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By K. E. Brown, F. W. Jessen

It is deemed feasible to store atomic reactor fuel wastes in salt dome cavities when the differential pressure acting on the cavity does not exceed 3,000 psi and the temperature is less than 400°F. Tests at pressure increments of 1,000 psi were conducted on a 2-in, cylindrical cavity contained in a 6-in. long X 6-in. diameter salt core. The cavity exhibited stability under pressures up to 3,000 psi and temperatures to 300°F. At temperature ranges of 100 to 400°F and with pressures of 5,000 psi and above continuous deformation of the cavity resulted. Initial movement of the salt was observed at all pressares. This was evidenced by vertical deformation and cavity size reduction. INTRODUCTION The development of storage techniques for safe disposal of atomic reactor fuel wastes is a goal of the Atomic Energy Commission. Salt domes have been used successfully for storage of liquefied petroleum gases and hence should prove to be suitable for liquid waste products of atomic reactors. The purpose of the following work is to evaluate the practicability of utilizing salt dome cavities for such wastes and to de-termine particularly the effect of pressure and temperature on a salt cavity. This work represents a part of a comprehensive study covering the over-all problem of storing re- actor fuel wastes in salt domes and/ or salt beds. The problem of accurately determining the stresses which exist in rocks in the earth's crust is indeed a complex one. Theoretical studies have differed widely in many of the basic assumptions about the physical properties of the rock itself. Some solutions of problems in underground stress analysis assume that rock is elastic, homogeneous and isotropic in character; others assume that rock possesses plastic, viscous, elasticovis-cous properties, or a combination thereof. Experimental evidence has shown that salt possesses plastic, viscous and elasticoviscous properties.' Recent work has shown that salt becomes more plastic with temperature.' APPARATUS AND PROCEDURE For experimental study a cylindrical cavity was chosen for analysis since it provided maximum volume per unit of depth with constant diameter. This particular cavity shape, allowing for irregularities, can be easily leached out of salt domes in field operations. For the purpose of approaching underground storage conditions as nearly as possible, the following factors were considered: (1) control of pressure and temperatures, (2) measurements of temperature gradients throughout salt mass, and (3) measurement of structural stability of salt cavities. In order to obtain triaxial pressure on the salt cavity without resorting to the application of fluid pressure directly on the salt mass, a system was designed whereby the same effect could be produced through axial loading only. By cutting a core to close dimensions to fit a steel test cylinder and applying axial pressure, a lateral pressure was obtained. This equipment is shown in Figs 1 and 2. A steel cylinder was made from

By P. Rama Rao, D. H. Sastry, T. R. Anantharaman

Deformation (a) and twin (a') stacking-fault densities in cold - wor ked filings of fcc silver alloys containing 2, 4, and 6 wt pct Sb have been estimated from Debye-Scherrer peak shifts and asymmetries, respectively. Evidence has been obtained for a possible strain-induced transformation in a two-phase alloy containing 9.0 zut pct Sb. Using Fourier analysis of X-ray line profiles, the effective particle sizes and root mean square strains have been determined for the three fcc alloys in the cold-worked condition. The values of the compound fault probability 0.5a + a') evaluated from peak shifts and asymmetries are in reasonable apeement with those derived from peak broadening. COLD work of fcc metals may produce both deformation and twin (also called growth) stacking faults, which can be detected and quantitatively estimated by X-ray diffraction techniques.' In addition to peak broadening, deformation faults produce a shift in peak positions and twin faults an asymmetry of the Debye-Scherrer reflections. Besides stacking faults, a reduction in the size of coherently diffracting domains and distortion within these also broaden the powder reflections of deformed fcc metals and alloys. Fourier analysis' of peak shapes provides quantitative infor- mation about the effective domain sizes, lattice strains, and also the compound fault probability (1.5a + a'), where a and a' represent deformation-and growth-fault density, respectively. The Paterson method2 for determination of a from peak shifts, the warren3 or wagner4 method for evaluation of a' from peak asymmetries, and the Fourier analysis for quantitative estimates of all the contributory causes of peak broadening have, in the course of the last decade, been extensively applied to obtain data for a number of deformed fcc metals and alloys. One of the useful results of earlier measurements of a in several solid solutions is a correlation between a and the solute valence as well as electron concentration per atom (p) in alloy systems based on a particular solvent. The work of Smallman and westmacott5 on Cu-Zn, Cu-A1, and Cu-Sn, of Foley, Cahn, and ~aynor' on Cu-Si and Cu-Ge, and of Rama Rao and Anantharaman7 on Cu-Si and Cu-Sb has revealed a common trend in the increase of a with p and also a with valency of the solute. Cahn and Davies' and Ad-ler and wagnerg have observed similar trends for silver alloys containing cadmium, indium, and tin. Data on a for silver alloys with a pentavalent solute element are not, however, available. It is primarily to fill this need that the present investigation on Ag-Sb alloys was undertaken. Utilizing the Patersonand Warren methods, values of a and a', respectively, have been determined in three silver alloys containing 2, 4, and 6 wt pct Sb. By Fourier analysis of the broadening of Debye-Scherrer peaks, effective particle sizes, lattice strains, and the compound fault probability (1.5a +a') have also been evaluated.

Jan 1, 1967

By L. Albert, A. M. Cameron, J. W. Gullick

International Nickel has 15 operating mines in Canada, and. 3 mines which are being maintained on a standby basis. Annual production is about 20 million dry short tons of ore, most of which is from underground mines, the remainder from open pit mines. Although some of the underground mines are served by ramps for trackless vehicles, most men, materials and ore are hoisted in vertical shafts. In addition to shafts and raises for ventilation and service purposes which are not equipped with hoisting plants, there are now 28 vertical shafts at Inco mines served by 47 hoists. Of these, 22 shafts and 36 hoists are located in the Sudbury District of Ontario, one shaft and two hoists are at the Shebandowan mine in Western Ontario, and 5 shafts with 9 hoists are located near Thompson, Manitoba. This paper describes the components of Inco hoisting plants-which have been applied at different times and under different conditions, and indicates the rationale for the various designs. Some engineering and maintenance practices, considered advantageous, are also described. MINE SHAFTS Many of the shafts sunk in the early days of the Sudbury District, on what are now Inco properties, were inclined shafts. Of the identifiable shafts sunk between 1885 and 1920, thirteen were inclined and at least four were vertical. Since 1920, thirty vertical and two inclined shafts have been sunk. All of the inclined shafts have been abandoned for hoisting purposes but the majority are still in use as service or ventilation ways. Shafts at Thompson- and Shebandowan are all vertical. The vertical shafts range from 2 or 3 compartment exploration and development shafts to 6 compartment shafts for ore, man and material hoisting. They vary in depth from 1,050 to 7,137 feet and in addition to hoisting compartments all contain manway and service compartments. Prior to 1948, unlined rectangular shafts with timber sets and guides were used. The rectangular cross-section was favoured since the rock is competent, the shape compatible with timber set framing, and since available shaft space could be used efficiently. Resistance to air flow was not of prime importance. Concrete linings were introduced for the first time at the Creighton No. 7 Shaft which was collared in 1948. All major shafts since that date have been concrete lined. Vertical set intervals were increased in steps, with the introduction of structural steel sets secured to the concrete linings. Whereas 7 foot set intervals were used with timber framing, this increased to 10 feet 6 inches with steel sets and more recently to 15 feet with steel and steel backed wood guides. With the advent of deeper shafts, or shafts which are expected to be deepened, the trend is to circular concrete lined shafts. The majority of shaft guides in service are of select Douglas Fir. This practice commenced as a component of timbered shaft design but continued as a

Jan 1, 1975

By James F. Watson, Jack L. Christian

The austenite to martensite solid state phase transformation in the 300 series austenitic stainless steels was studied over the temperature range of 788° to -423°F. The primary variables controlling this transformation are temperature, plastic strain, and chernical composition. Consequently, temperature effects were studied by measurements at 78", -100°, -320°, ad -423°F; plastic strain was controlled by cold-rolling the alloys varying amounts between 40 and 80 pet; and chemical compositions were varied by using 301, 302, 304 ELC, and 310 stainless steel alloys which exhibit varying degrees of austenite stability. The martensite contents of various alloys were correlated with several mechanical properties including yield and ultimate tensile strength, elongation, and notched tensile strength (Kt = 6.3). The yield and tensite strengths generally increased with increasing martensite content, but ) no correlation was observed between elongation and martensite content. The notched/un-notched ratio (Kt = 6.3) at -423°F exhibited a maximun at about 65 pct martensite. Attempts to correlate austenite stability (as measured by the Ms temperature) with both martensite content, and incremental increases in martensite content resulting from tensile testing were generally unsuccessful because the variable of cold Work could not be separated. As a result, only a general correlation of these variables was obtained. A (111) pole figure of austenite in a cold-rolled 301 stainless steel mas prepared an attempt to further investigate directionality as exhibited by mechan- ical properties. Yield strength was correlated with this pole figure, but ultimate and notched tensile strengths were not. 1 HE literature contains a wide variety of references pertaining to the austenite to martensite solid state phase transformation in austenitic stainless steel, see references 1 through 18. Discussions of the crystallography and kinetics of this transformation are given in references 2, 7, 8, 'and 11, while further discussions of austenite stability and hardening mechanisms are given in references 1, 3, and 10. Two important processing variables involved in this reaction are the degree of cold work and the temperature at which cold work is performed, and the nature of these effects are outlined in references 4 and 12, and 15 and 16 respectively. The latter two references deal primarily with cold rolling conducted at subzero temperatures. The effect of chemical composition on the austenite to martensite reaction was treated in detail by Eichelman and Hull, and they determined an equation for the M, temperature of this type of steel which included the effects of nickel, chromium, carbon, nitrogen, silicon, and manganese.5 Previous studies have shown that austenitic stainless steels of AISI types 301, 302, 304ELC, 310, and 301-N exhibit mechanical properties between 78" and -423"~ that are microstructure dependent, and the dependence of this microstructure on chemistry, temperature, tensile stress, and prior deformation is in general accord with previous investigations which have measured the effects of each of these variables on structure, or more specifically, the austenite-to-martensite transformation.lg The behavior of fully stable alloys, such as ATST type 310, at cryogenic temperatures is

Jan 1, 1962

By James W. Halley

"Fine-grained" steels have been standard products for many years. This paper describes an investigation of the effects of some of the more common grain-growth inhibitors used to produce these steels. The properties, response to heat-treatment, and the deoxidation practice necessary to produce "fine-grained" steels have been the subjects of numerous papers. It has been shown that a substantial quantity of aluminum is necessary to produce this type of steel, and it has been indicated that titanium and zirconium produce a similar result. For some time there was a lively argument between those who felt that aluminum refined the grain and those who felt that aluminum oxide was the effective agent. This argument subsided with an apparent victory for those promoting aluminum oxide. Most of the work on grain size has divided steels into two broad classifications based upon McQuaid and Ehn's original carburizing test at 1700°F or some similar test at this temperature. If the steel coarsened at this temperature, it was coarse grained, while if it did not it was he grained. Quantitative data concerning how much a certain amount of aluminum raised the coarsening temperature and affected the properties, all other factors being constant, are very meager. Similar information regarding titanium and zirconium is practically nonexistent. The published information on aluminum yields some quantitative information. Epstein, Nead and Washburn' gave the first quantitative information on the amount of aluminum necessary to prevent grain growth at 1700°F. They showed that the addition of a pound to a pound and a half of aluminum per ton (0.05 to 0.075 per cent) would assure steel that would not coarsen during a McQuaid-Ehn test at 1700°F. Herty, McBride and Houghg reported that a steel containing 0.020 per cent acid-soluble aluminum coarsened between 1742° and 1832°F., while steels containing 0.016 per cent and 0.017 per cent acid-soluble aluminum coarsened between 1652° and 1742°F. compared with similar aluminum-free steels that started to coarsen at 1580°F. Herty3 also showed a marked improvebent in low-temperature notched impact resistance with increasing acid-soluble aluminum up to 0.030 per cent. McQuaid4 compared the coarsening of steels containing 0.011 to 0.204 per cent total aluminum and found a maximum coarsening temperature at 0.098 per cent aluminum. He had no samples between 0.015 and 0.098 per cent aluminum. Houdrement6 found a marked decrease in coarsening temperature with aluminum contents above 0.50 per cent. Kinzel, Crafts and Egan6 found that aluminum and zirconium improved the low-temperature impact resistance of a low-alloy steel, while titanium was detrimental. Sims7 shows that sulphur increases the coarsening temperature and that the maximum coarsening temperature is between 0.04 and 0.06 per cent added aluminum. The fact that the coarsening temperature passes through a maximum with increasing aluminum content appears to be well established. The aluminum content that

Jan 1, 1947

By James W. Halley

"Fine-grained" steels have been standard products for many years. This paper describes an investigation of the effects of some of the more common grain-growth inhibitors used to produce these steels. The properties, response to heat-treatment, and the deoxidation practice necessary to produce "fine-grained" steels have been the subjects of numerous papers. It has been shown that a substantial quantity of aluminum is necessary to produce this type of steel, and it has been indicated that titanium and zirconium produce a similar result. For some time there was a lively argument between those who felt that aluminum refined the grain and those who felt that aluminum oxide was the effective agent. This argument subsided with an apparent victory for those promoting aluminum oxide. Most of the work on grain size has divided steels into two broad classifications based upon McQuaid and Ehn's original carburizing test at 1700°F or some similar test at this temperature. If the steel coarsened at this temperature, it was coarse grained, while if it did not it was he grained. Quantitative data concerning how much a certain amount of aluminum raised the coarsening temperature and affected the properties, all other factors being constant, are very meager. Similar information regarding titanium and zirconium is practically nonexistent. The published information on aluminum yields some quantitative information. Epstein, Nead and Washburn' gave the first quantitative information on the amount of aluminum necessary to prevent grain growth at 1700°F. They showed that the addition of a pound to a pound and a half of aluminum per ton (0.05 to 0.075 per cent) would assure steel that would not coarsen during a McQuaid-Ehn test at 1700°F. Herty, McBride and Houghg reported that a steel containing 0.020 per cent acid-soluble aluminum coarsened between 1742° and 1832°F., while steels containing 0.016 per cent and 0.017 per cent acid-soluble aluminum coarsened between 1652° and 1742°F. compared with similar aluminum-free steels that started to coarsen at 1580°F. Herty3 also showed a marked improvebent in low-temperature notched impact resistance with increasing acid-soluble aluminum up to 0.030 per cent. McQuaid4 compared the coarsening of steels containing 0.011 to 0.204 per cent total aluminum and found a maximum coarsening temperature at 0.098 per cent aluminum. He had no samples between 0.015 and 0.098 per cent aluminum. Houdrement6 found a marked decrease in coarsening temperature with aluminum contents above 0.50 per cent. Kinzel, Crafts and Egan6 found that aluminum and zirconium improved the low-temperature impact resistance of a low-alloy steel, while titanium was detrimental. Sims7 shows that sulphur increases the coarsening temperature and that the maximum coarsening temperature is between 0.04 and 0.06 per cent added aluminum. The fact that the coarsening temperature passes through a maximum with increasing aluminum content appears to be well established. The aluminum content that

Jan 1, 1947

SEC. 1. The membership of the Institute shall comprise six classes, namely: 1. Members; 2. Honorary Members; 3. Senior Members; 4. Associates; 5. Junior Members; 6, Rocky Mountain Members. All shall be equally entitled to the privileges of membership, excepting that no member of any class whose residence shall be outside of the United States of America, except those residing in Mexico and Canada, shall be entitled to vote. SEC. 2. Members: A person to be eligible for election or transfer into the class of Member must be at least 27 years of age and must have had at least six years' employment in the practice of engineering, mining, geology, metallurgy, or chemistry, during at least three years of which he must have held positions of responsibility in one or more of these fields. This shall not apply to Rocky Mountain members. SEC. 3. Honorary Members: Each candidate for Honorary Membership must be recommended by at least ten Members and must be elected by ballot at a meeting of the Board of Directors by the unanimous vote of the Directors present; provided, however, that the number of Honorary Members shall at no time exceed twenty. Such election shall automatically act as a transfer of such member from any other class of membership to which he may belong. SEC. 4. Senior Members: The Directors may in their discretion place on the Honor Roll of Senior Members and transfer from any other class of membership other than Honorary Membership the name of any Member or Associate who has reached the age of seventy years and who has paid dues for thirty years, and for good reasons in • special cases may vary the age or number of years. SEC. 5. Junior Members: A person to be eligible for election to the class of Junior Member must be qualified through education or experience to hold a subordinate position in engineering, mining, metallurgical, geological or chemical work. He shall be not more than twenty-seven years of age at the time of election. The rights of the present Junior Associates shall continue until the same expire by limitation of time. SEC. 6. Associates: A person eligible for election or transfer into the class of Associate shall be one, who, in the opinion of the Committee on Admissions and the Board of Directors, is suitable for such election or transfer by reason of his interest in or connection with mining, geology, metallurgy, or chemistry. SEC. 7. Rocky Mountain Members: Membership in a special class of life memberships without annual or other dues and without right to receive the printed literature distributed to other members or the right and privilege of wearing the official badge of technical members, is given to all members of the Rocky Mountain Club at the time of consolidation. These special Rocky Mountain memberships shall be claimable by any member of the Rocky Mountain Club, not already a member of the Institute,

Jan 1, 1929

By Howard A. Meyerhoff

IN western Argentina, in the Province of Mendoza and the Territory of Neuquén, there is a series of solid bitumen deposits which are claimed to be the most extensive in the world. In a linear belt 500 km (311 miles) long, roughly paralleling the axis of the Andes but occupying a variety of positions with respect to the orogenic structures, more than 100 outcrops have been reported. Approximately 70 of these have been prospected, and some mining has been planned or attempted for about half this number. Limited tonnage, high sulphur content, inaccessibility, labor shortages, and disastrous fires have taken a heavy toll among these operations, and in 1947, only 4 deposits were being actively mined, but 8 more were being explored and developed with a view to placing them in production. Inasmuch as these deposits are currently yielding a substantial fraction of Argentina's limited domestic supply of solid fuels, a brief account of the deposits and of the small industry based upon them will provide a commentary on Argentina's fuel problem. CHARACTER OF SOLID BITUMENS The solid bitumens of Mendoza and Neuquén are petroleum derivatives, as is indicated by the following ultimate analysis: [CONSTITUENTS PER CENT Hydrogen6.8 Carbon82.4 Nit.rogenI.6 Oxygen2.7 Sulphur2.2 Ash4.3 100.0] In physical appearance they bear a close resemblance to coal and are commonly called carbón. Their specific gravity has an extreme range of 1.10 to 1.777 in 36 samples weighed, but the average is between 1.14 and 1.24. The material is black, lustrous to dull, with blocky to sub-conchoidal fracture. Though brittle, the purest material is softer than anthracite and exhibits slightly more cohesion. This bituminous material is generally known as asfaltita or asphaltite, but definitive chemical studies indicate that the term asphaltite or grahamite can be applied correctly only to the exceptionally pure material of the Auca Mahuida district in Neuquén. Some other name should be used for the chemically variable bitumens found in all other districts, and the designation "asphaltitic pyrobitumen" has been proposed by Carlos A. S. Piscione.1 At the present time these bitumens, are finding favor as solid fuel and as a gas-producing medium. They yield heat values of 8650 to 9750 cal,2 or 14,500 to 16,750 Btu per pound, and in consequence of their excellent heating qualities and the scarcity of imported coal, they are marketed as far away as Buenos Aires at a price of go to 95 pesos, or approximately $22 to 23, per metric ton. In burning, the material swells and decomposes into solid, fluid, and gaseous fractions which burn violently though nonexplosively. Its combustibility has been responsible for disasters and virtually complete destruction at Mina La Esperanza and Mina Santa Marta, and for minor mishaps elsewhere, as at Mina Escondida in August 1947.

Jan 1, 1947

By John W. Crutchfield, Emil F. Bowers

This paper presents the initial results of a pressure maintenance by water return project in the Lower Pawelek reservoir of the Falls City Field, Karnes County, Southwest Texas. Production is from the Lower Wilcox section at an average depth of 6,030 ft on the downthrown side of a "down to the coast" fault. Of particular interest is that the reservoir oil is high gravity and low viscosity but far undersaturated with gas at the original reservoir pressure and temperature. The early pressure and production history indicated only a partial or limited water drive, and this fact, coupled with the characteristics of the reservoir fluid and artificial lifting problem.. made advisable the water injection program to maintain reservoir pressure. The paper includes a description of the reservoir and reservoir fluids, the pressure and production history, a description of the water injection facilities, and the results of reservoir calculations involving Hurst's unsteady state flow and the material balance equations for calculation of water influx and prediction of the pressure-production relationship at various rates of withdrawal. The reservoir calculations and initial field results indicate that reservoir pressures in excess of 85 per cent of original pressure can be maintained by control of water injection even at very high rates of production. INTRODUCTION The purpose of this paper is to report the initial results of a pressure maintenance by water return program in the Lower Pawelek reservoir, Falls City Field, Karnes County. Texas, and to calculate an empirical relationship between pressure and production in this limited or partial water drive reservoir. Location The Falls City Field (Fig. 1) is located two miles north of the town of Falls City in Karnes County, Texas, and is approximately 50 miles south-southeast of San Antonio on U. S. Highway 181. Discovery and Development The initial block of leases was taken after a surface geological survey identified a major east-west fault and closure on the south or downthrown section. The discovery well, Southern Minerals Corporation's Joe F. Bartosch 1; was completed at 4,650 ft in the Lower Bartosch sand in December, 1944. Drilling was continuous until December, 1946, and resulted in the discovery of 12 oil reservoirs between 4,600 and 6,050 ft in the Middle and Lower Wilcox formations. These are shown on the cross-section (Fig. 2). As of the current date there are 64 oil wells completed as follows: Southern Gulfshore Magnolia Minerals Oil Petroleum Reservoir Corp. Co. Corp. Total Upper Bartosch 0 0 0 0 Lower Bartosch 16 0 0 16 Shulze 7 0 1 8 Upper Faires 2 0 0 2 Lower Faires 4 0 0 4 Upper Moczygemba 0 0 0 0 Lower Moczygemha 2 0 0 2 Olinick 2 0 0 2 Upper Martinez 0 0 0 0 Lower Martinez 5 0 0 5 Upper Pawelek 0 0 1 Lower Pawelek 27 1 3 31 Total 65 2 4 71

Jan 1, 1950

By Oliver Bowles

Slate is a fine-grained rock that has a more or less perfect natural cleavage, permitting it to be split readily into thin, smooth sheets. According to a definition established by the American Society for Testing Materials, slate is "a microgranular crystalline rock derived from argillaceous sediments by regional metamorphism and characterized by a perfect cleavage entirely independent of original bedding, which cleavage has been induced by pressure within the earth." Origin and Character Slates have originated from sedimentary deposits consisting largely of clay. Such materials were carried by streams and deposited in seas or lakes in successive horizontal beds. Other materials such as calcium carbonate, sand, or gravel may have been deposited over the clays, and the pressure of these superimposed materials gradually consolidated the materials below into bedded deposits of shale. Shale is a stratified rock formed by recrystallization of the clay plus adsorbed potash into the mica, sericite. Shale has neither the strength nor the splitting properties of slate since the mica is fine and the flakes unoriented, and the quartz grains do not interlock. Many shales, however, have been subjected to intense metamorphism. Mountain-building forces ever at work in the earth's crust may crumple, squeeze, and fold the shales. At the same time, accompanying high pressures and temperatures changed the original minerals into mica, interlocking quartz, or chlorite. Both mica and chlorite consist of minute tabular sheets or flakes with perfect cleavage. The intense pressure and movement tend to orient these flakes in parallel positions normal to the direction of pressure. The parallelism of the flakes results in the tendency to split with ease in one direction, which has been termed "slaty cleavage," and which is the outstanding characteristic of slate. Actually, the cleavage is through the cleavages of the individual micaceous plates. The original bedding of the slate is often folded and contorted, and the slaty cleavage commonly cuts across it at sharp angles. Composition MINERALOGICAL COMPOSITION The most abundant mineral in most slates is white mica (sericite), a hydrous silicate of potash and aluminum. It is present in very small flakes recognizable only with a microscope. Small grains of quartz (silicon dioxide) are also abundant and are regularly distributed among the mica flakes. The mica-like mineral, chlorite, is usually present in considerable quantities. Chlorites are of several kinds, the more common being hydrous silicates of aluminum and iron or magnesium. Other common minerals present in small quantities are rutile, clay, hematite, pyrite, carbonaceous matter, feldspar, zircon, tourmaline, calcite, dolomite, and siderite. The general range of the chief constituents is about as follows: Sericite, 38 to 40; quartz, 31 to 45; chlorite, 6 to 18; hematite, 3 to 6; and rutile, 1 to 1 ½ pct. CHEMICAL COMPOSITION Results of many analyses indicate that clays, shales, and slates differ little in chemical composition. The changes that take place during metamorphism are confined largely to rearrangement of the elements into new minerals with accompanying changes in physical properties such as hardness and cleavage. The general

Jan 1, 1960

By W. E. Hess

AT the Vesta mines of Jones & Laughlin Steel Corp. on the Monongahela River, 35 miles south of Pittsburgh, JCM Joy continuous miners and 6-SC shuttle cars are used for pillar extraction in the Pittsburgh seam. The entire output of the mine goes into coke production for blast furnace operation. Because this coal, which has high metallurgical qualities, is close to the place where it is ultimately used and because supplies with these properties are being rapidly depleted, complete recovery is essential wherever possible. With the introduction of the continuous miner into the Vesta mines, therefore, it was necessary to plan a project which would facilitate development and provide for future total recovery. Use of continuous miners under varying conditions has necessitated changes with regard to roof action, method of development, method of pillar attack, and innovations in mining. In the following pages these changes will be discussed. Prior to the introduction of continuous miners the Pittsburgh seam at the Vesta mines averaged 72 in. of mineable coal, overlain with 12 to 16 in. of draw-slate, which weathers on exposure to air and is not firm enough to be held in place. Immediately overlying the drawslate is a wild or rooster coal, very irregular in thickness, varying from 12 to 30 in. Above the rooster coal the strata consists predominately of shales or lime shales and sandstone. Just before the introduction of continuous miners in the Vesta mines, however, a definite change took place in the character of the roof structure overlying the coal seam, see Fig. 1. Thickness of laminated roof immediately overlying the rooster coal varies with localities and introduces a problem in roof control. In fact, a set of eight main entries being developed in Vesta NO. 4 mine entered an area with so difficult a roof structure that management considered abandoning the section altogether. In this locality the roof consisted of thin alternate layers of roof coals and shales, with an occasional irregular seam of slick slate, the whole being laced with slips, and the coal itself being cut by clay veins and spars. In an effort to overcome some of the difficulties, roof bolting was introduced in 1949. This made it possible to recover coal from an area which otherwise would have been economically unmineable. There have been very few roof bolt failures. When conventional machines were replaced by continuous miners there were changes in the development of butt entries. By these changes it was hoped: 1—to speed development, 2—to provide as much solid coal as possible for the machine on retreat, 3—to avoid the necessity of crossing entries already driven, 4—to facilitate ventilation, 5—to provide a roadway for shuttle car haulage which would eliminate the necessity of running through doors or curtains, 6—to provide better roof control by increasing the distance between the center lines of entries, and 7—to provide more coal when the machine was in position to load.

Jan 1, 1956

By Nelson B. Russell, Donald C. Bond

The effects of various antifreeze agents on the formation of hydrogen sulphide hydrate have been studied. On a molar basis the relative lowering of Tm (the maximum temperature at which solid hydrogen sulphide hydrate can exist in equilibrium with the given solution) for the various agents is: sodium chloride 1.00, calcium chloride 1.71, methyl alcohol 0.57, ethyl alcohol 0.68, ethylene glycol 0.73, diethylene glycol 0.73, sucrose 0.87, dextrose 0.71. On a weight basis the relative lowering of TM is: sodium chloride 1.00, calcium chloride 0.91, methyl alcohol 1.08, ethyl alcohol 0.89, ethylene glycol 0.69, diethylene glycol 0.41, dextrose 0.23, and sucrose 0.15. Introduction The solid hydrate of hydrogen sulphide, H2S.6H20,1 is in equilibrium with water and liquid hydrogen sulphide at 85°F at a partial pressure of hydrogen sulphide approximately equal to the vapor pressure Of hydrogen sulphide at that temperature. At temperatures below 85°F, in the presence Of excess hydrogen sulphide (greater than the amount required to react with the water present) the hydrate is in equilibrium with liquid hydrogen sulphide at approximately the vapor pressure Of hydrogen sulphide. In the presence of excess water, the solid hydrate is in equilibrium with liquid water, containing dissolved hydrogen sulphide, at a pressure lower than the vapor pressure of liquid hydrogen sulphide at the same temperat~re.2 The lower the temperature, the greater the difference between the decomposition pressure of the hydrate and the vapor pressure of liquid hydrogen sulphide at that temperature. Conditions conducive to the formation of solid hydrogen sulphide hydrate may exist in wells producing a gas containing a high percentage of hydrogen sulphide under high pressure, or in the lead lines from such wells. This paper gives the results of tests made with various antifreeze agents which might be useful in decomposing the hydrate or in preventing the formation of the hydrate of hydrogen sulphide Apparatus Fig , shows a diagram of the apparatus used. The sample under observation was contained in a glass tube 34 cm long, having a 6 mm id and a wall thickness of 2.5 mm. The lower end of this tube was closed- with a plug of butyl rubber. The upper end of the tube was connected to a diaphragm gauge, with a stainless steel diaphragm, a manometer, a vacuum pump, and a cylinder of hydrogen sulphide, The tube was held nearly horizontal, with the liquid resting along the lower side of the tube. It was found that when the tube was held in a vertical position, it broke with ex. plosive violence as soon as it became filled with solid hydrate. Presumably this was caused by the expansion which occurred on solidification of the hydrate. The difficulty disappeared when the tube was held in an inclined position during the tests.

Jan 1, 1949

By John P. Dolan, Charles A. Einarsen, GiIman A. Hill

his paper discusses how the following formation characteristics can be determined mathematically from drill-stem test pressure charts: true formation pressures, effective permeability of the entire section tested, well productivities, wellbore damage, and possible detection of barriers (faults, pinchouts, etc.). This paper also presents a practical method for im-mediate determination of eflective permeability and well-bore damage from successful double shut-in pressure tests. A list of recommendations for improving the reliability of drill-stem rest pressures is also presented. INTRODUCTION A drill-stem test is a temporary completion designed to sample the formation fluid and to establish the possibility of commercial production. Early pressure recording devices were used merely to verify proper operation of the testing tool. Until recently the accuracy of the pressure gauges has been insufficient for any reliable quantitative use of the recorded pressures. In view of the need for more reliable formation evaluation and as a result of the recent interest in exploration work employing the concept of hydrodynamic en-trapment,1,2,3,4 better pressure recording gauges are now in use. These devices can record pressure within 1 per cent above 1,000 psig and can detect differential pressures as low as 1/2 psig. In addition to formation pressure, several other reservoir characteristics can be determined from DST charts; namely, well productivity, formation permeability, wellbore damage, and the possible existence of bar- riers (faults, pinchouts, changes in permeability, etc.). This paper presents a practical method to interpret DST pressure charts for formation pressures and many other reservoir properties, a method that has been developed in analyzing approximately 4,000 DST charts during the last five years. The techniques used are a composite of published articles on drill-stem testing5,6, together with well-known pressure build-up analysis methods. THEORY It has been shown7,8 that the following equation may be used for analysis of pressure build-up curves: Po - Pw = 162.6 qu/kh log ( to + ?/?) .....(1) When this equation is applied to the curves obtained in drill-stem testing, the assumptions and boundary con-ditions are more nearly realized than in conventional flow and build-up tests on producing wells. Zak and Griffen9 have recently discussed in detail the use of this equation in analyzing DST charts. One of the problems with DST curves is the lack of reservoir data for precise analysis. Therefore, it is necessary to develop empirical rules and field methods for analyzing DST charts in quantity. For this reason, the empirical methods presented in this paper have been developed, and their derivation is found in Appendices A and B. METHOD USED FOR ACCURATE READING OF DRILL-STEM TEST CHARTS In order to apply the pressure build-up theory to DST charts, it is necessary to obtain a digitized expression for the pressure and time data recorded by the pressure gauge. These data may be either provided by the service company or interpolated by projecting a

Jan 1, 1958

By Ernest R. Dickie

BRIEFLY, the ore body of the Bagdad Copper Corp., Bagdad, Ariz., is a monzonite porphyry carrying copper values fairly evenly distributed from the surface down through the primary zone. Tabular in shape, the ore body extends over several hundred acres. The oxidized zone averages about 60 ft thick and assays about 0.50 per cent copper. Under this is the chalcocite zone, or the zone of secondary enrichment, which averages about 100 ft in thickness and assays about 1.4 per cent copper. Under this zone lies the primary zone which extends in places to a known depth of about 150 ft and averages in grade about 0.60 per cent copper. Several years ago the block-caving method of extracting the ore from the chalcocite zone was installed. Then it was necessary to draw only about 250 to 300 tons per day, the capacity of the old milling plant. Early in 1943 the new 2500-ton concentrator was put into production. Drawing ore from the developed stopes in sufficient tonnage to furnish the mill at capacity proved unsuccessful in more ways than one. First, drawing the ore at an accelerated rate caused a large amount of dilution, there-by lowering the grade, and also caused funneling through to the surface. The surface material was detrimental to milling metallurgy and development of these stopes was costly per ton of ore extracted. The grade of ore decreased. making it impossible to maintain a mill head of over 0.9 per cent copper. The

Jan 1, 1947

By H. R. Nelson, Martin S. Maier

The diffusion of metals plays an important role in many metallurgical processes,lJ as, for example, the formation of alloys by the annealing of mixed powders, the homogenization of segregated alloys, and the carburization of steel. Other equally important phenomena, such as recrystalli-zation and creep, are based on the movement of like atoms. Although much is known concerning the effects of diffusion in metals, very little is known concerning the details of the physical processes involved. 1,3-5 The data that should give most information concerning the processes involved in diffusion are those on the diffusion of like atoms—self-diffusion. Very little is known concerning the rates of self-diffusion of metals because before the discovery of artificial radioactivity in 1934 there was no means of detecting the movement of the diffusing atoms through the solvent metal except in the elements possessing natural radioactive isotopes—lead, thallium and bismuth. Using the induced radioactivity of the element as an indicator, atoms can now be "tagged" so that the rate of diffusion of the radioactive isotopes of an element (tagged atoms) into the stable isotopes of the same element can be measured. As the radioactive and stable isotopes of an element differ only in nuclear structure, it is quite certain that all isotopes of an element diffuse at the same rate. Two techniques have been used to measure self-diffusion. In one, a thin layer of radioactive metal is deposited on the face of the solvent metal, and the decrease in the surface activity as the radioactive atoms diffuse into the solvent metal is measured. This method has been used to determine the coefficients of self-diffusion of lead6-l0 and bismuth10 using natural radioactive isotopes, and of gold1' and copper12 using artificial radioactive isotopes. In the other technique a thin layer of the radioactive metal is deposited oil one face of the solvent metal, and after diffusion has occurred the specimen is cut in thin sections to obtain the concentration-penetration curve. This method has been used to determine the coefficients of self-diffusion of gold,13 zinc,14-l6 silver," and copper.17 The second method was used for the work described in this paper. Materials The coefficients of self-diffusion were determined in two grades of polycrystalline copper and in single crystals of copper. Commercial OFHC copper was used in the form of disks I in. in diameter by % in. thick. Samples of high-purity copper described by Phillips and Skinnerl9 were supplied through the courtesy of the American Smelting and Refining Co. The high-purity copper rod was used without remelting by cross-rolling to a thickness of 0.075 in., from which disks 7/8 in. in diameter were cut. A good grade of bus-bar copper was used to make the single crystals.

Jan 1, 1942

By Martin S. Maier, H. R. Nelson

The diffusion of metals plays an important role in many metallurgical processes,lJ as, for example, the formation of alloys by the annealing of mixed powders, the homogenization of segregated alloys, and the carburization of steel. Other equally important phenomena, such as recrystalli-zation and creep, are based on the movement of like atoms. Although much is known concerning the effects of diffusion in metals, very little is known concerning the details of the physical processes involved. 1,3-5 The data that should give most information concerning the processes involved in diffusion are those on the diffusion of like atoms—self-diffusion. Very little is known concerning the rates of self-diffusion of metals because before the discovery of artificial radioactivity in 1934 there was no means of detecting the movement of the diffusing atoms through the solvent metal except in the elements possessing natural radioactive isotopes—lead, thallium and bismuth. Using the induced radioactivity of the element as an indicator, atoms can now be "tagged" so that the rate of diffusion of the radioactive isotopes of an element (tagged atoms) into the stable isotopes of the same element can be measured. As the radioactive and stable isotopes of an element differ only in nuclear structure, it is quite certain that all isotopes of an element diffuse at the same rate. Two techniques have been used to measure self-diffusion. In one, a thin layer of radioactive metal is deposited on the face of the solvent metal, and the decrease in the surface activity as the radioactive atoms diffuse into the solvent metal is measured. This method has been used to determine the coefficients of self-diffusion of lead6-l0 and bismuth10 using natural radioactive isotopes, and of gold1' and copper12 using artificial radioactive isotopes. In the other technique a thin layer of the radioactive metal is deposited oil one face of the solvent metal, and after diffusion has occurred the specimen is cut in thin sections to obtain the concentration-penetration curve. This method has been used to determine the coefficients of self-diffusion of gold,13 zinc,14-l6 silver," and copper.17 The second method was used for the work described in this paper. Materials The coefficients of self-diffusion were determined in two grades of polycrystalline copper and in single crystals of copper. Commercial OFHC copper was used in the form of disks I in. in diameter by % in. thick. Samples of high-purity copper described by Phillips and Skinnerl9 were supplied through the courtesy of the American Smelting and Refining Co. The high-purity copper rod was used without remelting by cross-rolling to a thickness of 0.075 in., from which disks 7/8 in. in diameter were cut. A good grade of bus-bar copper was used to make the single crystals.

Jan 1, 1942

By Reed W. Hyde

Some information has been published on the sulphur dioxide concentration of gases from D wight-Lloyd machines sintering lead ores but most of this relates to the customary practice in which the charge is relatively high in lead and is sintered in two passes Over the machine. It is thought, therefore, that the following observations on sintering in one pass a low-lead, low-sulphur charge may be of interest. The data were obtained in sintering lead sulphide notation concentrates mixed with minus 20.mesh oxide ore, the mixture carrying 16,8 per cent Pb and 5.3 per cent S. No returns were used because of the limited amount of sulphur available as fuel, but the mixture pelletized readily. to make a satisfactorily permeable bed on the sintering machine. As there was a possibility that the SO z might be utilized, a series of runs was made to determine the concentration ofSO2 in the gases at different points along the wind box, and the effect on this concentration of changes in the wind-box vacuum. Several runs each were made with vacuums (as measured in inches of water at the wind box) of (1 2 to 1 in., ('1 4 to 3% in. and (3) 6 to 5% in. The averages of the determinations from each series are platted On the graph (Fig, 1). The vertical scale shows the percentage by volume of SOz (as determined by the Reich method), and the horizontal scale shows the time e~pressed as percentage of the total sintering time. Sintering was considered to be complete when no SOz could be detected in the gases. In all cases the depth of bed was 4 1/2 in. and the weight of dry charge per square foot of hearth area was 45.2 lb. Other data are given in the table. The volume of gases corresponding to the several points along the wind box at which SO2 determinations were made could not be measured, but has been calculated from the SO2 Content and sulphur elimination on the assumption that the volume is uniform throughout the sintering period. This is not strictly true, as with a given vacuum the volume usually will drop slightly toward the middle of the period and increase toward the end. However, this made is a reasonable approximation. several interesting features are brought out by these data. In general, the published data on SO 2 concentration in sintering under ordinary conditions show a high SO2 concentration in the early stages followed by a rapid decrease in concentration and with relatively little SOz in the gases from the last half or third of the operation. In other words, the greater part of the sulphur elimination is obtained in the first third or half of the period. In these runs, the 6-in. vacuum, which approximates the usual operating condition, gives a curve that follows this customary pattern. Maximum SO2 concentration is obtained 234 min. after

Jan 1, 1944

By R. J. Mellen

IN a chloridizing roasting and cyanidation plant treating a manganese-silver ore, certain experiments pointed out the way to an improvement in roasting conditions by a change in the degree of crushing. The following short article is simply a description of the various changes that were necessary throughout the plant and of the results obtained. There is nothing unusual in these changes but they are of some interest in that they constitute a good many of the alterations usually resorted to in increasing production. In the beginning, the changes were brought about by an experiment that was not successful, but which clearly pointed to means of improving the operation, and is an example of the fact that often experimental work may result in failure and still be of value in arriving at correct fundamentals. This experiment also shows how far a metallurgist can get out of line in his deductions, particularly in a field with which he is not thoroughly familiar. The ore at the plant of Compania Minera Kildun y Anexas is crushed through a 13-in. gyratory, a 3-ft. standard Symons cone crusher and a set of 16 by 42in. rolls, the latter in closed circuit with two Tyler 400-mesh screens. The crushed ore is then conveyed to two fine-ore bins. These provide storage and feed for two 6 by 90-ft. kiln roasters, which discharge directly to coolers. The coolers in turn discharge to a 6 by 12-ft. ball mill, where the ore is ground for cyanidation. Laboratory and pilot-plant tests previous to the mill construction had indicated that minus ¼ -in. crushing was satisfactory for roasting, although results did not vary greatly with some coarser or considerably finer crushing. The plant was designed for a capacity of 300 tons per day on this basis. After a few months operation, the initial operating problems were overcome and the anticipated milling rate was reached but the results being obtained were not really good. The plant seemingly did not have much excess capacity in any department but the roasting operation was the most critically limiting factor both as to tonnage and the results obtained. If this could be improved there seemed some possibility of improving the total tonnage rate to some extent. Segregation of the calcine into various screen sizes showed that the coarser sizes ,were not giving the expected recoveries. In all of the laboratory and pilot-plant testing, any limiting conditions as to the percentage of salt used, temperatures, feed rates, etc., had affected the fine material more than it had the coarse, and even in making laboratory tests at this time the mill conditions could not be exactly duplicated. Further laboratory testing indicated that a segregation of the crushed ore into a coarse and fine product for roasting might result in some improvement in roast-

Jan 1, 1945

By R.K. Matheson

The Nivloc mine is 9 miles west of Silver Peak, Esmeralda County, Nevada. It has been operated by Desert Silver, Inc., since the summer of 1937. The cyanide mill treats 190 tons of silver-gold ore per day. In March of this year the main operations were about 2000 ft. west of the main shaft, which served as the downcast air passage for the mine. The mechanical blowers in use were sufficient to ventilate the west end at that time, but it was foreseen that as the mining area was extended ventilation would be an increasingly difficult problem, especially during the summer months. It was decided, therefore, to continue a raise from a point 100 ft. above the highest mining level (440-ft. level) to the surface; a distance of 443 ft. at approximately 65°. From this point 40 ft. of drifting would be necessary to connect with a shallow shaft on the surface. (Fig. I.) All the raise was serviced from the 440-ft. level; costs and progress obtained in the initial 100 ft. are included in the following pages. Choice of Raise A raise in this locality would have four objects: (I) to provide an ample supply of fresh air to the working places; (2) to explore the vein from the highest mining level to the surface; (3) to provide waste fill for stopes in the area, and possibly provide a transfer for waste from the surface; (4) to provide an emergency exit in the west end of the mine. Before the mining of an ore shoot in the eastern part of the mine, a similar raise had been driven for purposes of ventilation and exploration. This raise and subsequent mining showed the vein to be noncommercial for a distance of about 250 ft. below the surface. However, in the area under operation, old surface workings showed occasional values. Thus from a development standpoint the raise was an excellent prospect. The raise chosen was a three-compartment, stulled raise having two chutes and a manway. The manway inside timber measured 4 ft. 8 in. by 4 ft. 6 in.; it carried air and water pipes, a ladderway and a slide for a timber skip. The two chutes would be used for drawing off the muck and the cleaning of the bulkhead, as well as for ore and waste passes, which would be especially useful for lateral exploration from the raise. The fairly large cross section of the raise (6 by 15 ft.) would provide a large quantity of waste for backfilling, and more could be obtained from a glory hole on the surface. Time Schedule As speed in completing the raise was the main requisite, three shifts were considered, but two shifts were decided upon when it was found that three would necessitate considerably larger fans than those available for clearing the raise. The main diffi-

Jan 1, 1943

By R. K. Matheson

The Nivloc mine is 9 miles west of Silver Peak, Esmeralda County, Nevada. It has been operated by Desert Silver, Inc., since the summer of 1937. The cyanide mill treats 190 tons of silver-gold ore per day. In March of this year the main operations were about 2000 ft. west of the main shaft, which served as the downcast air passage for the mine. The mechanical blowers in use were sufficient to ventilate the west end at that time, but it was foreseen that as the mining area was extended ventilation would be an increasingly difficult problem, especially during the summer months. It was decided, therefore, to continue a raise from a point 100 ft. above the highest mining level (440-ft. level) to the surface; a distance of 443 ft. at approximately 65°. From this point 40 ft. of drifting would be necessary to connect with a shallow shaft on the surface. (Fig. I.) All the raise was serviced from the 440-ft. level; costs and progress obtained in the initial 100 ft. are included in the following pages. Choice of Raise A raise in this locality would have four objects: (I) to provide an ample supply of fresh air to the working places; (2) to explore the vein from the highest mining level to the surface; (3) to provide waste fill for stopes in the area, and possibly provide a transfer for waste from the surface; (4) to provide an emergency exit in the west end of the mine. Before the mining of an ore shoot in the eastern part of the mine, a similar raise had been driven for purposes of ventilation and exploration. This raise and subsequent mining showed the vein to be noncommercial for a distance of about 250 ft. below the surface. However, in the area under operation, old surface workings showed occasional values. Thus from a development standpoint the raise was an excellent prospect. The raise chosen was a three-compartment, stulled raise having two chutes and a manway. The manway inside timber measured 4 ft. 8 in. by 4 ft. 6 in.; it carried air and water pipes, a ladderway and a slide for a timber skip. The two chutes would be used for drawing off the muck and the cleaning of the bulkhead, as well as for ore and waste passes, which would be especially useful for lateral exploration from the raise. The fairly large cross section of the raise (6 by 15 ft.) would provide a large quantity of waste for backfilling, and more could be obtained from a glory hole on the surface. Time Schedule As speed in completing the raise was the main requisite, three shifts were considered, but two shifts were decided upon when it was found that three would necessitate considerably larger fans than those available for clearing the raise. The main diffi-

Jan 1, 1943