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By Robert Ammon

The milling practice at Mascot, at present, consists of dry crushing to % in., jigging, fine grinding, and flotation. The ore arrives at the mill from two mines, No. 1 mine shaft being located in the mill proper, and No. 2 mine shaft being situated about 2000 ft. east of the mill. No. 2 mine is served by a coarse-crushing plant and aerial tramway to the mill bins. In jigging the No. 2 mine ore, 50 per cent. of the mill feed is rejected by the jigs as a jig tailing ranging from 5/8 in. to 10 mesh; the balance of this ore, after removing a jig concentrate, is ground to flotation feed. In jigging No. 1 mine ore, approximately 80 per cent. of the mill feed is rejected as a coarse jig tailing, and the balance is ground to flotation feed without the removal of a jig concentrate. The difference in the treatment is due to the fact that No. 2 mine ore carries a higher content of zinc than that from No. 1 mine. Of the total concentrates produced, 30 per cent. is made on jigs and 70 per cent. by flotation. At present, 1350 tons of No. 2 ore and 800 tons of No. 1 ore are treated each 24 hours. The recoveries for a typical week are as follows: Heads.......................... 3.55 Jig concentrates................ 61.30 Jig tails......................... 0.80 Flotation concentrates........... 60.60 Flotation tails................... 0.09 Extraction..................... 86.50 The associated minerals are a straw-colored sphalerite, dolomite, calcite, pyrite, quartz, and a small amount of greenockite, which occurs as a stain and sometimes small crystals along with sphalerite, but is comparatively rare. Pyrite is present in small quantities, usually less than 0.5 per cent.; quartz occurs in bands a few inches thick. A complete analysis of the ore is as follows: PEr Cent. Per Cent. Calcium carbonate.............. 48.11 Zinc sulfide.................... 5.43 Magnesium carbonate........... 35.36 Cadmium sulfide............... Nil Silica.......................... 8.73 Lead sulfide................... Nil Alumina....................... 1.04 Iron oxide...................... 1.33 Total....................... 100.00

Jan 1, 1925

By M. Gensamer, B. Lustman

A recent paper described, by means of stereographic pole figures, the preferred orientations produced by cold-rolling low-carbon steel of automobile-body grade. It was found that the surface of the sheet exhibited a deformation texture significantly different from the texture possessed by the inside of the sheet. Efforts to produce the surface texture by buffing a sample exhibiting the normal or inside texture, using a copper lap, have not been successful. In view of the data collected by Kurdjumow and Sachs2, it was suggested that the surface texture might be the result of partial recrystallization of the normal or inside texture. The recrystallization texture of cold-rolled mild steel was studied by Goss3, but he did not use the pole-figure method, so that his conclusions were in need of verification. For these reasons the work reported in here was undertaken. It had two objectives: (1) to study the general behavior of cold-rolled mild steel on recrystallization below the critical temperature Al, and (2) to attempt experimentally a demonstaration of the genesis suggested above for the surface texture in cold-rolled steel. The material was that used by Gensamer and Mehl, cold-rolled to 97.5 per cent reduction in thickness. Fig. 1 is an X-ray pinhole photo-gram of this material, taken with the beam transmitted perpendicular to the plane of the sheet, using a molybdenum target at 35 kv. with no filter. It is typical of the inside texture, but in it the deformation texture is better developed than in the samples that showed a well developed difference between the inside and surface samples. Specimens from this sheet, which was 0.005 in. thick, were annealed for 5 min. at various temperatures in a lead bath. After about 0.0005 in. had been etched from both surfaces, X-ray photograms were prepared, using the technique described in the earlier paper1. Figs. 2 to 6 are photograms of the annealed sheet, at 566, 571, 577, 580 and 650" C. (1050' to 1202" F.), in order. Fig. 2 shows that 5 min.

Jan 1, 1937

By M. R. Geer, H. F. Yancey

Coal washing assumed an important role in the mining industry of the Pacific Northwest long before washing practice became firmly established in the Appalachian field. A Scaife washer was operated in the state of Washington in 1887; and by 1927, the fist year for which complete statistics were compiled, nearly a third of the state's coal production was washed, in comparison with only about 5 pct for the country as a whole. Mechanical preparation was adopted in Alaska in 1922 in a plant constructed by the United States Navy at Chickaloon to provide bunker coal for naval forces operating in northern waters. Many of the coal beds mined in Washington and Alaska contained more impurities than those mined elsewhere, and this circumstance contributed to the early interest in mechanical cleaning. A much more important factor, however, was the inclination of the coal beds. With steeply pitching beds, hand sorting by the miner at the face is impossible; consequently, all material must be loaded and dealt with on the surface, just as is now proving the case with mechanical mining of flat beds. Thus "full-seam" mining afforded the same stimulus to the early development of coal washing in the Northwest that it is providing under mechanization in the rest of the country today. Objeet and Scope ' This report is intended to provide a general summary of the status of coal-washing practice in Washington, Oregon, and Alaska—an area in which some of the coals mined are more difficult to wash than those mined elsewhere in the country. Detailed flowsheets of plants, statistical data on production, and cost figures have been omitted, largely for the sake of brevity but also because such information is principally of only local interest. On the other hand, considerable washability data have been included because they are Washington At present Washington leads all other states in the percentage of its total production that is cleaned mechanically, 82.1 pct in comparison with the national average of 25.6 pct in 1945.' The washing problems encountered in this state are highly variable because the mountain-building forces that created the Cascade Range caused such intense folding and faulting of the beds in some fields that the rank of the coal was increased to anthracitic, while in other fields more distant from the mountains the beds were relatively undisturbed, and the rank of the coal ranges down through subbituminous to lignitic. The complicated and burdensome schedule of sizes prepared in most coal fields has never been required by the Washington market. With but few exceptions, lump is prepared on a round-hole screen of about 3 in. size, egg is about 3 to 1 5/8 or 1½ in., and nut is 1 5/8 or 1½ in. to either 1 or ¾ in. Some mines screen no finer than ¾ in. and market a slack coal of that top size, but generally the coal is screened at about ¼ in. to give a stoker coal of ¾ or 1 to ¼ in., and a ¼ in. to 0 "buckwheat." During recent years the market for lump coal has largely disappeared, and lump is crushed to supplement the pro-

Jan 1, 1950

By Milton F. Goodrich

The idea of using water-only cyclones to scalp clean coal from the feed to other devices has recently been gaining in popularity. 1,2,6 An indication of this popularity is that water-only cyclone scalping is already employed in a few commercial plants. Water-only cyclones scalp the feed to a heavy media vessel in the Warwick plant,3 to Deister tables in the Rowland plant,4 and to a heavy media cyclone circuit in the Amherst Coal Company's McGregor plant.5 Despite this trend, we found surprisingly little information in the literature on the performance of water-only cyclone scalping. Therefore, we undertook a study of its technical merits, limiting our investigation to a water-only cyclone scalper in conjunction with heavy-media cyclones. We felt this to be a reasonable combination of equipment, since both processes operate well on the same size coal. First, we developed simple mathematical relationships revealing general characteristics of the system. Next, we compared the clean coal yield and quality obtainable by the system on four different coals with that obtainable by heavymedia cyclone circuit alone and water-only cyclones middlings recirculation circuit alone. This was done by using computer program to simulate the operation of the three systems. Finally, we considered the effect that implementation of the system would have on a typical preparation plant circuit. The System The system chosen for study [(shown schematically in Fig. 1)] comprises two stages: a primary stage (PI) consisting of water-only cyclones, and a secondary stage (P2) consisting of heavy media cyclones. The water-only cyclones separate raw coal at a low specific gravity, producing a relatively small amount of high quality clean coal and a refuse containing much good coal. This refuse is then recleaned in the secondary stage, where the sharper separation of heavy media cyclones, operated at a higher specific gravity, reclaims additional clean coal. Finally, the clean coal from both stages is combined as the system's clean coal, while the refuse from the secondary stage becomes the system's refuse output. Theoretical Analysis Two assumptions lie behind the following theoretical analysis. The first assumption is that partition curves can represent not only the performance of each stage but also the performance of the system as a whole. These curves show partition number as a function of specific gravity, where partition number is either the percent or fraction of a raw coal with a given specific gravity reporting to clean coal. The second assumption is that these curves are independent of the raw coal characteristics. At any specific gravity the system's partition number is solely a function of the partition numbers of each stage. For convenience, we use the following variables in describing this relationship: g = specific gravity, PS(g) = the partition number of the system at specific gravity g, P1(g) = the partition number of the primary stage at specific gravity g, P2(g) = the partition number of the secondary stage at a specific gravity g. For this system, we can show algebraically that the relationship is: [Ps (g) = PI (g) + (1- P1(g)) P2 (9)(1)] With this equation the system's partition curve can be determined if the partition curves of each stage are known. By definition, the partition numbers P1(g) and P2(g) must each be less than, or equal to, one. Placing these constraints on Eq. (1), we arrive at the relations: [Ps(g) % P1(g) and Ps(g) % P2(g)(2)] Thus, at any specific gravity the system's partition number is always greater than that of both stages. [Figure 2] shows typical partition curves. Curve I represents the primary stage, which consists of water-only cyclones operating at a separating gravity of 1.3. (The separating gravity or the

Jan 1, 1979

INTRODUCTION Mineral explorationists range throughout the world looking for mineral deposits which can subsequently be made into mines. Mines are not found! Mines are made, usually at great cost and not often without significant risk. The process of making a mineral deposit into a mine involves a great deal more than a complete appraisal of the deposit. The making of a mine demands detailed feasibility studies of many factors, any one of which, if shown to be negative, could prevent the deposit from becoming a successful mine. Total feasibility analysis involves at least three major fields of study: (1) economic analysis, (2) financial analysis, and (3) political and environmental analysis. Economic analysis is essential for proper decision making at many stages in the evaluation of a mineral deposit. It is the yardstick by which the proposed mining project and alter- native investment opportunities are measured. Economic analysis is concerned with costs and potential profits and can be applied at various stages in the evaluation of a mineral deposit to determine if the project merits the additional funds necessary to bring about each succeeding stage. The nearer any deposit approaches to becoming a mine, the more carefully the economic analysis should be made. Financial analysis has reference to how, where, and at what cost the investment funds can be obtained to finance the proposed mine project. Political and environmental analysis has reference to the many intangible factors that can affect the making of a successful mine. These factors loom larger today than ever before, due to the profusion of government regulations and various environmental constraints now imposed upon the mineral industry both by law and by the weight of public opinion. Ultimately, many of these intangible factors can be resolved back into the economic analysis as an added cost to mine development, but more and more often they have become insurmountable obstacles to any development of certain deposits. The role of financial analysis and the analysis of intangible factors must never be underestimated when making a feasibility study upon which to base investment decisions; however, the emphasis of this text will be upon economic factors. THE PARAMETERS OF MINERAL ECONOMIC ANALYSIS The two major elements of economic analysis are: 1) The potential earnings or annual generated cash flow of the project. 2) The investment costs necessary to realize such earnings. If the difference between the two elements is large enough, i.e., if the return on investment from the cash flow is high enough to attract investors, the project can be said to be financially feasible. The mineral economic analysis must take into consideration many parameters, some of which are known or knowable within certain limits, and others of which are unknown or unpredictable within certain limits and for which estimates must be made. Because both known and estimated parameters enter into the same analytical computations, it is important for the analyst to realize the nature of each when considering the results of an analysis. Some parameters such as the grade of mining block, although estimated, are said to be known within mathematically defined limits. Other parameters, such as the future price for minerals produced, etc., can only be estimated or "guestimated" within reasonable limits for which no precision can be assigned. Known or Knowable Mineral Economic Parameters Known or knowable parameters are: 1) Tonnage and grade of valuable minerals for the total mineral body and for individual mining blocks. 2) Metallurgical recovery of valuable mineral. 3) Minimum operating life of mineral de- posit at any specified production rate. 4) Royalty schedule, property payments, etc. 5) Sales-purchasing agreements, smelter schedules, marketing contracts, etc.

Jan 1, 1980

By J. K. Edgar

FOR a number of years the production and use of super-purity aluminum (better than 99.99 pct) has been steadily increasing. High-grade lots of such aluminum show certain outstanding characteristics not found in lower grades containing principally silicon, iron and copper as impurities. One such characteristic is the so-called "self annealing," that is, the recrystallization of the heavily cold worked metal at room temperature.' Interest therefore is aroused as to the effects of the impurities. It is well known that the amounts of silicon and copper which occur in aluminum of better than 99.95 pct are completely soluble, at least at elevated temperatures (above 200°C). Iron, on the other hand, is known to be fairly insoluble in solid aluminum. Thus, a knowledge of the limits of this solubility would be of considerable use in studying the characteristics of the high purity metal. All investigators have agreed that iron has very low solubility in solid aluminum. Rosenhain, Archbutt and Hanson2 reported traces of an aluminum-iron phase in the "purest samples of aluminum obtainable." Gwyer and Phillips3 reported that an iron constituent was visible under the microscope in metal containing ".0088 per cent iron." Roth1 found that in the solid state the solubility of iron in high purity aluminum was 0.01, 0.01, 0.017 and 0.022 pct at 225, 320, 500 and 600°C respectively. Dix5 working with alloys containing less than 0.03 pct impurities, concluded from microscopic and chemical analyses of eutectic areas that the eutectic composition was close to 1.7 pct iron. Dix placed the eutectic temperature at 655°C'. Archer and Fink,6 using high purity aluminum-iron alloys, found that the hypoeutectic liquidus intersects the eutectic horizontal at 1.7 pct. This is in fair agreement with the investigation of the aluminum-iron system by Gwyer and Phillips who placed the eutectic at 1.89 pct iron with a freezing point of 653° PREPARATION AND CHEMICAL ANALYSIS OF ALLOYS Electrolytically refined aluminum (99.991 pct aluminum) and a high purity aluminum-iron alloy (11.45 pct iron) were used in the preparation of the alloys for this investigation. To aid in controlling the iron content of the experimental alloys, an alloy of lower iron content was prepared first from these materials. The high purity aluminum was melted and heated to about 700°C in a plumbago crucible coated with alundum cement and the aluminum-iron alloy was then added. The alloy was thoroughly stirred, fluxed and skimmed and then chill-cast into notch bars. The analyses of the three materials are given in Table I. The alloys for the determination of the solid solubility of iron in aluminum were prepared by melting 600 g total of the aluminum plus the required quantity of the aluminum-iron alloy (No. 92531). These

Jan 1, 1948

By J. O. Edstrom, G. Bitsianes

Parabolic rate constants were determined for the formation of wiistite by the solid state reaction between magnetite and iron. The reaction was diffusion controlled and inert marker studies indicated that the mass transport through the wiistite layer was accomplished by means of iron migration. Relationships between rate constants and self-diffusivities are discussed. The transport capacity for iron through dense wustite layers was found to be sufficient to carry on reduction, even in gaseous reduction processes. REDUCTION of iron oxides has been the subject of many investigations. Most of the work, however, has been done on rather crude material and there have been difficulties in correlating the data on a quantitative basis. The reverse process of the oxidation of iron has been studied more thoroughly and usually by starting from the pure metal. As a result, the mechanisms of oxidation are now fairly well known.It has generally been found that the oxidation of metals follows the so-called parabolic 1aw;Q hat is, when the oxides are formed as dense layers, their thicknesses are proportional to the square root of the reaction time. This parabolic law follows directly from Fick's first law of diffusion The flux, J, is defined as the quantity of diffusing substance passing per unit time through unit area of a plane at right angles to the direction of diffusion. This flux is proportional to the concentration gradient of the diffusing substance, and the factor D, known as the diffusion coefficient, is introduced as the proportionality factor with dimensions of (length)'/time. If the flux is measured as the increase of thickness of the oxide layer (Ax) per unit time (t) and the assumption is made that the concentrations of the reactants at the phase boundaries of the layer are independent of time, Eq. 1 reduces to Integration of this expression yields the parabolic function ax = k . t. [31 The product of the diffusion coefficient and the concentration difference across the oxide layer is included in the constant k. When iron is exposed to a highly oxidizing atmosphere, the oxide phases are formed normally in a topochemical fashion, i.e., the interfaces between the phases maintain parallel positions to the original surface of the specimen. Above 570°C, the phases are orientated usually in the order of iron, wiistite, magnetite, and hematite, a condition in conformance with the requirements of the Fe-0 system. Below 570°C, the wiistite phase is unstable and has not been found as a normal constituent in the oxidation products of iron. There is a strong probability that continuous layers of the solid oxides are formed as the specific volumes of the phases increase in direct order from iron to hematite.' , ' Regarding mass transport through dense solid oxide layers, Wagner' has postulated that diffusion in oxides generally may be interpreted as migration processes of ions and electrons. There must be normally a concentration gradient across the growing oxide layer for diffusion to occur. This condition requires that there be deviations from the ideal stoi-chiometric composition of the oxide and accordingly deviations from the strict order of an ideal lattice. Such "lattice defects" include interstitial ions, cation and anion vacancies, quasi-free electrons, and electron holes and are decisive for all migration processes.' Cations, anions, and electrons all may have some mobility in the oxides but the movement of one of the particles generally far exceeds that of the others. Previous work', 1,2,7 has shown that during the oxidation of iron the migration through the wiistite layer, and probably also the magnetite layer, is confined to the movement of iron ions. Migration through hematite, however, has been found to take place by oxide ion movement. These behaviors are in direct agreement with the type of vacancies present in the respective oxides. Reduction of Iron Oxides The gaseous reduction of iron oxides, like the oxidation of iron, takes place in a topochemical fashion at distinct interfaces between the appearing phases."-'" Porosities might be expected in these reaction products, since their specific volumes are less than that of the starting material. Frank and van Der Merwe" have shown, however, that it is theoretically possible for nonporous layers to form if the degree of misfit in specific volume is less than about 15 pct. These aspects on porosity formation

Jan 1, 1956

By O. W. Simmons, L. W. Eastwood, C. M. Craighead

Binary alloys of titanium with silver, lead, tin, nickel, copper, beryllium, boron, silicon, chromium, molybdenum, manganese, vanadium, iron, and cobalt were studied. One-half-pound ingots of the alloys were prepared in an arc furnace, employing a water-cooled copper crucible, an argon atmosphere, and a water-cooled tungsten electrode. The half-pound ingots were fabricated by forging at 1700°F in air to 1/4-in. slab, followed by hot rolling at 1450°F to 0.060-in. sheet. Tensile properties, minimum bend radii, hardnesses, response to heat treatment and aging treatment, and phase relationships were determined for these alloys. EARLY in 1947, as one phase of the evaluation of materials for Air Force Project RAND, Battelle successfully arc melted Bureau of Mines titanium and obtained a few basic properties of unalloyed titanium and several titanium-base alloys. The low density, excellent resistance to corrosion, and high tensile properties of these materials stimulated a great deal of interest among metallurgists and designers seeking better materials of construction. As a result of this early work, Battelle, under contract with the Air Materiel Command, Wright-Patterson Air Force Base, has continued extensive studies of titanium alloys. The result of the first year's work is described in three papers. The present one deals with binary alloys, the second with ternary alloys, and the third paper with quaternary alloys. Melting Method The arc furnace for melting titanium and other refractory metals was developed at Battelle Institute under Air Force Project RAND. During the present work, considerable improvement has been made in the design and operation of this furnace for making small ingots of titanium and titanium alloys. The improved furnace design is illustrated by fig. 1, which shows a cross-sectional view with the dimensions of various parts and their relationship to one another. The essential features of the furnace are the inert argon atmosphere, the water-cooled copper crucible, and the water-cooled tungsten elec- trode. The melting crucible is formed by spinning 0.064-in. annealed copper sheet. A groove for an O-ring seal is spun into the flange of the crucible. A fiber gasket between the crucible flange and the water-cooled brass cover provides electrical insulation. The brass cover is fitted with a sight glass, an opening for the water-cooled electrode, and a tube through which the material is charged. Direct current is used for melting, with the positive electrical terminal connected to the water jacket and the negative terminal to the electrode. A typical heat is made in the new furnace as follows: Approximately, 0.20 lb of titanium under 2 mesh per in. and the alloy addition, if any is to be used, are placed in the bottom of the crucible. The cover is clamped on the crucible so that the O-rings make a tight seal. The remainder of the charge is placed in a glass bottle, and this is connected by a rubber hose to the charging tube of the furnace. The crucible and the charge are evacuated with a mechanical pump to a pressure below 50 mm of mercury, or less if desired, and held at this reduced pressure for 5 min. Hot water is circulated through the jacket during the evacuation period to assist in outgassing the crucible. Following the evacuation, tank argon of 99.92+ pct purity is flushed through the crucible for 5 min and then adjusted to give a positive pressure of 1 to 2 psi. During subsequent operation, an argon pressure regulator introduces only enough argon to compensate for the small leakage which occurs through a mercury trap. Re-evacuation and reflushing the melting chamber with argon produced a negligible reduction in contamination. Two 550-amp generators, connected in parallel, constitute the power source. Normally, about 800 amp are used for melting alloys which are not extremely refractory. The generators are set for 90 v open circuit and 200 amp. The arc is struck and the electrode is withdrawn to produce an arc of 23 to 26 v. This voltage is maintained while the electrode tip is moved slowly around a circle 1 in. smaller than the diameter of the ingot. The current

Jan 1, 1951

By T. C. Atchison, W. I. Duvall

Over the past ten years a series of investigations have been conducted to determine some of the pnysical processes involved in breaking rock with confined concentrated charges. Detailed discussions of many of these investigations have been published elsewhere.1-6 Laboratory experiments made by other investigators using Hopkinson pressure-bar techniques have shown that solid materials are fractured in tension by the reflection of an incident compressive stress pulse at a free surface.7-12 In these tests a small charge of explosive is placed in contact with one end of a bar. Detonation produces some plastic flow and crushing of the bar near the charge and generates a compressive stress pulse that travels along the length of the bar. At the free end of the bar the compressive stress pulse is reflected back into the bar as a tensile stress pulse. If the tensile strength of the bar is exceeded during this reflection process, a tensile fracture normal to the length of the bar is produced, and the broken end of the bar moves forward with a constant velocity equal to the average particle velocity trapped in the broken fragment. The new surface formed by the fracture becomes the new free end of the bar that reflects the remaining portion of the incident compressive stress pulse. This process is repeated any number of times until all of the incident stress pulse is reflected. Hino has demonstrated this kind of breakage for three rock types—marble, granite, and sandstone." He has defined a blastibility coefficient, B, as the ratio of compressive strength, C, to tensile strength, T, thus: C The quantity B is the maximum number of slabs that can be produced by reflection breakage. Normally fewer slabs are produced because of loss of energy as the stress pulse travels through the rock. Fig. 1 illustrates reflection-type fracture for a triangular compressive stress pulse. The number of slabs produced by the reflection breakage process is the first whole number less than the ratio of the peak stress of the incident pulse to the tensile breaking stress of the solid. Thus the number of slabs is given by the thickness of each slab is given by and the total length of rock broken is given by N = number of slabs S - peak stress in incident pulse T = tensile strength of the rock F = fall length of incident stress pulse h = thickness of each slab D = total length of rock broken During the reflection process the particle velocity at the free surface is twice the particle velocity in the incident stress pulse. Thus the velocity with which the broken fragments move forward is given by where v, = velocity of broken fragment, and v = average particle velocity contained in that portion of the incident pulse trapped in the broken fragment. Results of these laboratory experiments cannot be applied directly to rock blasting where the explosive charge is placed in a drillhole. In laboratory tests the charge is unconfined and in contact with the rock in only one direction. In a drillhole additional confinement is offered by the rock surrounding the charge and by the stemming placed above it. This additional confinement may be enough to allow the explosive gases to do additional work on the rock during their expansion. Other writers have discussed possible effects of gas expansion'on rock breakage.13-10 However, very few experimental data are available to determine to what extent expansion of the gases is responsible for rock fragmentation. The USBM has studied the physical processes involved in breaking rock with confined concentrated charges by using simple crater tests breaking to one free surface. Crater tests have been performed in four rock types: granite, sandstone, marlstone, and chalk. Table I gives some physical properties of these rocks. Fig. 2 shows plan and section drawings of two typical crater tests and illustrates some of the test variables measured. For these tests the charge was placed at the bottom of the drillhole and primed with an electric cap. The hole was stemmed to the collar with sand and the charge detonated. Size and shape of the crater were measured after it was cleared of broken rock. As a given charge size was buried deeper in a drillhole, the crater depth usually was equal to or

Jan 1, 1960

By F. P. Dewey

The subject of this paper does not embrace the consideration of ways and means for the increase of analytical accuracy, or the question what could or should be attained in that direction. I desire simply to call attention to the degree of accuracy exhibited in actual every-day practice. In estimating this, little weight will be given to the evidence afforded by the agreement of duplicate or multiple determinations by the same chemist; for I am convinced that such agreement is a delusion and a snare. Nor will special importance be attached to the agreement of two or even three analyses in special cases, or to the agreement between two methods practiced by the same analyst. I propose to compare the results obtained by several chemists, working upon the same sample and by various methods, in order to exhibit, as I have said, the actual condition of practice. The available material for illustrating this phase of the question is unfortunately scanty; but something has been done; and I hope, by calling attention to some of the work in this line, to stimulate further work in the same direction by inducing others to prepare suitable samples and submit them to various chemists who are competent and willing to make the necessary determinations, and fully describe the methods they employ. I draw most of my illustrations from the Transactions of the Institute, the Proceedings of the Association of Official Agricultural Chemists, and from personal experience. Manganese in Steel.—In May, 1881, Mr. William Kent presented a paper to the Institute entitled " Manganese Deter-

Jan 1, 1897

By AIME AIME

A JOINT meeting of the Canadian Institute of Mining and Metallurgy and the American Institute of Mining and Metallurgical Engineers was held at Spokane, Wash., and Cranbrook and Kimberley, B. C., on Oct. 13 to 16, 1926, with a total attendance at one or another or all of the meetings of about 300 members, ladies and guests. At the opening meeting, on Oct. 13, in the Davenport Hotel, Spokane, about 175 were present. The morning was occupied with addresses of welcome by M. H. Sullivan for the Columbia Section, who called on Mayor Charles A. Fleming, President John F. Davies for the Chamber of Commerce, and E. R. Hannibal, president of the Associated Engineers, each of whom heartily welcomed the visitors. Philip N. Moore, former president, and special representative of President Samuel A. Taylor, replied for the A.I.M.E., and Lewis Stockett, president, and Charles A. Banks, chairman, British Columbia Division, Canadian Institute of Mining and Metallurgy, and Henry M. Parks for the visiting A.I.M.E. sections, made appropriate responses on behalf of those they represented. The members and ladies had luncheon with the Associated Engineers. Philip N. Moore addressed the assembly on "The Opportunity of the Engineer"; an excerpt from his address appears on another page. At the afternoon meeting, with Lewis Stockett in the chair, Frank M. Smith was asked to make a brief statement of the proceedings of the Silver Producers Association in Denver. Supplementing his talk by placing a well-prepared statement on "Silver and the Plan of the Royal Commission on Indian Currency and Finance" in the hands of those present, he narrated what had been undertaken by the Silver Producers Association and what they hoped to accomplish.

Jan 1, 1926

J. MURRAY RIDDELL, Chairman Guy N. BJORGE, Vice-chairman Alluvial Mining CLINTON BERNARD , H. R. NORSWORTHY BENERE H. GRANT O. B. PERRY

Jan 1, 1938

By K. G. Davis, P. Fryzuk

VALUES for the equilibrium distribution coefficient, ko, are required for solidification studies. The procedure generally adopted1-3 is to progressively solidify alloy rods over a range of growth rates, under conditions where a high degree of convective mixing is present in the melt. The measured effective distribution coefficients are plotted as a function of growth rate, and the extrapolated value at zero growth rate is taken to be the equilibrium ko. The method is limited by the experimental difficulty in obtaining steady growth rates less than approximately 10-4 cm per sec. Extrapolation from rates greater than this can lead to inaccuracy. In the present investigation, an alternative method for determining ko, previously used by the authors for the Ag-Sn system,' was applied to alloy of thallium in tin. Specimens were grown under conditions where convective mixing was absent, and experimental concentration-distance curves in the region of the initial transient were compared with curves calculated for trial values of ko. The ko which gave the best fit was taken as the correct value. To obtain the calculated curves, a value for the diffusion constant in the liquid near to the melting point was needed. This was determined by the capillary-reservoir method. Diffusion Constants. Details of the experimental procedure used by the authors in the capillary-reservoir technique have been described in previous papers,5'6 the only variation being in the use of Tl204, a ß emitter, as tracer, in contrast to previous work where ? emitters were used. The activity of a ß-emitting sample, as measured in a Geiger counter, is proportional to the surface area. To measure the concentration of thallium along the capillary after diffusion had taken place, the solidified capillary (1 mm diam) was cut into 2.5-mm lengths and each length compressed between parallel steel plates to form a disc approximately 5 mm in diam. The discs were placed in a flat-bottomed dish, covered by a thin lead shield with a central aperture, and counted in a 0 counter with fixed geometry. Each disc was etched lightly in a dilute solution of nitric and hydrochloric acids a short time before counting, to eliminate sur-

Jan 1, 1968

By Joseph W. Richards

ALTHOUGH bad grammar, the above query is probably, at the present moment, good sense. Why was the Institute started and why does it continue to exist? The small group of men who worked out the original idea felt keenly the necessity of mining engineers and metallurgists getting together in order to discuss their professional problems. That necessity still exists but it does not strike the present generation of engineers so keenly because they have always had the Institute: they do not know what it is to be without it. The Institute is a great graduate school of mining and metallurgy. All of its members who have something to say worth saying are the teachers; all the rest may be learners or critics. No university or technical school can match the experience and information which the authors of its papers possess, no institution of learning has a body of potential listeners so extensive and so vitally interested.

Jan 1, 1921

By Ralph Nichols

THE mines are near the town of Gilmore, in the Texas mining district. This district was organized in 1880. The present producing mines are near the terminus of the Gilmore & Pittsburg railroad. This railroad connects with the main line of the Oregon Short Line at Armstead, Mont. The distance from Gilmore to Armstead is 70 miles. The town of Gilmore is near the head of the Lemhi river at an elevation of 7,000 ft., while the mines are from 400 to 1,300 ft. higher. The range of mountains upon which the mines are situated forms the water shed for the Little Lost river on the west and the, Lemhi river on the east. The winters are severe and the snow fall is heavy, but the mines ship right through the year. There is an abundance of timber suitable for all mining purposes in the immediate vicinity. Water is piped into the camp from a lake about 2.5 miles distant. This lake is about 1,400 ft. above the town of Gilmore. The water is exceptionally good for boiler and domestic purposes. The mining companies now producing are. the Pittsburg Idaho Co., Ltd. ; the Latest Out Mining & Smelting Co:, Ltd. ; and the Gilmore Mining Co., Ltd. Besides the above properties, there are a number .of prospects in different stages of development. The producing mines of the district are fissures in limestone, with a general northeast-southwest course .and a dip to the west of from 40° to 70°. These fissures are at about right angles to the bedding of the limestones. The veins are regular and the ore shoots strong, both laterally and in depth, so far as developed. The ore shoots have a rake or pitch to the south. The character of the formation is shown in Fig. 1, a section of the Latest Out mine.

Jan 11, 1913

By Raymond L. Orr

Calorimetric measurements have been made of the heats of solution of gold axd indium in a number of liquid tin-rich alloys at a temperature of 705°K. Relative partial molar enthalpies of gold were determined as a function of dilute solute concentrations in the binary Au-Sn system and in the ternary Ag-Au-Sn and Au-In-Sn systems. Relative partial molar enthalpies of indium were determined for dilute solute concentrations in the In-Sn, Ag-In-Sn, and Au-In-Sn systems. Through application of the interaction coefficient concept introduced by Wagner and more recently extended by Lupis and Elliott, ualues have been obtained for the Au-Au, Ag-Au, Au-In, and Ag-In enthalpy interaction coefficients in liquid tin. The results are inte@reted in terms of solute atom distributions through comparisons with the predictions of dilute solution models. INTEREST in the thermodynamic behavior of dilute liquid alloys stems from two primary sources. From a practical viewpoint, it is often of importance to know or to be able to predict the effect that one solute will have on the thermodynamic properties of the other solutes in a multicomponent system. Studies of dilute solutions may also be rewarding from the theoretical point of view. Some of the difficulties arising from the more complex interactions possible in concentrated solutions are avoided, leading to easier interpretations in terms of solutibn models and bonding energies. For example, in a binary alloy the limiting values of the partial molar properties of the solute represent the case for which each solute atom is completely surrounded by atoms of the solvent, and no other interactions are possible. Solute - solute interactions in dilute solutions are conveniently treated by the interaction coefficient concept of ~agner.' Using a Taylor series expansion for the logarithm of the activity coefficient, In yi, of a component, i, in a solution consisting of dilute solutes with atomic fractions xi, xj, xk, and so forth, in a solvent, s, and neglecting the second- and higher-order terms, Wagner obtained the expression: Where Wheyz is the limiting value of yi in the pure solvent, s. The interaction coefficients, e:, e:, and so forth, are defined by: Since yi is related to the excess partial molar Gibbs energy of component i by GfS = RT In yi, E: is re- ferred to as the "Gibbs energy interaction coefficient". Lupis and Elliott~ extended Wagner's treatment to the relative partial molar enthalpy, aHi, and the excess partial molar entropy, qS. They defined the enthalpy interaction coefficient as: and the excess entropy interaction coefficient as: leading to expressions similar to Eq. [I]: The three interaction coefficients are related by? The self-interaction coefficients, E:,qf,and at, represent the effects of interactions between atoms of component i in solvent s; €3, q{, and 03 represent the effects of interactions between i and j atoms in solvent s; and so forth. From the Maxwell-type relationships between partial molar quantities, it may be shown1'2 that e{ = ej, r}\ =tj), and ai - crj. Experimental determination of the interaction coefficients requires extremely precise measurements of the appropriate properties as functions of solute concentrations in very dilute regions. Few such data exist for intermetallic alloys, even for binary systems, because of experimental difficulties. This is especially true with respect to enthalpy data for dilute alloys. Examination of the compilation of data for binary alloys by Hultgren, Orr, Anderson, and ~elle~~ reveals that determination of limiting values of &fii(Xi =0) often requires extrapolations to infinite dilution from xi = 0.05 to 0.10. Direct measurements of the enthalpy interaction coefficient, qi, for multicomponent systems are virtually nonexistent. This quantity is of course subject to direct calorimetric measurement, but its determination must be limited to cases where high experimental precision is possible. The heats of solution of gold and indium in liquid tin can be measured with relatively high precision, which makes determinations of 73 involving these metals as the measured solutes experimentally attractive. This paper presents the results of such determinations of the Au-Au, Ag-Au, Au-In, and Ag-In enthalpy interaction coefficients in liquid tin. MODELS FOR DILUTE-SOLUTION BEHAVIOR Random-Solution Behavior. In the quasi-chemical treatment of solutions, only nearest-neighbor bonds are considered, and a composition independent value

Jan 1, 1967

By D. N. Dietz

A method for determining average reservoir pressure is presented, which is simpler to apply than that devised by Matthews. Brons and Hazebroek. For bounded reservoirs, identical results are obtained if stabilized-flow conditions prevail. The present method yields inferior results in the transient state. The method can, with a slight modification, also be used for water-drive reservoirs. INTRODUCTION In the method proposed by Matthews, Brons and Haze-broek' for determining average reservoir pressure in a multi-well reservoir, the cumulative production (the production time of each well) enters twice: once when the buildup is plotted against In (t + t)/At to arrive at p*, and a second time when the correction, p*-p is determined with one of the several formulas for differently shaped drainage areas. Once a steady state has been attained, however, the previous production history should be immaterial. The same pressure distribution could have been arrived at after different cumulative productions of the individual wells. In principle, therefore, it should be possible to determine average pressures without referring to cumulative productions. In the following paragraphs an expression is presented for the difference between the pressure in a producing well and the average pressure of its drainage area. Then the build-up time needed to overcome this difference is indicated. CASE OF A CIRCULAR DRAINAGE AREA AND A CENTRAL WELL PRESSURE DISTRIBUTION BEFORE SHUT-IN The general differential equation of radial flow (see Muskat, Eq. 10.2) may be written In the steady state, the rate of production of the well is equal to the rate of expansion of the fluid contained in the the drainage area; thus

Jan 1, 1966

By K. Barron

The glass insert stressmeter, or photoelastic stressmeter, is an instrument designed to determine stress changes occurring in rocks. It has several potential advantages over other such devices in that it is a biaxial device, it is simple and it is cheap to make. The object of this study was to assess the behaviour of the meter under biaxial loads and to examine some of the problems associated with measurement and interpretation of the fringe pattern. This assessment has been carried out by comparing theoretical and laboratory behaviour of the meter. It has been shown that: a) There are certain optimum measuring points in the meter at which measurements should be made for best accuracy. The position of the points depends on the ratio, n, of the biaxial stresses. b) The meter&apos;s sensitivity can be assumed to be independent of the rock modulus E provided that the rock modulus is < 2.5 x 106 psi and not, as previous workers have assumed, if E < 5 x 106 psi. c) There are several methods of separating the principal stresses; some of these are unsatisfactory. A new method is proposed and discussed. d) The meter sensitivity decreases as the ratio of the biaxial stresses approaches unity; the accuracy thus also decreases. e) The axes of symmetry of the fringe pattern give an excellent indication of applied stress direction. f) Laboratory calibrations are in good agreement with theory. The glass insert or photoelastic stressmeter is an instrument designed to determine stress changes occurring in rocks. It has several potential advantages over other such devices in that it is a biaxial device, it is simple and it is cheap to make. This stressmeter has been described in detail.l,2 Basically it is a hollow glass cylinder that has at one end a light source and filters producing circularly polarized light. The glass cylinder and source are grouted into a borehole in rock and the face of the cylinder is viewed through a quarter wave plate and an analyser. If, after installation, a stress change occurs in the rock, then a distinct colored isochro-matic fringe pattern is seen in the meter. The fringe order at any point in the meter can be determined by counting the fringes and measuring the fractional fringe orders by the Tardy compensation method. The relation between the fringe order at any point and the applied stress change can be determined either by calculation or by calibration. The object of this study was to assess the behavior of this stressmeter under biaxial loading conditions and to examine some of the problems of interpretation of the fringe pattern. THEORY Hiramatsu et a14 have determined the stress distribution in a hollow cylindrical inclusion in an elastic host material when subject to uniaxial stress. Using this solution, the biaxial case of stress p in the x direction and stress q in the y direction (Fig. 1) can be solved by superposition. This has been done in Appendix 1. It is shown that the principal stress difference (s1 -s2) at any point (r,?) in the cylinder is given by:

Jan 1, 1965

By R. Q. Shotts

Data are presented which show that fractions of varying densities from the same coals are oxidized at different rates by nitric acid. From oxidation data, the approximate quantity of "bright" and "dull" components may be calculated. Definite relationships between oxidation rate and rank are shown for density fractions from at least two bituminous coals. IN 1945, the Alabama State Mine Experiment Station initiated a program on the determination of fusain in certain coals of the state.1 Because of the lack of equipment and personnel necessary for microscopic examination, the chemical method2 was used exclusively. During studies of the distribution of fusain in fractions of various sizes and densities on a medium volatile coal, the author noted a higher rate of oxidation of the nonfusain part of the coal in the lower density fractions and the difference in rate became more marked the shorter the oxidation period used. Since durain and other "dull" components of coal have been reported3,4,5 to have higher densities than vitrain and the bright components, it appeared probable that the observed differences in rates of oxidation in the different density fractions were due to variations in petro-graphic composition. Such effects have been noticed by other workers.2,3 In order to obtain further evidence on this point, several Alabama coals of different rank have been separated into three or more fractions on the basis of density and their rates of oxidation determined, using the procedure specified for the determination of fusain, except that measurements were made at short time intervals as well as for the longer periods specified in the regular procedure. Significant differences in the rates of oxidation of the nonfusain part of the various fractions have been observed, and it is believed that studies of this type may make possible the identification of individual petrographic components or lead to methods for their semiquan-titative estimation. Properties of the Coals: Analyses and some properties of the whole coals studied are given in table I. The range in rank of the coals is not the widest possible within the state, but samples of the low volatile coals that occur in thin beds in the northeastern part of the state, were unavailable. The Clements bed is not, at present, considered workable. The coal is of medium volatile rank, quite friable, and very strongly swelling. Analyses of different size fractions show that the larger sizes are of higher rank than the smaller ones. Material passing the No. 50 sieve (all sieve sizes refer to the U. S. Standard Series) is close to high volatile A in rank, except for the size passing a No. 200 sieve, in which fusain is partially concentrated. Because of the variation in rank of the sizes of Clements coal, oxidation studies were made of two widely different size ranges. The + 1-in. material was crushed on rolls to about 1/8 in. and separated into five fractions by float-and-sink treatment in a benzene-carbon tetrachloride mixture. The — No. 4 + No. 8 and the — No. 4 + No. 50 size ranges tested overlap but the latter size was studied last with the expectation that it would include more oxidizable material than the first size. This proved not to be the case. Yields and properties of the specific gravity fractions oxidized are shown in table II. Milldale bed coal probably represents, as nearly as it is possible to do so, the "average" coal of the state. The material passing a No. 4 sieve was the only size range oxidized. It was divided into four specific gravity fractions. As in every case the heaviest fraction, which was essentially rock, was discarded. The Black Creek is the most extensive in area of

Jan 1, 1951

By R. Q. Shotts

Data are presented which show that fractions of varying densities from the same coals are oxidized at different rates by nitric acid. From oxidation data, the approximate quantity of "bright" and "dull" components may be calculated. Definite relationships between oxidation rate and rank are shown for density fractions from at least two bituminous coals. IN 1945, the Alabama State Mine Experiment Station initiated a program on the determination of fusain in certain coals of the state.1 Because of the lack of equipment and personnel necessary for microscopic examination, the chemical method2 was used exclusively. During studies of the distribution of fusain in fractions of various sizes and densities on a medium volatile coal, the author noted a higher rate of oxidation of the nonfusain part of the coal in the lower density fractions and the difference in rate became more marked the shorter the oxidation period used. Since durain and other "dull" components of coal have been reported3,4,5 to have higher densities than vitrain and the bright components, it appeared probable that the observed differences in rates of oxidation in the different density fractions were due to variations in petro-graphic composition. Such effects have been noticed by other workers.2,3 In order to obtain further evidence on this point, several Alabama coals of different rank have been separated into three or more fractions on the basis of density and their rates of oxidation determined, using the procedure specified for the determination of fusain, except that measurements were made at short time intervals as well as for the longer periods specified in the regular procedure. Significant differences in the rates of oxidation of the nonfusain part of the various fractions have been observed, and it is believed that studies of this type may make possible the identification of individual petrographic components or lead to methods for their semiquan-titative estimation. Properties of the Coals: Analyses and some properties of the whole coals studied are given in table I. The range in rank of the coals is not the widest possible within the state, but samples of the low volatile coals that occur in thin beds in the northeastern part of the state, were unavailable. The Clements bed is not, at present, considered workable. The coal is of medium volatile rank, quite friable, and very strongly swelling. Analyses of different size fractions show that the larger sizes are of higher rank than the smaller ones. Material passing the No. 50 sieve (all sieve sizes refer to the U. S. Standard Series) is close to high volatile A in rank, except for the size passing a No. 200 sieve, in which fusain is partially concentrated. Because of the variation in rank of the sizes of Clements coal, oxidation studies were made of two widely different size ranges. The + 1-in. material was crushed on rolls to about 1/8 in. and separated into five fractions by float-and-sink treatment in a benzene-carbon tetrachloride mixture. The — No. 4 + No. 8 and the — No. 4 + No. 50 size ranges tested overlap but the latter size was studied last with the expectation that it would include more oxidizable material than the first size. This proved not to be the case. Yields and properties of the specific gravity fractions oxidized are shown in table II. Milldale bed coal probably represents, as nearly as it is possible to do so, the "average" coal of the state. The material passing a No. 4 sieve was the only size range oxidized. It was divided into four specific gravity fractions. As in every case the heaviest fraction, which was essentially rock, was discarded. The Black Creek is the most extensive in area of

Jan 1, 1951