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By Nicholas J. Grant, Rolf Nordheim, Bill C. Gissen

The system Cb-Re was examined in detail utilizing pure metals, careful melting techniques, and heat treatments. Metallographic and X-my methods were utilized for phase identification. In addition to soli-dus determinations, the composition limits and mode of formation of the intermetallic compounds 0 and X were determined. The binary constitution diagram Cb-Re has been the subject of a number of investigations.'-7 Two intermetallic phases have been described and tentative diagrams have been proposed. It was the aim of this investigation to obtain a complete and more accurate diagram through the use of purer alloys and improved techniques. The diagram worked out by Knapton7 was published after our experiments were concluded. EXPERIMENTAL METHODS The starting materials were Re powder of 99.95 pct purity, supplied by the Chase Brass and Copper Co., and Cb rondelles of 99.6 pct purity, supplied by the Electro Metallurgical Co. Typical analyses are given in Table I. The alloys were prepared from compacted Re powder and Cb rondelles; for critical compositions master alloys were used. The melts varied from 5 to 50 g and were melted nonconsumably in an Heraeus arc furnace, under ti-tanium-gettered laboratory grade argon. It was necessary to invert the buttons and remelt, followed by crushing and further remeltings, the number of which depended on the nature of the fracture of the alloy. The weight of each alloy was carefully checked after each melting operation. It was found that significant weight losses occurred only when loose Re powder was used. These losses amounted to a maximum of 3 wt pct in such instances, but did not occur after the charge was once molten, regardless of the composition. Typical loss of weight after 4 remelts, including crushing, amounted to 0.2 wt pct. Oxygen pickup during successive melting cycles did not appear to be significant, since the weight gain values never exceeded 0.01 wt pct. Control of the composition of the alloys was not too difficult; however, the problem of homogenization was a far greater one. Optimum alloying conditions were finally achieved by a combination of master alloy production and repeated crushing and remelting cycles. The compositions of the alloys utilized to determine the constitution diagram are listed in Table 11. The weight balance method was used ultimately as the sole analytical method, since repeated checks indicated that it was accurate to 0.2 pct, if the weight losses were small. The results were double-checked by a metallographic study of the homogeneity of the samples, since this was presumed to be the more serious factor. Heat treatments at 1200°C or lower were accomplished by sealing the specimens in Vycor tubes, heating in a Globar furnace, followed by air cooling. This was considered to be an adequate quench because of the very low temperatures involved. For temperatures of 1200" to 1900°, alloys were homogenized and heat treated in a tungsten-filament resistance vacuum furnace. From 1900" to 2750" a tantalum-tube resistance furnace was utilized, operating under a vacuum of 105 mm Hg. The temperature measurements up to 1200" were made by means of platinum-platinum 10 pct Rh thermocouples, with an accuracy of 5°C. For temperature measurements greater than 1200°C, optical pyrometers were used. They were calibrated against a standardized instru-

Jan 1, 1962

By J. L. Lummus, Frank O. Jones

A field procedure for determining graphite in drilling mud is presented which is sensitive to 0.25 lbs/bbl and accurate to 20 per cent. The method, utilizing oil flotation principles, is suitable for use in all oil-free water base muds except those containing ligno-sulfonates, lignites, or surface active agents. INTRODUCTION Graphite was proposed as a drilling lubricant as early as 1940, but has not been used extensively until recently.' Graphite has been substituted for oil in certain instances to avoid oil contamination of core samples since it has been found that graphite does not interfere with routine core analysis. A major difficulty in using graphite is the lack of a satisfactory field test for determining the concentration of graphite in mud. DISCUSSION OF METHOD Several possible methods of analysis were investigated, including chemical separation, density fractionation, radioactive tracers, and oil flotation.', ' The investigation indicated that a satisfactory method for field use could be devised using an oil flotation method. It was found that toluene shaken with a highly diluted mud sample would recover sufficient graphite to be apparent in the oil phase. Refinement and standardization of technique made it possible to detect as little as 0.25 lb/bbl graphite and to estimate graphite content to within 20 per cent by using a visual comparison method. The test is limited to oil-free water base muds. Quebracho, phosphates, or barytes in the mud do not interfere with the test but surface active agents, lignosulfonates' and lignites evidently render !he graphite water-wet so that none may be recovered by oil-flotation. PROCEDURE Instructions for determining graphite in mud: 1. Collect representative one gallon sample.- from mud stream. Transfer immediately 350 ml portions of this Sample into four mixing cans labeled A, B, C, and D. Measure out graphite (same as is used to treat mud) for addition to cans as indicated below, using either a specially designed scoop or a balance. Can: A B C D Grams Graphite to Add: none 2 4 6 Stir samples B, C, and D for 15 minutes with propeller type mixer. 2. Clean and rinse well with water four 250 ml volumetric flasks labeled I, 11, 111, and IV. Using a 5 ml hypodermic syringe, transfer the indicated volume of samples from cans A, B, C, and D into designated flasks as follows: From Can: A B C D To Flask: I II III IV Add This Volume of Mud: 4 ml 2 ml 2 ml 2 ml 3. Measure 100 ml fresh or distilled water to each flask and swirl gently to disperse mud. Add exactly 5 ml toluene to each flask. Stopper and shake each

Jan 1, 1953

By John Henry Harden

THIS shaft, located in the northern anthracite coal-field about 2300 feet southwest from the court-house at Wilkes-Barre, in the County of Luzerne, Pa., is the property of the Lehigh & Wilkes¬Barre Coal Company. The sinking and cribbing to the solid rock (see isometrical view on Plate IX) was conducted by Mr. W. W. Kenrick. Dimensions of the earthwork to the foundation of the cribbing, 51' 1/2" long, 17' 4" wide, 31' deep. Total amount of excavation in cubic yards, 998.5. Dimensions within the permanent cribbing, 45' 4" long, 11' 16" wide, 31' deep, divided into six parts by buttons 8" x 12", each division, 7' x 11’ 16". Permanent cribbing 12" x 12"; temporary cribbing 8" x 10"; backing, 3" thick ; clay puddle to ex¬clude surface water, 12" thick; conductors, 8" x 5"; lining, 1" thick from top to bottom of the shaft. In the latter part of the year 1871, Mr. H. Murray & Son contracted with the company for the continuation of the work at the rate of four hundred dollars per yard in depth, to a point 100 feet below the Hillman seam of coal ; for the succeeding 100 feet the price to be increased twenty-five dollars per yard, and so on until the bottom rock of the Baltimore seam of coal was found. The size of the shaft within the cribbing timbers is 45 feet 4 inches by 11 feet 6 inches, having six divisions of equal dimensions, 7 feet by 11 feet 6 inches, giving room for what is known as the" standard diamond" car, whose capacity when level full is 100 cubic feet, but when loaded by the miner 4 inches above the top contains 112 cubic feet.* Two divisions of this shaft are intended for hoisting coal from the Hillman seam, two from the Baltimore seam, one for ventilation, and one for pumps. The works were planned for two pairs of direct-acting, high-pres¬sure engines, twenty-four boilers, and for an estimated output of 2500 tons of coal per day from the two seams. * Size of the "standard diamond" car, 8' 10” x 4' 3" x 2’ 8” = 100.1.4; loaded 4 inches above the top, 8' 10” x 4’ 3” x 3’ “0"= 112.7.6.

Jan 1, 1877

By J. Parke Channing

IT IS most appropriate for mining engineers and in fact for all engineers to perfect themselves in leader-ship, because in the last ten years there has been a growing realization on the part of capital that there is no man who by capacity and training is better fitted to manage industrial production than the engineer. In high places the engineer and particularly the mining engineer is coming into his own. We have Mr. Hoover as Secretary of Commerce, Mr. Hammond as the head of the Federal Fact-finding Coal Commission, and Prince Caetani as Royal Italian Ambassador, all mining engineers. More and more in all industries is the manager more likely to be an engineer than he was five, ten, fifteen or twenty years ago, and the feeling is that the man who has had a technical training, whether from going to a school or from burning the midnight oil, makes a better manager than the man who has not. We find in mining that the best man for a manager is one who has had engineering, training. The manager who comes up from the office, from the purchasing department or from the transportation department may make a fairly good manager but he is never so good as the man who has had a technical training. He cannot speak the language of his staff.

Jan 5, 1923

By C. D. King

The key to maximum production of ingots for the war effort is maximum production of pig iron. For any given furnace and ore, the most important single influence on blast-furnace production is the quality of the furnace coke. It is generally recognized that undoubtedly deterioration of Pittsburgh coking coals will be accelerated by the depletion of some of the better grades of coal; the extension of mine mechanization, with attendant degradation of coal quality; and the opening of new coal fields, some of which are known to be fairly high in sulphur. All of these factors, contributing as they do to higher ash and sulphur, and not infrequently to inferior quality of coke, made it desirable that an extensive study be undertaken to determine the effect of washing some coals that closely approximate the average chemical composition of coking coals to be expected in the future by the subsidiaries of the United States Steel Corporation in the Pittsburgh district. It was deemed advisable to study the improvement obtained by washing such coals and to determine the need for supplementing the existing coal washery at Clairton to assure continued maximum production of pig iron from presently available furnace stacks. The existence of the coal washer at Clairton made it possible to conduct tests on coals used in raw form and on similar coals after washing. This study covers an investigation of the quality of furnace coke as affected by washing coking coals, and is confined to the use of 100 per cent high-volatile Pittsburgh seam coals from Fayette and Greene Counties in the manufacture of furnace coke. It covers the improvement in chemical and physical properties as well as the degree of uniformity in coke obtained by washing these coals, and includes the effect on blast-furnace production and practice. The test period involved is four months; the plants, Clairton by-product coke plant, and the Carrie No. 1 and No. 2 blast furnaces at Rankin, Pa., Carnegie-Illinois Steel Corporation. Clairton By-product Coke Plant and Coal Washer The Clairton by-product coke plant carbonizes 30,800 net tons of coal daily in 22 batteries containing 1482 ovens. Three oven sizes are used, commonly known as wide, medium, and narrow types, recognizing the difference in oven width, These ovens also differ in height and length. The present coal washer has a daily capacity of approximately 14,000 tons, leaving 16,800 tans to be carbonized in raw form. Washed coal is used on all the narrow ovens and on three of the six batteries with ovens of medium width, raw coals being used on all the remaining ovens. The customary arrangement is

Jan 1, 1944

By A. S. Odeh

ABSTRACT A simplified model was employed to develop mathematically equations that describe the unsteady-state behavior of naturally fractured reservoirs. The analysis resulted in an equation of flow of radial symmetry whose solution, for the infinite case, is identical in form and function to that describing the unsteady-state behavior of homogeneous reservoirs. Accepting the assumed model, for all practical purposes one cannot distinguish between fractured and homogeneous reservoirs fram pressure build-up and/or drawdown plots. INTRODUCTION The bulk of reservoir engineering research and techniques has been directed toward homogeneous reservoirs, whose physical characteristics, such as porosity and permeability, are considered, on the average, to be constant. However, many prolific reservoirs, especially in the Middle East, are naturally fractured. These reservoirs consist of two distinct elements, namely fractures and matrix, each of which contains its characteristic porosity and permeability. Because of this, the extension of conventional methods of reservoir engineering analysis to fractured reservoirs without mathematical justification could lead to results of uncertain value. The early reported work on artificially and naturally fractured reservoirs consists mainly of papers by Pollard,l Freeman and Natanson,2 and Samara.3 The most familiar method is that of Pollard. A more recent paper by Warren and Root showed how the Pollard method could lead to erroneous results,4 Warren and Root analyzed a plausible two-dimensional model of fractured reservoirs. They concluded that a Horner-type pressure build-up plot of a well producing from a fractured reservoir may be characterized by two parallel linear segments. These segments form the early and the late portions of the build-up plot and are connected by a transitional curve. In our analysis of pressure build-up and drawdown data obtained on several wells from various fractured reservoirs, two parallel straight lines were not observed. In fact, the build-up and drawdown plots were similar in shape to those obtained on homogeneous reservoirs. Fractured reservoirs, due to their complexity, could be represented by various mathematical models, none of which may be completely descriptive and satisfactory for all systems. This is so because the fractures and matrix blocks can be diverse in pattern, size, and geometry not only between one reservoir and another but also within a single reservoir. Therefore, one mathematical model may lead to a satisfactory solution in one case and fail in another. TO understand the behavior of the pressure buildup and drawdown data that were studied, and to explain the shape of the resulting plots, a fractured reservoir model was employed and analyzed mathematically. The model is based on the following assumptions: 1. The matrix blocks act like sources which feed the fractures with fluid; 2. The net fluid movement toward the wellbore obtains only in the fractures; and 3. The fractures' flow capacity and the degree of fracturing of the reservoir are uniform. By the degree of fracturing is meant the fractures' bulk volume per unit reservoir bulk volume. Assumption 3 does not stipulate that either the fractures or the matrix blocks should possess certain size, uniformity, geometric pattern, spacing, or direction. Moreover, this assumption of uniform flow capacity and degree of fracturing should be taken in the same general sense as one accepts uniform permeability and porosity assumptions in a homogeneous reservoir when deriving the unsteady -state fluid flow equation. Thus, the assumption may not be unreasonable, especially if one considers the evidence obtained from examining samples of fractured outcrops and reservoirs. Such samples show that the matrix usually consists of numerous blocks, all of which are small compared to the reservoir dimensions and well spacings. Therefore, the model could be described to represent a "homogeneously" fractured reservoir. la this paper, the fundamental equation of flow

Jan 1, 1966

By A. E. Dwight

Lattice parameters were determined for eighteen equiatornic alloys of the CsCl-type structure, ten of which were previously un-reported. It was found that fomation of the CsCl-type structure in binary alloys of the transition elements is largely dependent on position of the elements in the periodic table. The relative size of the two elements was not found to be a controlling factor. A recent paper by Beck, Darby, and Arora1 corre-lates the occurrence of CsC1-type ordered structures with the position of the constituent elements in the periodic table for the first long period. It was also suggested that a definite increase in relative bond strength between unlike atoms occurred when, in binary alloys of iron-group elements, the other component is changed from a chromium-group element, to a vanadium-group element, to titanium. A later paper by Philip and Beck2 noted that the lattice contraction increased in the order CrFe, VFe, and TiFe. It was also noted by Philip and Beck2 that the lattice contractions of CsC1-type alloys decreased in the order: TiFe, TiCo, and TiNi, which is an apparent reversal of the contractions expected from the position in the periodic table. It was suggested that the increasing lattice contraction is an indication of increased stability, i.e., greater A-Bbond strength. The present investigation was carried out to determine whether the relation of the position in the periodic table to the formation of the CsC1-type structure was also correct for alloys involving the second and third long-period elements. A systematic search was made for CsC1-type structures among equiatomic alloys and for those found, the lattice contraction was determined. EXPERIMENTAL TECHNIQUE The elements Y, Gd, Ti, Zr, Hf, V, and Cb are designated the A group and the elements Mn, Re, Fe, Ru, Os, Co, Rh, Ir, Ni, Pd, Pt, Cu, Ag, and Au are designated the B group. Equiatomic alloys were prepared for 57 AB combinations. The alloys were arc melted in a multicrucible furnace3 in buttons ranging from 5 to 20 g. Chemical analyses were not made, as the charge weights agreed closely with those of the buttons after melting. The alloy buttons were homogenized at 800° to 900°C. Metal log raphic and X-ray specimens were prepared and heat treated at temperatures from 600° to 1200°C. Specimens for X-ray diffraction were usually ground to a powder in an agate mortar; however, needle-shaped solid specimens were used when the alloy was sufficiently ductile to permit their preparation. Diffraction patterns were taken with a Straumanistype Debye-Scherrer camera using filtered Cu or Co radiation. The lattice parameters were obtained in A by plotting the calculated a0 values against the cos29/sin 0 + cos2?/? function and extrapolating linearly to ?= 900. Metallographic control specimens were polished on cloth wheels with diamond paste and etched with various phosphoric and nitric acid reagents. RESULTS The eighteen equiatomic alloys listed in Table I gave evidence of a cubic structure with two atoms in the unit cell, although two of these cubic structures exist only at elevated temperatures and transform to a tetragonal structure on quenching. Nine of these eighteen alloys gave diffraction patterns with super-lattice lines showing that the structure is of the CsC1-type. The lack of superlattice lines in patterns of the other nine alloys may be attributed to the small difference in atomic scattering power of the components. Metallographic study indicates the occurrence of nine narrow single-phase fields at the AB composition. Any or all of these nine may also have a CsC1-type structure. The VFe alloy was found to have a CsCl-type structure by Philip and Beck2 through the use of CrKa radiation (for which the scattering factor of V is anomalously low), whereas the Cu radiation used

Jan 1, 1960

The President was authorized to appoint. a liaison committee to cooperate with the American Mining Congress in respect to the establishment of a Secretary of Mines and Mining in the President's Cabinet. He appointed Edwin Ludlow, Chairman; Mark L. Requa, and J. E. Johnson, Jr. It was resolved to accept the invitation from the Chicago Section to hold the. 120th meeting of the Institute in Chicago during the week of Sept. 15, 1919. The action of the Executive Committee in appointing Past President E. Gybbon Spilsbury as representative of this Institute to visit France upon the invitation of the Société des Ingénieurs Civils de France was approved. The final budget for 1918 was approved. Tile President was authorized to appoint a committee, of three members, on the Development of Activities of the Institute. The committee was requested to report at the February meeting of the Institute, and the Secretary was requested to invite each local section to have a member present to discuss the subject. The President appointed J. W. Richards, Chairman; P. N. Moore, and Allen H. Rogers. Thirty-six members, ten associates, and five junior associates were elected; one change of status from associate to member was granted; two members were reinstated; and eight resignations were accepted. An extension of time was granted to nine members. The dues of eight members were suspended on account of their being in active service.

Jan 2, 1919

Four years of intensive exploration effort by Newmont Mining Corp. came to a climax May 27 in Carlin, Nevada, when local, state and federal officials joined with company officers at a dedication ceremony saluting the start-up of the firm's wholly- owned subsidiary, Carlin Gold Mining Co. Before an invited audience at the plant site, Newmont's President Plato Malozemoff and Nevada's Senator Howard D. Cannon pushed a lever and watched the new melting furnace rotate until a brilliant orange-red stream of molten gold poured forth from its mouth. Two 1200-oz bricks of bullion were poured that day, and they are the first of many that will lift this mine into the number two slot among U.S. gold producers and fourth among all gold mines on the North American continent. Although Newmont Mining first entered the gold mining business in 1929 in Grass Valley, California, the Company had operated no gold producers since the well-known Government directive L208, issued in 1942, closed down U.S. gold mines for the duration of World War 11. Now it is back in the business in a big way-a $10 million mine and plant capable of producing 200,000 oz. per year. All the gold from the Carlin operations is earmarked for the U.S. Mint in San Francisco.

Jan 7, 1965

By C. G. Dunn, J. L. Walter

Secondary recrystallization to the (110) [001] texture in high-purity silicon iron occurs if low-oxygen material is annealed in a nonoxidizing atmosphere. Any departure from these conditions results in a growth of (100) oriented grains. The nature of the matrix and secondary recrystallization structures and textures and the nature of grain boundary interactions during growth show that the low gas-metal interfacial energy of the (110) surfaces provides the driving force for growth of these grains. A type of grain growth, characterized by a driving force which derives from energy differences of {hkl} surfaces at the gas-metal interface, has been treated in recent papers.&apos;-7 Secondary recrystallization to the cube text!:: in high-purity silicon iron provides one example. The present paper also deals with a surface energy driving force but the texture that results by secondary recrystallization is not the cube texture; it is a texture in which the (110) plane is in the plane of rolling and the [001] direction is in the direction of rolling. The phenomenon described in this paper is different from the impurity (dispersed phase)-controlled secondary recrystallization process in which the (110) [001] oriented grains grow under the action of grain boundary driving forces.8-12 It is also different from tertiary recrystallization,2 which also produces the (110) [001] texture in high-purity silicon iron, since the matrix textures and grain sizes are different. Finally, it is unlike any other form of secondary recrystallization reported in the literature. The possibility of obtaining the (110) [001] texture in high-purity silicon iron became clear in a study of the effect of impurity atoms on the energy relationships of (100) and (110) surfaces. In this study Walter and Dunn6 observed the migration of (100)/(110) boundaries, i.e., boundaries between two grains, one of which has a (100) plane and the other a (110) plane, respectively, parallel to the plane of the sheet specimen. At 1200°C the (100)/(110) boundaries advanced into (100) grains in a vacuum anneal, then reversed their direction and migrated into (110) grains in a subsequent anneal in impure argon. Finally, the direction of migration reversed once again with (110) grains growing into (100) grains in a second vacuum anneal. These results were explained in terms of a change in concentration of oxygen atoms at the gas-metal interface during the anneals. Thus, oxygen atoms were added to the surface during the anneals in impure argon to the point where ?100, the specific surface energy of the (100) oriented grains, was lower than ?110, the surface energy of (110) oriented grains. In vacuum, however, the oxygen concentration at the surface was lowered to the point where ?110 < ?100. Concerning the possibility of secondary recrystallization in high-purity silicon iron with a low initial oxygen concentration, the observed effect of adsorbed oxygen atoms has indicated6 that a good vacuum anneal would favor the rapid growth of matrix grains with the (110) plane in the plane of the sheet much more than grains in the (100) orientation. The growth of only (110) oriented grains of course would depend upon y110 being less than ?hkl, where hkl refers to any plane different from (110). The present paper is concerned with the application of the above ideas to secondary recrystallization to the (110) [001] texture in high-purity silicon iron. The matrix and secondary recrystallization textures and structures are defined and discussed. Observations of growth of nuclei for secondary recrystallization and of boundary interactions are included to provide direct information on the surface energy relationships between (110) and other (hkl) surfaces. EXPERIMENTAL PROCEDURE As before, 2,4-6 high-purity iron and silicon were melted and cast in vacuum to provide an alloy containing 3 pct Si with less than 0.005 wt pct impurities. The oxygen content of the ingot was lower than in previous ingots, being approximately 3 ppm (by weight). The carbon content of this ingot may have been slightly higher than was found for previous ingots. The same rolling and annealing schedule used previously2 was followed in this study to obtain samples 0.012 in. (0.3 mm) thick. These samples were electropolished prior to annealing. After rolling and polishing, the oxygen content of the material was approximately 6 ppm; material used in the previous studies contained about twice this amount of oxygen.

Jan 1, 1961

By John A. Wolfe

The brightest spot in the mineral economy of the Philippines is the rapid development of reserves and production of copper. If economic conditions remain favorable and the political climate is encouraging, production of this metal should increase at a rate of 25% per year for several years, long enough to rank this country as one of the major copper producers of the world. In 1965 the copper content of the Philippine concentrates was approximately 60,000 tons. It should total about 80,000 tons in 1966. A major part of the increase will come from the rich Lepanto mine located in Mountain Province, North Central Luzon. Daily production of ore increased from 1200 to 2000 tons at the end of 1965. The ore, consisting of enargite and luzonite, averages nearly 3% copper and 0.12 oz. of gold per ton. It is an under- ground operation with a cutoff grade of 2%. The ore is associated with extensive silicification under a volcanic cap rock. Further expansion to increase ore production to 2600 tpd is under way.

Jan 4, 1967

By W. Fletcher, R. C. Schenk

This paper outlines a well proven and highly successful approach for eliminating many of the problem and frustrations often experienced in plant startup. The approach was developed over a period of 25 years during which some 90 plants were commissioned. This paper discusses startup from the contractor’s perspective and highlights the requirement for detailed startup planning, careful coordination during the project final construction phase, and thorough checkout of equipment prior to introduction of feed. Performance of these services has demonstrated a valuable opportunity exists for the contractor to assist in the training of the plant operating personnel.

Jan 1, 1983

By K. J. Gill, P. Chiotti

Thermal, metallographic, and vapor pressure data were obtained to establish the phase diagram for the thorium-zinc system. Four compounds corresponding to the stoichiometric formulas Th2Zn, ThZn2, ThZn4, and Th2Zn17 were observed. The melting points of these compounds under constrained vapor conditions were found to be 1055°, 1105°, 1095°, and 1015°C respectively. Four eutectics exist in this system with the following eutectic temperatures and thorium contents in wt pct: 1040°C, 89 Th; 945°C, 79 Th; 1045°C, 55 Th; and 995°C, 35 Th. The standard free energy, enthalpy, and entropy of formation for the compounds were determined from zinc vapor pressure data. ALTHOUGH the thorium-zinc system has been of technical importance in the commercial production of thorium and more recently of interest to some proposed pyrometallurgical separations processes for reactor fuels or blanket materials no detailed investigation of this system has been made in the past. Some previous work has been reviewed by Hanssen.1 The existence of four compounds, Th2Zn17, ThZn4, ThZn2, and Th2Zn has been established and their crystal structures have been studied by X-ray diffraction techniques.2-4 Smirnov et al.,5 determined the solubility of thorium in zinc and investigated the two-phase region between Th2Zn17 and zinc. They calculated some of the thermodynamic properties for these alloys from emf measurements on the cells Th/ThCl4(fused)/Cl2, C and Th2Zn17(s)+ Zn (l)/ThCl4(fused)/Cl2, C. The possibility of reaction between the anode materials and fused thorium tetrachloride to produce a lower valent thorium salt would have to be resolved in order to evaluate the results obtained. The purpose of the present investigation is to establish the phase diagram for the thorium-zinc system and to determine the standard free energy, enthalpy, and entropy of formation for the compounds formed. MATERIALS AND EXPERIMENTAL PROCEDURES The metals employed in the preparation of alloys were Bunker Hill slab zinc, 99.99 pct pure and Ames Lalboratory thorium. The major impurities in the thorium in parts per million were oxygen—1250, carbon—350, iron—100, nitrogen-75, and silicon—70. The metals in the form of several small massive pieces, total weight about 30 g, were cleaned with acid, washed with water and acetone, and dried just prior to their use for alloy preparation. Tantalum containers were found to be suitable for all the alloys studied and were used throughout this investigation. The metals were sealed inside of small 1-in.-diam tantalum crucibles by welding on preformed tantalum covers. A small thermocouple well made from 1/8-in.-diam tantalum tubing was welded in the bottom of each crucible. These crucibles were enclosed in

Jan 1, 1962

By Oscar Aguilar, Jin-ichi Nakamura, Kazuo Nakamura

Michiquillay deposit is porphyry copper type with over 500 million tons of 0.70% Cu ore. In order to reflect the true nature of the orebody as to shape and value distribution and to compensate for insufficient data in the outer perimeter and lower portion, "Triangular Interpolation Method" is used in the computation, wherein block grades are determined by "Plane Regression". The results thus obtained are compared with the data from additional investigation and are found to match well with each other. Aside of total Cu, assays of Fe, S, MoS2 and geological data such as alterations, rock types, faults, mineralogical compositions are also stored in the computer. By this system, their distribution and relation to alteration become clear and classification of ore into types affecting recovery become possible. The outputs of this system make clear the characteristics of the orebody and furnish valuable geological data on the orebody.

Jan 1, 1976

By Henry Louis

The present paper comprises two parts, which have but little con nection with each other. The first is practically a continuation of the investigation on the specific gravity of gold liberated by means of acids from its alloys, on which subject I had the honor of presenting a brief note to the Institute in 1893 (Trans., xxii., 117); the second describes a series of experiments on the behavior of amorphous gold under the action of heat. By the kindness of Prof. W. C. Roberts-Austen, C.B., F.R.S., chemist to the Royal Mint, I have been permitted to conduct these experiments in the mint assay-office, all the splendid resources of which institution were freely placed at my disposal. I desire at the outset to express my obligations to Prof. Roberts-Austen and to all the staff of the mint assay-office for their kind assistance at all stages of my experiments. I.—The Specific Gravity of Gold in Alloys. In the note already referred to, I published my discovery that the gold left on dissolving out the silver from an alloy of gold and silver (which I shall here term " residual amorphous gold ") has a higher specific gravity than ordinary molten gold. I also pointed out that I had worked on very small quantities of material and with balances and weights on the accuracy of which I could not implicitly rely, and that in so far my numerical data were inconclusive, although I had but little doubt of the correctness of the broad fact. The experiments now recorded were made under the most advantageous circumstances and with various refinements. Particularly, I found that the ordinary form of specific-gravity bottle was responsible for several small sources of error; and I accordingly designed an improved form (Journal of the Society of Chemical Industry, April 30,1894, vol. xiii., p. 322), which I used for this work. The

Jan 1, 1895

By O. B. Bennett

WHEN diamond drilling was introduced in the Rhokana mines in 1939 it was used principally for pillar removal and for completion of the upper portions of shrinkage stopes which were being affected by increasing pressure. This method of drilling long blastholes proved so successful that it was extended gradually to cover stoping, pillar recovery, and hanging cave work. BY 1949 virtually all the ~roduction of Mindola and Nkana was being obtained by this method. At the present time 87,500 ft are drilled each month by the 80 diamond drills in daily operation. Responsibility for control and issue of diamond drilling equipment and crowns, as well as tabulation of all performance figures, was taken over by a sPecially formed Roto drill department, which also investigated the problems encountered with this new method. To assist this department a fully equipped test chamber, Fig. 1, was established underground where performances of various types of machines and equipment could be studied under conditions as nearly uniform as possible. Since the establishment of this department, which was eventually taken over and incorporated into the study department, considerable experimental work has been done on every aspect of the subject. The problems can be classified broadly under four headings: improvement of drilling equipment, crown design, machines, and stoping layouts. One of the major problems with drilling equipment has been to eliminate vibration. Owing to flexing of rods in the hole, severe friction is set up on the back end of the &apos;Ore barrel and On any high spots in the rods, inducing harmonic vibration in the string of rods and causing the crown to chatter against the face. This not only causes premature crown failure but also reduces penetration speeds and increases wear on the machines and rods used. In the early days, when only holes of EX size were drilled, vibration was largely overcome by periodic greasing of rods and core barrel during each run, but with the change-over to the larger BX hole it became obvious that application of grease by hand was inefficient and time-consuming, and attempts were made to perfect a self-lubricating core barrel. A series of these core barrels was made up and tested and a number of the latest type were used under normal operating conditions, but although footages up to 120 ft were drilled without refilling the overall performance was inconsistent, and the idea was shelved in view of the success of the stabilizer rods referred to later in this paper. At the same time tests were made with barrels 5 ft and later 6 ft long instead of the normal 2 ft. Although a slight improvement was noticed, greasing was still necessary. It was found that rod vibration increased as the core barrel became worn, and in an early test chamber experiment crowns drilled with a worn core barrel averaged 95 ft with a diamond loss of 4.76 carats, whereas the same type of crowns with a new barrel averaged 228 ft with a diamond loss of 3.13 carats. until then all BX drilling had been done with B-sized rods, but during a test on a string of BX-sized rods it was noticed that vibration was negligible. Because of the larger surface area of metal bearing on the sides of the hole, however, the friction and resistance of rods of this size rendered them impracticable on any but the most powerful of the machines, The use of stabilizers spaced evenly along the rods was the next logical step, and for this B couplings, see Fig. 2, were set with three tungsten carbide inserts 1 in. long placed around the periphery equidistantly and at an angle of 45" with a right hand lead. These were placed immediately behind the core barrel and then at 12-ft intervals, as it was found that vibration still occurred when the stabilizers were more than 15 ft apart. The effect of these stabilizers was immediately noticeable; holes were drilled with a minimum of vibration, penetration speeds were improved, and as it was no longer necessary to grease the rods there was a marked decrease in the overall drilling time for each hole. While tests were being made with the stabilizer comeb periodic were taking place with a set of tapered threaded rods, and because there was marked improvement in efficiency it was decided to incorporate the stabilizers and tapered threading in all new rods ordered for Rhokana. The feature of these rods is that only four full turns are required to tighten the coupling as against nine for the present type of B rods. Also, as they are self-centering it is virtually impossible to crossthread them. Each rod has a male 5" tapered Acme thread, Fig, 3, on one end and a female at the other, so that separate couplings are unnecessary, and every fifth rod has an

Jan 1, 1955

Reserves of Lake Superior Manganiferous Iron Ores Appendix-to paper by CARL ZAPFFE, presented at the Cleveland Meeting and issued, as Pamphlet No. 1664-C, with MINING AND METALLURGY, May, 1927. The author has made several test calculations on the basis of manganese units that are and may be involved. These are summarized here, as follows: On the basis of ore produced since 1902, the average annual requirement of manganese units would be 91,000 tons; on the basis of the production of the past 4 or 5 years, the requirement would be 145,000 tons; on the basis of the year 1925 only, the requirement would be 155,000 tons. Applying the percentage of increase for the next 20 years, as determined by the plat shown as Fig. 3, the total requirement for 20 years would be 3,254,000 to 3,572,000 tons, if one used either 145,000 tons or 155,000 tons as a starting point. We have the following interesting facts: (1) The Michigan ores contain only about one-half the manganese that the Cuyuna ore does; (2) the Cuyuna district now produces one-half the ore required; (3) the known tonnage of Cuyuna reserves is greater than the speculative reserve offered for Michigan; (4) at the present rate of production considered by tons, Michigan ores ought to be exhausted in 20 years or less while Cuyuna ores would last up to 44 years. We should now reduce the estimated reserves to a basis of manganese, units. Using the tonnages listed in Combinations 2, 3, 4 and 5, the reserves total 3,118,225 tons, 3,781,995 tons, 5,041,995 tons and 5,560,342 tons, respectively. Therefore, needing at the most 3,572,000 tons, as shown above, either Combination 3, 4 or 5 will not only suffice but will leave an excess-enough for 1 or 2 years for Combination 3, 9 or 10 years for Combination 4, and 12&apos; to 13 years for Combination 5. This means that for Combination 5, without using tonnages obtained from beneficiating or from mixing poor ores, the reserves are ample for about 33 years.

Jan 5, 1927

By D. W. Jessup

ONLY during the past few years, have the old-fashioned methods for treating ores and byproducts progressed to any marked degree. The advent of multiple-hearth roasting, blast roasting, the baghouse, and the Cottrell plant indicate this change. It is the ability to roast and smelt sulfide ores and concentrates that gave this later advance. Further investigations that will solve the problem of efficient handling of flotation concentrates and their attendant ills are of the utmost importance. With future developments, the metallurgist will be able to smelt more economically and to treat still more refractory ores, making the early methods appear small by comparison. GROWTH OF SMELTING Well within the memory of pioneer smeltermen, the ore was charged directly into the blast furnace as it came from the mine;, only the richer, ores were smelted, the lower grade and sulfide ores were undesirable. Matte was, at first, thrown away and only two products) were desired-slag and lead bullion. Scores of the early smelting companies failed because of improper reduction of ores, high costs, short- furnace campaigns, inability to treat sulfide ores, and ores containing-zinc. With the failure of weaker companies, consolidation eliminated many others, until today the field is dominated by a comparatively few organizations, each strong financially. Formerly, a few thousand dollars would erect a small plant, whereas the modern smelter costs over a million dollars. Ore and products to which formerly no attention was paid are now treated with comparative ease and at a good profit. Mechanical improvements have modernized smelting, furnace tonnages show an increase, and metal recoveries have made enormous strides. Even though costs have decreased, they have not dropped in proportion to the improvements, because of the higher wages, greater costs of fluxes and coke, more undesirable ores smelted, etc.

Jan 8, 1925

By W. L. Remick

WHEN the senior author of this paper presented an article on Fine Coal Cleaning by the Hydrotator Process,1 at the February, 1927, meeting, that process had been developed in the anthracite region only to clean what is known as "slush" or "silt." This size constitutes all of the material passing the finest screen used at any particular colliery. Some use a 3/32 -in. round-hole screen and others use a 1/16 -in. round hole. Practice varies throughout the anthracite region, but the tendency seems to be toward the finer screen.

Jan 1, 1929

By A. T. Tye

Regarding the results of preferential flotation at Cananea, Weinig has stated that "The concentrating mill of the Cananea Consolidated Copper Company at Cananea, Mexico, furnishes an excellent example of modern flotation practice. The complete figures are given because they present so forcibly the advantages of modern selective flotation."l L. 0. Howard notes that "Perhaps the outstanding example so far published of improverrients in results through the use of selective flotation is that at Cananea, Mexico."" The following notes indicate the conditions under which the results were obtained at Cananea; also the effect of preferential flotation at the concentrator and at the smelter. In the majority of cases the full benefit of selective flotation is not fully realized, as the concentrates are shipped from the mill to smelters owned by other companies. The saving made at a certain concentrator by this method of flotation may be well understood but the saving resulting at the smelter from the results of preferential flotation at the mill is not so apparent, as the data are contained in the smelting records of some other company. Even at the smelter the great gains following preferential flotation may not be noted, as a variety of ores and concentrates from various sources are received and the difference from the changed concentrator operations of a certain mill may be obscured. As noted by Professor Weinig and Professor Howard, the results obtained at Cananea brought about a veritable metallurgical revolution at the smelter. There are probably plants at which a better preferential separation is being made than at Cananea. Many ores contain barren iron pyrite which, when depressed, does not carry copper into the tailing. Some concentrators use clean uncontaminated water for milling and in others the ore is ground much finer than at Cananea. All of these items are concerned in perfect separation. The details of preferential flotation at Cananea, therefore, are not presented on account of the perfection of the

Jan 1, 1928