile, and perhaps soda-lime feldspar and garnet deposits, basic pegmatites are of little economic importance. Consequently in this paper, as in common parlance, the term pegmatite generally relates to coarsegrained acidic rocks or what is aptly called giant granite. Available data indicating the size and importance of the production and trade in specified pegmatite minerals are summarized in Table I. Geological Features Much of the latest thinking on the economic geology of pegmatites is now available in a 115-page monograph' by a group of experts who participated with geologists of the Federal Geological Survey in the widespread wartime investigations. Doubtless the most significant feature of the monograph is indicated by the title, The Internal Structure of Pegmatites, but it also contains a vast amount of other new information and includes the assimilated concepts of many earlier writers, whose works are given in a comprehensive list of references. The shape, size, attitude, and continuity of many pegmatite bodies is controlled by the structure of the older rocks in which they occur. If the older rocks are easily penetrated, e.g., biotite schist, most of the pegmatites in a given district will be found outside the parent granite mass as exterior pegmatites. Marginal pegmatites are more prevalent if the older rocks are massive, unsheared, and sparingly jointed. Networks of pegmatites are abundant in highly-jointed rocks. In strongly foliated schists the bodies are usually lenticular, whereas in highly-folded areas they assume tear drop, pipe or pod-like, bow-shaped, or sinuous forms. Jahns2 recognizes five major shape classes: l—dikes, sills, pipes, and elongate pods; 2— dikes, sills, pipes, and pods with bends, protuberances, or other irregularities; 3-—trough-or scoop-shaped bodies with or without complicating branches; 4—bodies with the form of an inverted trough or scoop; and 5—other bodies, including combinations of the above and miscellaneous shapes. Many pegmatite deposits are scarcely big enough to be recognizable as such. Most of them, in fact, are small tabular deposits less than 4 in. wide and usually without economic concentrations of minerals. On the other hand, some pegmatites are more than a mile long and over 500 ft wide. The ratios of length to breadth range from 1 : 1 to 1 : 100. Although the vertical dimension bears no invariable relationship to strike length, tabular deposits or large lenses are often symmetrical enough to show nearly as much continuity down dip as in their horizontal extension, and some pipes or pods are amazingly persistent in the vertical plane. Small pegmatites often string along a fairly definite trend line; in a given district major bodies may lie roughly parallel, and where only a few of them do not, the erratically disposed bodies generally differ in composition from those conforming to the regular pattern. This does not apply, however, in all districts. Characteristically, pegmatite veins pinch and swell or split into branches. When they pinch out entirely it is often possible to find a new body by prospecting the extension of the strike or dip, but the chances of finding a hidden deposit are ordinarily too uncertain to justify much subsurface prospecting. Diamond drilling may yield valuable information as to the continuity of known deposits whose upper portions are well-exposed. Some deposits, in fact, can be proved up for hundreds of feet by surface trenching and then intersected by drill holes at various depths like any other vein-like deposit. Others twist and branch, apparently defying all efforts to explore them short of actual mining.

Jan 1, 1954

By E. V. Kleber, V. F. Novy, R. C. Vickery

The constitutional diagram has been determined in part for the sgstem Gd-Fe. Seven intermetallic compounds have been found at compositions corresponding to the following Gd-Fe ratios; 2:3, 1.2, 1:3, 2:7, 1:4, 1:5, and 2:17. The GdFe, com-pound has a congruent melting point above 18000C, whereas all of the remaining compounds melt incongruently, A eutectic exists at 13 wt pct Fe at a temperature of 860°C. ALTHOUGH an increasing amount of information is becoming available on the metallurgy of pure rare earth metals, relatively few studies have been made on alloy systems of the rare earths with the transition metals; iron, nickel, and cobalt. Even less data are extant on gadolinium alloys although it could be expected that such alloys might possess interesting structural and magnetic properties. In this paper, we report investigation of the gadolinium-iron alloy system. Studies on systems containing nickel and cobalt will be presented in subsequent papers. Previous data on gadolinium-iron compounds have been reported by Endter and Klemm1 and Jep-son and Duwez.2 Both reports agree on the existence of the GdFe2 phase, which crystallizes with the C-15type structure (face-centered cubic). This is in accord with results of the present work. EXPERIMENTAL The gadolinium metal used was prepared by met-allothermic reduction of high purity gadolinium fluoride. A typical analysis of the metal thus produced indicated the principal impurities to be oxygen, 0.24 pct, and tantalum, 0.2 pct. The iron metal used for alloying was a reagent grade material. Alloys were prepared by arc melting mixtures of the respective metals. Appropriate quantities of the two component metals were carefully weighed, blended and compacted. The compacts were then melted in an arc furnace consisting of a water-cooled copper hearth plate and a water-cooled tungsten-tipped counter electrode contained in a brass vacuum chamber. The furnace was evacuated to an absolute pressure of less than 1 with the leak rate established as being less than 20 per min. The furnace was then back-filled to a pressure of 25 mm of Hg, with a 75 pct He-25 pct Ar mixture. Arc melting at 25 v and 300 amp was maintained for about 30 to 60 sec until the entire visible area of the sample became molten, by which time surface tension had drawn the molten metal pool into a large bead. After allowing the furnace to cool, the alloy button was inverted and the melting process repeated. A series of melts was made covering entire composition ranges from 100 pct Gd to 100 pct Fe. All samples were examined metallographically in the as-melted condition. As the phase diagram became rationalized, and differential thermal analysis had given indications of temperature arrests, annealing was carried out in vacuo or in argon. Generally, compositions containing more than about 60 pct Gd were heat treated at 750°C and alloys containing less than this quantity of gadolinium were annealed at 1000°C. Prolonged annealing times were of the order of 100 to 200 hr. Metallographic examination was repeated on the annealed materials. Although little material loss was noted during alloy preparation, chemical analyses were made on all specimens prepared. The following standard, analytical techniques were used. Samples were dissolved in hydrochloric acid; insoluble material was filtered off, ignited and fused with potassium pyro-sulfate. The cold melt was leached with dilute acid and the solution thus obtained added to the main sample solution. Gadolinium was always determined gravimetri-callv as oxide from oxalate ignition. Iron was removed from an aliquot sample by electrolysis on a mercury cathode at approximately six volts. Gadolinium in the iron-free electrolyte was then precipitated as oxalate which was subsequently ignited to oxide. Iron was determined volumetrically in an aliquot solution by reduction with stannous chloride fol-

Jan 1, 1962

By Robert W. Elayer

In August 1973, Fluor Utah, Inc. was awarded a contract by the Office of Coal Research (now part of the Energy Research and Development Administration), Department of the Interior, for the examination of the unique problems of large-scale surface coal mining to supply coal conversion complexes. The project consists of two phases: (I) data base development and organization and the formulation of a mathematical computer modeling system for surface mining; and (II) computer manipulation of the data base in accordance with the model to formulate several typical preliminary mine and preparation plant designs. Phase I is scheduled for completion by the end of July 1975 and Phase II is scheduled to be complete by the end of January 1977. This paper briefly discusses the effort for only one aspect of this project-the foundation for geologic modeling developed in Phase I. The Phase I task of geologic modeling can be broken down into two subtasks: (1) identify areas containing sufficient resources of surface-minable coal to support at least one large-scale mining complex, and (2) characterize the coal deposits in these areas for computer storage and retrieval. Identification of Resource Regions The basic production design criteria for large-scale surface mining, as stated in the contract, is the production of 75,000 tpd, or more, of bituminous coal. Based on a constant total Btu delivered to the preparation plant, daily production requirements for the different coals are as follows: Bituminous Coal 75,000 tons Subbituminous Coal 99,500 tons Lignite 143,500 tons Although surface mining on this scale is presently non-existent, it is deemed that future operations, in order to meet the increased energy demands, will have to look forward to much larger scale mining complexes. These tonnage requirements are meant to be used only for the delineation of general areas suitable for large-scale surface mining, termed "Strippable Coal Resource Regions." They were determined largely from the latest complete detailed coal resource estimates of the U.S. Bureau of Mines. Nine strippable coal resource regions, containing over 100 billion tons of strippable coal resources, were identified [(see Fig. 1)]. The delineation of these regions was based entirely upon estimates of surface-minable coal resources and the production design criteria for large-scale surface mining. If surface-mined coal was combined with underground production, or if smaller production rates were specified, then other areas which were omitted here would be included in the regions. Changing the outline of the regions, however, would not add any new mining situations to be studied in Phase II. Characterization of Coal Deposits The second subtask involved the general characterization of the coal deposits contained in the strippable coal resource regions. The geological information gathered

Jan 1, 1976

By Eugene F. Poncelet

GLASS squares compressed on edge by steel jaws in poor contact with them developed jagged "partial-contact" cracks caused by the formation of local tensile stresses. Compressed by steel jaws in perfect contact, they developed smooth "release cracks" during release of the pressure, caused by the formation of local tensile stresses during the release of the pressure. All these cracks were parallel to the pressure. A Microflash photograph of a disintegrating specimen under sufficient pressure reveals a network of fractures roughly normal to each other, also "release cracks" and a disintegration cloud. The Griffith theory is amended to account for the formation of a first crack. A new theory, based on the theory of thermal agitation and wave propagation, is proposed to account for the progress, velocity and forking of cracks. The network of fractures is shown to have been caused by reflection at a free boundary of pressure pulses emanating from a first crack. Postulating equal distribution of energy in the pulses emitted on either side by a crack in particles, the smaller fragments are shown to continue fracturing preferentially while some of the coarser fragments remain as residual pieces. As comminution of the smaller fragments proceeds, the solid is reduced to a collection of residual particles of smaller and smaller sizes, accounting for the disintegration cloud. EXPERIMENTAL PROCEDURE Apparatus A crushing device was built, consisting of a steel frame and two polished, hardened steel jaws to crush plate-glass samples. This device was placed between the plates of a hydraulic press (Fig. I). For centering the pressure, the upper jaw was guided vertically and provided with a rounded top. Later it was replaced by a steel jaw ground out of a 2 1/2-in. steel ball, the lower surface being ground to an optical flat (Fig. 2). Four rollers were used to eliminate side thrust. Screws and centering blocks served to center the specimen in the crushing device. Hard steel shims placed under the shoulder of the upper jaw limited the "follow through" of the jaw after failure of the specimen. Stress distribution was first examined with a polariscope mounted on an optical bench, using monochromatic light. It was found convenient to mount polaroid, quarter wave and monochromatic plates directly on the device itself. A Microflash unit* was used as light source for the photographs. A gold-paint pattern was fused on the plate-glass specimens and connected into the spark-

Jan 1, 1944

By L. G. Sprague

The fact that this latest and most modern of the Universal Atlas Cement Company's plants at Northampton, Pa., is the fifth to be built on these same properties, and their development has been coincidental with the development of the portland cement industry in the United States, makes it appropriate to sketch briefly the history of the successive steps. In 1889 the original Atlas Portland Cement Co. started its first plant at Coplay, Pa., which is just across the Lehigh River from the present operation. Here the first successful rotary kiln was operated. These kilns were 6 ft. in diameter and 60 ft. long; they used oil for fuel until 1895, when the company developed the first burner for pulverized coal. Shortly thereafter the Atlas company acquired property on the Northampton side of the river and built its plants Nos. 2, 3 and 4 within the following 10 years. During the same period the production of portland cement in the country was increasing rapidly; owing largely to the development of the rotary kiln and improved grinding machinery, which lowered the cost and improved the quality of cement. The general tendency has been to increase the length of the kiln and grind both raw material and finished product Liner. In 1930 the Atlas Portland Cement Co. and the Universal Portland Cement Co. united to form the Universal Atlas Cement Co., and the same policy of improving the product has been continued. Because of the large amount of research carried on by manufacturers and consumers, including government agencies, there is a demand for cement of different chemical and physical characteristics for different kinds of structures. Hence many engineers in recent years have been asking for cements of certain definite chemical analyses, which they think are best for their particular jobs. This has compelled the manufacturer to make several different types of cement by variations of chemical composition. These are generally made by adding silica, iron and lime to the standard mixture, which necessarily increases the cost of the product. Therefore the Universal Atlas Cement Co., in planning the modernization of its Northampton plant, gave this particular phase of the problem considerable study. Research and experimentation have developed the practicability of making a wide variety of differently proportioned raw materials for cement by a process of subtraction rather than by addition. With such process, it is possible to discard the undesirable aluminates and silicates from the ample supply of impure limestone available from the company's property rather than by adding high-grade limestone obtained from distant sources. This procedure requires mineral separation by centrifugal sedimentation and froth flotation. Since the chemical and mineral constituents of Lehigh Valley "cement rock," together with the methods of separating them, have been thoroughly described in published

Jan 1, 1948

By L. G. Sprague

The fact that this latest and most modern of the Universal Atlas Cement Company's plants at Northampton, Pa., is the fifth to be built on these same properties, and their development has been coincidental with the development of the portland cement industry in the United States, makes it appropriate to sketch briefly the history of the successive steps. In 1889 the original Atlas Portland Cement Co. started its first plant at Coplay, Pa., which is just across the Lehigh River from the present operation. Here the first successful rotary kiln was operated. These kilns were 6 ft. in diameter and 60 ft. long; they used oil for fuel until 1895, when the company developed the first burner for pulverized coal. Shortly thereafter the Atlas company acquired property on the Northampton side of the river and built its plants Nos. 2, 3 and 4 within the following 10 years. During the same period the production of portland cement in the country was increasing rapidly; owing largely to the development of the rotary kiln and improved grinding machinery, which lowered the cost and improved the quality of cement. The general tendency has been to increase the length of the kiln and grind both raw material and finished product Liner. In 1930 the Atlas Portland Cement Co. and the Universal Portland Cement Co. united to form the Universal Atlas Cement Co., and the same policy of improving the product has been continued. Because of the large amount of research carried on by manufacturers and consumers, including government agencies, there is a demand for cement of different chemical and physical characteristics for different kinds of structures. Hence many engineers in recent years have been asking for cements of certain definite chemical analyses, which they think are best for their particular jobs. This has compelled the manufacturer to make several different types of cement by variations of chemical composition. These are generally made by adding silica, iron and lime to the standard mixture, which necessarily increases the cost of the product. Therefore the Universal Atlas Cement Co., in planning the modernization of its Northampton plant, gave this particular phase of the problem considerable study. Research and experimentation have developed the practicability of making a wide variety of differently proportioned raw materials for cement by a process of subtraction rather than by addition. With such process, it is possible to discard the undesirable aluminates and silicates from the ample supply of impure limestone available from the company's property rather than by adding high-grade limestone obtained from distant sources. This procedure requires mineral separation by centrifugal sedimentation and froth flotation. Since the chemical and mineral constituents of Lehigh Valley "cement rock," together with the methods of separating them, have been thoroughly described in published

Jan 1, 1948

By Basil B. Zavoico

Production of crude oil in the U.S.S.R. during 1940 is estimated at 222,600,000 bbl., as compared with the revised figure for the preceding year of 220,866,000 bbl., an increase of 0.79 per cent, and at least 20 per cent under the planned output. Baku district (Transcaucasus) remained the outstanding producing area, accounting for 72.33 per cent of the country's production, and this relative importance of the Apsheron Peninsula oil fields will probably continue lor many years to come. In the North Caucasus the Grozny-Dagestan area further extended its decline and the current production in this district amounts to only 6.06 Per cent Of the total U.S.S.R. output, as compared with 35.76 Per cent in 1931 The production of the Kuban-Maikop district remained virtually ullchanged from the preceding year, amounting to 7.64 per cent of the country's total. The Caucasian Province, therefore, in I940 produced 86.03 Per cent of the total production of the U.S.S.R. The promising new Permian Basin district, called Ural-Volga, and referred to sometimes as the "Second Baku," increased its production appreciably during 1940 and the output of that area amounted to 8.31 per cent of the country's total, as compared with 6.35 per cent during 1939. Production in the secondary districts of Russia continued unchanged though some new discoveries in the Emba salt-dome basin suggest that the output of that district may show some increase during the coming years. The drilling operations in Russia during 1940 reached a stage of virtual breakdown and the original footage plans were fulfilled only to the extent of around 50 per cent, while exploration drilling schedules were fulfilled only to the extent of about 25 to 35 per cent. This condition was due to the obvious lack of modern drilling rigs, of good drill pipe and of strong casing. Extended breakdowns occurred with deeper drilling, particularly because with few good drilling rigs available speed was being attempted with rotation rates that the inferior drill pipe and bits could not stand. Twistoffs and endless fishing jobs resulted. The difficulties became so serious a handicap to the Soviet oil industry that it was found imperative to order American drilling equipment and pipe in quantity, a policy that was never favored in the U.S.S.R because it requires expenditure of gold or exchange in major volume, and such exchange usually is reserved only to less bulky key items or to orders directly connected with armaments; also, oil-field machinery is very bulky and the prevailing lack of bottoms makes the transport of the indicated equipment a major problem. But little progress was similarly made in other branches of the Soviet oil industry. NO new major pipe-line construction work was undertaken, consequently the railroads continued to be overburdened with traffic that could be much more effectively handled by pipe lines. No new refinery construction of any magnitude was initiated, though several plants that had been in course of construction for several years were completed during 1940 and operation 554

Jan 1, 1941

By Basil B. Zavoico

Production of crude oil in the U.S.S.R. during 1940 is estimated at 222,600,000 bbl., as compared with the revised figure for the preceding year of 220,866,000 bbl., an increase of 0.79 per cent, and at least 20 per cent under the planned output. Baku district (Transcaucasus) remained the outstanding producing area, accounting for 72.33 per cent of the country's production, and this relative importance of the Apsheron Peninsula oil fields will probably continue lor many years to come. In the North Caucasus the Grozny-Dagestan area further extended its decline and the current production in this district amounts to only 6.06 Per cent Of the total U.S.S.R. output, as compared with 35.76 Per cent in 1931 The production of the Kuban-Maikop district remained virtually ullchanged from the preceding year, amounting to 7.64 per cent of the country's total. The Caucasian Province, therefore, in I940 produced 86.03 Per cent of the total production of the U.S.S.R. The promising new Permian Basin district, called Ural-Volga, and referred to sometimes as the "Second Baku," increased its production appreciably during 1940 and the output of that area amounted to 8.31 per cent of the country's total, as compared with 6.35 per cent during 1939. Production in the secondary districts of Russia continued unchanged though some new discoveries in the Emba salt-dome basin suggest that the output of that district may show some increase during the coming years. The drilling operations in Russia during 1940 reached a stage of virtual breakdown and the original footage plans were fulfilled only to the extent of around 50 per cent, while exploration drilling schedules were fulfilled only to the extent of about 25 to 35 per cent. This condition was due to the obvious lack of modern drilling rigs, of good drill pipe and of strong casing. Extended breakdowns occurred with deeper drilling, particularly because with few good drilling rigs available speed was being attempted with rotation rates that the inferior drill pipe and bits could not stand. Twistoffs and endless fishing jobs resulted. The difficulties became so serious a handicap to the Soviet oil industry that it was found imperative to order American drilling equipment and pipe in quantity, a policy that was never favored in the U.S.S.R because it requires expenditure of gold or exchange in major volume, and such exchange usually is reserved only to less bulky key items or to orders directly connected with armaments; also, oil-field machinery is very bulky and the prevailing lack of bottoms makes the transport of the indicated equipment a major problem. But little progress was similarly made in other branches of the Soviet oil industry. NO new major pipe-line construction work was undertaken, consequently the railroads continued to be overburdened with traffic that could be much more effectively handled by pipe lines. No new refinery construction of any magnitude was initiated, though several plants that had been in course of construction for several years were completed during 1940 and operation 554

Jan 1, 1941

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

A perspective is presented of the steel plant sulphur distribution and elimination problem from coal to liquid steel ready for teeming, giving distributions of sulphur over a range of coke sulphur content, and some methods of sulphur control, in the blast furnace, external desulphuriza-tion between blast furnace and open hearth, distribution between fuel, slag, and metal, and methods and limitations of control of sulphur in the open hearth furnace. AS a part of the 1951 AIME symposium on sulphur in steelmaking, it was thought that a discussion of the distribution of sulphur throughout the whole series of operations, from coal and ore to finished steel ingots, might have some value in giving a perspective on the whole problem. The following discussion is an attempt to present such an overall picture. The order is that of the actual plant operations, beginning with a very brief consideration of the coking process. Sulphur in Coal and Coke Since by far the largest source of sulphur entering the steelmaking cycle is in the coal used to make coke for the blast furnace, it would seem reasonable to eliminate some of it, either from the coal, or the coke, or during the coking process. This has appeared impracticable up to the present, at least, for two main reasons: the low activity of the organic sulphur in either coal or coke, and because of price limitations involved in treating a low cost material such as coke. A variable portion, usually M or less, of the sulphur is present in coal in the form of pyrites or similar compounds, and a large part of this sulphur may be removed in the coal washery. Most of the sulphur, however, is normally present as "organic" sulphur, intimately associated with the coal structure. Its distribution prevents any separation by mechanical means. Its low activity makes improbable ' any rapid chemical removal, although hydrogen will remove sulphur from both coal and coke. Thus, prolonged recirculation of coke oven gas in the coking process would tend to leave a smaller percentage of the total sulphur in the coke residue. Table I shows a typical distribution of sulphur from coal into products in the coking process. As the sulphur in the coal increases, the sulphur in the coke tends to increase in about the same proportion. Sulphur in the Blast Furnace The best picture of the situation in the blast furnace is provided by a sulphur balance of raw materials entering, and of products leaving, the furnace. The difficulties in accurate weighing and sampling of the variable solid materials entering this process, and the number of hours required for the raw materials to descend through the furnace under variable operating conditions, make it difficult to obtain an accurate balance. However, balances made over periods of weeks or months tend to average out some of these uncertainties. Table I1 presents three typical sulphur balances similar to a number that the writers have calculated. In most of these the slag volume calculated from the sulphur balance is, in some instances more, and in other instances less, than the value corresponding to the best input and output balances of the other slag constituents (lime, silica, alumina, etc.). Probably the greatest source of error in these calculations is the sulphur content of the slag. Despite some possible inaccuracies the balances of Table II show rather definitely the following points: 1—That 87 to 95 pct of the total sulphur input is in the coke and 95 to 97 pct of the total sulphur output is in the slag. Also, that if any sulphur leaves the furnace with the gas it is relatively small, amounting to a possible 1 pct or less. 2—At the lower sulphur coke level of 0.86 pct the total amount of sulphur charged is 15 Ib of sulphur per ton iron increasing to 26 lb per ton at the higher sulphur, intimately associated with the coal struc-rare burdens containing sulphur-rich ores will the total sulphur burden fail to be nearly proportional to the content in the coke used. 3—The 7 to 9 pct of the total sulphur input from the limestone of furnaces B and C is due to the relatively high sulphur content of the stone, 0.226 and 0.265 pct, respectively. In the case of furnace A, the sulphur content of the limestone was only 0.06 pct which resulted in only 3 pct of the total sulphur input coming from this source. It is rather interesting to compare the sulphur balances of a typical ferromanganese furnace with

Jan 1, 1952

By J. F. Olk, E. D. Spaulding, R. E. Thurmond, G. A. Komadina, R. W. Hernlund, J. A. Journeay

THE Pima pit is a 1700x1400-St oval, the long T axis parallel to the strike of the orebody. The north side of the pit is carried as a final pit slope that coincides with the footwall of the orebody. The south side and east and west ends of the pit are working slopes continually being stripped back toward the final slopes. An inclined roadway extending from the natural ground surface on a 5 pct grade down to the north- east corner provides access to the pit. This road enters the pit 130 ft below the natural surface and continues as a pit ramp on a 5 pct grade to the 3150-ft bench (roughly the base of the alluvial cover). At this point the ramp system is steepened to a 12 pct grade and continues to the pit bottom. The 5 pct grade is maintained in the alluvial section to facilitate scraper haulage out of the pit. In addition to this main access ramp, temporary working ramps in the alluvial areas allow shorter routes to dumps. These are left on top of working benches and do not change the overall working slopes. Below the base of the alluvium, haulage is by truck down to the skip loading point or up from the bottom to the skip loading point.

Jan 4, 1958