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By Horace Pops

ThE shape of the primary solid-solubility limits in copper-rich Cu-Si-Zn alloys has been discussed recently1 in terms of the atomic size effects and the electron concentration, e/a. Although these limits are parallel to lines of constant e/a near the Cu-Zn binary system, they deviate and become sharply restricted near the Cu-Si binary system. Hence, it appears that the primary solid solutions with higher silicon contents are restricted by the presence of the intermediate phases, [, X, and p, which follow them in the ternary system. owever, a similar restriction in solid solubility might also result from an interaction in the copper matrix between silicon and zinc. Evidence for such possible interactions has been observed, for example, in the Al-Mg-Si systion.A where a contraction of lattice spacings in the ternary solid solution appears to correspond to an electrochemical interaction between the two solute elements, magnesium and silicon. The observed contraction is associated with restricted solid solubility. It was also shown1 that the extent of the primary solid solubility did not follow the trends predicted by the lattice-distortion oncept." This again suggested that the solubility limits are strongly affected by the stability of the intermediate phases, I, 1, and X . In order to test the above hypotheses, lattice-parameter measurements of the fcc @-phase alloys were determined within the ternary system at a constant electron concentration (e/a = 1.30) and also at a constant zinc content. Alloys of predetermined composition were prepared from high-purity metals and melted and cast under reduced pressures of helium in quartz capsules. Since the weight losses were very slight and did not exceed one part in 2000, the nominal compositions listed in Table I were accepted without chemical analyses. The ingots were homogenized

Jan 1, 1964

By Murray F. Hawkins

One problem of reservoir engineering is the early estimation of the size of newly discovered reservoirs. Often these reservoirs are the expansion type in which sizeable pressure drops occur incident to the production of small fractions of the oil reserve, above the bubble point, and a material balance equation may be solved to find the reserve. The early form of this equation' considered only the compressibility of the oil. Later Hall2 showed that the compressible nature of the formation rock is a significant factor in production above the bubble point, and he presented a new equation which includes a formation (rock) compressibility term. This note presents a further extension of the material balance equation as applied to volumetric, unsat-urated reservoirs above the bubble point by including a term for the presence and compressible nature of the interstitial water. DERIVATION OF EQUATION During a decline in average reservoir pressure, ?p ?p, — p1 the initial reservoir volume V,, will decrease to some value V, due to the reduction in pore volume as a result of a decline in the hydrostatic pressure within the formation, and v, = V,, [l — cr ?p]......(1) The formation compressibility c, is expressed in terms of change in pore volume per unit pore volume per psi. Simultaneously the connate water within the formation expands from an original volume V,, = S,V,, to a new volume V,, V, = s,v,, [I + c,npl.....(2) The difference [Vp — V,] is the hydrocarbon pore volume at pressure p and may be equated to the volume of the remaining oil, or [N — N,] B, = V,,, [I - C~A;] - S,.Vpi [l + cwnjl.........(3) But the initial oil in place in terms of the initial reservoir volume, connate water, and formation volume factor is Ilividing Equation 3 by Equation 4 eliminates V,, to give,

Jan 1, 1956

By F. F. Aplan

INTRODUCTION The production of coal is a major part of the mining industry. In recent years the production of clean coal has exceeded 600 million tons per year and is reported to have reached 660 million tons in 1977. Some data on the magnitude and utilization for the year 1975 is given in Table I. In that year 648 million tons of soft coal were produced and of this 267 million tons or 41% approximately was produced by mechanical cleaning. Of the 374 million tons of raw coal sent to the preparation plant, 28.6% or 107 million tons was discarded as refuse. In the same year approximately 6 million tons of anthracite were produced and nearly all of this was cleaned. Table I also gives coal utilization statistics for 1975. Excluding the approximately 20% that goes to the coke ovens or to export, the vast bulk of our production is used for steam power generation and a lesser amount for captive production of industrial power and process steam. While much of the nation's coal (e.g. strip mined western, sub-bituminous coal) may be removed from the ground in relatively pure form by modest selective mining techniques, other coal requires preparation in order to remove ash and sulfur constituents and produce an acceptable steam or coking coal. Preparation is often a valuable adjunct to mining in that rather non-selective, low cost, bulk mining techniques may be used underground while the preparation plant is used to reject refuse at low cost above ground. The principal reason for the use of coal preparation is to convert raw coal into an acceptable commercial product by removing the high ash and high sulfur refuse material.

Jan 1, 1979

By Charles H. Snow

In vol. xxxviii. (1889) of the American Journal of Science, under the heading " Pseudomorphs of Native Copper after Azurite from Grant County, New Mexico," Mr. W. S. Yeates describes a most interesting mineral form. His account is, that a material having the appearance of native copper, but brittle and much lighter than native copper (its specific gravity being but 4.15), was submitted to him by the United States National Museum for examination. The substance was reported to him as occurring in masses, varying in weight from one ounce to seventy pounds, and composed of tabular crystals, variously grouped. The specimens, when submitted, ex hibited on the outride a covering of clay; and they were subse quently found to consist of particles of the same clay, intimately mixed with others of native copper, producing the aspect, without the weight, of pure copper. The forms of the azurite crystals, found at Bisbee, Arizona, having been suggested by the specimens, actual goniometer-measurements were made; and the two crystallizations were thus proved to be identical. To these statements I may be permitted to add some further in formation, derived from personal examination, and illustrated with a suite of characteristic specimens collected by myself (Fig. 2).* Grant County, New Mexico, in which these specimens were found, is a region noted for variety of mineral species. The sixth edition of Dana's Mineralogy mentions many of the silver, copper, and lead minerals, and also turquois, aragonite, wulfenite, barite, fluorite, vanadinite, etc. I have noticed also green and red garnet, molybdenite, native alum, smithsonite, bornite, etc., some of which presented peculiar modes of occurrence. The green garnet, for instance, was in masses formed of separate crystals, each of pea-size and not unlike that vegetable in tint and smooth opaqueness. The crystallization was

Jan 1, 1893

C. W. GOODALE, Butte, Mont. (member of the committee)- (communication to the Secretary*).-In the discussion of papers presented at the New York meeting of the Institute in February, 1915, and at the Arizona meeting of September, 1916, I referred to the activities of the Bureau of Safety of the Anaconda Copper Mining Co., and I am now submitting a few thoughts by members of the staff of that Bureau. TORGUS H. OAAS, Safety Engineer, Butte, Mont.-The report of the Committee on Safety and Sanitation has described different organizations and methods of promoting safety work. The effectiveness of these methods would aid the beginner in his campaign for safety in choosing the kind of organization and methods best adapted for his field. This can he obtained only by a discussion by those who have long promoted this kind of work, and have met with successful results. As the field is new, the different organizations promoting safety work have not established a common basis of rating for their fatal, serious and slight accidents, so that it is now difficult to get a comparative basis from which to judge the merits of the different systems. C. W. GOODALE.-The U. S. Bureau of Mines adopted, at the beginning, as a basis for its yearly summaries, the number of accidents. per 1,000 men employed, and it is now (quoting from its report for 1915, Technical Paper 168) making the following "classification of injuries:" "Many of the States now have. compensation laws, and in order to conform with their classifications, the bureau's classification of serious and slight injuries for 1915 is on a 14-day (2-week) basis instead of 20 days, as in previous years. The new classification of injuries includes three types, as follows: 1. Fatal 2. Serious (time lost, more than 14 days) (a) Permanent disability Total Partial

Jan 6, 1917

Discussion of the paper of C. A. HEBERLEIN, presented at the New York meeting. February, 1915, and printed in Bulletin No. 98, February, 1915, pp. 497 to 504. H. D. PALLISTER, State College, Pa. (communication to the Secretary).-Mr. Heberlein gives some valuable information concerning a branch of mining and metallurgy which has been neglected during the last few years. His results with fine grinding and Deister tables suggest the possibility of the mining and concentration of a number of low-grade cinnabar deposits. Can Mr. Heberlein give us some idea as to the loss of quicksilver during this concentration? The advances in the metallurgy of quicksilver have been few and far between in recent years. Most miners prefer to go on in the old way with the Scott furnace and the large brick condensers. The large brick condensers are expensive to build and to keep in repair, and on account of the thick walls are extremely slow in cooling the gases; and further, the joints between the bricks offer passages for the quicksilver vapors unless great care is used in building the condensers.. The escape of vapor at the top of the furnace is another source of loss and I see the people at the Oceanic mine have overcome this in a very ingenious way. I recently returned from the Terlingua quicksilver district of Texas, where the only operating company, the Chisos Mining Co., is operating a 20-ton furnace of the Scott type. They had considerable difficulty with getting a good extraction, clue to too high heating in the furnace with its .resulting diffIculties. Recently they have been running with ashpit and fire doors closed and have been able to keep the heat lower, thus getting much better results and practically verifying the results obtained at the Oceanic mine.

Jan 5, 1915

By William Metcalf

ONE of the first questions that naturally occurs to one who handles steel is," Why does steel harden ?" To answer this question the chemist and physicist have devoted much thought and experiment, and while they have enriched our store of knowledge of the nature and composition of steel, it cannot be said that they have yet solved the problem involved in its hardening. In this communication I desire to direct attention to a line of investigation yet new, in the hope that it may receive the attention of those who have the time and ability to pursue it. I refer to magnetism. There are so many parallels between the action of magnetism on iron and steel and their behavior under different processes of manipulation that it would seem as if there must be some connection between the forces at work in the different cases, if, indeed, there be not an identity of force in all of these cases. Magnetism, as we know it, may be merely the visible manifestation of forces, which may serve as a guide to lead us to an exact knowledge of the structure, the strains, and the resistances with which we have to deal in iron and steel. If a magnet be brought close to a pile of iron filings, the little pieces of iron will fly to the magnet and attach themselves to it, with their longer axes lying in the direction of the force. If molten cast iron be poured into an iron mould,-a chill,-the iron upon crystallizing will form needlelike crystals, with their long axes normal to the surfaces of the mould. If the iron is at the proper temperature and the quantity small, compared to the mass of the mould, the whole of it will be chilled, or set, in' the hard needlelike crystals

Jan 1, 1881

UNIVERSITY OF ALABAMA University, Ala Mining and Metalliaigical Society* G L Roberts, President T G Patterson, Vice-President B B Copeland, Treasurer K F Stark, Secretary G T Bator, Faculty Sponsor F H Crockard, Counselor

Jan 1, 1949

By C. Bohdanowicz

The most important decline in drilling activities during the past three years was in the Boryslaw-Tustanowice-Mraznica fields and at Bitkow in Eastern Poland. A small drilling increase in some fields of the Jaslo and Drohobycz districts, notably Wankowa, Ropienka and Rypne, cannot compensate in new production for the decline in the large fields.

Jan 1, 1934

By P. E. Queneau

This study confines itself to an examination of the two million square miles of the northern hemisphere which the geographer defines as the terrestrial Arctic, a land region in which the mean temperature of the warmest month is below 50°F-in effect, the area lying north of the tree line. The engineering and human problems there are of a different order of magnitude than is generally true of the sub-Arctic. Whether in Dawson Creek or Matanuska, in Mo i Rana or Skelleftea such problems are of the same general order as in Hibbing or Noranda. The same is not true in Kong Oscar's Fjord or Lancaster Sound.

Jan 7, 1961

By John M. Lovejoy

CURRENT production of petroleum on such a vast scale presents many interesting problems- the solutions of which are important not only to those directly interested in the business, but to the nation as a whole. The industry is relatively young. Until the discovers of the Lucas gusher at Spindletop in 1901 petroleum production had not exceeded 60,000,000 bbl. per year compared with 1,100,000,000 bbl. in 1936. In 36 years then, the modern oil business, production, transportation, refining, and marketing has developed. I will first present a brief history. of oil production as a background for consideration of the present situation.

Jan 1, 1937

By Eugene S. Perry

The oil and gas industries in Montana continued through the year of 1937 with but little change. No new fields were discovered, and production of old fields did not vary greatly. No outstanding wildcat wells were drilled. Perhaps most notable were extensions in the western part of the Kevin-Sunburst field. Acid treatment of wells in the Kevin and Pondera fields continued with gratifying results.

Jan 1, 1938

By J. Spotts McDowell

Preparation and Use.-Magnesite is an important refractory in open-hearth, heating, and electric furnaces for steel-making and in many of those employed in the metallurgy of copper and lead. It is sold in the form of brick, finely ground "furnace magnesite" for brick-laying, and dead-burned grains for making and repairing furnace bottoms. The latter are a mixture of granules varying in size from pieces of about 5/8 in. (1.6 cm.) diameter to very fine but sandy particles. Dead-burned magnesite results from calcining 'the crude or lightly burned mineral at a temperature that will not merely drive off practically all the CO2, but will also cause sintering of the particles. During this process the pieces shrink considerably and become hard, dense, and inert to atmospheric moisture and C02; under-burned material, on the other hand, will hydrate on exposure to the air. A small percentage of ferric oxide seems to be necessary for the production of a satisfactory sinter; from 4.5 to 8 per cent. in the dead-burned grains is considered the most desirable amount. Dolomite has been little used for brick-making in the United States, but it is prepared for use in the granular condition calcined or "double-burned" and is the principal ingredient of several materials offered for sale under various trade names for refractory purposes. Dolomitic refractories are almost wholly confined to the open-hearth and electric furnaces, where they are used for fettling and as substitutes for magnesite. Much more magnesite and dolomite are used for basic open-hearth steel-making than for all other refractory purposes. The hearth of the furnace is usually built up of magnesite brick and dead-burned grain magnesite so laid that the brick base is protected by a working bottom of the granular material. The latter is sintered into place in layers 1 to 2 in. (2.5 to 5 cm.) thick to a total depth of 12 to 18 in. (30 to 45 cm.) at the center of the furnace. After each heat, burned dolomite is thrown against the banks as high as it will stick, and all holes in the bottom are filled. At most plants such holes, at the end of each week, are also

Jan 2, 1919

By F. Ernest Brackett

Much new data on the flow of gases have been discovered by recent experiments by the United States Bureau of Mines and others. Although additional investigation is still desirable, the information now collected seems sufficient to attempt a tentative coördination of the friction factors. This study is confined entirely to fluid friction when the density throughout the duct is so nearly constant that it may be so considered. In the subsequent discussion the following notation is used: v = mean velocity of fluid in duct, feet per second. v' = mean velocity of fluid in duct, meters per second. R = hydraulic radius of duct, feet. R' = hydraulic radius of duct, meters. S = the hydraulic slope. P = pressure absorbed by friction, pounds per square foot. p = pressure absorbed by friction, pounds per square inch. p' = pressure absorbed by friction, kilograms per square centimeter. A = sectional area of duct, square feet. O = perimeter of duct, feet. L = total length of duct, feet. L' = total length of duct, meters. d = diameter of a cylindrical duct, inches. w = density of the fluid, pounda per cubic foot. g = specific gravity of the fluid compared with water = w % 62.4. C, k and B, coefficients for formulas using English measures. C' = metric coefficient, v' = C'R'S. n = coefficient of roughness for English or metric units. The basic formula for the flow of all fluids at constant density appears in the three following forms, which may be denominated: (a) The slope formula, v = CRS (b) The ventilation formula, " n = (c) The pipe formula, C is Kutter's coefficient, which is extensively applied to the flow of water in ducts and channels of all kinds; E is Atkinson's coefficient, used principally with air in ventilation problems; B is a coefficient conveniently adapted to studying the flow of water, gases and vapors in pipes. 312

Jan 1, 1927

By A. M. Kivari

OPERATIONS at the cement plant of the Permanente Corporation, in the hills about 45 miles south of San Francisco and 12 miles west of San Jose, are interesting to the members because of the adoption of modern hydrometallurgical methods for handling and beneficiating raw materials. Closed-circuit grinding, classification, hydro-separation, flotation, agitation, and thickening are familiar terms at this cement plant. These methods make it possible to process refractory limestone into suitable materials for the manufacture of modern cements. Construction of the plant, having originally a design capacity of 7000 bbl. of cement per day, was started in June 1939, and the first clinker was ground on the following Christmas Day. At present the capacity is about 12,000 bbl. of finished cement per 24 hr., most of the production being low-heat cement.† About 3130 tons of limestone, 620 tons 0f clay and 100 tons of gypsum are required daily for this output. Occasionally iron oxide is added to the clay to obtain the desired composition. This, together with the gypsum, is purchased from outside sources. Up to the present time, the corporation has produced low-heat, normal, moderate heat, moderate-sulphate-resisting, high-early-strength, oil-well, modified-oil-well, and plastic cements. Most of the present production is for Shasta Dam, near Redding, Calif., the total demand being 5,800,000 bbl, of cement. Following are some of the cement requirements of the Shasta Dam specifications: [PERCENT Chemical composition: Loss in ignition2.25 Insoluble residue0.75 SO32.25 MgO5.00 Uncombined lime in clinker (CaO)1 .25 Percentage of iron oxide to alumina ratio1.5 Compound composition: Tricalcium silicate (3CaO.Si02)Not more than 3,5 Dicalcium silicate (2CaO.Si02)Not less than 40 and not more than 65 Tricalcium aluminate (3CaO.Al2O7)Not more than 7 Tetracalcium aluminoferrite (4CaO.A12O3.Fe203) Not more than 0 Fineness:] The average specific surface in the bin sampled shall be not less than 1800 sq. cm. per gram and no individual sample less than 1600 sq. centimeters. Compressive strength: The average compressive strength-in pounds per square inch of three mortar cubes composed of one part cement and 2.75 parts fine testing sand, by weight, shall be equal to or greater than the following:

Jan 1, 1941

By S. S. Nosseir

INTRODUCTION General Quebec Cartier Mining Co. (QCM) , Canadian subsidiary of U . S. Steel Corporation, operates two iron ore concentrators at the North Shore of Quebec, Canada. In its own operation at Mount-Wright, over 45 million tons per year of crude ore are ground, fully autogenous in six 32' x 11-1/2 Dominion Engineering mills. At Fire Lake/Lac Jeannine operation, property of Sidbec Normines Incorporated (SNI), over 11 million tons per year of crude ore are now ground, fully autogenous, in twelve 18' x 5' Hardinge mills. The Lac Jeannine concentrator was originally constructed by Q.C.M. in 1960 and. was since operated, in beneficiating the Lac Jeannine ore deposit until fully depleted in 1977, at an annual capacity of 18 million tons of crude per year. In 1976, the new company, Sidbec Normines, was incorporated by Sidbec Corporation, British Steel Corporation and Quebec Cartier Mining Co., in proportional holding of approximately 51 to 41 to 8 percent respectively. Since the acquisition of the Fire Lake ore deposit and Jeannine concentrator by the new company, QCM has continued managing the mine and concentrator operaions for S idbec Normines Inc . Flow-Sheet Design Jeannine concentrator comprises twelve identical mill lines grinding the ROM from 8" top size to 14 mesh nominal size range. Driven by 600 H.P. motor at 13.4 RPM corresponding to 73% critical speed, each mill is operated in a closed circuit with two 5' x 12' vibrating screens dressed with 2.5 x 25 mm slot opening cloth. The ground product or screen undersize at an average grade of 33.4% Fe is beneficiated by gravity separation to produce 66.5% Fe concentrate grade at 42.7% weight recovery. At the rougher stage, each mill line con-tains 128 Humphrey 13- 1/2" pitch spirals which produce a final tailing, middling and concentrate products. The rougher concentrate - so produced - is classified by 64 Humphrey pocket sizers into an underflow, which is a final concentrate product and an overflow which is further upgraded by 64 cleaner spirals. The cleaner concentrate together with the pocket sizer 's underflow are dewatered on a 13'[6] horizontal filter. The rougher middling together with the cleaner tailing are recycled in a closed circuit with the spirals. The rougher tailing after dewatering by hydrocyclones is disposed of by a

Jan 1, 1982

By R. D. Carpenter, R. R. Smith, E. E. Sougstad

The Union Carbide Nuclear Co., in treating low-grade tungsten and molybdenum-copper ores at the Pine Creek mill near Bishop, Calif., employs flotation for the initial extraction of the valuable mineral constituents. These minerals consist of scheelite, molybdenite, and several of the copper sulfides including chalcopyrite, bornite, and possibly chalcocite. They occur in a gangue consisting mainly of quartz, pyroxene, and garnet, with minor amounts of calcite and fluorite. This assemblage of minerals presents an unusual grinding problem because of their diverse physical characteristics. Feed for the circuit is prepared from -6-in. ore by screening on a rod deck screen to remove the -5/8 - in. material. The+5/8-in. fraction is crushed in open circuit with a 5 1/2-ft Symons shorthead crusher operated at a closed- side setting of approximately 3/8 in. The screen undersize and the conecrusher discharge are then combined to constitute the feed to the grinding circuit. The ore is ground in 6 by 5-ft grate discharge ball mills operated in closed circuit with 54-in. spiral classifiers. The classifier overflow goes directly to the flotation section of the plant. Studies of the flotation circuits have shown that the maximum recovery of the tungsten, molybdenum, and copper minerals occurs in the particle-sue range from 100 mesh to approximately 400 mesh. Further, practice has shown that the ore minerals are effectively reduced to this size range if the total ore is ground to -48 mesh. The general objective of the grinding circuit then is to reduce the feed to -48 mesh with the production of a minimum amount of slimes. The purpose of this investigation was to ascertain how well the grinding circuit was carrying out this objective. The method used was to sample the various products in the grinding section, size these products accurately, and analyze the resulting sue distributions. SAMPLE PREPARATION AND SIZING TECHNIQUES Representative samples of the different grinding circuit products, namely, new feed to the ball mill, ball mill discharge, classifier overflow, and classifier return, were taken over a 24-hr period. The samples were then dried without being filtered to avoid loss of slimes. The percentages of solids in the different products were the following: new ball-mill feed, 95%; ball mill discharge, 82%; classifier overflow, 52%; and classifier return, 84% solids. Particle sizes were determined on each product by wet-screening an

Jan 1, 1962

By A. G. Buyers, J. Chilton, W. E. McKee

STUDIES in the high-temperature separations chemistry of thorium-uranium fuels are complicated by the corrosive nature of these molten metal systems at 1700°C. Separations research pointed toward the removal of fission products from irradiated alloys has included a frozen salt-bed technique which minimizes container reaction with molten metal. The procedure consists of high-frequency induction melting of the metal sample which is supported by a vertical column of unmelted salt. The molten alloy melts its immediate salt environment and as it slowly falls down through the column, many of the fission-product poisons are transferred into the fused salt environment. The molten salt-metal zone is encompassed at all times by unmelted salt, preventing container contact. EXPERIMENTAL Experimental methods were separated into three parts: salt-bed preparation, melting, and analytical evaluation. A) Bed Preparation—The procedure for casting calcium-fluoride briquettes was to compress the powdered salt at 5,000 to 10,000 psi and place this compact in a graphite crucible. This assembly was inserted into a vacuum system and surrounded by an alumina crucible as well as molybdenum, tantalum, or stainless-steel foil thermal-radiation shields. The pressure was reduced and the crucible heated under vacuum to about 800°C with the fore pump operating. Argon was then admitted to about 0.8 atm and heating continued until the CaF2 had melted 5 min at the melting point seemed satisfactory). Apparatus used included a 1-in. graphite crucible with 1/8-in. walls, 1-in. ID in a 2-in. induction coil, and a 6-kw Ajax high-frequency converter. B) Melting—Calcium-fluoride briquettes, 1 in. in diam and 1/2 to 1 in. thick, were prepared and vertically stacked to obtain the desired column length. An irradiated alloy sample (nut = 10") was placed on top of the salt column and the assembly was connected to the vacuum system. After evacuation of 5 to 10 , purified argon was admitted and the system flushed twice. The final argon pressure was about 0.8 atm. The sample was heated by induction at a power of about 6.5 kw. It then melted its way down through the salt and freely dropped into a small alumina crucible. The molten metal is followed during its downward course by the induction coil which is mounted upon a power-driven machine-threaded shaft. Experimental conditions used in the foregoing work

Jan 1, 1960

By Hans W. Schreiber, David W. Neuhaus

INTRODUCTION The Sunshine Mining Company's 16-to-1 silver deposit and mine project are located 384 kilometers (240 miles) southeast of Reno and 360 kilometers (225 miles northwest of Las Vegas, near Silver Peak in Esmeralda County, Nevada, at an elevation of about 2,300 meters (7,000 feet). The deposit was being developed, in 1980, with an anticipated mining and milling rate of 150,000 tons per year (500 tons per day). Mining is by underground, blast-hole open stoping with rubber-tire haul-age out of the mine to a straight cyanide plant about 5.5 kilometers (3.5 miles) from the mine. The product is a gold-silver dore which may be further refined at Sunshine's new refinery in Kellogg, Idaho. The cash required for investment was obtained through the sale of $25 million worth of 6 l/2h silver-indexed bonds. Some $15.0 million were directed toward development of the 16-to-1 Mine. These bonds, which were underwritten by Drexel Burnham Lambert of Chicago in April of 1980, are redeemable on or after April 15, 1985 or payable at maturity on April 15, 1995 at the greater of either $1,000 or the market price of 50 ounces of silver. Behre Dolbear acknowledges the kindness of the Sunshine Mining Company (Sunshine) for allowing the presentation of the subject material for this Case Study. APPLICABILITY AS CASE STUDY Reasons for Choosing 16-to-1 The reasons for choosing the 16-to-1 Mine as a case study follow: 1. The project is simple and compact. 2. The mine is small -- thus relevant to other contemplated precious metal operaions in the western states. 3. The areas of risk are readily identifiable. 4. A change to project financing was under consideration. 5. A substantial decline in the price of silver was experienced during development. 6. Behre Dolbear undertook one of the first feasibility studies of The 16-to-1 and thus has some familiarity with the project. Historical Summary The following history serves as a setting: 1. Late 1920s -- minor "production"(?) via shallow pitting. 2. Mid-1930s -- neighboring Nivlok Mine successfully put into production. 3. Late 1950s (?) -- Callahan Mining Company acquired rights and drilled -- results inconclusive; Callahan turned property back to Arthur Baker, who had originally interested Callahan in 16-to-1 on the basis of its similarity with Nivlok Mine.

Jan 1, 1985

By P. W. Davies, B. Wilshire

PREVIOUS investigations on the effect of stress changes on the high-temperature creep and fracture behavior of metals have been confined mainly to the testing of complex alloys.172 Most of these alloys are structurally unstable during creep because of such effects as recrystallization, overaging, and so forth, so that the results are of limited application. The only pure metals on which observations have been made appear to be lead3,4 and aluminum5 tested under conditions which allow rapid recovery to take place. The present work describes the effect of a single stress change on the creep and fracture properties of a metal in which recovery is very slow. A Ni 0.1 at. pct Pd alloy was used so that grain growth and recrystallization did not occur during creep testing. Specimens of this material were heat-treated to give a uniform grain size of 10 grains per mm. Creep tests were carried out at 500°C using Denison T.47 machines; details of the creep-testing procedure have been given elsewhere.' All specimens were tested at a constant initial stress of 5 tpsi which gives, in an uninterrupted test, a minimum creep rate of ~10-3 pct per rnin and a life of -7000 min. Specimens were loaded under these standard conditions and after 1500 min, when secondary creep was well-established, the stress was changed to either 8, 7, 6, 4, 2, or 0 tpsi for a period of 180 min. After this time the original load was reapplied and the tests continued to fracture. The results are shown in Fig. 1. No detectable creep occurred during the period of reduced load, even at 4 tpsi, and on reverting to 5 tpsi the creep rate returned to its original value immediately. Increasing the stress to 8, 7, or 6 tpsi resulted in an instantaneous extension followed by a period of transient creep leading to a new steady creep rate. On returning from these higher stresses, the resulting creep rate at 5 tpsi was very low compared with the original value. Furthermore, the higher the stress during the stress change, the lower was the final creep rate. It is significant that the rupture life is comparatively unaffected by any of these stress changes.

Jan 1, 1965