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By John Chipman, Raymond L. Orr

Recently reported high -temperatlnre thermal data have been incorporated in a reuieu' of the therinody-r~ar,zic propesties of iron. The recent data permit mot corlsisterlt and reliuble choices to be made for many of the pvoperties, resnlling in bettev eslablished tabrtlations of the thermodynamic functions. Values of Cp, h°t- - H°298. s°t - S2sa, and (G°T -H298/T are tabulated for -the solid and liquid phases of iron in both stable and melastable regions. Consistency of the tabulations has been maintained to the precision necessr~vg to yield the Gibbs energy change belween the bcc (a, 8) and fcc (y) phases of iron at temperatures pertinenl to alloy studies. THE importance of the metal iron has stimulated many investigations of its thermodynamic properties at elevated temperatures. Reviews and summaries of the data have appeared from time to time. Among these the 1951 paper of Darken and smith' deserves highest recognition for the quality of its criticism and for the precision and thermodynamic consistency of its tabulations. These and the more recent table of Kaufman, Clougherty, and weiss2 reproduce the Gibbs energy of the (a, 6)--Y transformation with sufficient sensitivity for alloy studies. Other recent compilations of equal validity3-' make no attempt to do this. Since this Gibbs energy change is not greater than 22 cal per g-atom over the temperature range 1100° to 1800°K, it is obviously desirable to tabulate values of (G°t - H298)/T to four places of decimals even though the absolute accuracy would not warrant such precision. The necessity for reconsideration of the problem is emphasized by the recent publication of several important researches on the enthalpy and heat capacity of iron. The older work is well-covered in the reviews cited and will be mentioned here only when actually employed in the tabulations. GAMMA IRON Transition temperatures are taken from the selec -tions of Elliott and Gleiser,5 based on the work of Boulanger7 and earlier workers. For the a-y (A3) transition at 1184°K we retain the value 215 cal per g-atom adopted by Darken and smith.' It has received confirmation in the recent work of Dench and Kubaschewski,' and of Braun8 using adiabatic calori-metry, and of Wallace, Sidles, and Danielson10 using a pulse heating technique. The enthalpy data of Olette and ferrier" and of Anderson and Hultgren12 at higher and lower temperatures may be extended to 1184°K to give the following values which are adjusted to the required value of ? 1184 Ha,1184 Ha 298 = 8030 cal per g-atom [11 hr,1184-ha298 = 8245 cal per g-atom [2]

Jan 1, 1968

By C. F. Floe, N. J. Grant, A. Joukainen

A CCORDING to Guertler,¹ Smith and Hamilton were the first to study the Cu-Ti alloy system, but because of the presence of large amounts of impurities their data are inconclusive. Hensel and Larsen² in 1930 and Kroll³ in 1931 determined the high copper end of the diagram. In 1939 Laves and Wallbaum,4 in a short summary of compound determinations in titanium binary alloys, mentioned the existence of a face-centered cubic compound, Ti2Cu, with 96 atoms per unit cell, and also a compound, TiCu3, which was said to have a deformed hexagonal close-packed structure in which ordering had been observed. The existence of a compound, TiCu, was also found but the crystal structure was not determined. Craighead, Simmons, and Eastwood determined the effect of copper on the allotropic transformation in titanium. The investigation was limited to alloys containing from 0 to 5 pct Cu, using titanium of fairly high impurity content. Experimental Procedure The methods used for determination of the diagram included metallographic examination and X-ray and thermal analyses. The system was first worked out using the higher purity "Process A" sponge titanium and oxygen-free copper. The high titanium region was next determined more accurately using iodide titanium and vacuum-melted, oxygen-free copper. Preparation of Alloys: The alloys were prepared by arc melting under prepurified helium atmos- pheres in water-cooled copper crucibles. The high purity iodide titanium alloys were prepared by melting three compositions along with a specimen of unalloyed titanium in a single charge. A multiple crucible was used which consisted of a copper plate with four cup-like depressions. The unalloyed titanium was melted first and acted as a getter for any oxygen and nitrogen in the melting chamber. Since these elements harden titanium, the hardness of the standard specimen was an index of the highest probable contamination of the alloy melts. The sponge titanium alloy melts weighed from 100 to 200 g each and the iodide titanium melts from 20 to 25 g. In order to insure greater homogeneity most of the alloys were melted, turned, and remelted twice. The larger buttons made from sponge titanium were sectioned or crushed before remelting. These alloys up to 30 pct Cu were further homogenized by forging at 925°C. They forged readily up to 20 pct Cu but numerous cracks were formed in the alloys containing from 20 to 30 pct. Above 30 pct Cu the sponge titanium alloys were homogenized by heating for 200 hr at 850°C under prepurified helium. Similarly, the iodide-grade titanium alloys containing up to 3 pct Cu were homogenized at 1000°C for 24 hr and all others at 880°C for 200 hr.

Jan 1, 1953

By A. T. Beckwith

This paper concerns measurements of low-velocity air currents and investigations on mine ventilation by means of chemical smoke. The chemical smoke used is produced without flame and at ordinary mine temperatures, and therefore presents no ignition hazard in gaseous mines. Mining men who have worked in open-light mines are familiar with the use of tobacco smoke in determining the direction of air travel and approximately measuring low-velocity air currents. The use of chemical smoke is merely a refinement on the use of tobacco smoke and the flameless low-temperature source of chemical smoke has the additional advantage that it can be used in gaseous mines. Ventilation smoke apparatus, Fig I, is a simple device for producing dense white smoke. The apparatus as shown consists of a rubber aspirator bulb which is used to force air through a glass "smoke tube" containing a chemical substance that produces a dense white smoke on contact with air. The apparatus was developed by engineers of the U. S. Bureau of Mines.' It is inexpensive, weighs less than 3 oz, is easily carried in a pocket, and requires no skill to operate. Smoke apparatus is invaluable in practical mine ventilation work for: (I) measuring low-velocity air currents, (2) detecting air leakage, (3) observing airflow characteristics. This paper is written to show the merits of the apparatus, and is based upon findings through its extensive use, during the past six years, in mines operated by the Lehigh Navigation Coal Co. Inc. The Apparatus Certain chemicals, such as stannic chloride (SnCl4) and pyrosulphuric acid (H2S2O7) produce "white smoke" when exposed to air. Stannic chloride gives off vapors which on contact with moist air cause the formation of solid hydrates, of which SnCl4.5H20 is the most important. Pyrosulphuric acid decomposes when exposed to air with the evolution of sulphur trioxide (SO3) fumes, which become white upon contacting moisture in the air. A chemical, such as one of the above, is the active substance in a smoke tube. The glass tubes are hermetically sealed during manufacture, and are made ready for use by merely breaking off the tips at the conveniently located file cuts. When the smoke tube is connected to the aspirator, depressing the aspirator bulb forces air into contact with the chemical, and smoke discharges from the open end of the tube. A white solid (a hydrate if stannic chloride is used, or sulphur trioxide if pyrosulphuric acid is used) may occasionally form a cake in the outlet opening of a smoke tube. A small wire such as an ordinary paper clip can be conveniently employed to dislodge the cake should it seal the tube. When the assembled apparatus is not

Jan 1, 1949

By J. R. Long, C. E. Armantrout, A. H. Roberson, T. R. Graham

The preparation and fabrication of copper-manganese-zinc alloys and the evaluation of their engineering properties have for some time been an integral part of a research program of the Federal Bureau of Mines. In this project, the use of electrolytic manganese, along with high purity copper and zinc, has permitted the preparation of alloys which contain a minimum of impurities which are capable of producing detectable effects in the properties studied. The initial work on the physical properties of wrought alpha solid solution alloys and a general outline of the alpha solid solution area for alloys containing up to 40 pct manganese have been presented in previous reports. 1,2,3,4,5 The evidence accumulated in these investigations clearly showed that additions of manganese to the copper-zinc alloys improve the strength properties of the brasses without seriously reducing their characteristic ductility and that considerable quantities of manganese could be added to the copper-zinc alloys without exceeding the mutual solubilities of manganese and zinc in copper. In continuation of these studies, additional data have been obtained in a more extensive investigation of the phase relationships of the copper-manganese-zinc ternary system. The present report delineates the solid phase boundaries of the ternary system from the copper-zinc binary to the manganese-zinc binary for alloys containing up to 50 pct zinc. Summary The alloys used in this investigation were conditioned by working schedules designed to homogenize the structures prior to heat treatments required for the equilibrium phase sludy. The treated alloys were examined by metallographic methods including hardness and X ray data for confirmation. The composite data of the investigation are summarily presented in a series of isothermal sections of the ternary system at 1100, 1200, 1300, 1200, and 1500°F. The salient features of the system as outlined by these data are: (1) the continuous alpha solid solution which extends across the diagram to the gamma-manganese phase of the manganese-zinc system at temperatures of 1200°F and higher, (2) the beta phase of the copper-zinc system similarly forms a continuous solid solution with its counterpart the beta field of the manganese-zinc binary, (3) at 1100°F (593°C) these alpha and beta fields are greatly restricted by complications arising from the allotropic transitions of manganese and the decreasing solubility of manganese in the alpha solid solution. The alpha + beta fields intersect the alpha + alpha manganese field to produce the beta + alpha manganese field and two adjacent three phase fields consisting of alpha + beta + alpha manganese and gamma + beta + alpha manganese. Typical microstructures representative of the various phase areas at each temperature are also presented. Review of Previous Work The copper-manganese-zinc alloys have periodically been the subject of numerous investigatioris of both prac-

Jan 1, 1950

By Benjamin Lustman, Robert F. Mehl

The study of the high-temperature oxidation of pure metals, intensively pursued experimentally since the pioneer work of Pilling and Bedworth1 and supplemented by the recent theoretical work of Wagner2 is now well advanced. The important field of low-temperature oxidation—i.e., oxidation extending only through the temper color range of thicknesses with the formation of thin oxide layers in the range from 0 to 1000-2000 ?ngstroms—has not, however, reached so satisfactory a state, owing largely to the experimental difficulties that attend the formation and study of thin films; most of the work has been performed on abraded specimens whose surface irregularities approximated the thickness of the films formed. It has been pointed out that the rates of oxidation of different crystal faces are different and that the difierence may be related to the orientation relationships that obtain between the metal crystal and the superimposed oxide crystal; it has also been suggested that this orientation dependence should manifest itself in rates of oxidation and perhaps also of corrosion in aggregates bearing preferred orientations.3'4 An attempt to measure the rates of oxidation of different crystal faces of alpha iron demonstrated this difference in rate,4 but the dependence of rate on orientation was not simple; the method of measuring thickness, consisting of a color comparison with barium stearate multi-films of known thickness, was not wholly satisfactory. For these reasons it was felt that a study of the rate of oxidation of single crystals prepared with surfaces as smooth as possible, using a sensitive method giving true film thickness, should be profitable. Copper was chosen because of the ease of reduction of the primary oxide film and because the orientation relationships between metal and oxide and the composition of the oxide films formed at low temperatures are known; for thin films formed at low temperatures, Cu2O alone occurs.22-24 Experimental Procedure Various methods of surface preparation were attempted in order to obtain a surface that would oxidize uniformly and repro-ducibly. Electrolytic polishing by the method of Jacquet5 and electrolytic and mechanical polishing followed by annealing in vacuum and in purified hydrogen at various temperatures were tested. The procedure finally adopted consisted in a fine metallographic polish followed by an anneal in purified hydrogen at temperatures of 900° to 1000°C. for 12 hr. or more. The surface so produced was very bright and smooth, with no trace of specular etching (though some grain-boundary etching was observed); indeed, fine scratches produced by the final polish were entirely removed by the high-temperature anneal. After this treatment, the specimens were cooled in hydrogen to room temperature and placed in the oxidation furnace, where

Jan 1, 1941

By D. H. Woodard, G. W. Geil, W. D. Jenkins

In a series of papers, the authors and their associates have shown the influence of four important factors on the technical cohesion limit.3-l6 By "technical cohesion limitv is meant the technically determin-able resistance to fracture. The four importarlt factors are the stress system, plastic deformation, temperature, and the strain rate, The influence of any one of these factors may be represented by a curve of cohesion limits, although only one point on the curve may represent actual fracture. As the term "fracture stress" has come into general use with the same significance as technical cohesion limit, the authors are here using the shorter term with the understanding that it does not always refer to actual fracture. The previously mentioned series of papers contains ample evidence that plastic extension increases the fracture stress of any polycrystalline metal, In a recent paper of the series,l5 moreover, it has been shown that the effect of plastic deformation on the fracture stress cannot be attributed to reorientation of internal flaws but is a work-strengthening effect, similar to the effect on the flow stress, $ Plastic deformation of steels, however, may not only cause work hardening, but may result in additional hardness because of strain aging. Some evidence that strain aging affects not only the flow stress but also the fracture stress, has been presented in a recent paper.14 In that paper, a study is made of the influence Of Plastic deformstion at room temperature on the fracture stress of ferritic steels at -188°C. The study is based on the results Of two-stage tests. In the first stage each specimen was extended a predetermined amount at room temperature; in the second stage the specimen was tested to fracture in liquid air. In most Of these tests the specimen remained in the testing machine during the interval of about one hour between the first and second stages, it was kept under a tensile load of about 500 lb and maintained at liquid air temperature. The effects of various amounts of .plastic deformation on the fracture stress were thus determined and curves were obtained to represent the influence of plastic deformation at room temperature on the stress at brittle fracture in liquid air.14 Two specimens Of ingot iron, however, were removed from the testing machine at the end of the first stage and several days elapsed before the specimens were tested to fracture in liquid air, As shown in Fig 8 of that paper," the points representing fracture of these specimens are far above the curve representing the results of two-stage tests without a rest interval between the stages. The evidence thus indicated that the fracture stresses of these two specimens were increased not

Jan 1, 1949

By G. W. Geil, D. H. Woodard, W. D. Jenkins

In a series of papers, the authors and their associates have shown the influence of four important factors on the technical cohesion limit.3-l6 By "technical cohesion limitv is meant the technically determin-able resistance to fracture. The four importarlt factors are the stress system, plastic deformation, temperature, and the strain rate, The influence of any one of these factors may be represented by a curve of cohesion limits, although only one point on the curve may represent actual fracture. As the term "fracture stress" has come into general use with the same significance as technical cohesion limit, the authors are here using the shorter term with the understanding that it does not always refer to actual fracture. The previously mentioned series of papers contains ample evidence that plastic extension increases the fracture stress of any polycrystalline metal, In a recent paper of the series,l5 moreover, it has been shown that the effect of plastic deformation on the fracture stress cannot be attributed to reorientation of internal flaws but is a work-strengthening effect, similar to the effect on the flow stress, $ Plastic deformation of steels, however, may not only cause work hardening, but may result in additional hardness because of strain aging. Some evidence that strain aging affects not only the flow stress but also the fracture stress, has been presented in a recent paper.14 In that paper, a study is made of the influence Of Plastic deformstion at room temperature on the fracture stress of ferritic steels at -188°C. The study is based on the results Of two-stage tests. In the first stage each specimen was extended a predetermined amount at room temperature; in the second stage the specimen was tested to fracture in liquid air. In most Of these tests the specimen remained in the testing machine during the interval of about one hour between the first and second stages, it was kept under a tensile load of about 500 lb and maintained at liquid air temperature. The effects of various amounts of .plastic deformation on the fracture stress were thus determined and curves were obtained to represent the influence of plastic deformation at room temperature on the stress at brittle fracture in liquid air.14 Two specimens Of ingot iron, however, were removed from the testing machine at the end of the first stage and several days elapsed before the specimens were tested to fracture in liquid air, As shown in Fig 8 of that paper," the points representing fracture of these specimens are far above the curve representing the results of two-stage tests without a rest interval between the stages. The evidence thus indicated that the fracture stresses of these two specimens were increased not

Jan 1, 1949

By Benjamin Lustman, Robert F. Mehl

The study of the high-temperature oxidation of pure metals, intensively pursued experimentally since the pioneer work of Pilling and Bedworth1 and supplemented by the recent theoretical work of Wagner2 is now well advanced. The important field of low-temperature oxidation—i.e., oxidation extending only through the temper color range of thicknesses with the formation of thin oxide layers in the range from 0 to 1000-2000 ?ngstroms—has not, however, reached so satisfactory a state, owing largely to the experimental difficulties that attend the formation and study of thin films; most of the work has been performed on abraded specimens whose surface irregularities approximated the thickness of the films formed. It has been pointed out that the rates of oxidation of different crystal faces are different and that the difierence may be related to the orientation relationships that obtain between the metal crystal and the superimposed oxide crystal; it has also been suggested that this orientation dependence should manifest itself in rates of oxidation and perhaps also of corrosion in aggregates bearing preferred orientations.3'4 An attempt to measure the rates of oxidation of different crystal faces of alpha iron demonstrated this difference in rate,4 but the dependence of rate on orientation was not simple; the method of measuring thickness, consisting of a color comparison with barium stearate multi-films of known thickness, was not wholly satisfactory. For these reasons it was felt that a study of the rate of oxidation of single crystals prepared with surfaces as smooth as possible, using a sensitive method giving true film thickness, should be profitable. Copper was chosen because of the ease of reduction of the primary oxide film and because the orientation relationships between metal and oxide and the composition of the oxide films formed at low temperatures are known; for thin films formed at low temperatures, Cu2O alone occurs.22-24 Experimental Procedure Various methods of surface preparation were attempted in order to obtain a surface that would oxidize uniformly and repro-ducibly. Electrolytic polishing by the method of Jacquet5 and electrolytic and mechanical polishing followed by annealing in vacuum and in purified hydrogen at various temperatures were tested. The procedure finally adopted consisted in a fine metallographic polish followed by an anneal in purified hydrogen at temperatures of 900° to 1000°C. for 12 hr. or more. The surface so produced was very bright and smooth, with no trace of specular etching (though some grain-boundary etching was observed); indeed, fine scratches produced by the final polish were entirely removed by the high-temperature anneal. After this treatment, the specimens were cooled in hydrogen to room temperature and placed in the oxidation furnace, where

Jan 1, 1941

By J. Doig, A. A. Bauer, F. A. Rough

AN investigation of the phase relationships between the 6 phases of the uranium-molybdenum and uranium-titanium systems was conducted by studying alloys ranging in composition from uranium-31.5 at. pct Mo to uranium-34 at. pct Ti. A ternary cut showing these phase relationships is presented upon the basis of thermal-analysis data obtained and upon the basis of metallographic and X-ray examination of heat-treated and quenched specimens. Isothermal ternary sections, inferred from phase relationships determined between the two 6 phases, are presented for the uranium-molybdenum-titanium system. In the compositional range investigated, a y-ura-nium solid solution exists at elevated temperatures. At about 900°C the 6-U2Ti phase, with limited molybdenum solubility, forms. Epsilon, a molybdenum-titanium-rich solid solution appears at the same temperature and appears in all phase regions that exist below this temperature except in the terminal 6-phase regions. The y phase is stable to approximately 660°C in titanium-rich alloys; the y and 6-U,Ti phases decompose at this temperature to a uranium in a peritectoid reaction. The y phase in molybdenum-rich alloys of the ternary cut studied decomposes in an apparent eutectoid reaction, at about 593°C, to the 6-uranium-molyb-denum and a-uranium phases. The e phase while present, presumably plays a negligible role in these reactions. Titanium was found to markedly decrease the stability of retained y in these alloys. Age-hardening characteristics developed in titanium-contain ing uranium-molybdenum alloys quenched from 900°C and aged at 500°C are ascribed to decomposition of Y to the a and E phases. INTRODUCTION Alloys of uranium with molybdenum, titanium, and zirconium, respectively, are of great interest because of their potential use as reactor fuels. As a result, this interest extends to the constitution of these alloys, particularly in view of their uniqueness of constitution. In these systems, alone, among known uranium-alloy systems, an intermediate phase forms by direct decomposition of the high-temperature y-uranium phase. As a prelude to possible future alloy development, an investigation has been undertaken of the phase relationships that occur between the intermediate phases of these systems. These phase relationships have already been reported for the intermediate 6 phases of the uranium-zirconium and uranium-titanium systems.' The results of an investigation of the phase relationships between the 6 phases of the uranium-molybdenum and uranium-titanium systems are reported here. The phase relationships that occur in the binary systems pertinent to this investigation are shown in Fig. 1. Phase nomenclature appearing throughout this report is as shown in these diagrams; the crystal structure of the phase is given in Table I. EXPERIMENTAL PROCEDURE Alloys were cast and fabricated to flat plate. Specimens were heat treated and examined metal-

Jan 1, 1959

By N. P. Peterson

The Castle Dome copper deposit is of the porphyry type and occurs in a body of quartz monzonite intruded into the pre-Cambrian formations and possibly into the lower part of Paleozoic limestones. The hypogene sulphide minerals include pyrite, chalcopyrite, and a little molybdenite, sphalerite, and galena. Pyrite and chalcopyrite are disseminated or occur in small closely spaced veins. The hypogene mineralization shows distinct zoning in the distribution of pyrite and chalcopyrite, so that the mineralized area can be subdivided into pyritc and chalcopyrite zones. Supergene enrichment, though not extensive, played an important part in the formation of the ore body, the chalcopyritc zone being the pro-tore. Several thin, gently dipping, diabase sills occur in the mineralized area, and the copper content of the rock in and near these sills is substantially higher than average. The formation of the ore body was caused by the combined effects of three forces: I. Zoning of the hypogene mineralization, 2. Richer copper metallization associated with the diabase sills, and 3. Supergene enrichment. Introduction The Castle Dome copper deposit is about 5 miles west of Miami, Arizona, and about 3 miles west of the Miami-Inspiration ore body. The Castle Dome open-pit mine is on the south flank of Porphyry Mountain, a prominent landmark of the region. In 1941 the deposit was being explored by the Miami Copper Co., but the exploration program was cut short when the decision was made to develop the property as a war project. The Castle Dome Copper Co., Inc., was organized as a wholly owned subsidiary of the Miami Copper Co., and arrangements were made with Defense Plant Corp. for developing and equipping the mine and for construction of a concentrating plant. Construction and development work was begun in January 1942, and milling operations started June so, 1943. Since that date, the mine has been in continuous operation by the Castle Dome Copper Co, under a lease agreement with Defense Plant Corp. producing copper at an average rate of 4,000,000 lb per month. To January I, 1946 the total production amounted to 121,088,000 lb of copper,4,271 oz of gold, and 211,755 oz of silver. A detailed study of the copper deposit and the surrounding area was undertaken by the U. S. Geological Survey in June 1943. Other members of the Survey participating in the work were C. M. Gilbert, G. L. Quick, and J. P. Albers. The results of this work will be published as a comprehensive report by the Geological Survey. General Geology The Globe quadrangle, which includes the Castle Dome area, was studied by F. L. Ransome1 of the Geological Survey in 1901 and 1902. His maps and descriptions were

Jan 1, 1949

By N. P. Peterson

The Castle Dome copper deposit is of the porphyry type and occurs in a body of quartz monzonite intruded into the pre-Cambrian formations and possibly into the lower part of Paleozoic limestones. The hypogene sulphide minerals include pyrite, chalcopyrite, and a little molybdenite, sphalerite, and galena. Pyrite and chalcopyrite are disseminated or occur in small closely spaced veins. The hypogene mineralization shows distinct zoning in the distribution of pyrite and chalcopyrite, so that the mineralized area can be subdivided into pyritc and chalcopyrite zones. Supergene enrichment, though not extensive, played an important part in the formation of the ore body, the chalcopyritc zone being the pro-tore. Several thin, gently dipping, diabase sills occur in the mineralized area, and the copper content of the rock in and near these sills is substantially higher than average. The formation of the ore body was caused by the combined effects of three forces: I. Zoning of the hypogene mineralization, 2. Richer copper metallization associated with the diabase sills, and 3. Supergene enrichment. Introduction The Castle Dome copper deposit is about 5 miles west of Miami, Arizona, and about 3 miles west of the Miami-Inspiration ore body. The Castle Dome open-pit mine is on the south flank of Porphyry Mountain, a prominent landmark of the region. In 1941 the deposit was being explored by the Miami Copper Co., but the exploration program was cut short when the decision was made to develop the property as a war project. The Castle Dome Copper Co., Inc., was organized as a wholly owned subsidiary of the Miami Copper Co., and arrangements were made with Defense Plant Corp. for developing and equipping the mine and for construction of a concentrating plant. Construction and development work was begun in January 1942, and milling operations started June so, 1943. Since that date, the mine has been in continuous operation by the Castle Dome Copper Co, under a lease agreement with Defense Plant Corp. producing copper at an average rate of 4,000,000 lb per month. To January I, 1946 the total production amounted to 121,088,000 lb of copper,4,271 oz of gold, and 211,755 oz of silver. A detailed study of the copper deposit and the surrounding area was undertaken by the U. S. Geological Survey in June 1943. Other members of the Survey participating in the work were C. M. Gilbert, G. L. Quick, and J. P. Albers. The results of this work will be published as a comprehensive report by the Geological Survey. General Geology The Globe quadrangle, which includes the Castle Dome area, was studied by F. L. Ransome1 of the Geological Survey in 1901 and 1902. His maps and descriptions were

Jan 1, 1949

By Joseph J. Yancik, Leonard L. Felts, George B. Clark

LITTLE information has been published concerning the actual or useful amount of energy obtained from explosives when they are used for blasting. To provide more data on this subject, 8-in. neet cement cubes were blasted in a steel plate box and the breakage energy evaluated by comparison with drop crusher results using Rittinger's,1,2 Kick's: and Bond's' theories. Theories by Bond and Wang," Lineau," Hatch,' Roller,' and others0-11 were not considered applicable. Those used in this investigation may be expressed mathematically as follows: Rittinger Theory E = KrS2, pet,—--------------- [1] 40 d1 d0 Kick Theory TV 1 1 E = =K, [S pct. log1/d+ 100 log d0] [2] 40 d1 L. L. FELTS, Junior Member AIME, was formerly a Graduate Assistant, University of Illinois. G. B. CLARK, Member AIME, was formerly Professor of Mining Engineering, University of Illinois, and is now Chairman, Dept. of Mining Engineering, Missouri School of Mines and Metallurgy. J. YANCIK, Junior Member AIME, is a Graduate Assistant, Missouri School of Mines and Metallurgy. This paper, TP 4213A. is revised from a thesis submitted by L. L. Felts in partial fulfillment of the M.S. degree in Mining Engineering, University of Illinois, 1954. Discussion may be sent (2 copies) to AIME before May 31, 1956. Manuscript, Dec. 27, 1954. Bond's Third Theory E = Kb [S pct. log1/d1+ 100 log d0] where E — energy used in breakage, d, = average diameters of screen fractions of broken cement particles, pct, = percent weight of diameter d;, and 2 pct, = 100 pct. The above equations were employed to determine values of Kr, Kk, Kb for particle size distributions similar to those obtained from the blasting curves. These values were in turn used as a means of approximating the specific energy of the blasted material. The efficiency of crushing by blasting was then determined by dividing this specific energy by the maximum available work of the explosive per gram of cement. Knowledge of the failure of solids under static loads is developing very rapidly. Their failure under impact loads has been investigated and reported in the literature only to a limited extent. Theories of failure' hnder static loading include 1) the maximum stress theory, 2) the maximum strain theory, 3) the theory of constant elastic strain, 4) theory of maximum shearing stress, and 5) Mohr's theory of strength. All of these have elements of usefulness in establishing an applicable theory for blasting. For highly strained and compressed material such as that surrounding a detonating charge of high explosive, the Mohr theory and its implications appear to be the

Jan 1, 1957

By Walter E. Duncan

The ores treated at the concentrating plant of the Mahoning Mining Co. at Rosiclare, Ill., come largely from the blanket replacement deposits of the northeastern part of Hardin County, Illinois, and consist of complex mixtures of galena, sphalerite and fluorspar in a gangue of quartz, limestone and sandstone. The several minerals are so intimately inter-grown that they are not liberated at the sizes usually treated by gravity concentration in the usual fluorspar plant. For this reason, an all-flotation method of concentration was developed in cooperation with the station of the U. S. Bureau of Mines at Rolla, Mo.1 This flotation separation made possible the commercial production of four coyentrates: (I) lead, (2) zinc, (3) acid-grade fluorspar and (4) metallurgical-grade Auorspar.2-3 Although fluorspar had been floated in Canada in 19214 and a plant at Rosiclare had been floating fluorspar since March 1929, fluorspar flotation was given considerable impetus by the Mahoning developments, since some features of the separation were quickly adopted by other fluorspar companies. The Bureau of Alines5 reports that in the 10 years preceding 1938 only 50,552 tons of fluorspar flotation concentrates were produced in the United States. The rise from this almost nominal figure to 153,750 tons (largely acid grade) in one year (1944)6 is indicative of the activity in fluorspar flotation occasioned by the war. The industry thus proved to be in a position to expand and keep up reasonably well with the requirements imposed upon it by such developments as occurred in high octane aviation gasoline alkylation, in the use of Freon as a refrigerant and as carrier in the Aerosol "bug bomb," and in the atomic bomb project as well as in the expansion of other activities requiring hydrofluoric acid; for example, the aluminum industry. Despite the fact that the greater portion of acid-grade fluorspar produced in the last few years went directly into war uses, the future of the acid-grade field still appears bright. As a result of the activity in fluorine chemistry during the war, it seems likely that in the relatively near future almost every phase of human endeavor will find some fluorine compound involved. The other three concentrates produced deserve some attention in these introductory remarks. The lead concentrates first produced contained both zinc and fluorspar but these did not interfere with their marketing. As will be pointed out later, the quality has steadily improved until today it is a relatively high-grade product. Although some zinc smelters accept concentrates containing limited quantities of fluorspar, it is a very objectionable impurity. Since flotation methods fail in the complete removal of the fluorspar, chemical methods were developed for

Jan 1, 1948

By Walter E. Duncan

THE ores treated at the concentrating plant largely come from the blanket replacement deposits of the northeastern part of Hardin County, Illinois, and consist of complex mixtures of galena, sphalerite and fluorspar in a gangue of quartz, limestone and sandstone. The several minerals are so intimately intergrown that they are not liberated at the sizes usually treated by gravity concentration in the usual fluorspar plant. For this reason, an all-flotation method of concentration was developed in cooperation with the station of the U. S. Bureau of Mines at Rolla, Mo.1 This flotation separation made possible the commercial production of four concentrates: (I) lead concentrates, (2) zinc concentrates, (3) acid-grade fluorspar concentrates and (4) metallurgical-grade fluorspar concentrates.2,3 Although fluorspar had been floated in Canada in 1921 4 and a plant at Rosiclare had been floating fluorspar since March 1929, fluorspar flotation was given considerable impetus by the Mahoning developments, since some features of the separation were quickly adopted by other fluorspar companies. The Bureau of Mines5 reports that in the 10 years preceding 1938 only 50,552 tons of fluorspar flotation concentrates were produced in the United States. The rise from this almost nominal figure to 153,750 tons (largely acid grade) in one year (1944)6 is indicative of the activity in fluorspar flotation occasioned by the war. The industry thus proved to be in a position to expand and keep up reasonably well with the requirements imposed upon it by such developments as occurred in high octane aviation gasoline alkylation, in the use of Freon as a refrigerant and as carrier in the Aerosol "bug bomb," and in the atomic bomb project as well as in the expansion of other activities requiring hydrofluoric acid; such as, for example, the aluminum industry. Despite the fact that the greater portion of acid-grade fluorspar produced in the last few years went directly into war uses, the future of the acid-grade field still appears bright. As a result of the activity in fluorine chemistry during the war, it seems likely that in the relatively near future almost every phase of human endeavor will find some fluorine compound involved. The other three concentrates produced deserve some attention in these introductory remarks. The lead concentrates first produced contained both zinc and fluorspar but these did not interfere with their marketing. As will be pointed out later, the quality has steadily improved until today it is a relatively high-grade product. Although some zinc smelters accept concentrates containing limited quantities of fluorspar, it is a very objectionable impurity. Since flotation methods fail in the complete removal of the fluorspar, chemical methods were developed for

Jan 1, 1946

By S. Floreen, H. W. Hayden

The effects of additions of iridium, rhodium, ruthenium, and platinum on the mechanical properties of Fe-C alloys were determined. Each alloying element significantly improved the toughness. The results support the general correlation that substitutional alloying elements that stabilize austenite are beneficial to toughness, but the reason for this improvement is not obvious at this time. The effects of substitutional alloying elements on the toughness of steel has long been a subject of both practical and theoretical interest. Despite the obvious importance of this subject, there has been little progress toward understanding the mechanisms by which alloying elements influence the fracture behavior. In some cases the deleterious effects of alloying elements have been shown to be due to mechanisms such as embrit-tlement of prior austenite grain boundaries or inhibition of cross slip in the ferrite lattice. In other cases, however, the reasons for the change in properties are not clear. This is particularly true for elements, such as nickel, that improve the resistance to brittle fracture. Several investigators'-3 have pointed out that there is a correlation between an alloying element's phase diagram with iron and its effect on toughness. Those substitutional alloying elements that stabilize austenite generally are beneficial to toughness, while ferrite-stabilizing elements generally are detrimental to toughness. The word "generally" should be emphasized, because exceptions to this rule can be found. Nevertheless, this correlation does offer a starting point at which to try to understand the effects of alloying elements on toughness. Thus far only two austenite-stabilizing elements, manganese and nickel, have been considered. In order to test the validity of this correlation several other austenite-stabilizing elements were evaluated. The noble metals, iridium, rhodium, ruthenium, and platinum, have binary phase diagrams with iron similar to the Fe-Ni and Fe-Mn diagrams. That is, each element expands the austenite field, and also higher concentrations of each element produce martensitic transformations at low temperatures. Presumably then, each of these elements should improve the toughness of steel. The present study was conducted to determine if this were so. EXPERIMENTAL PROCEDURE The alloys were made as l-lb air-induction melts using electrolytic iron and commercially pure noble metals. Each heat was initially deoxidized by a carbon boil, and further carbon was added to attempt to produce a 0.05 wt pct C content in each alloy. Before pouring, a final refining addition of 0.2 wt pct A1 and 0.1 wt pct Ti was made to each heat. The alloys were cast in a copper mold to produce ingots approximately 1 in. in diam and 5 in. long. The chemical compositions of the alloys tested are given in Table I. The residual nitrogen, aluminum, and titanium contents were all relatively constant at 0.003, 0.03, and 0.02 pct, respectively. The carbon

Jan 1, 1968

By C. C. Hatch, C. R. Larsen, A. L. Mular

INTRODUCTION As part of a general project to simulate a concentrator, the Brenda Mines Limited secondary crushing plant, located near Peachland, B. C., was sampled during the summer/winter of 1975 and during the summer of 1978 to develop a digital crushing plant simulator specific to the Brenda operation. The purpose of this simulator was to study the influence of manipulable process variables on important responses, to permit study of alternative circuit configurations if desired and to perform off-line optimization. This chapter describes the methodology employed to develop the simulator (1,2) along with typical applications thereof. SECONDARY CRUSHING PLANT FLOWSHEET Figure 1 is a schematic diagram of the Brenda Mines Ltd. secondary crushing plant circuit. The 41,000 tonne coarse ore stockpile(a) is -17.8 cm oversize (3B belt) from two 2.4 m x 6.1 m double deck vibrating screens (not shown). The screens are fed by 2 conveyor belts operated in series with two 198 cm apron feeders that handle discharge from a 152 cm x 226 cm primary gyratory crusher(520 KW). The primary crusher (not shown) treats up to 2725 tonne/hr of run of mine ore. Minus 1.9 cm undersize from the primary screens (#19 belt), referred to as the primary fines, joins secondary screen undersize (#18 belt) for conveyance to fine ore bins (not shown). Feed from the coarse ore stockpile is transferred by means of vibrating feeders and two 120 cm conveyors (#6 and #7) to transfer belts #8 and #9 (Figure 1). The belts feed two secondary 2.1 m standard heavy duty cone crushers (bl and b2). Each crusher is equipped with a modified coarse mantle liner and a standard coarse bowl liner. Mantle throw is 9.2 cm; nominal close side set is 3.2 cm. Crusher .countershafts (443 rpm) are driven by single 260 KW synchronous motors (710 rpm). Discharge from the secondary cone crushers joins discharge from four 2.1 m tertiary heavy duty shorthead cone crushers (fl, f2, f3 and f4). With the exception of the cavity (modified medium mantle liner and standard coarse bowl liner; close side set is nominal .84 cm) and motor (224 KW), these machines are similar to the secondary crushers. The combined discharge is transferred from a 183 cm conveyor to #14 belt which empties into a 1452 tonne secondary screen surge bin (c). Vibrating feeders beneath the bins feed five 2.4 m x 6.1 m double deck vibrating screens (dl, d2, d3, d4 and d5) sloped at 200 from the horizontal. The upper decks are punched plates with 2.7 cm x 7.6 cm slots. Lower decks are woven wire (1.27 cm diameter) sections with

Jan 1, 1982

By Walter E. Duncan

The ores treated at the concentrating plant of the Mahoning Mining Co. at Rosiclare, Ill., come largely from the blanket replacement deposits of the northeastern part of Hardin County, Illinois, and consist of complex mixtures of galena, sphalerite and fluorspar in a gangue of quartz, limestone and sandstone. The several minerals are so intimately inter-grown that they are not liberated at the sizes usually treated by gravity concentration in the usual fluorspar plant. For this reason, an all-flotation method of concentration was developed in cooperation with the station of the U. S. Bureau of Mines at Rolla, Mo.1 This flotation separation made possible the commercial production of four coyentrates: (I) lead, (2) zinc, (3) acid-grade fluorspar and (4) metallurgical-grade Auorspar.2-3 Although fluorspar had been floated in Canada in 19214 and a plant at Rosiclare had been floating fluorspar since March 1929, fluorspar flotation was given considerable impetus by the Mahoning developments, since some features of the separation were quickly adopted by other fluorspar companies. The Bureau of Alines5 reports that in the 10 years preceding 1938 only 50,552 tons of fluorspar flotation concentrates were produced in the United States. The rise from this almost nominal figure to 153,750 tons (largely acid grade) in one year (1944)6 is indicative of the activity in fluorspar flotation occasioned by the war. The industry thus proved to be in a position to expand and keep up reasonably well with the requirements imposed upon it by such developments as occurred in high octane aviation gasoline alkylation, in the use of Freon as a refrigerant and as carrier in the Aerosol "bug bomb," and in the atomic bomb project as well as in the expansion of other activities requiring hydrofluoric acid; for example, the aluminum industry. Despite the fact that the greater portion of acid-grade fluorspar produced in the last few years went directly into war uses, the future of the acid-grade field still appears bright. As a result of the activity in fluorine chemistry during the war, it seems likely that in the relatively near future almost every phase of human endeavor will find some fluorine compound involved. The other three concentrates produced deserve some attention in these introductory remarks. The lead concentrates first produced contained both zinc and fluorspar but these did not interfere with their marketing. As will be pointed out later, the quality has steadily improved until today it is a relatively high-grade product. Although some zinc smelters accept concentrates containing limited quantities of fluorspar, it is a very objectionable impurity. Since flotation methods fail in the complete removal of the fluorspar, chemical methods were developed for

Jan 1, 1948

By J. O. Mack, J. B. Russell

In recent years, solders containing 20 per cent tin with no bismuth or cadmium have been developed by a few companies, and personnel have been properly instructed in their use. In addition, since the tin shortage, a number of solders containing less than 10 per cent Sn, and some with no tin, are being used in special applications; however, a solder of this type has yet to be developed that is satisfactory for general use. Table i.—Composition Ranges Prescribed for a Commercial Substitute Solder Element Per Cent Tin............................ 18.0-20.0 Bismuth (maximum)a............ 0. 80 Silver.......................... O. 10-1. 25 Bismuth plus silver (minimum).... 0. 85 Antimony (maximum)............ o. 50 copper (maximum),,,,,,,,..,,,, 0. I3 Other elements, each (maximum)6 0.03 Lead........................... balance a When silver is present in excess of I per cent, the Bi content shall be 0.50 per cent or over. b Total not to exceed 0.10 per cent. The Purpose of this investigation was to develop an alloy containing a maximum of 15 per cent tin, with no cadmium or bismuth, having solder properties equivalent to or better than those of substitute solders now specified containing 18 to 20 per cent tin (Table I), or even approaching the properties of the standard 40 to 60 tin-lead solder. The wartime restriction of bismuth and cadmium, elements that have beneficial effects on fluidity, solidification range and strength, makes it increasingly difficult to develop a general-purpose solder containing a maximum of 15 per cent tin. In the course of this investigation, 11 elements were used in various combinations with the basic I5-85 tin-lead alloy. These elements were antimony, silver, copper, indium, thallium, lithium, calcium, barium, potassium, sodium and arsenic. In all, 116 low-tin solders were investigated, with the idea of developing a solder that will give satisfactory working properties, specifically in regard to "wetting" of copper, brass and iron, solidification range, capillarity and shear strength of a soldered joint. The relative importance of these various properties plays an important role in the search for substitute solders, since it is almost impossible to retain all of the desired qualities of a high-tin solder and still make drastic changes in the alloy composition. Because of the expected uses of a general-purpose solder, it is necessary that it have properties, especially in regard to wet-tability, that will enable it to be used on the metals in common use. Basic Considerations The maximum liquidus temperature of a solder is determined by the particular application involved and the temperature that can be obtained by the source of heat. If a soldering iron is used, it must not only

Jan 1, 1945

By E. J. Duggan

THE mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. DESCRIPTION OF ORE The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chalcopyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the dose association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. HISTORY Construction of the first milling unit was started in 1917 and it was operated during parts of 1918 and 1919. In the postwar years, there was almost no demand for the product and production was at a standstill until 1924. The Company's research campaign to develop the use of the metal began to show results in that year. From 1926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and' 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for molybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill 10,000 tons per day, but it was not long before 12,500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. To increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. CRUSHING There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. 1) and No. 2, of 15,000 tons capacity (Fig. 2).

Jan 1, 1942

By E. J. Duggan

THE mine and mill of the Climax Molybdenum Co. are at Climax, Colorado. Climax is on Fremont pass directly on the Continental Divide, at an elevation of 11,400 feet. DESCRIPTION OF ORE The Climax ore-bearing rock is essentially an altered and highly silicified granite, fully half of the gangue being quartz. Molybdenite is the only mineral of economic consequence and most of it is intimately associated with quartz in the form of fine veinlets and stringers. Other minerals are molybdite, pyrite, and chalcopyrite. While the molybdite may be readily extracted by hydrometallurgical methods, its content is too low to make such operations profitable. The pyrite content ranges from 2 to 5 per cent and that of the chalcopyrite from 0.03 to 0.05 per cent. There is not enough copper to make its recovery attractive and one of the metallurgical problems is the elimination of it and the pyrite from the product. On account of the close association of the molybdenite with the quartz and the necessity of eliminating the pyrite and chalcopyrite from the final product, fine grinding followed by flotation is the only feasible method of concentration. HISTORY Construction of the first milling unit was started in 1017 and it was operated during parts of 1918 and 1910. In the postwar years, there was almost no demand for the product and production was at a standstill until 1024. The Company's research campaign to develop the use of the metal began to show results in that year. From 1926 to 1932 production ranged from 500 to 1200 tons per day. In 1931 an additional mill section was built and production was gradually increased until in 1934 and 1935 about 3000 to 4000 tons was milled per day. This tonnage overtaxed the milling facilities, and as the demand for molybdenum was growing constantly, additional milling equipment was needed. In 1936 and 1937 another crushing plant, six additional mill sections and the necessary auxiliary plants were constructed and brought into operation. This plant was designed to mill I0,000 tons per day, but it was not long before 12,500 tons was being treated during periods of peak demands. During the summer of 1941, to meet defense needs, the mill has been required to treat 15,500 tons per day. This has been accomplished with little plant expansion and at no loss in recovery, as will be described later. To increase capacity still further, one more primary section is now being built. This will bring capacity to 18,000 tons per day, which can be stretched to 20,000 tons at a small sacrifice in recovery. CRUSHING There are two crushing plants: No. I, of 5000 tons capacity (flowsheet in Fig. I) and No. 2, of 15,000 tons capacity (Fig. 2).

Jan 1, 1942