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By George Krauss, Seth R. Thomas

One complete cycle, allstenite to martensite to aus-tenite, of Martensitic transformation was induced in a commercial 18 Cr-8 Ni stainless steel. Transmission electron microscopy showed that the reverse transfarmation, martensite to austenite, perpetuated the dense defect substructure of the martensite produced by the direct transformation, austenite to 'martensite, and reproduced the retained austenite matrix orientation. A general precipitation of lathlike M,,C, particles was noted to be present in, and is suggested as cause for, regions of martensite that Izad formed after a complete cycle of transformation. The habit plane of the precipitate particles was found to be within 5 a plane previously established as the habit of nitride precipitates of similar morphology in bcc iron. The direct transformation of austenite (y) to martensite (a) in metastable 300 series stainless steels has received considerable attention in the past,'-8 and the martensite product given special emphasis in recent years,9"*u but little information has been published about the reverse martensitic transformation during which bcc a martensite is converted to fcc y austenite."-" The reversal mechanism and the resulting crystallographic relationship between parent and product austenite are of fundamental concern. The reversal product, fully austenitic, and hereafter referred to as reversed austenite, exhibits a defect structure far different from its virgin counterpart. Some studies have been reported in which the structure of reversed austenite in iron nickel alloys is compared to annealed austenite with respect to austenite strengtheningM and Stabilization.' The martensite in these Fe-Ni alloys, however, shows a basic difference from that produced in stainless steels. Structurally, the Fe-Ni martensite is profusely twinned," while the (low-carbon) Fe-Cr-Ni martensite contains a high density of tangled dislocations and is apparently twin-free.I0 reediis" has recently shown in an Fe-16 Cr-12 Ni stainless alloy that this typical martensitic defect structure and the reverse transformation directly influence the structure of the reversal product. He found that much of the reversed austenite had returned to the orientation of the virgin austenite and that fine twins were formed in the austenite after the reverse transformation. This investigation was undertaken in order to describe as completely as possible the structural changes produced by a cycle of transformation in a typical 18-8 stainless steel. EXPERIMENTAL PROCEDURE The 301 stainless steel under investigation (17.44 Cr, 7.75 Ni, 0.80 Mn, 0.60 Si, 0.40 Mo, 0.40 Cu, 0.08 C, 0.03 N, bal Fe)was received as 0.060-in.-thick

Jan 1, 1968

By Avnulf Muan, Egil Aukrust

Activity-composition curves jor cobalt oxide and manganese oxide in solid solutions with spinel-type structure have been determined by studying the equilibrium between the spinel phase and a coexisting solid-solution phase in the tewzperature range 1100" to 1400"C. Within limits of experimental error, the system displays the behavior of a regular solution. The excess integral free energy of mixing is represented by the equation . The standard free energy of formation of cobalt manganite, CoMn,O,, from the oxide components COO mzd Mn2OS at 1200°C is calculated to be -8.6 kcal. AMONG the technologically and theoretically most important oxide phases are those which crystallize with the same structure type as the mineral MgO . A120,. Although the name spinel originally was used only for this mineral, the term is generally used today to designate the whole group of oxide phases with this same structure type. It is in this wider meaning that the term is being used in the present paper. The previous literature on spinels is very extensive. There are two aspects of spinels which have attracted particular attention: the structure, including the nature and distribution of the cations present, and the physical properties (electric, magnetic). The status of knowledge in these fields has been well-summarized in recent review articles.1'2 By contrast, thermodynamic properties of spinels are virtually unknown. Delineation of the activity-composition curve over the whole composition range of a solid-solution series with spinel-type structure is the objective of the present work. The system at elevated temperatures (>1000"C) constitutes such a series. The activity-composition curves for this system can be calculated from a study of the equilibrium between the above spinel solid solution and a solid solution (Co,Mn)O with sodium chloride-type structure. Data for the latter solid solution have become available recently.3 The system, in contact with metallic CO at 1200°C, obeys Raoult's law. The determination of the activity-composition curve for the spinel phase is based on the following relations. Consider the reduction-oxidation equilibrium for cobalt ions according to the equation: Because it is impossible in practice to measure single activities of electrically charged constituents, we will treat the equilibria in terms of electrically neutral components. Eq. [I] can then formally be written as follows: where and the activities of cobalt oxide in the (Co,Mn)O and the spinel solid-solution phases, respectively, K is a constant at constant temperature, and Pa is the oxygen pressure of the gas phase. The numerical value of K can be determined from a determination of the oxygen pressure P& for the equilibrium coexistence of Cos04 and COO in the binary system Co-0. Under these circumstances be set equal to 1, and hence K = [Po Eq. [3] can therefore be written

Jan 1, 1964

By J. M. Hodge

This paper represents a start on the problem of choosing alloy steels for a given application on the basis of toughness as indicated by the room temperature notched bar impact value at the desired hardness level, and as such, is very useful. I believe, however, that the authors will agree that it is only a start on this problem. The authors state that neglecting temper brittleness or other possible effects of the tempering temperature level, the impact values at a given grain size level will be determined principally by the hardenability. This is, of course, true inasmuch as the hardenability determines the microstructure, but it serves to emphasize the broad field that hardenability now covers, since this study apparently includes mi¬crostructures ranging from full martensite to coarse pearlite. Furthermore, the + or - 20 ft-lb spread in the correlation with hardenability, which certainly limits the usefulness of these results for many applications, indubitably reflects the wide range of microstructures corresponding to a given value of hardenability in terms of the difference between full martensite and asquenched hardness.

Jan 1, 1947

By J. W. Harden

THE observations here following are the result of an examination made as to the feasibility of extending the South Mountain Railroad from Pine Grove to Waynesborough, and of the commercial value of the ore through which such extension would be made to pass. The road at present built and in operation has a junction with the Cumberland Valley Railroad, half a mile east of Carlisle, and extends southward to Pine Grove, a distance of 172 miles. The extension proposed will add to it near 27 miles, making its total length 442, more than 23 of which will run through, or in close proximity to, beds of ore. Along the route taken there are a number of openings at which the ore is gotten, of which the Craighead bank, Huntzinger's bank, Lowrey's bank, Wenkoop's bank, the Henry Clay bank, the Fuller

Jan 1, 1873

By G. R. Kotler, W. A. Tiller, W. C. Johnston

Experiments carried out on 100 g samples of molten tin and Sn-Pb alloys showed that the magnitude of the field strength, H, does not effect the supercooling, aTn, at which the initial nucleus of solid forms. However, the nuwlber of nuclei forming per cc, n, is found to increase greatly with H for fixed ATN and to increase with ATN for fixed H. The value of n increases greatly with the presence of solute and is found to he indefiendent of the cooling rate of the simples. Further, the most effective time for stirring to enhance n is immediately following the initiation of recalescence. The effect of inducing nucleation with the aid of a mechanical impact both in the presence and absence of stirring has also been studied. All proposed mechanisms for the enhancement of n by physical means seem to be inadequate in accounting for these results. IN any real nucleation experiment there are two effects influencing the results that would not be predieted from classical homogeneous nucleation theory. These may be separated into chemical and mechanical effects. The chemical effects are twofold. 1) Nucleation catalysts are present which enhance the nucleation frequency of new grains over that predicted by classical homogeneous nucleation theory.' 2) Solute will restrain the growth of nuclei so that many others may form.2,3 The mechanical effects are also two-fold. 1) Vibration (sonic or ultrasonic) of the freezing liquid enhances the nucleation frequency of new grains over that given by the chemical effects alone.4-6 2) Macroscopic fluid-flow effects produce the same effect as l).'7' It is also possible to influence the nucleation experiment by certain physical effects such as the application of electric and magnetic fields.9 From classical homogeneous nucleation theory one would predict the following relationship between the nucleation rate, N1 (number of nuclei formed per unit volume per unit time), and bath supercooling, AT, for an isothermal experiment.

Jan 1, 1963

By P. W. George

THE rapid increase in cost of discovering new orebodies by churn drilling in the Tri-State district has led to some attempts to lessen the expense by using geophysical methods. Electrical prospecting was tried three or four years ago, but no successful results were recorded. In 1926, the Federal Mining and Smelting Co. engaged the North American Exploration Co., of Houston, Texas, to make some tests with the torsion balance. At first, it was thought that the most favorable conditions for discovery of ore deposits with the torsion balance would be in locations where the ore was known to occur at rather shallow depths, FIG. 1.-CROSS-SECTION OF DUENWEG LEASE. therefore the first trial was made at Duenweg, east of Joplin, where some zinc ore had been developed at a depth of 130 ft. The gravity gradients at 61 stations on a tract of about seventeen acres indicated plainly one large zone of maximum gravity and a smaller narrow zone of minimum gravity. Fourteen churn-drill holes, half of which struck ore, show that all the ore found was located outside the zones of maximum and minimum gravity. In the zone of maximum gravity the limestone beds were closest to the surface, and the first chert beds encountered were at the greatest depth. The results (shown in Table 1) indicated that the gravity of the limestone was greater than the gravity of the fractured chert together with its orebodies, and probably also greater than the bouldery soil and clay on top of the limestone. A cross-section of the Duenweg lease is shown in Fig. 1.

Jan 1, 1928

By Jan M. Mutmansky, Gordon H. Walrod, Y. J. Wang

The selection of a shaft size for ventilation purposes is a classic problem of concern to mining engineers. This presentation is concerned with the sizing of conventionally-sunk circular shafts with concrete or shotcrete linings. Costs of sinking shafts of three different depths, from 150 m to 300 m (500 ft to l000ft), and six different diameters, from 4.27 m to 7.32 m (14 ft to 24 ft), were obtained through questionnaires to shaft contractors. The primary objective of the economic model presented is the balancing of detailed annual capital and operating costs to achieve the minimum overall cost of the shaft installation over its life. The economic model is similar to previous analyses but treats several of the individual cost categories in an updated fashion. Included in the paper is a detailed analysis of the effects of varying power costs on optimal shaft size. Also included are a statistical analysis of current shaft costs and comments on the choice between shotcrete and concrete linings.

Jan 1, 1980

By M. E. Wadsworth, R. L. Pearson

THE dissolution of UO, in carbonate solutions can occur only il the tetravalent uranium is oxidized to the hexavalent state. The carbonate system is of particular interest because it provides a means whereby the oxidized uranium may be dissolved by forming the complex ion UO,(CO3)3-1. In the presence of oxygen the over-all reaction may be written UO+1/2O,4CO3T+2HCO2UO,(CO4)11+H2O [1] UO, was used in this study because it could be obtained in high purity and also because it is representative of the most refractory of the primary uranium minerals. The mineral ulrichite is the natural counterpart of the synthetic UO, used. The kinetics of the leaching of UO, should be similar for any of the uraninite type minerals. A pressurized system was used not with the hope it could be extended to commercial systems as a primary purpose but simply to provide the means whereby important temperature and pressure parameters could be varied for the evaluation of the kinetic processes. The importance of such pres-

Jan 1, 1959

By J. C. Bartlett

(Read at the Wilkes-Barre Meeting, May, 1877.) THE following discussion was suggested by an experiment of Mr. Krom, the manufacturer of air-jigs, to illustrate the superiority of air over water as a medium of concentration. The paper is written in the interest of no system of concentration, but simply to test the experiment, and, perchance, to add something to the general fund of information on the subject. To speak of testing an experiment by a theoretical discussion, may seem a misuse of terms, but the theories concerning falling bodies and the resistance of fluids are pretty well crystallized into laws, which may properly be used to show where experiments which seem to refute them were improperly performed, and that instead of refitting they only corroborate. It is well known that a sphere of galena 1/8 inch in diameter, and a sphere of quartz 1/2 inch in diameter, are equal falling in water ; that is, these two spheres, being placed together in a column of water at rest or in motion, will practically remain together, falling or rising together, or remaining in suspension. Mr. Krom, to show that these

Jan 1, 1878

By Kwong Yung Kwang

This mine, which is sometimes called Tong colliery, from its director, Tong King Sing, is about 80 miles northeast of Tientsin (see map, Fig. 1). Fig. 2 shows a vertical transverse section through the coal-beds. These beds have a dip to the south of about 45' and are interstratified with sandstone, shale, indurated clay and fire-clay. Of the

Jan 1, 1888

By Linn Bradley

F. H. WILLCOX, Pittsburgh, Pa. (communication to the Secretary*). -We must keep in mind, in balancing the savings-to be anticipated by the most efficient combustion of gas, the best heat absorption by stoves or boilers, and by heat conservation of the thermal equivalent carried in. blast-furnace gas-against the cost in operation, maintenance, depreciation and interest charges of the equipment required to realize the savings, that in boilers and hot-blast stoves we are basing our economies upon two different, combustion phenomena. In a boiler, evaporating water at a heat efficiency of 70 per cent., approximately 47,550 B.t.u. is consumed per boiler horsepower. Taking the blast, furnace gas cited in the above paper, we are using about 38 ¾ lb. of cold-washed gas per boiler horsepower. Using the hot-cleaned gas, the same equivalent heat is furnished by 34 1b. of the gas , This leaves of the 38 ¾ lb. required in the case of cold-washed gas, 4 ¾ lb. or 6,650 B.t.u. available for raising more steam when dry-hot gas is used, which means that an equivalent amount of coal firing can be dispensed with. Per ton of iron, this figure 2/3 X 12,000/38.75 X 4.75 = 980 lb. gas or 1,372,000 B.t.u. or 100 lb. coal. At $1.50 per ton of coal, the net saving per ton of pig iron produced by the use of dry-hot over wet-cold cleaning will, therefore, be about $0.06 2/3. At a stove heating 8,500 lb. of air per toil iron from 150° F., one-quarter of the dry-hot cleaned gas, or 3,000 lb., would heat the air to 1,450° F. 3,000 X 1,400 X 0.70 = (8,500 X 0.255) = 1,300. The cold-wet gas would heat the same amount of air to 1,280° F. 3,000 X 1,225 X 0.70 = (8,500 X.0.255) = 1,130.

Jan 4, 1917

By N. C. Gay, P. J. Van Der Veever

The Klerksdorp gold mining district, South Africa, experiences a relatively high level of seismicity with many seismic events of magnitude greater than 4. These large events appear to have foci close to fault planes, most of which displace the strata by several hundred metres over a wide area. To help understand the mechanism of these large mining- associated seismic events, in situ stress measurements were made at 4 sites using "CSIR doorstopper" strain relief techniques. The sites were chosen close to, or some distance from fault planes and other structures. The results show that at all 4 sites a relatively large horizontal stress is oriented approximately north- west. The magnitude of this stress ranges from 0,78-1,O of the vertical stress. A marked horizontal stress anisotropy, varying from 13-30 MPa is also present. The dominant structural features are large, north-east striking normal faults which extend across the entire length of the goldfield. Displacements across these faults vary from a few centimetres to several hundred metres. Bedding plane faults across which large displacements may occur are also common as are north and north-west striking dykes. The majority of the faults strike at approximately 90º to the maximum horizontal stress and at all sites a large compressive stress plunges in the direction of dip of the strata and fault planes. Thus a significant shear stress may act across suitably oriented planes causing them to be in a state of potentially unstable equilibrium. This abnormal stress situation would exist along the entire length of the fault

Jan 1, 1982

By K. G. Davis, P. Fryzuk

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

Jan 1, 1968

By W. T. Isbell

In the Parkes process of lead refining, after desilverization has been completed by means of the addition of zinc, there remains in the lead from 0.5 to 0.6 pct zinc. At this stage in the refining operations virtually all impurities are reduced to refined lead specifications except zinc and antimony. All excess copper, nickel and cobalt have been removed by the zinc additions. The silver remaining should not exceed 0.1 oz per ton. It then remains only to remove the zinc and antimony in order to produce refined lead. At the Herculaneum plant of the St. Joseph Lead Co., practice formerly was to complete the refining by means of oxidation in a reverberatory furnace. This process consisted of heating the lead to 1400O to 1500OF and allowing it to come in contact with excess air passing over the surface of the lead and periodically agitating the bath of lead with air. This procedure not only resulted in the loss of zinc but of a considerable quantity of lead converted to the oxide in the form of dross and fume. The fume, collected in the baghouse, together with the dross was returned to the blast furnace, thus adding appreciably to the circulating load on the blast furnace and refinery. In addition, operation and maintenance of the reverberatory furnace were expensive and required the performance of disagreeably hot and somewhat hazardous work. The process of dezincing with high vacuum was developed at this plant as a substitute for the oxidation process and finally led to lower operating costs, the practical elimination of dross and fume and the direct production of metallic zinc, which is very suitable for re-use in the desilverization process. Essentially this process consists of subjecting the lead to high vacuum, in a kettle provided with means for heating, a bell-shaped member open at the lower end so constructed that it can be immersed in the molten lead, and means for evacuating the interior of the bell and for sealing an agitator. Metallic zinc is thus produced and collects on the upper cooled part of the bell member. This metallic zinc is periodically removed and re-used in the desilverization process. The process is quite simple. It is conducted at a low temperature, thus permitting the use of iron equipment, and shows a material saving in refining costs over the old method. Equipment and Its Uses Degincing Machine.—The equipment employed in this process consists of a dezincing machine, a steel kettle of 120 tons capacity, a vacuum system and equipment for handling the dezincing machine (Fig 1). The dezincing machine consists of a compact assembly, suitable for handling with an overhead crane, consisting of a treating chamber and the agitating mechanism. Treating Chamber.—The treating chamber is a bell-shaped shell (4), similar in shape to a hemispherical kettle with a lid

Jan 1, 1949

By C. F. Seaman, R. G. Clay

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

By R. G. Clay, C. F. Seaman

The mines of The Tennessee Copper Co. are in the Ducktown Basin, in southeastern Tennessee. The ore is a heavy sulphide consisting principally of chalcopyrite, pyrite and pyrrhotite and in places running as high as 95 per cent sulphur and iron. Operating mines are the Burra Burra and Eureka. Development ' of the Boyd ore body was started in 1940. Late in 1939 the decision was made to change from fuse and cap to all electric blasting and at the same time electric cap lamps were installed to replace carbide. These changes were made to improve the safety and working conditions in the mines. Both the Bureau of Mines and explosives manufacturers recommended electric blasting as a safety measure. Until that rime, electric blasting had been used only for shaft sinking and in places where it was absolutely essential for safety. Very few men had had experience in this type of blasting, as the places where it was required were those that also required direct supervision. It was necessary for the mining staff to work out methods, procedure, and equipment to be used. In this they were ably assisted by representatives of the explosives manufacturers. Many mistakes were made and it is the purpose of this paper to describe the original plan and installation, show the major difficulties encountered. the remedies applied, and present the practices now being followed. Comparison of costs of the two methods will also be discussed. About six months were required to make the change and eliminate fuse and cap blasting entirely. The first electric blasting stations were installed in November 1939 and the last fuse was taken underground in May 1940. A very unusual condition forced the use of fuse for a short time in one heading after May, but that will be discussed later. Original Equipment and Installation A power line is generally recommended as a source of current where permanent blasting stations are used. The plan worked out for electric blasting was to tap 220-volt alternating-current lines when they were available and use 250-volt direct current from the trolley wire at all other places. A blasting machine would be available at each shaft for places where there were no regular blasting stations and could be used if a regular station should fail. The equipment decided on was No. 16 duplex reinforced lamp cord for lines to within about 50 ft. of the working place and No. 20 lead wire from that point to the shots. An open double-pole, double-throw switch with slate base was put in the circuit, so connected that the down position would close the line to the shots 2nd the up position would connect it to the Iine from the power contactor. This switch was located in the circuit as close to the blast as was safe (should the shots explode when it was placed in the shooting

Jan 1, 1943

AVERY great profit is derived from removing the gilding and retrieving gold, without destroying the works of silver or other metal. If this method did not exist, the greater part of the gold that is put on certain works would be lost. Even though you might think to scrape it off patiently with a sharp scraper, if it had been drawn with draw-plates, or if it were at the bottom of objects where the scraper could not enter easily, it would be utterly in vain. Therefore, I find that to do this it is necessary to use one of three methods. The first is when it is found on a thin object. Take a large crucible, fill it with mercury, and heat it on a little furnace. When you see that it begins to boil and simmer, put your work in it and stir it around, shaking and rubbing it well in this, and finally rub it with a scratching tool until you see that the mercury has drunk up and absorbed all the gold. Then, having done this, take the said mercury and pass it through an untanned leather purse, and in the bottom of this you will find all the gold like sand. Put this in the bottom of a crucible or in something else over a charcoal fire and evaporate it. If the objects are of a thicker form the gilding is removed from them in another way, by covering them with powdered sulphur and lighting a fire on them. When you see them burning well with the fire all around, beat them in a clean bowl of wood or something else and collect everything that separates with the sulphur from the work; this will be gold in the form of a leaf wherever it has been touched by the fire. The gold is then extracted from this burned sulphur by purifying it in a cupeling hearth with lead. The third method is to touch the gold that you wish to take off the works with an aqua fortis made from saltpeter, rock alum, vitriol, sal ammoniac, and some verdigris.* This acid dissolves and eats the gold by virtue of its corrosive power. Having thus touched the gold with this acid and having saturated the latter with a quantity of gold, put it in a cucurbit and evaporate it over a little furnace as is done with silver, collecting the acid and the spirits if you wish.

Jan 1, 1942

By Wilson D. Michell

The Magma copper vein trends east-west, dips 70" south, and cuts through a 6000-ft thickness of limestones, quartzites,. shale, diabase, and schist. The vein is itself a fault with a horizontal offset of 500 ft on the formation contacts in the two walls. There are two principal faults of large displacement and numerous smaller steep and flat faults that complicate the prediction of ore occurrence. A study of the geology, its systematic recording to facitliate ready reference, and graphic methods of interpretation of mineralization have aided the exploration, development, 2nd mining of the ore. In making predictions of the distribution of mineable ore bodies and their shape in critical areas, studies are made of "mineralization contours" and "mineralization trends," which are developed on a longitudinal section of the vein from variations in the concentration of the combined base metal content as shown by assays of available openings in the mine. Introduction This paper describes an example of geological work applied directly to mining and exploration problems in a deep vein deposit, and it is presented as being of interest especially to those engaged in working more or less similar veins. In 1944 to 1945 the writer spent a year as assistant to J. K. Gustafson in a thorough geological examination of the Magma mine. Much of the writer's subsequent independent work described in the present paper had its inception in ideas originated by Gustafson during that examination. Objectives of Geological Work at Magma Mine The objectives of geological work at Magma, as at operating vein mines of a similar nature, can be summarized as follows: 1. Direct assistance to the mining department by means of predictions as to the size, shape, grade, and location of ore bodies; by solution of fault problems and irregularities in the veins; by prepara-?ion of detailed ore-reserve estimates in specific areas; by keeping a complete record of the geology and assay results in the mine; and by guiding development work. In short, geology can aid in the realization of the prime purpose of the mining department, which is that of producing the greatest amount of ore in the most economical and expeditious manner possible. 2. Guidance for exploration, especially diamond drilling, in undeveloped areas of the mine. 3. Contributions to the understanding of the general geology of the deposit and its regional setting, including ore genesis, structure, and mineral relationships. Magma Ore Deposit The east-west trending Magma vein, dipping about 70" south, cuts through a 6000-ft thickness of concordant limestones, quartzites, shale, diabase, and schist, named in descending order. These

Jan 1, 1949

By Wilson D. Michell

The Magma copper vein trends east-west, dips 70" south, and cuts through a 6000-ft thickness of limestones, quartzites,. shale, diabase, and schist. The vein is itself a fault with a horizontal offset of 500 ft on the formation contacts in the two walls. There are two principal faults of large displacement and numerous smaller steep and flat faults that complicate the prediction of ore occurrence. A study of the geology, its systematic recording to facitliate ready reference, and graphic methods of interpretation of mineralization have aided the exploration, development, 2nd mining of the ore. In making predictions of the distribution of mineable ore bodies and their shape in critical areas, studies are made of "mineralization contours" and "mineralization trends," which are developed on a longitudinal section of the vein from variations in the concentration of the combined base metal content as shown by assays of available openings in the mine. Introduction This paper describes an example of geological work applied directly to mining and exploration problems in a deep vein deposit, and it is presented as being of interest especially to those engaged in working more or less similar veins. In 1944 to 1945 the writer spent a year as assistant to J. K. Gustafson in a thorough geological examination of the Magma mine. Much of the writer's subsequent independent work described in the present paper had its inception in ideas originated by Gustafson during that examination. Objectives of Geological Work at Magma Mine The objectives of geological work at Magma, as at operating vein mines of a similar nature, can be summarized as follows: 1. Direct assistance to the mining department by means of predictions as to the size, shape, grade, and location of ore bodies; by solution of fault problems and irregularities in the veins; by prepara-?ion of detailed ore-reserve estimates in specific areas; by keeping a complete record of the geology and assay results in the mine; and by guiding development work. In short, geology can aid in the realization of the prime purpose of the mining department, which is that of producing the greatest amount of ore in the most economical and expeditious manner possible. 2. Guidance for exploration, especially diamond drilling, in undeveloped areas of the mine. 3. Contributions to the understanding of the general geology of the deposit and its regional setting, including ore genesis, structure, and mineral relationships. Magma Ore Deposit The east-west trending Magma vein, dipping about 70" south, cuts through a 6000-ft thickness of concordant limestones, quartzites, shale, diabase, and schist, named in descending order. These

Jan 1, 1949

By F. D. Rosi

The slip and twinning behavior in extended titanium crystals were studied in some detail. The formation and appearance of coarse kink bands are discussed. Their crystallographic geometry was determined by X-ray analysis. A phenomenological interpretation of the complexities in kink band development is also presented. The critical resolved shear stress, coefficient of shear hardening, and plane of fracture were determined for several crystals extended at room temperature. THE slip and twinning elements observed in the room-temperature deformation of titanium were enumerated in a previous paper1 in which considerations were advanced regarding the nature and selection of these elements and their effect on the known mechanical properties of this metal. The present study concerns the crystallographic, microscopic, and mechanical aspects of flow in relation to the slip and twinning elements, and includes a prediction of slip systems, nature of slip and twin markings, inhomogeneities of plastic flow, and stress-strain characteristics. Unless otherwise noted, arc-melted titanium sponge (99.77 pct) was used in these experiments. The method of production of crystals, their dimensions, and the surface preparation for micrographic examination have been reported.&apos; For obtaining stress-strain characteristics, only those crystals which traversed the entire width of the specimen and were at least 8 mm in length were used. Tensile deformation of the crystals was performed with conventional grips for sheet specimens and a constant-stress loading beam, designed after the method of Andrade and Chalmers.&apos; The specimens were loaded by allowing sand to flow from a reservoir into a bucket suspended from the longer end of a balanced 6:1 lever arm at a rate controlled to load the specimen approximately 2 kg per min. Strain measurements were made using the Baldwin SR-4, bonded, resistance-wire strain gage, Type A-8, which permitted a reading accuracy of 2 microinches per inch. The formulas used in the evaluation of shear stress and shear strain, as in deriving the coefficient of shear hardening, are given by Schmid and Boasv n terms of the original orientation and change in length of the crystal. For calculating the critical resolved shear stress, the standard equation was used. The crystallographic nature of the unpredictable slip observed in a number of specimens was determined by the single-surface X-ray method of analysis as described in ref. 1. Experimental Results Prediction of Slip System: It was reported&apos; that room temperature slip in titanium takes place predominantly on {10i0} and in a <1120> direction, giving three potential slip systems. Hence, it should be possible to predict the operative primary system in the manner used by Taylor and Elam&apos; for alu- minum (i.e., the one with the greatest component of shear stress in the direction of slip). The stereo-graphic construction in Fig. 1 shows that this is true. In all cases, slip was found to occur initially on the (0110) plane and, in a number of cases, in the [21f0] direction. (The direction of slip was not determined for all orientations.) It follows that duplex slip can be expected when two slip systems are geometrically equally favorable for slip, as demonstrated in Fig. 2. In both crystals slip took place simultaneously on the (1700) and (olio) prismatic planes, as indicated by Fig. 2a and b. In the case of crystal B the movement of the specimen axis with increasing extension was toward the (10i0) direction, which represents the stable end orientation resulting from alternate slip on the [2ii0] and [1120] directions. This is analogous to the duplex slip process in face-centered cubic metals." Unpredictable Slip: In a number of crystals whose orientations were well within the operation of a single slip system, secondary slip occurred on planes not predicted by the criterion of maximum shear stress. Examples are shown in Fig. 3. In addition to

Jan 1, 1955