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By Enid C. Plante, K. L. Sutherland

Practical metallurgists are unanimous in stating that oxidation of mined sulphide ore adversely affects separation of the constituent minerals under standard conditions in a mill. Frequently, the need for different operating conditions in different plants is attributed either to the extent of oxidation of the mineral surfaces or to soluble oxidation products. The investigations reported in this paper study the effect of oxidation on the primary condition necessary for flotation, namely that of adhesion of an air bubble to the mineral. Oxidation Products from Sulphide Minerals Weinig and Carpenter1 state that the relative rates of oxidation of some sulphides are: FeAsS > FeS2 > CuFeS2 > ZnS > pbS > Cu2S tetrahedrite. They also show that oxidation is rapid at first, that sulphides of older geological formations are more stable than those of recent date and that mixtures oxidize more rapidly than any one mineral alone. The more rapid oxidation of mixtures is borne out by results of Gottschalk and Buehler2 who found that PbS, zns and cus oxidized 8 to 20 times faster when pyrite was present. Only the end products of oxidation, sulphate and the metal ion, were identified in these tests. Gaudin3 states that pyrite is difficult to oxidize in the laboratory. Winchell,4 using pyrite containing 99.98 pct FeS2, allowed water to percolate through a bed of this ground material. At the end of three months only a trace of Fe+++and So4- were found. Even after 10 months the quantity of Fe+++ was only 7.8 mg and SO4- 25.5 mg per liter of percolating solution. Car-michael5 confirmed Winchell's results on the slow oxidation of pyrite. Usually the oxidation of sulphide surfaces has been conducted in neutral or acid solutions, the acidity being caused by the oxidation products. Eliseev and Nagir-nyaka found that treatment of pure pyrite with lime gave only thiosulphate, but considered it possible that thiosulphate was obtained not by oxidation of the pyrite itself, but by oxidation of calcium disul-hide. Mitrofanov7 suggested that thiO~ulphate was formed by the following process: oxidation of sulphide to give free sulphur (this frequently happens with marcasite), followed by oxidation of the sulphur to sulphur dioxide and reaction of the sulphur dioxide with sulphur to give thiosulphate. Ferric and ferrous hydroxides, or similar compounds, were shown to be present. The Research Staff of the United Verde Copper Co.8 treated ore from the United verde mine with lime. Thiosulphate, sulphide and polysulphides were stated to be

Jan 1, 1949

By E. Seidl

THE basic material for testing aluminum, copper and zinc is a cast metal, principally in the form of wire bars (Figs. 1 to 3) and plates for the produc- tion of sheets or strips, (Figs. 4 to 7). Experience has proved that the chemical and crystallographic proper-ties of such cast metal exercise a certain influence when it is being worked, which is usually done while the metal is hot. In a chemical sense the slight impurities, which even the purest industrial metal contains, have a disturbing effect. They settle either along the outside limits of granulation, as in the case of aluminum and copper, or else in the interior of the crystal, as in the case of zinc; but frequently are also found accumulating at certain points (formation of the blowholes and skin) (Fig. 7). The crystallographic character of the metals is either that of uniformly fine granulation, or else there occurs a certain distinct casting texture marked by sharply defined cooling conditions and consisting of crys-tals of a pronounced stalk-like or blade form. Since we generally distinguish between a crystal texture and a crystal structure, the term crystal texture is used here for the arrangement of the crystals relative to each other and their microscopic composition; crystallic structure is used to mean the radioscopically ascertain-able structure of the crystal. The influence of the impurities and of the crystalline character on the working of the metals has been fre-quently pointed out in practice as well as in science.

Jan 11, 1927

By Klaus Schwerdtfeger

The rules of solution of oxygen from H2O-H2-He gas and of nitrogen from N2-H2 gas in shallow melts of liquid iron were measured at 1610o and 1600o C, respectiuely. Concentration profiles were detemined in the liquid iron. Tire rate data indicate that the solution process is controlled by diffusion in the iron melt. The diffusivities for oxygen and nitrogen in liquid iron, as calculated from the present data, are DFe-o = (12 ± 3) < 10-5 sq cm per sec and DFe-N = 11 ± 2) X 10-5 sq cm per sec at the temperatures employed. AN attempt was made by Shurygin and Kryukl to measure the diffusivity of oxygen in liquid iron. In their experiments a silica disc was rotated in liquid iron containing oxygen, and the rate of formation of liquid iron silicate was measured. Assuming that the rate of dissolution of silica is controlled by diffusion of oxygen in the iron, the oxygen diffusivity was computed from the rate data giving Dfe-0 = 6.1 X 5 sq cm per sec at 1600°C. Although this value seems to be of the right order of magnitude, there is no proof of the correctness of the assumptions involved in the interpretation of these rate data. The oxygen concentration in the iron at the iron-iron silicate interface was taken to be that in equilibrium with the silica-saturated silicate melt. That is, it was assumed that no concentration gradient existed in the liquid silicate. This is a questionable assumption, unless it is proved that the thickness of the silicate layer is very much smaller than that of the diffusion boundary layer in the iron. Furthermore, Shurygin et al.1 used the Levich equation2 to interpret their rate data. This equation was derived for mass transfer between a solid disc and a single-phase liquid. The hydrodynamic and diffusion boundary layers in the iron stirred by a disc, via coupling of the silicate melt, may be appreciably different from those predicted by Levich&apos;s derivations. In the present work the diffusivities of oxygen and nitrogen in liquid iron were measured at 1610" and 1600oC, respectively. EXPERIMENTAL METHOD Iron melts contained in high-purity gas-tight alumina crucibles were reacted with H2O-H2-He gas for the determination of the oxygen diffusivity and with N2-H2 gas for the determination of nitrogen diffusivity. At the end of the reaction period, the samples were quenched in a cold H2-He gas stream at the top of the furnace. Oxygen or nitrogen contents in the iron were determined by chemical analysis. Two different types of diffusion experiments were perforxed. To determine concentration profiles, a few rate measurements were made using 4-cm-deep melts. The solidified samples were sliced into discs and each disc was analyzed for oxygen or nitrogen. In another series of experiments, oxygen or nitrogen was diffused into shallow melts (about 0.5 to 1 cm in depth) and the total sample was analyzed to obtain an average concentration of the diffusate. In most experiments, 4- to 5-mm-ID alumina crucibles were used. Some experiments were also made in smaller (3 mm) and larger (7 mm) diam crucibles. This variation in diameter caused no difference in the reaction rate, within the limits of experimental uncertainty. To promote the establishment of a stable density profile in the melt, all the samples were suspended in the lower end of the hot zone so that the top of the melt was hotter by a few degrees. Molybdenum wire resistance heating was used. The reaction tube of the furnace was a gas-tight recrystal-lized alumina tube. In most experiments the furnace was heated by an ac power supply. To check the possibility of inductive stirring, some experiments were carried out in a dc operated furnace, with essentially the same results. The temperature of the furnace was controlled automatically in the usual manner. The temperature was measured with a Pt/Pt-10 pet Rh thermocouple and is estimated to be accurate within ±5°C. The iron used was prepared by melting and vacuum-carbon deoxidizing electrolytic "Plastiron" in a zir-conia crucible. The main impurities are: Si 0.004 pct P, S <0.002 pct Cr 0.005 pct N 0.001 pct Zr 0.002 pct O 0.003 pct Mn 0.004 pct C 0.002 pct The gas composition was controlled by constant pressure head capillary flowmeters. Oxygen was removed from the gas mixture by passing it through columns of platinized asbestos (450°C) and anhydrone. Selected H2O contents were obtained by passing the purified gas through oxalic acid dihydrate-anhydrous oxalic acid mixtures held at constant temperature in a water bath. Water vapor pressure data for the oxalic acid dihydrate-anhydrous oxalic acid equilibrium were taken from the 1iterature.3 The flow rate used was about 1.5 liters per min. The whole system was checked for tightness at regular intervals.

Jan 1, 1968

By R. V. Wheeler

A matter of much concern at present to the coal-mining industry of Great Britain is the effect of the working of the Coal Mines Act, 1930. This Act, which received the Royal Assent on Aug. 1, 1930, contains five parts: (1) production, supply and sale of coal; (11) the reorganization of the coal-mining industry; (111) hours of work; (IV) the Coal Mines National Industrial Board; (V) general. The first three parts are of major interest and discussion can be confined to them. Part I.—Production, Supply and Sale of Coal Under Part I of the Coal Mines Act, 1930, provision is made for central and district schemes to regulate the production, supply and sale of coal by the fixation of standard tonnage, quota and minimum prices. By Nov. 1, 1930, all the district schemes came into force with the exception of that for Scotland, where the majority of the coal-owners did not see their way to accept the amendments to their draft scheme demanded by the Mines Department. A compulsory scheme for Scotland was thereupon prepared by the Mines Department and put into force on Nov. 24, 1930. It is administered by the coal-owners of the district. The essence of the schemes is the allocation of tonnage for each district by a Central Council, and the allocation of quota for each pit in a district by an Executive Board for that district. Up to Dec. 31, 1930, the various bodies of coal-owners engaged in administering the schemes were employed in work preliminary to their full operation. Standard tonnages had to be determined and coals had to be classified. In consequence, it was not found possible to put the regulation of output into force until Jan. 1, 1931. The determination of minimum prices proved to be an even longer process, and only in a few districts were minimum price schedules operative during the first quarter of this year. The working of the schemes during the March quarter, 1931, has been reviewed in an official publication by the Mines Department,&apos; which can be summarized as follows:

Jan 1, 1932

By Boyd Dudley

THE material presented in this paper is an abstract of a thesis submitted by the writer to the faculty of the Massachusetts Institute of Technology, as part requirement for the degree of Master of Science. The investigation was undertaken at the suggestion of Prof. H. 0. Hofman, and it was conducted, under his direction, in the laboratories of the Mining Department of the Institute. The purpose of the work was, first, to study some of the reactions that have been considered as a part of the mechanism of the chloridizing roast, and second, to study the conditions under which nickel oxide and copper oxide may be chloridized. While the production of chlorine gas by various reactions during the chloridizing roast may not be primarily useful in the chloridation of metal oxides and compounds, it nevertheless serves a useful purpose in preventing the dissociation and decomposition of chlorides once they have been formed; for example, the dissociation of cupric chloride into cuprous chloride and chlorine. Therefore it was thought that a study of some chlorine-producing reactions would prove of interest, and the first series of experiments dealt with this phase of the subject. The object of the latter part of the work was to show the possibility of treating heavy sulphide ores, such as those of Sudbury, Ontario, containing much iron and small amounts of copper and nickel, by chloridizing and removing the copper, leaving an iron-nickel oxide product suitable for smelting in the blast furnace for nickel-bearing pig iron. The chloridizing of the copper would, of course, be preceded by a roast for the removal of sulphur, the sulphur oxides being available for the manufacture of sulphuric acid, if under the local conditions this product would be of commercial value.

Jan 12, 1914

By D. S. Tannhauser, I. Bransky

Measurements of the electrical conductivity, the thermal electromotive force, and the deviation from stoichiometry by thermogravimetry were made on ferrous oxide (wüstite) single crystals as well as on poly-crystalline samples in the temperature range of -900" to 1330°C. In the region defined by T 2 1000°C and 1.05 5 O/Fe 5 1.10 the electrical conductivity was shown to be proportional to the sixth root of- the oxygen partial pressure and to be linearly dependent on the deviation from stoichiometry. The absolute Seebeck coefficient of single and polycrystalline wüstite was shown to be independent of temperature in the above region and a p to n transition of the Seebeck coefficient was observed at O/Fe = 1.09 in polycrystalline wüstite. Analysis of the results showed that the predominant defects in wüstite in the above region were doubly ionized cation vacancies. The analysis also ruled out the existence of complexes of the type postulated by Roth, except that below 1000°C there is some evidence of &apos;association of defects. Of the three oxides of iron (FeO, Fe3O4, Fe2O3) only FeO (wüstite) can exist within a wide range of the ratio O/Fe. This has been shown by several investigations of the phase diagram using for instance chemical analysis of quenched samples&apos;-4 or thermogravimetric measurements. 5,8 Darken and Gurry&apos;s1 data for example in the temper- ature range 1100o to 1400°C show that the lowest O/Fe ratio obtainable for FeOl +. is about 1.045 and that the existence range extends to O/Fe = 1.20 at 1400oC. On the basis of detailed analysis of thermogravimetric measurements made by Vallet and Raccah6 the existence of three varieties of wüstite at elevated temperatures was suggested by these authors and by Kleman. It is well-established that wüstite has a NaCl structure with vacant cation sites.3, 8 From a neutron-diffraction study of wüstite samples quenched to room temperature Roth9 concluded that defects in wüstite are mostly complexes composed of two cation vacancies and one interstitial cation in a tetrahedral site. The atomic arrangement in the vicinity of such a defect is similar to that in magnetite, where each occupied tetrahedral site is surrounded by four vacant octahedral positions disposed at the corners of a tetrahedron, so arranged that each vacancy is shared by two tetrahedral cations. Thus there is an average of two vacancies per interstitial. Recent X-ray diffraction measurements made by Smuts10 on quenched wüstite samples confirmed the results of Roth&apos;s neutron-diffraction experiments. Roth&apos;s model of defects in FeO can be regarded as a solid solution of Fe3O4 in FeO; salmon1&apos; used this defect model to calculate the activity of Fe3O1 in FeO at high temperature and claimed good agreement with the results obtained experimentally by Darken and Gurry.&apos; Each missing cation in wüstite is accompanied by two trivalent cations, thus preserving charge neutrality. The "holes" which differentiate between the Fe 2+ and Fe3&apos; states of the ions in octahedral positions can overcome the attractive field of the cation vacancies, and are then free to move through the crystal as charge carriers. A similar picture holds for COO, NiO, andMnO.l2, 13 Localization of the charge carriers in transition

Jan 1, 1968

By M. King Hubbert

DuRing the past summer the writer was asked by the Illinois State Geological Survey to make a study of some of its economic geological problems with regard to the applicability of geophysical methods of prospecting to their solution. The following four problems were suggested for investigation: (1) the fluorspar area of Southern Illinois, (2) oil structures in the south central part of the state, (3) the lead and zinc-bearing area of northwestern Illinois, and (4) the water supply and gravel deposits of the glacial drift. The prime requirements were to find a means of obtaining a maximum of useful information with a fairly modest outlay of expenditure for technical staff and equipment. The method of procedure adopted was that of making a preliminary study of each of the problems named by means of available office data and published reports with supplementary field reconnaissance where necessary. Preliminary Studies Fluorspar.—The fluorspar deposits of southern Illinois as shown by Weller,1 Bastin2 and others are of two main types: (1) vein deposits along steeply dipping normal fault planes, and (2) "blanket" deposits occurring practically horizontally along a certain sedimentary horizon. The vein deposits are composed mostly of calcite and fluorspar in sheetlike veins of thicknesses, where mined, ranging from about 2 to 12 ft. The surrounding topography is in a mature state of development with a relief of about 100 ft. or more. Any geophysical prospecting for this ore would have to depend either upon the physical properties of the ore and associated minerals or upon the properties of the geologic structure containing them. Fluorspar has a density of 3.18, which is somewhat greater than 2.7, that of calcite. Its

Jan 1, 1934

By Max Heberlein, R. P. E. Hermsdorf

The electrolytic refining of metals for the removal of undesirable impurities has become a recognized necessity in the nonferrous field. Copper, lead, zinc, nickel, silver and gold have been produced in this manner for years. The United States Metals Refining Co., realizing that considerable quantities of lead and tin could be made available to the trade if a satisfactory method of purification of these metals could be established, decided to investigate the possibilities of electrolytic solder. In cooperation with Dr. Edward F. Kern, of the School of Mines at Columbia University, the company developed a process of electrolytic refining of lead-tin alloys using a tin-lead fluosilicate electrolyte. A pilot plant was installed in November 1927, and the commercial unit was put in October 1929. Since that time thousands of tons of electrolytic solder have been produced in this plant and consumed by all branches of industry that use solder. Older Methods of Producing Solder In the early days practically all solder was made by melting together virgin lead and tin in the proper proportions to meet specifications; then the solder was cast into the commercial shapes required. Today the base of most solders comes from scrap or secondary materials. Ordinarily solder is produced from secondary materials either by sweating solder-bearing scrap or by smelting lead and tin-bearing drosses, residues and similar materials. When solder is produced from scrap by sweating, clear solder is collected directly. It is then given some kind of kettle refining and usually blended to specification with virgin metals before it is cast into finished shapes. In producing solder from drosses and residues, the materials are usually first given a reducing smelting treatment from which a crude solder is obtained. This crude solder may be refined in several ways, but usually the first step is a furnace liquation to remove the bulk of the copper, which is followed by one of the more or less complicated methods for the removal of antimony and other impurities.

Jan 1, 1937

By J. Versluys

A well can flow exclusively through the casing or exclusively through a tubing but can also flow partly through a casing and at the top part through a tubing. The main principles of the flowing of wells under these conditions were explained in the author&apos;s former paper.&apos; Practically all factors that influence the regular flow are dealt with therein. The question of irregular flow, however, will be discussed briefly here. As a rule, an irregular flow must be avoided, as it may be assumed that the efficiency is thereby very small, while on many fields it is injurious to the sand of the oil-bearing layer. Periodical flowing often occurs with rising mixtures of gas and liquid. The phenomenon can be explained if it is accepted that two conditions are possible in the mixture; namely, the foam condition and the mist condition. If there is no stabilizer in the liquid, the foam condition can exist only when the liquid forms a greater portion of the volume than the gas. The mist condition, on the other hand, can exist only when the liquid occupies a smaller part of the volume than the gas. The foam condition occurs when there are gas bubbles in the liquid mass, while in the mist condition drops of liquid are disseminated in a gas-filled space. For sake of simplicity, it has been assumed that the mist condition prevails when 50 per cent. or more of the volume is occupied by gas, while with less than 50 per cent. gas the foam condition will arise. For distilled water and for oils which do not tend to form stable foams this is probably approximately true. If this does not hold good, however, two cases are possible: (1) The fields of the two conditions overlap or (2) they do not meet. If the fields of the conditions overlap there is an intermediate field where either of the two conditions can exist, depending on the history of the mixture. Suppose the foam condition prevails first and the gas

Jan 1, 1931

By R. V. Norris

The removal of mine-water by hoisting in tanks instead of pumping, while somewhat a reversion to the methods of the ancients, has come very rapidly into favor in the anthracite region of Pennsylvania during the past few years. In fact, so much so, that at the present time there are at least eight large collieries at which all the water is hoisted, and six more plants are in preparation. The particulars of the plants in operation and under construction, as far as decided on, are given in the tables on pages 107 and 108. Besides these, the Delaware, Lackawanna and Western R.R. Co. is constructing a hoist of large capacity for a depth approximating 500 feet, to be located in the Keyser Valley, near Scranton. The hoists will be driven by electric motors, geared directly to the sheaves. The Lehigh Valley Coal Company is planning a large hoist for the Packer collieries in the Shenandoah district, but no work has been done on this as yet. The earliest regular hoisting, I believe, mas done by means of semi-cylindrical tanks (Fig. 1) at the Nanticoke collieries of the Susquehanna Coal Company. These tanks were attached under the regular shaft-carriages, taking in water through six large clack-valves in the bottom, and discharging through an endgate opened by a lever which was operated by a guide-piece on the shaft head-frame. These tanks were designed July, 1880, by the late J. H. Bowden, one of the very early members of the Institute. Similar tanks are still used in emergencies at the Nos. 1, 2, 3 and 6 shafts of the company. The objections to. their use were: that water could only be hoisted during the night-shift, or when the shafts were not in use for hoisting coal, thus requiring a very large sump, and greatly limiting the water capacity of the plants; that the alternate wetting and drying

Jan 1, 1904

By Wilbur Nelson

IN 1839,1 Gerard, Troost, the first, State Geologist of Tennessee, reported the occurrence of zinc ores in East Tennessee, in connection with the iron ores at Embreeville; in 1844,2 he refers to the zinc deposits of East Tennessee, especially the apparently rich deposits occurring, in Jefferson and Claiborne Counties, while his ninth annual report is a ,description of the zinc deposits of Jefferson County and a description of the most improved processes by which zinc is extracted from its ores. Although. Tennessee has long known of the occurrence of its zinc deposits and has recognized their value, only during the last few years have these ores been successfully worked. As far as can be learned, prior to the systematic development of the Mascot district in 1911, no large-scale or successful operation of any of the zinc deposits was carried on. But in that year, after several companies had made considerable progress in developing the Mascot area and had proved. a relatively extensive body of low-grade sulfide ore, the American Zinc Co. entered the field. It introduced modern methods of prospecting and mining the ore and from the first was successful. Too great credit cannot be given this company for its pioneer work in a previously unimportant field. Throughout the history of its operations, the American Zinc Co. has confined its efforts to underground mining of the sulfide ore, sphalerite. This company has found that successful results in zinc mining in East Tennessee can be achieved only by expending large sums in prospecting and blocking out bodies of pay ore in sufficient magnitude for large-scale operations. At Mascot are located the mines of the American Zinc Co. of Tennessee, the mines being situated along the Southern Railway in the Holston River Valley, Knox County, about 13 miles northeast of Knoxville. Considerable mining and prospecting had been done by several companies prior to the acquisition of this property by the, American Zinc Co. Previous to 1900 only open pits, which were worked for oxidized ores, were operated. and the production was small. In 1900 the

Jan 10, 1924

By W. Vickers

Single-point rotating cantilever fatigue tests have been carried out at 550" to 650°C on beryllium produced by a variety of fabrication routes. All the specimens gave similar plots of stress against number of cycles to failure but failure occurred at a lower stress range in the hot-pressed material than in the extruded material. A discontinuity or "kink" appears to exist in the curves and this is considered. A true endurance limit was not evident up to 10 cycles. When Brush hot-pressed warm-extruded beryllium was tested at 600oC a relationship where Nf = number of cycles to failure, ?ep = plastic strain range, and c = constant, was shown to be applicable for values of ?ep > 6.3 x 10-4. Further for this moterial and temperature and for frequencies between 150 and 4500 cycles per min, the variation of number of cycles to failure with frequency obeyed the expression N = Af0.44 where N =number of cycles to failure, f = frequency, and A = constant. Examination of the fracture surfaces indicated that they were generally intergranular and in all cases the higher the stress the more irregular was the fracture. Microexamination has shown that grain boundary cavities and fine intergranular cracks were associated with the fractures. REVERSING or cyclic stresses set up in fuel elements can lead to failure by fatigue. Fatigue stresses may be developed by instability of the fuel element in the coolant flow leading to rattling against the channel wall or fin vibration. Thermal cycling also sets up reversing stresses due to differential expansion of fuel and can. Fatigue resistance is of some importance therefore in the evaluation of canning materials. At the time the work described in this paper was initiated, beryllium was being considered for the can and endcaps of an advanced gas-cooled reactor fuel element. With this in mind, beryllium obtained from different manufacturing routes was tested on single-point rotating cantilever fatigue machines at temperatures the same as those envisaged in the reactor to obtain appropriate fatigue data. The results obtained indicated a discontinuity in the S-N curves coinciding with the yield stress in the specimens. Although in some cases the data is insufficient to warrant a detailed quantitative analysis, a qualitative analysis of this effect was thought to be justified. 1) EXPERIMENTAL DETAILS The materials used are given in Table I together with the production route and available analyses. The specimen shape is shown in Fig. 1. The fatigue specimens were machined slightly oversize, after which they were given a stress-relieving anneal of 1 hr at &apos;750° C in vacuum. This was followed by an etch in sulfuric-phosphoric-chromic acid mixture down to the finished dimensions given in Fig. 1. Etching served to remove the interference colored oxide film formed on the surface during annealing and very small recrystallized surface grains too. The specimen was mounted in the grips of a cantilever machine and rotated inside a furnace controlled at the testing temperature. The specimen was loaded in increments as the speed was increased until the required load was applied at 5000 cycles per min. A revolution counter was used to record the total number of cycles.

Jan 1, 1964

By A. W. Gray

Page CRUSHING STRENGTH...................... 659 The Black Dynamometer....................659 A Standardized Procedure for Crushing Strength Tests...... 660 Influence of Height of Test-piece upon Apparent Strength ..... 662 Influence of Packing Pressure and of Packing Time........ 662 Application of the Logarithmic Law...............665 Influence of Amalgamation Variables.............. 665 Influence of Temperature................... 669 The Transition Region Near 70" C............... 671 Influence of Packing Pressure at Various Temperatures ...... 673 Influence of Age upon Alloy and Amalgam........... 676 DIMENSIONAL CHANGES DURING THE HARDENING OF AN AMALGAM .......................... 678 The Black Amalgam Micrometer............... 678 Thermostated Dilatometer Designed by the Author........ 681 Graphical Representation of Reaction Expansions......... 683 The Typical Reaction Expansion Curve ............. 684 Influence of Packing Pressure................. 686 Influence of Trituration Time................. 687 Influence of Temperature................... 688 Influence of Mercury-alloy Ratio................ 689 Influence of Packing Time................... 689 Influence of Annealing the Alloy................ 691 Influence of Size of Alloy Particles...............691 Extension of the Typical Expansion Curve........... 693 Influence of Thermal Softening and of Reamalgamating a Hardened Amalgam.......................... 634 Curves Obtained with a Variety of Dental Alloys ........ 694 Copper Amalgam ...................... 694 PRACTICAL APPLICATIONS OF THE RESULTS.......... 695 Phenomena which reveal chemical and physical processes that probably go on in many alloys at temperatures not far removed from their melting ranges are observable in dental amalgams. These amalgams should. therefore. be of even more interest to the metallographer, the metallurgistl the chemist. and the physicist than to the dentist who uses them for saving decaying teeth. For this reason I am here de-

Jan 1, 1919

By J. Kostuik, M. J. de Bastiani

The method of mining the uranium orebody at Denison mine in the Elliot Lake district of Ontario is discussed. Details of the development of haulage-ways and rooms are giuen, as well as methods of ore haulage. The milling and beneficiation stages are noted briefly. Located in the Elliot Lake mining district of Canada, the Denison mine lies between Sudbury to the east and Sault Ste. Marie to the west, Fig. 1. This area is overlain by Huronian sediments filling a basin in the Keewatin basement complex. Subsequent folding of these sediments developed a synclinal structure, known locally as the Quirke Lake Trough, which forms the northern portion of the now famous big Z structure, the geological trademark of this region. Plunging to the west and extending nine miles from the granite outcropping in the north to the Keewatin lava contact to the south, the syncline has dips of 10" to 45" on the north limb and 10" to 20" on the south limb, Fig. 2. The Denison orebody, located on the north limb of the syncline, lies between 800 and 3000 ft below the surface. The main ore zone is within 50 ft of the Missisagi-Keewatin contact (see Fig. 3), and it varies in thickness from 6 to 32 ft. This ore zone has a composite dip of 19" to the south and is composed primarily of two pyritized conglomerate beds separated by a band of pebbly quartzite stratum. The upper conglomerate bed of the main zone varies from 2 to 15 ft in thickness, while the lower conglomerate varies from 2 to 12 ft. Over extended areas, mining is carried out over the entire section comprising the two conglomerate layers and intercalated quartzite; in other areas, the conglomerates are selectively mined, depending on the thickness of the quartz-ite. The ore zone contains approximately 6 pct sulfides and 80 pct silica. The predominant uranium-bear ing minerals are brannerite, uraninite, and an unidentified mineral (probably thucolite). Some of the uneconomic metallic minerals identified to date are pyrite, pyrrhotite, and chalcopyrite. MINE LAYOUT Shafts: Underground exploration and development (Fig. 4) were performed through two shafts 1/2 mile apart. The No. 1 shaft, sunk to a depth of 1856 ft with the main station established at the 1600-ft ore level, is a 250-sq ft, 5-compartment shaft capable of handling 4500 tpd. Two 5x5-1/2-ft compartments serviced by a 12-ft diam double-drum hoist, are used for hoisting 190-cu-ft bottom-dump skips in balance. A 30-man cage, offset by a counterbalance, operates within a 5-1/2x11-1/2-ft compartment. The fifth compartment is used as a manway, pipe, and cable-way. This shaft is also provided with an intermediate pumping station used during the initial mining phase as well as loading-pocket and spill-pocket installations and pumping facilities. The 480-sq ft No. 2 shaft, located down-dip from No. 1 shaft, has a 2776-ft depth. Serving as the main hoisting shaft with a maximum capacity of 9000 tpd, it has eight compartments, four of which are used for hoisting ore in 10-ton capacity bottom-dump skips. These four compartments, measuring 5x6-ft in section, are serviced by two 12-ft diam double-drum semi-automatic hoists. A 50-man cage is operated in a 15x8-ft compartment, and it is counterbalanced by a 12-ton weight. Two compartments are used for ventilation, manway, electric cables, and pipes. A station was constructed at the 1325-ft horizon of the No. 2 shaft for relay pumping. The main sta-

Jan 1, 1961

By W. E. C. Coxe

MANY of you who are interested in the manufacture of iron and steel, have no doubt heard of the "Wheeler process for combining iron and steel." Mr. Wheeler has formed a company, styled the "Combination Trust Company," whose business is to grant licenses for the use of this patent process in producing what he calls "iron-clad steel." It has been proposed to apply it in the manufacture of car or wagon axles, also to shafting, having steel cores or centres covered with iron; to tubes of steel with an inner and outer skin of iron ; to structural purposes, such as girders and beams; bridge and tie-bolts, where the strength of the iron and the rigidity of the steel would be desirable. In answer to a request from Mr. Wheeler, permission was accorded him to make some experiments at the rolling mill of the Philadelphia and Reading Coal and Iron Company, at Reading, Pennsylvania, with a view to manufacturing some steel-headed rails. In making this kind of a rail two difficulties are encountered. If the iron is brought to the requisite heat to weld properly, the steel is likely to be burned ; if the steel is cared for, to prevent burning, the several layers of iron are not brought up to the proper welding heat. Mr. Wheeler claimed that the steel slab covered with iron which be furnished for the head of the rail could not be injured by any heat to which it might be subjected in the ordinary course of heating the rail-pile. He claimed, also, that by this method the porosity, which he thought was characteristic of all solid rails of Bessemer steel, would be overcome. Four steel slabs were furnished for this experiment, being Bessemer steel of .61 carbon, made at the Edgar Thomson Steel Works, and covered with iron in one of the Pittsburgh mills. The slabs were in section seven inches wide by three inches thick, and the iron covering was one-eighth to three-sixteenths inch thick. Each slab was placed on a pile of iron made up in the ordinary way for a rail fifty-six pounds per yard and thirty feet long. The heating was under the personal direction of Mr. Wheeler, and

Jan 1, 1879

By Ralph W. Smith

Development of the lime industry in Ste. Genevieve County began in a crude way in 1840. According to information furnished by the Missouri Bureau of Geology, in the early days small vertical kilns built of stone were used, and wood was the fuel. Modern development began in 1904 when one company introduced steel shells lined with firebrick. Within a few years two more companies erected similar plants, and in 1925 a fourth company began operations. Thus there arc today four lime companies at Ste. Genevieve, all operating at high capacity and with a gradually increasing production. Geology The limestone of Ste. Genevieve County, which has the chemical composition and purity required for lime burning, is the upper 80 or 90 ft, of Spergen formation. The Missouri Bureau of Geology, in its geological map of Ste. Genevieve County, gives the Spergen formation a total thickness of 160 ft., but only the upper portion is pure enough for lime burning, with an average raw-stone analysis of 98.7 per cent CaCo3; 0.35 SiO2; 0.3 R2O3; 0.35 MgCO3; 0.3 SO3. The Spergen, one member of the Meramac group, Iowa series, of the Mississippian system, underlies an area approximately 12 miles long north and south, by 5 miles wide east and west (Fig. 1) in Ste. Genevieve County. The entire eastern two-thirds of Ste. Genevieve County is sedimentary, the formations dipping gently toward the east on an average grade of about 4 per cent, or 2 1/2°. Thus the outcrops of the various sedimentaries run approximately north and south. A line running west from Ste. Genevievc crosses the outcrops of the sedimentaries, ranging from the top of the Mississippian to the Ozarkian geological system in a space of 12 miles. Thus with the gentle dip of 2 1/2° the 160-ft. thickness of the Spergen formation stretches out in its outcrop so that it is approximately one mile wide. The early quarry and lime developments at Ste. Genevieve mere in this easily obtained surface rock of the Spergen outcrop.

Jan 1, 1938

By Ralph W. Smith

Development of the lime industry in Ste. Genevieve County began in a crude way in 1840. According to information furnished by the Missouri Bureau of Geology, in the early days small vertical kilns built of stone were used, and wood was the fuel. Modern development began in 1904 when one company introduced steel shells lined with firebrick. Within a few years two more companies erected similar plants, and in 1925 a fourth company began operations. Thus there arc today four lime companies at Ste. Genevieve, all operating at high capacity and with a gradually increasing production. Geology The limestone of Ste. Genevieve County, which has the chemical composition and purity required for lime burning, is the upper 80 or 90 ft, of Spergen formation. The Missouri Bureau of Geology, in its geological map of Ste. Genevieve County, gives the Spergen formation a total thickness of 160 ft., but only the upper portion is pure enough for lime burning, with an average raw-stone analysis of 98.7 per cent CaCo3; 0.35 SiO2; 0.3 R2O3; 0.35 MgCO3; 0.3 SO3. The Spergen, one member of the Meramac group, Iowa series, of the Mississippian system, underlies an area approximately 12 miles long north and south, by 5 miles wide east and west (Fig. 1) in Ste. Genevieve County. The entire eastern two-thirds of Ste. Genevieve County is sedimentary, the formations dipping gently toward the east on an average grade of about 4 per cent, or 2 1/2°. Thus the outcrops of the various sedimentaries run approximately north and south. A line running west from Ste. Genevievc crosses the outcrops of the sedimentaries, ranging from the top of the Mississippian to the Ozarkian geological system in a space of 12 miles. Thus with the gentle dip of 2 1/2° the 160-ft. thickness of the Spergen formation stretches out in its outcrop so that it is approximately one mile wide. The early quarry and lime developments at Ste. Genevieve mere in this easily obtained surface rock of the Spergen outcrop.

Jan 1, 1938

By H. C. Goodrich

At the Utah Copper mine, steam shovels were first used, in 1906, for the removal of overburden, and in June, 1907, for the mining of ore. Prior to 1907, the ore came from underground development work and from stoping. The use of the steam shovel gradually increased until 1914; since then all material, both overburden and ore, has been loaded by power shovels operating on the various benches of the mine. During these 19 years, there has been a great improvement in open-pit mining, largely in the use of equipment having greater capacity and dependability. The history of the development of mechanical shovels and related equipment, as affecting the Utah Copper Co., would be interesting and would justify considerable space; but of greater interest at this time is the operation of the equipment now in use. The Utah Copper mine is in a great deposit of copper-bearing monzo-nite, which occurs as a prominent headland of the Oquirrh range of

Jan 1, 1925

By J. S. Levine

A series of four displacement experiments has been run in a large alundum core. Flow potential distribution in each liquid phase was measured continuously through oil-wet and water-wet capillary barriers, using a null-point pressure balance, and saturated distribution was determined from electrical resistivity measurements. A method has been devised for calculating the effective (and relative) permeability to each phase at any time and at any point in the flow system for the transient displacement process. Relative permeability points have been calculated and average curves drawn for oil displacing water and water displacing oil. The values of dynamic capillary pressure have also been determined. Data have been substituted into the generalized Buckley-Leverett fractional flow equation and a comparison between actual and theoretical breakthrough recoveries is shown. Displacement experiments have been performed in the laboratory for many years in an effort to determine the effect of a number of parameters on multiphase flow in porous materials. In the literature of petroleum technology, Wyckoff and Botset&apos; were among the first to develop the relative permeability concept and to show experimentally that oil, gas, and water can flow simultaneously through a porous medium in a precise way governed by certain fluid properties, by the relative saturation of each fluid, and by characteristics of the medium itself. Somewhat later, Leverett2 and Hassler, Brunner, and Deahl11 developed some theoretical concepts of capillary behavior in porous materials. The work of Buckley and Leverett3 presented a theoretical analysis of immiscible liquid displacement based on the equation of continuity and the Darcy equation for viscous flow of each fluid. Subsequent investigators attempting to verify the Buckley-Leverett equation have, in general, deliberately neglected the capillary pressure term or have performed experiments in such a manner that it could be neglected. Terwilliger, et al4 in 1951, obtained agreement between theory and experiment for the constant rate displacement of water by gas in a long vertical sand column. Rapoport and Leas5 have investigated the

Jan 1, 1955

By J. Nagy, D. W. Mitchell, E. M. Murphy

In 1957 research was initiated in the U.S. Bureau of Mines experimental coal mine near Pittsburgh, Pa., to study factors affecting foam generation and transport, to evaluate the effectiveness of high-expansion foam for controlling mine fires, and to develop techniques for applying the method under U.S. mining conditions. These investigations showed that high-expansion foam containing at least 0.2 oz of water per cu ft of foam is effective in controlling experimental underground fires burning coal, wood, and oil. Sometimes the fire was completely extinguished, but more often, it was brought under sufficient control to permit either a direct attack on the fire with a stream of water or loading of the hot material into cars. A progress report&apos; prepared in July 1958 summarized the initial achievements of the USBM experiments. Since then other phases of the foam-plug method for attacking fires have been studied in the laboratory and in the mine. Previous studies by British engineers&apos; of the foam-plug method for fighting mine fires indicated that high-expansion foam was effective in controlling experimental timber fires in an underground passageway. Their subsequent workx-1 pertained to the practical aspects of fighting large fires within a mining area with a foam-plug. CONTROLLING EXPERIMENTAL FIRES In the USBM tests foam was formed by spraying a dilute solution of a foaming agent on a metal or cotton net of 1/8 to 1/4-in. mesh. Air passing through the continuously wetted net forms bubbles of 1/2 to 11/2-in. diam and produces a honeycomb of foam that fills the passageway. Under the ventilating-air pressure, this light-weight plug moves forward through the passageways, around sharp corners, and over obstacles. as illustrated in Fig. 1. High-expansion foam was transported to a wood fire, an oil fire, and 13 coal fires. Figs. 3 and 4 show a typical coal fire before and after attack with foam. In 12 of the 15 experiments the fire was brought under control when the water content of foam was 0.2 oz or more per cu ft. A fire was considered controlled when the flames were quenched and observers could cross the area without wearing breathing apparatus or protective clothing. In the other three experiments, conducted when the water content was less than 0.2 oz per cu ft of foam, the flames were retarded but the fire was not controlled. Coal fires have been attacked successfully by foam introduced at points varying from 155 to 1010 ft from the fire. The time of burning in coal beds 10 in. thick ranged from 11/2 to 5 hrs or more. Most of the experimental fire beds were 15 ft in length. However, in one experiment a floor fire 25 ft long and 5 ft wide was constructed $5 upwind from another fire 15 ft in length; in another instance, the fire was 100 ft long and 5 ft wide. Foam was applied to the fires for periods ranging from 7 to 36 min. The time required for foam application depends on the extent of the fire, time of burning, water content of foam, foam velocity, and degree of fire control desired. In addition to the coal fires, foam was transported to a fire covering 45 sq ft, produced by 15 gal of oil burning in metal trays on the floor. The foam extinguished the oil fire in about 1 min. In one other test, the burning of 1100 lb of dry sawmill slabs stacked in open cribs 4 ft high and 16 ft long was brought under control by foam in 2 min. Composition of Gases in Return Air: In several of the experiments samples of the return air from fire zones were collected; composition of the atmosphere before, during, and after foam application was then determined. Because of condensation in the relatively cool sampling tube, the amount of water vapor was not determined. Analyses showed that concentration of carbon dioxide and combustible gases increased as the foam began passing over the fire. This resulted from the decrease in the volume of air when foam generation started and from the formation of gases when water reached the fire.* The quantity of gases generated would not be greater than that from an equivalent amount of water applied directly to the fire. The highest total concentration of combustibles (CO, CH1, and H2 mixture) obtained during the experiment was about 2 pct; this occurred 6 min after foam reached the fire. This atmosphere was nonex-plosive, but calculations show that if the air flow were reduced to about 5 fpm and if the rate of gas liberation from the fire remained constant, the mixture would be explosive. The use of foam on a fire in all probability would affect the normal ventilation of a mine. If the mine is gassy, this factor must be carefully considered before the foam is applied. APPLICATION OF THE FOAM-PLUG TECHNIQUE IN MINES Equipment and procedures for applying the foam-plug methods must be adapted to the prevailing conditions at a particular mine. Some factors to be considered in developing equipment are: size or extent of the mine, dimensions and number of entries, ventilation system, mining methods, haulage facilities, availability of water, amount of methane liberated, and existing fire-control apparatus. • In most experiments the initial air velocity of 200 fpm decreased to 50 to 100 fpm as the foam plug increased In length.

Jan 1, 1961