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By R. I. Jaffee, J. W. Roberts, D. N. Williams

DISPERSION hardening as an alloying process has aroused increasing interest in the past few years. This alloying procedure, in which an insoluble phase is dispersed randomly through a metal or alloy, is believed to be especially suited for the attainment of creep strength at elevated temperatures since the dispersed phase is metallurgically stable. Another advantage is that the physical properties of the matrix phase, such as electrical conductivity, remain relatively unaffected by alloying. These two characteristics resulted in interest in the applicability of dispersion hardening to copper and copper alloys. A study of copper alloys prepared by powder metallurgy methods has shown some strengthening from oxide dispersions.' However, powder metallurgy processing would not be economical for large scale production of dispersion hardened copper-base alloys. Therefore, the research described in this article was initiated to examine methods of forming dispersion alloys which would not require powder metallurgy processing. If powder metallurgy processing is to be avoided, a method must be found by which the dispersion material can be introduced into and dispersed through the melt before casting. When the dispersed phase is an oxide or similar nonmetallic, it is extremely difficult to form a satisfactory liquid: solid dispersion since successful mixing requires that the melt wet the dispersed phase. It is almost impossible to force a nonwetted particle beneath the melt surface. This difficulty can be avoided, however, if the oxide particles are produced in molten copper by reaction within the melt. The problem would then reduce to maintaining randomness until solidification. Four procedures were considered for producing an oxide particle within molten copper by reaction. In all cases, the intent was to react oxygen (0) with a reactive metal (X) to form an oxide (XO) within the copper melt. These reactions were as follows: 1) (Cu-X), + (O)g (Cu)1 + (XO)S [1] 2) (Cu-X)1 + (Cu2O)s (Cu)x + (XO)S [2] 3) (Cu-OK)1 + (Cu-X)1 (Cu)1+(XO)9 [3] 4) (Cu-OK)1 +(X)s (Cu)1 +(XO)s [4] The reactions were carried out in an 8-lb melt of 99.98 pct Cu maintained in the vicinity of 2200oF by induction heating. Where possible the amount of the reactants was controlled to give 1 to 5 vol pct oxide. Mechanical stirring during the reaction period was used to minimize agglomeration of the reaction product. The presence of a reaction product was determined metallographically by study of an ingot cast from the melt at the end of the reaction period. Reaction [I] was studied by dissolving aluminum in the melt followed by forcing oxygen through the melt. The melts became quite sluggish during the reaction period suggesting that A12O3 was formed. However, the reaction product rose out of the melt rapidly, possibly due to the flushing action of the oxygen stream, and little dispersed phase was formed. Reaction [2] was examined by dissolving aluminum, silicon, or nickel i; the melt and then adding Cu20 (melting point 2250F). This procedure was unsuccessful. Although Cu2O was easily stirred into cop-

Jan 1, 1961

By J. D. Forrester

I have availed myself of the opportunity to prepare, and to give this talk because, as incumbent Chairman, I am directed by the Bylaws of the Council of Education to make an Annual Report of the Council activities which have transpired during the past year. In addition, and inasmuch as I have this captive audience, I am taking the liberty to express also some beliefs I hold about various other elements of Mineral Industry education. First, my report is that constructive progress has been achieved in the Council of Education during 1960-61. These advancements have been particularly notable through projects carried on by the Committee on Meetings and Programs, by the Student Chapters Committee, by the Accrediting Affairs Committee and by the Committee for Securing Better Representation of Mineral Industry Colleges. I believe each of these groups will report concerning details of their respective programs at the Business Meeting of the Council. On the other hand, not as much has been done in some areas of work as I had hoped would be the case when I took office as Chairman last February. Let me make clear that these delinquencies, for the most part, have resulted through my ineptness in handling of affairs; they are not caused by the failure to perform various assignments which I have given to council members.

Jan 1, 1961

By J. W. Mills

In 1956 the discovery of nickel on Jumbo Mountain, Snohomish County, Washington, focused attention on this part of the Cascade Range, far more renowned for its timber than for its mineral resources. Hand specimens assaying as high as 13 pct Ni encouraged Discovery Mines Inc. of Mount Vernon, Wash., to stake 12 claims in Township 31 North, Range 9 East of the Willamette Meridian. The property was leased to Climax Molybdenum Co. in January 1957. Trenching, sampling, and geologic studies were carried out during the late summer of 1957 when most of the snow had disappeared. Although the deposits are only about four miles south of the logging town of Darrington and about two miles. as the crow flies, southeast of the end of a good gravel road from that town, all camp supplies and equipment were air-dropped, owing to the extremely rugged nature of the terrain. GEOLOGY The rocks of Jumbo Mountain comprise a series of folded Tertiary sediments forming a belt one half to three quarters of a mile wide, trending northwest about parallel to the strike of the beds. This belt is sandwiched between two extensive in-trusives—gabbro on the northeast and quartz dio-rite on the southwest—and has been intruded by dikes of fresh and serpentinized dunite. Shear zones follow many of the dunite-sedimentary rock contacts, and in these the nickel mineralization is concentrated. A few transverse faults offset the sediments and ultrabasics. Sedimentary Rocks: Two kinds of quartzite make up the greater part of the sedimentary section. One is dark purplish gray, very fine-grained, and thin-bedded; the other is light gray to white, medium-grained, and well-bedded. Locally the light-colored quart lites contain grit and conglomerate horizons, rich in pebbles of white chert and very fine-grained quartzite. Intercalations of dark argillites near the summit of the mountain give many of the cliffs a striking banded appearance. Dark gray to black, aphanitic, thin-bedded argillites occur throughout the entire sedimentary belt. Beds vary from a few inches to a few tens of feet thick and some, near the summit, contain numerous fossile (willow?) leaves. Only the thicker beds were mapped. The sediments are quite fresh on the northeastern side of the belt and become more metamorphosed (hornfelsed) toward the southwestern side. Brown biotite is present in the metasediments in increasing amounts as the quartz diorite is approached. Near the intrusion, porphyroblasts of plagioclase, up to 1 in. across, testify to the soaking of sedimentary rocks by solutions emanating from the quartz diorite. Most of the sediments mapped lie along the southwestern limb of an anticline, the axis of which follows the crest of the mountain and plunges at a low angle to the northwest.

Jan 1, 1961

By Albert W. Deurbrouck

The percentage of coal cleaned mechanically has increased greatly in the past 30 years. This technological change in preparation methods, accompanied by mechanical innovations in most phases of coal mining, has enabled the industry to supply quality coals at competitive prices. The Bureau of Mines has contributed to this progress by conducting research on coal-preparation problems, by evaluating the performance of domestic-and foreign-made preparation equipment, and by recording advances as they occur in industrial coal-cleaning practices. The Bureau's work in this field is recorded in the more than 150 technical and scientific publications listed in this volume. These reports, published by the Bureau over the past half century, document the Nation's progress in a vitally important phase of fuel technology.

Jan 1, 1961

By W. W. Mullins

The development, by the mechanism of volume diffusion, of a grain boundary groove on an interface separating a solid phase and a saturated fluid phase is calculated under the following assumptions: 1) isotropy of the interfacial free energy 's, 2) applicability of the Gibbs-Thompson formula relating curvature and chemical potential, 3) a nearly planar groove surface. The groove profile is found to have a .fixed shape and linear dimensions that are proportional to (time)1/3. The constant of proportionality is evaluated and involves ?s in such a way that the latter my be evaluated from grooving kinetics. A grain boundary that intersects an interphase interface will tend to produce a groove along the line of intersection (Fig. 1). The ultimate motivation for the formation of the groove is the reduction in interfacial free energ that occurs as the grain boundary contracts. Smith has discussed the way in which the free energy of the interfaces determines their dihedral angles at the groove root and has given a detailed analysis of the effect of the dihedral angles upon the nature the microstructures developed in polyphase alloys. Herring2 has extended the considerations to the case in which the free energy of an interface depends on its orientation and has thus added torque terms to the condition for interface equilibrium. Although the grain boundary and the interphase interface may quickly achieve the proper equilibrium angles where they intersect, as determined by the balance of equivalent tensions and torques,"2 the groove will generally continue to grow (though with decreasing speed). Thermodynamically, this continued growth reflects the fact that the lowest free energy cannot be achieved until the entire grain boundary has been eliminated; mechanistically, the growth occurs in response to inhomogeneities of chemical potential that are caused by the constraint of fixed dihedral angles at the groove root. The latter point is illustrated in Fig. 1, which shows that atoms* near the groove are on the convex side of a cylindrically curved surface and therefore have a higher chemical potential than atoms on the flat interface beyond. Thus atoms tend to leave the curved groove shoulders and force the groove to deepen. When both phases are solid additional chemical potential gradient may arise from whatever strain fields are necessary to insure continuity of material across the interface. In the discussion that follows, this complication due to strain will not occur since we always assume one phase to be a fluid and the other a solid. The transport processes that may operate to permit growth of the groove are surface (interface)* diffusion, volume diffusion in the phase on either side of the interface, and evaporation-condensation when one phase is gaseous. (Also a melting-freezing process is responsible for groove enlargement on a solid-melt interface as discussed in Sec. VI). The theory of grain boundary grooving has been discussed in detail for the processes of surface diffusion and of one type of evaporation-condensation.3'4 The theoretical predictions for the mechanism of surface diffusion have been confirmed in an experimental study of grooving kinetics in copper.5 In this paper, the theory of groove development by volume diffusion will be presented. Volume diffusion can occur either by atomic migration in the solid S or by diffusion of S atoms in the fluid phase, or solvent, adjacent to the groove surface. As discussed in Sec. V, volume diffusion in the solid should be less important than surface diffusion in forming a groove. Apart from verifying this statement then, the main interest of the present analysis centers on diffusion in an external solvent. The solvent may be either a liquid (e.g., water, or-

Jan 1, 1961

By W. S. Landers

In contrast to the large fund of information available on the carbonizing properties of Eastern U.S. coals, the experimental data available for most Western U.S. and foreign coals are meager. To supply some self-consistent data, the Bureau of Mines conducted 220 low-temperature carbonization assay tests at 500° C. in a laboratory-scale carbonization apparatus. The results are included in this report. The U.S. coals tested are from Alaska, Arizona, Colorado, Illinois, Montana, New Mexico, North Dakota, Pennsylvania, Ohio, Texas, Utah, and Wyoming. Assays were also made on coals from Canada, Ecuador, France, Greece, India, Mexico, the Philippine Islands, Thailand, and Yugoslavia. In addition, one assay of pine sawdust and one assay of a peat from Ireland are reported. Prediction equations were developed relating the yield of products of carbonization obtained at 500° C. in the assay unit to the proximate and ultimate analyses of the coals. The equations for predicting the yields of tar plus light oil, char, and gas, and the heating value of the gas, are considered good, each equation having a multiple correlation coefficient of 0.90 or higher. The equation predicting volume of gas obtained upon carbonization is considered as fair, with a multiple correlation coefficient of 0.84. A reproducibility study of the assay made on 10 replicate samples of a subbituminous coal indicated that the standard deviation of the individual products was 0.31 percent or less.

Jan 1, 1961

By J. B. Gayle

Coke-pushing difficulties have been the subject of much controversy but relatively little scientific investigation. Nevertheless, continuing pushing difficulties can greatly decrease the life of a coke oven and should therefore be considered in the selection of the blend and operating conditions used for coke production. The present Federal Bureau of Mines investigation was undertaken to correlate results of previous investigators and to supplement these results to the limited extent practical by us of an experimental slot-type oven. To this end, pushing pressures have been determined routinely for Tuscaloosa oven tests conducted during the past 10 years. Nevertheless, the accumulated information remains incomplete in many respects, and in view of the recognized inadequacies of the data, publication is now undertaken only because of the scarcity of published information on this important problem and the discontinuation of carbonization studies with this oven.

Jan 1, 1961

By D. A. Pirttinen

The promotion of improved coal quality for thermal generation can benefit our national economy. The construction of electric generating facilities in Canada is expanding at a rapid rate to satisfy the demands of a rising population and new uses for electrical energy. With remaining economic water power resources in the process of development, generation from thermal sources continues to grow. The increase in thermal capability has been somewhat arrested by a de-pressed economy, widening of inter-provincial and state transmission networks and decisions to exploit certain water resources before high capital investment is further in-creased by rising construction costs. The introduction of extra high volt-age transmission lines will further delay thermal generation planning, but in due course, when the effects of restrictive factors have been dissipated, steam-electric generation from fossil fuels will rise to considerable proportions. In time, nuclear sources will contribute a growing share to thermal power, but, it is doubtful if nuclear participation will ever displace conventional fuels for steam rising, unless it can be adapted to operate efficiently for variable outputs in peaking service. It appears certain that fossil fueled

Jan 1, 1961

By W. K. Hardy, H. H. Stadelmaier, L. J. Huetter

The ternary system Co-Sn-C was studied in the cobalt-rich region. The liquidus projection ulas established and two fonr -phase reactions were found. An isothermal section at 950°C was worked out. In the region studied the following phases were observed: cobalt-rich solid solution, graphite, binary compound y -cobalt-tin, and a ternary compound. As a result of the limited solubility of carbon in liquid cobalt-tin alloys, the investigation was limited to the cobalt-rich region. Tin contents up to 32 at. pct were studied. Cobalt of 99.99 pct purity, tin of 99.97 pct purity and synthetic graphite of 99.93 pct minimum purity prepare the alloys. Carbon determinations of all alloys were obtained by gas volumetric analysis. Samples were investigated microscopically and by X-ray diffraction using CoK, and FeK, radiation. Cooling curves were obtained with a number of alloys to determine the temperature of the four-phase reactions. The thermocouple was calibrated with the melting points of high-purity silver and copper. In the region studied the following phases were observed: Cobalt-rich solid solution (face-centered cubic) Graphite. Binary compound y (solidifies congruently at composition Cop~n)'. Ternary compound k which is face-centered cubic with superlattice and nonferromagnetic.' In Fig. 1 the fields representing the projections of the liquidus surfaces were determined by microscopic investigation and are labled by their primary phase. In the field labled C graphite is the primary phase. During solidification in this field much of the primary graphite was expelled through the surface of the melt. It could not always be established microscopically whether the remaining graphite was primary or purely eutectic. Therefore it was not possible to determine accurately the boundary curve which drops from the binary eutectic el at 1309 "~~ to the point of the four-phase reaction at F. The ter-nary carbide is the primary phase in the field labled k. To the right of this field the primary graphite is consistently nodular whereas flake graphite is observed at tin contents somewhat below F. The ternary carbide does not appear to solidify congruently. The graphite is attacked in a peritectic reaction along the boundary between k and C. Thus no graphite is observed in alloys in the k field. In alloys right of the boundary between k and C the graphite nodules are located in the center of the k crystallites, further documenting the peritectic nature of the reaction with the graphite. A four-phase reaction was found at F and 1120 &3"C. The most likely reaction is L+ CY-Co + C — k. Below the temperature of the binary eutectic ez at 1112" a second four-phase reaction was observed at E and 1100 lt3"C. It is the ternary eutectic L — CY-Co + k + y. Figs. 2(a ) to (e) show microstructures of the ternary alloys in the as-cast condition. Fig. 2(a) shows the eutectic a-Co + C (with some primary graphite) with the carbide k in the grain boundaries. Fig. 2(b) shows primary cobalt (dark areas) and the eutectic a-Co + k which is formed along F-E.

Jan 1, 1961

By R. S. McClain

The Federal Bureau of Mines investigated residues from aluminum plants, to determine if carbon could be removed by flotation while recovering fluorine and aluminum compounds for recycling to the reduction process. This publication summarizes studies on samples representing three types of plant waste products, supplied by three aluminum-producing companies. Samples of flue dust, pot skimmings, and pot linings from the various plants all responded to a simple turpentine flotation scheme to reject 90 to 97 percent of the carbon. Aluminum compound recoveries ranged from 42 to 94 percent on the flue dust, were about 95 percent on the pot skimmings, and ranged from 75 to 89 percent on the pot linings. Fluorine recoveries from the same samples ranged from 35 to 63, 92 to 95, and 77 to 86 percent, respectively. The pot linings contained soluble salts of aluminum and fluorine which were recovered by precipitation with sodium aluminate and carbon dioxide gas. Composites of residues from each plant responded to flotation much the same as the individual samples. Aluminum is produced by the electrolysis of alumina in a molten cryolite bath. The electrolytic reduction cell consists of a carbon-lined box containing a pad of molten aluminum which serves as a cathode, a carbon anode, and an electrolyte of molten cryolite in which the alumina is dissolved. Alumina is reduced to aluminum at the cathode, and carbon is oxidized at the anode.

Jan 1, 1961

By R. B. McIntosh, S Ramanchuk

A brief description is given of the process used by Sherritt Gordon Mines Limited to recover cobalt metal from nickel-cobalt sulphides. A description of the equipment with the operating conditions is presented with an evaluation of the corrosion test results for various stainless steels and other acid resistant materials for the part of the process involving the sulphuric acid leaching of nickel-cobalt sulphide slurry under air pressure. SINCE 1954 the Sherritt Gordon ammonia leach process has been used at Fort Saskatchewan to recover nickel metal, copper sulphides and cobalt-nickel sulphides from nickel concentrate produced at Lynn Lake, Manitoba. 'To develop a process for recovering cobalt and nickel from the cobalt-nickel sulphides, studies were carried out in the Sherritt Gordon laboratory and pilot plant and, late in 1957, continuous acid leaching of sulphides was started in the refinery. In September 1958 the present equipment

Jan 1, 1961

By R. W. Wotring

Mr. Chairman, Officers, Members and Guests of the AIME, I want to express my appreciation for the privilege of being on your program at this meeting. Coal mining, in recent years, has become much more competitive than it formerly was. To stay in business, it has been necessary for coal mines to mechanize. This, in itself, has not been enough for after one has mechanized his mine, he must keep replacing the equipment with the latest type and model machines. We, at Olga, have been doing this very thing for the past several years. Olga No. 2 Mine is located at Caretta, West Virginia, in McDovell County, about twelve miles south of Welch, It is a shaft mine, operating in the No. 4 Pocahontas seam. The thickness in the area, we are presently mining, varies from five to six feet. In some areas the immediate roof is slate, ranging in thickness from a few inches to four and five feet; in other places, the sandstone lies on top of the coal, This sandstone is about twenty-five feet thick.

Jan 1, 1961

By R. J. Leary

Iron and steel scrap has accounted for 45 to 50 percent of the metallics used by the steel industry. In an electric furnace which utilizes a 100-percent cold charge consisting of about 99 percent scrap and 1 percent pig iron, sensible heat from molten pig iron is not present as it is when hot metal is charged to open hearths and converters. A study of means for more efficient utilization of iron and steel scrap in steelmaking indicated that the time and fuel requirements to melt scrap were important cost factors, particularly in arc-furnace steelmaking. A process to preheat scrap economically prior to charging into the arc furnace has been developed, utilizing efficiently the heat from combustion of fossil fuels as a substitute for electric-arc heating in the initial stage of arc-furnace steelmaking. Scrap is loaded into a portable vertical cylindrical furnace body that is open at the top and has a perforated dropout door at the bottom. In the heating operation, the furnace body containing scrap is placed upon a hollow base connected to a flue, and then the burner and combustion-chamber unit is lowered on top of the furnace body. The stream of hot-combustion products flows down through the bed of scrap imparting heat to it, and the cooled gases exhaust through the flue. After the heating operation the furnace body containing hot scrap is transported to the arc furnace for charging. Thus, the furnace body serves as a charging bucket for top-charged arc furnaces, enabling this preheating process to be integrated into existing material-handling patterns in many arc-furnace plants.

Jan 1, 1961

By Robert H. Merrill

This report summerizes the various studies made to determine the size and support of roofs in the upper shale formation, White Pine Copper Mine, Michigan. These studies were a cooperative effort of the White Pine Copper Company and the Federal Bureau of Mines. Before this investigation, the upper shale formation had been found to contain about 30 to 40 percent of the total mineralization in the ore zones. Investigators estimated that this shale could be economically mined by room-and-pillar methods, provided that the roof would stand over spans of 25 to 32 feet. The problem remained to determine the number of bolts required to safely support these spans. Unbonded fractures in the roof structure of the deposit complicated the use of laboratory methods for determining bolted roof support. Therefore, a method was devised to determine the in-place support which would consider the unbonded fractures. In this method the roofs of 3 rooms with spans of 20, 28, and 36 feet were intentionally overbolted as the rooms were mined. Following stability tests that sometimes exceeded 6 months, the bolts were selectively removed and roof sag, roof strain, and changes in bolt loads were measured concurrently. The number of bolts required for support of these spans was computed from these data.

Jan 1, 1961

By V. T. Dow

In April 1957 the Bureau of Mines instituted a reconnaissance of titanium-and zirconium-bearing heavy sandstone deposits in the Four-State area of Utah, Wyoming, New Mexico, and Colorado. This report summarizes the study. Eighty-two deposits or groups of deposits, containing an estimated 27,920,000 tons of titaniferous sandstone, were examined in the Four-State area; 16 in southern Utah, 19 in Wyoming, 33 in New Mexico, and 14 in southwestern Colorado. The 16 deposits examined in Utah contain an estimated 1,043,000 tons with an average grade of 17.98 percent TiO2, 5.67 percent ZrO2, 20.98 percent Fe, and 0.09 percent equivalent Th02. The 19 deposits investigated in Wyoming are estimated to contain 21,873,000 tons of titaniferous sandstone having a weighted grade of 5.22 percent 702, 0.55 percent ZrO2, 24.8 percent Fe, and 0.01 percent equivalent Th02.

Jan 1, 1961

By W. L. Dare

The Federal Bureau of Mines is publishing a series of studies on the methods and costs of mining uranium on the Colorado Plateau. This report describes the longwall-mining methods and techniques used at the Radon mine, Big Indian district, San Juan County, Utah. The tabular Radon ore body lies on the southwest flank of the Lisbon Valley anticline, strikes N. 10° W., dips 7° SW., and has an average thickness of 4-1/2 to 5 ft. The mine is developed by a three-compartment vertical shaft, a haulage level driven beneath the ore, and a system of drifts driven within and parallel to the strike of the ore.

Jan 1, 1961

By J. B. Clark, A. S. Yue

PfANN1 has shown that when a cylinder of molten binary alloy freezes directionally, the solute distribution can be described analytically by cg=k(l-g)k-1 [1] in which Cg is the solute concentration in the solid at the point where a fraction g of the original liquid has frozen and Co is the initial solute concentration. The term k is the effective distribution coefficient of the solute and is defined as the ratio Cx/CL where Cl is the solute concentration in the main body of the liquid. Both the slope and the intercept at -log (1 - g) = 0 of a logarithmic plot of Eq. [I.] provide a measure of k. Under conditions of freezing where diffusion in the solid is negligible and the solute concentration in the liquid is uniform, the effective distribution coefficient k, derived as above, is equal to the equilibrium distribution coefficient taken from the phase diagram. However, under ordinary freezing conditions, even though solid-state diffusion is negligible, the solute rejected from the freezing solid does not have sufficient time to diffuse uniformly throughout the main body of the liquid. A concentration gradient is formed in a boundary layer in front of the advancing solid-liquid interface as pointed out by Rutter and Chalmers.' Thus, under these conditions, the concentration of the freezing liquid in the boundary layer is higher than that predicted from the equilibrium phase diagram and consequently the effective distribution coefficient k, obtained from directional freezing experiments, is larger than the equilibrium distribution coefficient. The symbol k will designate the effective distribution coefficient in the discussion which follows. In this paper, the determination of the effective distribution coefficient is demonstrated for several dilute binary alloys and a polynary magnesium alloy. The effect of freezing rate on the value of the derived k is presented in detail. An example for the calculation of the solidus in a dilute binary solution using the experimentally determined effective distribution coefficient is given. MATERIALS STUDIED Binary magnesium alloys of copper, nickel, and zinc containing approximately 0.10 wt pct solute and a polynary alloy containing approximately 0.10 wt pct each of several elements were selected for study. The spectrographic analyses of these alloys are given in Table I. The analyses are believed accurate to within +5 pct of the stated value. EXPERIMENTAL PROCEDURE Extruded rods of the alloys, 3/8 in. in diam and 11 in. long were used for directional freezing studies. The alloy rods were melted horizontally in a high-purity graphite boat by a Nichrome-wound tube furnace which moved parallel to the length of the rod. The boat was supported inside a Vycor tube suitably connected to provide a protective atmosphere of dry SO2. For an experimental run, the following procedure was used. The alloy rod was cleaned in dilute HNO, and dried. The graphite boat containing the alloy rod was placed within the Vycor tube and the tube connected to the gas train' The 'ystem was purged for 20 min with dry argon and then a flow of dried SO, was maintained for the duration of the run. The movable furnace was positioned and either all or a desired portion of the alloy rod was melted. After attainment of steady-state temperature gradients the molten bar was solidified from one end to the other at a controlled rate by pulling the furnace along the length of the bar at the desired speed. At the conclusion of the run, the directionally frozen alloy rod was analyzed spectrographically at 1 1/4-in. intervals.

Jan 1, 1961

By Rudolph G. Wuerker

Charles T. Holland (Head, Dept. of Mining Engineeri*, Virginia Polytechnical Inst., Blacksburg, Va.) Mr. Wuerker has presented a very interesting discussion of the use of triaxial test methods for investigating the strength properties of rocks. Such methods, no doubt, eventually will develop considerable information of interest to those concerned with the design of mine layouts, particularly in the field of pillar design. From his discussion of my recent article, "Cause and Occurrence of Coal Mine Bumps" (Holland Mining Engineering 1958, p. 933-1002), it is evident that in one place at least I did not make my meaning clear to him and perhaps others. To clear the matter up I think it best to quote from the article, somewhat more fully than did Mr. Wuerker, as follows: "4) In actual operations — because rocksare not perfectly elastic, homogeneous, nor isotropic and because local yield does occur — the maximum stress as demonstrated by Phillips (Ref. 22, pp. 64, 65) and indicated by much experience in mining, does not occur at the walls of the opening but at a short distance inside the pillar. Furthermore, the maximum stress does not reach as great a value as theoretical considerations and laboratory experimental methods indicate.* Actual distance inside the pillar, measured from the wall, at which the maximum stress exists, has not been determined. Observations in many mines, however, indicate that this distance could have a mini-value of one to six or eight times the bed thickness and that it is probably affected by width and height of the opening, depth of cover, and relative values of the elasticity and plasticity of materials comprising the roof, floor, and coal seam. The actual value of the stress produced probably lies between the theoretical maximum and the average stress concentration that would be produced if the weight of the strata above the unsupported opening were evenly distributed over the pillars for a distance equal to the opening width." The footnote reference in the above quotation referred to the following: "*For example, the Pocahontas No. 4 coal bed in southern West Virginia is mined under cover up to 1800 ft thick. Development openings are driven 18 to 20 ft wide, and the bed is about 6 ft thick. According to the work of Panek, the tangential wall stress at mid-bed height under these conditions would reach values between 4000 and 5000 psi. Actual tests of 3-in. cubes of this coal show its compressive strength would be much less than this, perhaps as low as 400 psi. Yet the pillars usually show no evidence of failure in these headings. In this same bed at a depth of 800 ft, the author has seen an opening 225 ft between supports lying between two old groves approximately 1100 ft apart. According to the theoretical considerations, the stress in the pillar walls would have been about 18,000 psi, yet the pillar showed little or no evidence of weight. In view of these observations, it is clear that the wall stress does not attain the maximum values indicated by theory." (Underlining added to original wording.) By referring to Fig. 2A of my paper it will be noted that theoretically the maximum pillar stress would occur at the pillar wall, i.e., at the passageway surface of the pillar. Obviously this cannot be correct in the cases of stress ranging from 4000 to 18000 psi since the coal at the surface of the pillar is under no constraint and cannot have a strength much greater than 400 or 500 psi. Hence, my conclusion that the maximum stress does not occur at the wall but back in the pillar some distance from the wall. Since these stresses are pushed back in the pillar from the wall, it is also obvious that the loads transferred to the pillar from the opening will be spread over a greater area and hence Pillar stresses will not rise to the values postulated by theory and photoelastic experiment. Further since to visual inspection the coal along the pillar wall did not appear to be failed the conclusion was reached that the stress shift was caused by local elastic or plastic yield and by difference in the elastic modulus of the rocks composing the mine floor, mine roof, and coal bed. Later on under the heading "Strength of Mine Pillars" (pages 1000-1002) the effects of constraint is briefly described. Also a formula taking into account constraint is developed relating pillar strength to the uniaxial strength of coal and the L/T ratio of the pillar. Since my paper was written, reports of experiments conducted in South Africa (Denkhaus, et. al., 1959), in Sweden (Hast 19581, and in Canada (McInnes, et.al., 1959) reveal that the conclusion expressed relative to the existence of a low stress area existing around the edges of pillars and solid faces as described above is generally correct. But it seems possible that where the wall stress developed is less than the unconfined strength of the rock composing the pillar and where the roof, floor, and pillar

Jan 1, 1961

By W. A. Backofen, R. L. Fleischer

AN effect of orientation on the creep behavior of aluminum at 4.2 OK has been observed. Stress relaxation was measured in a hard-type tensile device after stopping the drive. From the spring constant of the device the load-time record could be converted to strain as a function of time. The strain sensitivity was 2-10-=, and the usual creep time of 1 min normally allowed a stress drop of about 25 g per sq mm. Further details of the apparatus and sample preparation appear elsewhere.' It should be noted that neither stress nor strain were constant, so that quantitative comparison with current creep theories will not be made. Shear strain (y) as a function of time (t) was consistent with logarithmic behavior: y = A In t, where A is a constant. In some cases t was made as large as lo3 sec. In the results the quantity A divided by t (the resolved shear stress, which is the driving force for creep) is plotted as a function of strain prior to creep. The quantity A/t is thought to indicate qualitatively the ease with which dislocations can move during creep. In Fig. 1 are drawn tensile stress-strain curves and A/T-strain measurements for single crystals of three orientations: l(a) and l(b) for the maximum resolved shear-stress orientation, l(c) for an orientation 2 deg from [211], and l(d) for a [111] stress axis. The following observations were made: 1) At low strains the A/t values are nearly the same for the three orientations. 2) As strains increase, A/t drops most rapidly for [Ill.] and less rapidly for [211], while for the maximum shear-stress orientation it is constant during easy glide and drops abruptly at the termination of easy glide. 3) In stage 11, A/t values are again approximately the same for the three orientations. These observations are regarded as evidence that 1) the initiation of slip produces obstacles of essentially the same effectiveness for all of these orientations, while 2) the rate at which slip obstacles are produced before stage I1 is fully developed is strongly controlled by the amount of slip on more than one system. This second statement is consis-

Jan 1, 1961

By L. Moyd, P. Moyd

The first commercially available portable gamma ray-neutron beryllium detector, the Rerylometer, was developed by the same group that developed the first practical portable scintillation counter, the Scintillo-meter. This group includes Professor George Brownell, of the Geology Department of the University of Manitoba in winnipeg, and his associates, K. I. Roulston, R. w. Pringle, and R. L. Funt, of the Physics and Engineering Departments of the University of Manitoba. A prototype model was tested in the field by Dr. Brownell during the summer of 1958. Commercial production was undertaken shortly thereafter by Nuclear Enterprises Ltd., of Winnipeg, and a license for manufacture and distribution in the U.S. was granted to Isotopes Specialties Co., of Burbank, California, a division of Nuclear Corporation of America. Early in 1959, we were retained by the 'Mining Division of Minerva Oil Co. to Legin a search throughout North America for minable deposits of beryllium, and to train members of their permanent staff to continue the program. The project included literature sur-veys, interviews with members of various federal, state, and provincial agencies, scanning of our own collection of suites of specimens from various mineral occurrences, and the examination of deposits in the field. The value of a beryllium detector to this program was obvious and arrangements were made to procure one as soon as possible, which resulted in our obtaining and using the first instrument that was produced commercially in the U. S. In addition to the commercially available Berylometer, Mr. William Vaughan and his associates, of the U. S. Geological Survey Geophysics Division, in Denver, have developed and tested several experimental instruments based on the same principles. DETECTOR Principles The detection of beryllium is based on the fact that "hard" gamma rays (1.6 MEV and geater), such as those produced by Antimony convert ordinary beryllium (Be9) to Be8 , with the release of a neutron. The neutrons are counted by a scintillator designed to favor neutron responses while suppressing those of gamma rays. 'The instrument is completely selective for beryllium. No other element can interfere. Instrunlentation The Rerylometer itself, which is cylindrical, is about 8 inches in diameter, 13 inches high, and weighs about 11 pounds. (Figs. 1, 2). The principal components include: 1. A disc of wax to moderate (or "thermalize") neutrons. 2. A boron-enriched polyester-zinc-sulphide fluorescent screen. This material produces strong flashes when struck by thermalized neutrons but only weak flashes from gamma rays. 3. A photomultiplier tube. 4. An electronic circuit, complete with one srnall 71/2-volt dry battery. The photomultiplier tube and electronic circuit convert neutron flashes to electrical impulses, but gamma ray effects are almost

Jan 1, 1961