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By J. H. Frye, R. M. Treco

ONE of the chief hindrances to an understanding of the hardness of solid solutions is the sparsity of suitable hardness data. There is great need of a large body of hardness data obtained from many different alloys of high purity by identical testing methods and corrected to correspond to a single grain size. A considerable amount of such data has been obtained by means of the Meyer analysis on copper-rich and silver-rich solid solutions. There are enough of these data to permit comparison of numerous alloys at equal solute concentration in these two solvents, but only in a few cases is it possible to construct hardness-concentration curves for any given alloy. Therefore additional hardness work on most of these alloys is desirable. The primary solid solutions of silver with cadmium, indium, tin, or antimony as solute are of special interest, since all elements are from the same period of the periodic table and since, as shown previously,1 the hardening effect of equal percentages of these solutes is proportional to the lattice expansion produced by them. This paper presents the relation between hardness and concentration in the silver-antimony system from very dilute to nearly saturated solutions of antimony in silver. EXPERIMENTAL PROCEDURE Alloys were made from silver, 99.99 per cent pure, and antimony, 99.975 per cent pure. Silver and antimony were placed in graphite crucibles that had an ash content of less than 0.75 per cent. These metals were melted and allowed to solidify in a vacuum, which, after melting was complete, was usually equivalent to 3 mm. Of mercury. A high-frequency furnace was used. The ingots so obtained were worked and annealed to lessen segregation and to control grain size. Annealing was done in evacuated and sealed glass tubes. Hardness measurements on specimens from these ingots were by means of the Meyer analysis.3,4 Using a 4-mm.. ball, impressions were made under seven loads from 25 to 400 kg. The diameter of impression made under each load was measured. This enabled calculation of the constants in the well-known formula of Meyer. L=ad¬n where L = load, d = diameter of impression, a = a constant, n = a constant The Meyer hardness is defined as P = L/(1/4pd2) or P = 4adn-2/p Meyer hardnesses corresponding to d = 1 and to d = 4 are of especial interest, since the first corresponds to comparatively slight plastic deformation and the second to the maximum plastic deformation producible by penetration of the ball. These

Jan 1, 1944

By A. D. Smigelskas, E. O. Kirkendall

THIS is the third paper in a series on the diffusion of zinc in alpha brass.1 At the time of the first-paper it was accepted that diffusion in a substitutional type of solid solution depended upon an equal interchange of solute and solvent atoms across the diffusion interface (or past a given lattice reference point). Discrepancies in the size of the areas under the diffusion curves on each side of the interface, if observed, were explained on some other basis than the fact that the solute atoms might cross the diffusion interface without an equal number of solvent atoms diffusing across the interface in the opposite direction. The second paper presented evidence indicating that zinc atoms in alpha brass diffuse across the interface more rapidly in one direction than the copper atoms diffuse across the interface in the opposite direction. This phenomenon forced the interface to move in the opposite direction of the diffusion of zinc atoms to provide space for the additional zinc atoms dissolved in the alpha brass. That the interface and columnar grain structure on one side would remain intact during this mass movement was not expected; however, it is gratifying that they did. While the evidence was sufficiently satisfying to the author, it was decided that additional experiments were in order to prove that zinc diffuses more rapidly than copper in alpha brass. It is the purpose of this paper to present the results of these experiments. PROCEDURE A bar of 70-30 wrought brass was selected, to avoid the relatively large volume change in going from beta to alpha brass as in the previous investigation. A bar 7 in. (18-cm.)* long and about ¾ in. (1.9 cm.) wide was surface ground and polished with abrasive paper on plate glass to 0.4184 ± .0002 in. (1.0625 ± 0.0005 cm.) thick. Six 0.005-in. (0.013-cm.) diameter molybdenum wires, chosen because of their very low solubility in copper and alpha brass, were stretched lengthwise along each of the two plane surfaces. Glass tubes placed at right angles to the wires every ½ in. held the wires tightly against the brass. The sample was then emersed in an acid copper sulphate plating bath and electroplated to a depth of about 0.020 in. (0.050 cm.) copper on all sides. The glass tubes were no longer necessary to hold the molybdenum wires in place and were removed. The sample was then electroplated to a depth of

Jan 1, 1946

By Richard McCaffery

THERE has been some discussion among blast-furnace operators regarding the relative merits of tuyeres of the converging type and tuyeres designed in an attempt to produce a diverging jet of air. An article1 on Progress in German Blast Furnace Practice indicates the interest in this subject abroad. This article illustrates a diverging type of tuyere which is shown here in Fig. 1A. The design appears to be rather poor as regards known properties of air flow. The adoption of this. tuyere, the article states, resulted in a decrease of blast pressure of about 4 lb. per sq. in. and an increase of the output of the furnace. If the improvements mentioned are due entirely to the change in tuyere design, the subject of tuyere design seemed worthy of investigation and this paper is a report of such work. Because of the gain in efficiency brought about by the use of a diverging tuyere in the German plant, it was thought desirable to investigate the shape of the air streams issuing from various-types of tuyeres to ascertain whether the gain was due to the tuyere itself or to the distribution of the air in the furnace. Since it is difficult to do experimental work on a full-sized furnace, or even on full-sized tuyeres, and because of the equipment available for the work, it was decided to carry out the work on scale models of tuyeres. Therefore, in this work, single tuyeres were used and the blast discharged into the atmosphere. The shape of the air stream was determined by means of a Pitot tube suitably supported in the air stream at various positions. The air was. furnished by a Roots blower driven by a 10-hp. direct-current motor having field control, and capable of discharging 250 cu. ft. per min. with a pressure on the discharge of 4 in. of mercury. A modern blast furnace with 21-ft. hearth requires about 50,000 cu. ft. per min. of free air at from 16 to 23 lb. per sq. in. pressure, depending upon the type of burden. This pressure is the total pressure required to force the air through the burden and through the tuyeres. Upon computation the pressure drop through the tuyeres is about 1 to 2 lb. per sq. in. By making the tuyeres to a one-fifth scale, and by using a pressure of 1.5 in. of mercury, the shape of the air streams should be

Jan 1, 1931

IN general in making such powders, all kinds of gravel, tuff, washed river silt, and similar earths whose grain is fine and lean by nature are good for this operation of casting, either by themselves or in mixture because they are disposed to receive the metals well on account of a certain dryness which they possess. Many kinds are also compounded by means of art. Of these, I shall proceed to tell of all those that experience [119] has taught me to recognize as good. But first I desire that we return to speak of the natural ones because they are real earths and can always be had wherever one may be, and in the necessary quantity. They please me greatly because they are easy to reduce on account of their disposition. A loam is made from these and mixed by beating with wool-cloth cut- tings, spent wash ashes, and horse dung. This is made into cakes and dried. These are then put to bake in a furnace or in some other way and, in fact, are baked very well. Then it is pounded and sifted with a fine sieve, or it is ground in a potter's color mill, or by hand on a porphyry with water, to the fineness that the craftsman desires, or as fine as he can make it. When it has been ground it is again drained free from water and dried out with fire. Then as much magistery of salt is taken as will moisten it. It is dried and pounded again and passed through a sieve. When it has been made as desired in this way it is moistened again before use with water, urine, or vinegar just enough so that it holds together when it is squeezed in the fist. Then, when it has been brought to this point, it is moulded as you will hear. Powders are also made of crushed brick, tripoli, vine ashes, tiles, and glazed drainpipes, or burned emery, calcined tin, straw, and of burned paper and horse dung as well as of young ram's-horn ashes and many other things. The goodness of all these depends on three things; namely, on receiving the metal well, on being so fine as to be almost impalpable, and on their being made .with a magistery that renders them hard and strong when they are dry. In addition to these powders I have used another when I needed it, and it has succeeded well every time I made it. To make this. I have taken I two parts of pumice and one of iron scale and ground them on the por-

Jan 1, 1942

By Henry S. Drinker

THE Butler Mine property is situated in the vicinity of Pittston, in the Wyoming coal-field of Pennsylvania. The coal has been worked out from the fourteen-foot or Baltimore vein for a number of years on part of the tract, the old chambers remaining open as when originally abandoned. This vein outcrops on the Butler property, and is nowhere more than sixty feet below the surface of the ground on the line of the cut-off to be described. Above the fourteen-foot vein and about, say midway, between it and the surface of the ground, there is another small vein or rider of coal, about twelve inches in thickness. The rock between the two veins is a carbonaceous slate, about twenty-five feet in thickness. Above the smaller vein is a bed of sandstone, varying in thickness, but generally not less than from eight to ten feet. This sandstone is also carbonaceous, but the proportion of carbon is much less than in the slate below. Both rocks have been analyzed, and before the next Institute meeting, the writer hopes to be able to furnish the exact proportions. The sandstone is overlaid with debris up to the surface of the ground. Some time during the spring of 1876, it was found that the coal left in the worked-out mine had been set on fire, and that the fire was gradually spreading. Inquiry developed the fact that the fire had originally been started by a woman, who had taken up her abode in an old tunnel leading into the workings. Owing to some delay in the early attempts made to extinguish the fire, it finally gained such headway that no ordinary means were found to be effectual. Finally, Mr. C. F. Conrad, civil and mining engineer, of Pittston, Penna., was called in as consulting engineer. He advised the making of a thorough cut through such portion of the old workings as had not been reached by the fire. It was found by survey that this cut could be limited to a length of about one thousand two hundred feet, as the vein faulted on both sides, and that the greatest depth of the cut would be about sixty feet. Subsequently it was decided to tunnel part of the way, thus saving the deepest portion of the cut. In 1856-7, there occurred a fire in the same mine. Two stone-walls in V-shape running from the apex of the V towards the outcrop, were built to cut the fire off, and these walls were carried as far as the vein had at that time been worked. After the walls were finished

Jan 1, 1879

By Marsh S. Watson, Robert A. Howard

The effect of depth of formation penetration by casing perforations has been studied by means of an electrolytic model. It is found that the relative productivity of perforated casing is profoundly affected by the depth to which the formation is penetrated. The first in. of penetration i: of especial importance. A scaling equation which was shown to agree with experimental results at zero penetration is used to calculate the relative productivity of a waled down casing with scaled down perforations. INTRODUCTION In an earlier piper' the authors showed that the point sink representation of casing perforations is inade- quate. It was Pointed out at that time that while the electrolytic model results were a marked improvement over point-sink calculations, they still left a great deal to be desired in that the electrolytic model represented a case not realizable in practice, namely, a clean hole through the casing but undisturbed formation on the other side of the casing. Obviously, to be effective, the perforation must penetrate a finite distance into the producing formation. It was the purpose of the present investigation to find how the relative productivity of perforated casing varies with depth of penetration. The term "depth of penetration" as used in this paper refers to the distance which the perforation penetrates the producing formation. The amount of casing and cement penetrated is not taken into account, a fact which is discussed more fully later. The "casing radius," therefore, is to be interpreted as the radius of the boundary between cement and formation wall. in cases where the casing is not in contact with the formation wall. APPARATUS The electrolytic model experimental setup was identical with that described in the previous paper' except that the platinum electrodes were now allowed to project from the lucite cylinder by varying amounts. When fewer than eight electrodes were used, the unused ones were covered with a coat of insulating material to minimize disturbances to the current flow-lines. Comparison runs made with the unused electrodes uninsulated, however, showed that the values of Q/Q0 so obtained did not vary by more than the experimental error from the values obtained with the unused electrodes insulated RESULTS The variation of relative productivity* with perforation density is shown for various penetration depths from 0 to 4.03 in. in Fig. 1. It will be seen that even one in. of penetration into the formation results in a huge increase in the relative productivity. Moreover, as-the depth of penetration increases, it becomes less and less advantageous to use large perforation densities, 80 pel. cent of open hole productivity being obtained with only three 0.5-in. diameter holes per foot when the penetration is 4.03 in.

Jan 1, 1950

SINCE I have mentioned the art of alchemy in. many parts of this treatise of mine, especially when 'came to the description of the practice of various operations,* I do not here intend to argue as to whether it may be called true or not. But in case it is true, inasmuch as it is seen in operation to be a result dependent on the actions and virtues of fires, I should not pass by in silence without speaking at all of this art here, nor would I wish to do so, especially since I think that you do not really know what it is except in so far as common report is heard everywhere to the effect that whoever carries it to completion makes silver and gold with it. Because of this belief, many seek it with much time, labor, and expense, as I know you have often heard related, if not elsewhere, by me in the chapter on gold that I have written in this volume. Now do not think that I wish to teach you this art here, for I am one of those who do not know it. Therefore I shall tell you here, so that you may know more, only. that those workers who so eagerly follow after and seek it proceed by just two pathways. One is that which takes its enlightenment from the words of the wise philosophers, by means of which they think to attain it. This they call the just, holy, and good way, and they, say that in this they are but imitators and assistants of Nature, indeed, manipulators and true physicians-of .mineral bodies, purging them of superfluities and assisting them by augmenting their virtue and freeing them from their defects. In this way they sometimes proceed to corrupt these bodies in order to be able to separate the elements they contain so as to reduce them, if possible, to their primary materials, as they call them, or sometimes they convert them into new substances by means of this art or by adding another spirit different from the first one. Thus they seek in this way to bring these materials to a certain point of corruption or to a separation of elements, or to remove or add spirits to things, or to make coarse materials subtle and sometimes to make subtle materials coarse. Therefore, as you can understand, these persons, with bridle broken, run a circular track night and day, without ever having rest; and surely I do not know whether anyone has ever arrived at the desired goal. Nevertheless, even though the goal of this art were emptiness, as many

Jan 1, 1942

Theoretically mobile primary crushers at the mine working face can be considered the ideal comminution system. A crusher which follows the ore can reduce materials handling costs because it eliminates about one ton of truck deadweight per ton of ore hauled to the crusher. It also eliminates the backhaul of this same deadweight ton. Development of mobile primary crushers has centered in Germany where many soft iron ore and limestone deposits are flat-lying and thus readily adaptable to the use of shiftable conveyors which easily followed the working face. These ores were easily crushed in impact- type mills which had extremely large receiving openings and therefore large reduction ratios. The feeding of such a unit was done with a front-end loader which dumped its load to an inclined apron feeder, which in turn fed the impactor. Under the crusher was an inclined conveyor which transferred the crushed ore to a belt wagon for delivery to the shiftable overland conveyor. Low headroom overhead eccentric single-toggle jaw crushers and short-shaft gyratories were also used on these units, which generally travelled on tracks. While these early mobile primary crushers were versatile and served their purpose, they generally suffered from low capacities. Manufacturers have quickly responded with larger crusher, feeders, and moveable supports, however; and today mobile crusher s have capacities of up to 1500 metric tons per hour, a figure comparable to the capacity of some 54 in. by 74 in. gyratories. An example of a quarry portable crushing plant is at the Bath cement plant of Canada Cement Lafarge, Limited. Two 12 cu yd front-end loaders are digging the stone and hauling it to the portable crushing plant which feeds a field conveyor system. The crushing plant is moved and the field conveyors extended whenever the length of the haul exceeds 300 ft. At the crushing plant, the loaders discharge the stone onto a pan feeder which feeds into the hopper of a Weserhutte-Hazemag compound crusher which combines primary and secondary rotors in the same housing. These are driven at 1200 rpm by 800 hp and 1,000 hp motors (600 and 750 kw). The maximum 3 ft feed (914 mm) is reduced to 3/4 in. (19 mm) at 800 tph. The Weserhutte steered walking mechanism, operated by hydraulic cylinders, takes five -foot strides. Another example is at Alcoa of Australia's operation at Huntley in Western Australia. At the mine face, front-end loaders load 50 mt rear-dump trucks that transport the bauxite ore to the mobile crusher which follows the mine operation. The crusher can turn 360° within its own length, walk at a speed of 34. 5 ft per minute (10.5 m per

Jan 1, 1978

By Henry S. Drinker

The Butler Mine property is situated in the vicinity of Pittston, in the Wyoming coal-field of Pennsylvania. The coal has been ' worked out from the fourteen-foot or Baltimore vein for a number of years on part of the tract, the old chambers remaining open as when originally abandoned. This vein outcrops on the Butler prop erty, and is nowhere more than sixty feet below the surface of the ground on the line of the cut-off to be described. Above the four teen-foot vein and about, say midway, between it and the surface of the ground, there is another small vein or rider of coal, about twelve inches in thickness. The rock between the two veins is a carbon aceous slate, about twenty-five feet in thickness. Above the smaller vein is a bed of sandstone, varying in thickness, but generally not less than from eight to ten feet. This sandstone is also carbonaceous, but the proportion of carbon is much less than in the slate below. Both rocks have been analyzed, and before the next Institute meet ing, the writer hopes to be able to furnish the exact proportions. The sandstone is overlaid with debris up to the surface of the ground. Some time during the spring of 1876, it was found that the coal left in the worked-out mine had been set on fire, and that the fire was gradually spreading. Inquiry developed the fact that the fire had originally been started by a woman, who had taken up her abode in an old tunnel leading into the workings. Owing to some delay in the early attempts made to extinguish the fire, it finally gained such headway that no ordinary means were found to be effectual. Finally, Mr. C. F. Conrad, civil and mining engineer, of Pittston, Penna., was called in as consulting engineer. He advised the making of a thorough cut through such portion of the old workings as had not been reached by the fire. It was found by survey that this cut could be limited to a length of about one thousand two hundred feet, as the vein faulted on both sides, and that the greatest depth of the cut would be about sixty feet. Subsequently it was decided to tunnel part of the way, thus saving the deepest portion of the cut. In 1856-7, there occurred a fire in the same mine. Two stone walls in V-shape running from the apex of the V towards the out crop, were built to cot the fire 06 and these walls were carried as far as the vein had at that time been worked. After the walls were finished

Jan 1, 1879

DeLamar, the largest silver mine in the United States outside the Coeur d’Alene District in Idaho, came into production in April 1977. It is a joint venture of Earth Resources Company and Canadian Superior Oil Co., with Earth Resources being the operator. The property is a consolidation of several old operations dating back to the 1800s; the better known of these are the DeLamar and Sommer-camp mines. DeLamar is situated in southwestern Idaho near the Oregon border in the Owyhee Mountains, 88 km (55 miles) southwest of Boise, Idaho. Heavy snowfalls and very low temperatures are common in the wintertime. Planned production was 77,500 kg (2.5, million oz) of silver and 5 13 kg (16,500 oz) of gold per year. Expected life at a milling rate of 1,540 mtpd (1,700 stpd) was 20 years, with an ore grade of 144 g Ag and 1.58 g Au per mt (4.2 oz Ag and 0.046 oz Au per st). Present production is at the rate of 1,900 mtpd (2,100 .stpd). The total cost to bring the mine into production was reported to be approximately $22 million, and the mill was reported to have cost about $10,000 per daily ton of design capacity. At DeLamar, naurnanite (Ag2Se) is the principal silver .mineral occurring with acanthite (Ag2S) disseminated through quartz veins and a silicified rhyolitic ash-flow tuff of Miocene age. The mineralization is relatively iron and manganese free and is readily amenable to cyanidation giving recoveries of 85% silver and 92% gold, which occurs free. When the plant was started, considerable difficulties were encountered in handling the wet sticky ore because of the high clay content. Following primary crushing, the ore is stockpiled. Due to the sticky nature of the ore, the vibrating feeders initially installed under the stockpile were replaced by a reclaim hopper and an apron feeder which was fed 24 hours a day by a front-end loader. The minus 200 mm (8 in.) ore is fed to a 5.5 m by 2.75 m (18 ft by 9 ft) autogenous mill followed by a 2.75 m by 4.57 m (9 ft by 15 ft) ball mill and cyclones. The mills are equipped with 750 kw and 520 kw (1,000 hp and 700 hp) motors. Both mills discharge to the same cyclone feed sump and a portion of the cyclone underflow is returned to the autogenous mill. Cyclone overflow, which is 20% plus 74 microns (200 mesh), is leached in four agitators in series for 72 hours (45 hours at the 2,100

Jan 1, 1981

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 Joseph M. Jacisin

Fatigue tests were conducted on 2017-T4 alumium in alloy in an u1lrcthig.h vacuum of 2 x 10-lo Torr and in air. The vatio of vacuum-to-air faligue life for this ~malerial varied Jrom 3.5:1 at a strain of 2.5 X 1O-3 ill. pegr in. to 8:l at a strum of 3.3 x X ill. per 171. Sotlze pvelivtlinary data were also obtained on the faligue life of the alumirnum alloy in a dvy-nitrogen atmosphere. A dark deposit observed on the fracture surfaces of the aiv-tested specimens was identified as oxides of tr~angarzese and magnesium This observalion, and the close correlalion betaveen results obtained in vaCuum and dry-nitrogen environtments, tend to confirm that air stvongly influences the fatigue mechanism by a covrosion process. THE fatigue life of certain metals is known to increase when the atmospheric pressure is reduced. Gough and Sopwith1-2 postulated that this is due to a joint chemical and mechanical process operating in air at the surface of the metal with oxygen and water vapor playing significant roles. Wadsworth and Hutchings3 observed that aluminum is mainly affected by the presence of water vapor. Very little information is available on the fatigue properties of structural aluminum alloys in a vacuum environment. This report describes work performed on 2017-T4 aluminum alloy in an ultrahigh vacuum (2 x 10-loTorr) and in air. Some preliminary data are also presented on the fatigue life of the alloy in a dry-nitrogen environment. 1) EXPERIMENT 1.1) Test Material. The material tested was commerciaT 2017-T4 aluminum alloy which, in addition to aluminum, has an approximate composition of 4 pct Cu, 1 pct Fe, 0.8 pct Si, 0.5 pct Mg, 0.5 pct Mn, plus smaller amounts of other elements.* Some typical naturally age to a substantially stable condition. Specimens were then machined from these rods and ground to a No. 8 microfinish. In order to minimize variation in material characteristics, specimens were made from rods procured from the same lot. 1.2) Apparatus. The apparatus and test specimen used in the fatigue tests are shown in Fig. 1. The device utilized a single-cantilever beam with rotating-bending loading. Bending stress was applied to the fatigue test specimen by placing a spacer beneath the bearing support, thereby providing a constant deflec-

Jan 1, 1968

By V. A. Zanadvoroff

The mine and concentrator of the Howe Sound Co., Chelan Division, are at Holden, in a remote section of the Cascade Mountains of northwestern Washington. Holden is at an elevation of 3200 ft., in the narrow valley of Railroad Creek. This creek drains into Lake Chelan, at a point some 40 miles up the lake from the town of Chelan. The concentrator treats 2000 to 2500 tons of low-grade copper-gold-zinc ore per day. Concentration is achieved by the differential flotation process, with recovery of 94 per cent copper and 82 per cent gold in the copper concentrate, followed by the recovery of 65 to 70 per cent zinc in the zinc flotation circuit. The final flotation tailing is deslimed, and coarse . sands, representing 60 to 70 percent of the total, are cyanided. The tonnage of tailing material amounts to 1900 to 2370 tons per day. General Considerations The property is within the Chelan National Forest. Railroad Creek is a fishing stream, being stocked regularly with trout from State hatcheries. Consequently, in planning a milling operation, the Howe Sound Co. was obliged to avoid any pollution of the stream by tailing material, and also destruction of valuable stands of timber. Moreover, the selected tailing- disposal sites should not mar the natural beauty of the valley. In addition to these requirements, the problem of tailing disposal at Holden was complicated by topography and climatic conditions. The valley is long and narrow, averaging 1200 to 1500 ft. in width. Its walls are flaring at a steep angle. There are no extensive flat areas suitable for deposition of tailing. All valleys tributary to Railroad Creek, also the well-marked paths of snowslides, must be avoided. Finally, the direction of prevailing winds and possible dust hazard arising therefrom dictated the necessity of selecting the tailing-disposal sites downstream from the mill building and two campsites. Heavy snowfall during the winter months of November to March, inclusive, impedes building of the retaining dike, making it necessary to prepare the needed storage space during the summer months. It is inadvisable, also, to do any dike building during the freezing weather, because the dike becomes unstable with the spring thaws, producing inevitable "sloughing" of the built-up borders. Tailing-disposal Sues The existing tailing-disposal areas are near the south wall of the valley. They were developed, and put to use in the following sequence: No. I. Area.—Covers approximately 10 acres. Built up to the height of 60 ft. and abandoned. Total length of dike border is 1900 feet.

Jan 1, 1947

By P. W. Bakarian, John C. McDonald

Three magnesium alloys were processed in various ways to exhibit a wide variation in the ratios of yield strength and tensile strength in the rolling direction compared to the cross-rolling direction. Basal-plane pole figures were prepared from several specimens. A fair correlation was observed between the property ratios and the ratio of percent of poles oriented in the cross-direction quadrants compared to the percent oriented in the rolling-direction quadrants of the pole figures. THE modes by which magnesium crystals deform are such as to lead to marked preferred orientations in all forms of wrought products. The effect of the active modes at room temperature, coupled with these orientations, is often to produce a lower compressive yield stress compared to the tensile yield stress. The phenomenon is due to easy twinning in compression, and has been well-related to the preferred orientations as described by basal-plane pole figures of the hcp magnesium single crystal. The literature contains some information on the variation with test direction of the tensile properties of sheet, but no quantitative relationship to the observed pole figures has been established. These effects have been reviewed in the few com- prehensive treatises in English which have been written on magnesium technology and metallurgy.1°3 Roberts presents several pole figures, but no tensile-property data in the plane of the sheet. Raynor has no pole figures, but quotes Frommer's4 results. The information in Beck's compendium more or less parallels the results of Jones and powell.5 In both a Mg + 1.5 Mn alloy and a Mg + 4.6 A1 alloy, cold rolling was found not to change the general situation of higher properties in the cross-rolling direction,* than in the rolling direction. Thomas'6 X-ray studies of Jones and Powell's specimens did not permit correlation of properties and preferred orientation. Ansel and Betterton7 report tensile and compressive properties on three other alloys as a function of degree of cold rolling, plus a recrystallizing anneal. As rolled, there was a change in anisotropy with degree of reduction, but all alloys were reasonably isotropic in tension when annealed. X-ray diffraction photograms illustrated the kinds of orientations present, but were not correlated with the properties. Bakarian8 has described a method of determining quantitatively a preferred orientation, and illustrated the method by applying it to pure magnesium and two of its alloys. This method was used in the present work as a basis for correlating the strength properties with the texture.

Jan 1, 1965

By Sylvan Z. Beer

Aluminum phosphide was synthesized and grown in solution of excess aluminum using a modified two-temperature technique. An inversion in the density difference between ALP and molten aluminum takes place at about 1100°. Dispersed nuclei can act as sources for further crystal growth in a manner resembling ripening. Advantage was taken of this phenomenon by introducing a stirring motion into the molten liquid. This was done with a split-phase 10-kH2 rotating magnetic field produced by four horizontal coils placed symmetrically about the furnace at the level of the molten metal. The method also prevents the formation of a surface film which would impede the reaction. By this technique vastly improved yields were obtained and flat crystals 5 by 5 by 0.5 mm were grown at lower temperatures and at greater rates than heretofore. THE family of III-V intermetallic compounds has received much attention since the prediction of their semiconducting properties in the early 1950s. Of the family, A1P has perhaps the highest band gap of those compounds normally possessing the zinc-blende structure. But it has been least studied because it is difficult to handle and grow. The handling difficulties have been dealt with elsewhere and it is sufficient to say here that, by various techniques, many of the difficulties can be overcome.' The difficulties encountered in crystal growth can be related to the refractory nature of A1P and in turn to its high band gap. Because of the anticipated high melting point of the material and the accompanying high phosphorus vapor pressure, the reported work on synthesis and growth was done by epitaxial synthesis-growth via chlorine2 and iodine3 transport and from solution of excess aluminum by the two-temperature technique.4'5 The transport-grown crystals take on the dimensions of the substrate but do not grow easily in thickness. In the usual two-temperature technique as applied on the synthesis of Alp, the temperature of the molten aluminum is kept at 1000° to 1250°C while the phosphorus is maintained at 1 to 20 atm. A thin film of aluminum phosphide and oxide form immediately, impeding the reaction and affecting the yield. Attempts to grow the Alp from such solutions by the methods used to grow Gap and InP from excess gallium and indium have not been successful. As discussed in the preceding paper,5 the yield can be greatly increased

Jan 1, 1969

By Richard H. Willey

EFFICIENT handling of large tonnages in open-pit mining demands, primarily, an abundance of well fragmented rock. To accomplish this, a drilling and blasting department composed of a well-knit, versatile group of workers and supervisors must be established. Some costs will remain relatively on the same level over a period of years, while others can be greatly lowered as will be shown. The Utah Copper Division, Kennecott Copper Corp., Bingham Canyon, Utah, has been ever alert for improvements in operating methods and equipment to gain increased efficiency and lower costs. BINGHAM CANYON MINE The greatest dimension of the mine is 8000 ft northeast-southwest with a width of 5000 ft southeast-northwest. From the lowest level, at an elevation of 6040 ft above sea-level, the mountain rises in 27 benches on the west side to W level (elevation 7800 ft) and 21 benches on the east side to Q level (elevation 7460 ft). At the inception of open-pit mining in 1906, the overcapping of leached porphyry had to be removed to reach a commercial grade of sulphide ore. The overcapping or waste now remaining is quartzite with intercalated limestone beds. The limestones are usually soft in nature except in the south-middle portion of the mine where they have been metamorphized along the contact zone. This area usually gives drilling and blasting operations a great deal of trouble. Here occurs the greatest amount of secondary blasting, as the well-established fracture planes cause the altered beds to break into large boulders. Several massive quartzite "ribs" occur in the overburden, and as a result, the uniform breakage is difficult to obtain. The major quartzite areas have been subjected to complex folding, fracturing and faulting because of the intrusion of the underlying stock. The monzonite porphyry, which carries the commercial values, presents no serious drilling problem because of its softness and extensive minute fracturing throughout the mass. A series of major slips or faults pass through the ore body which, although of considerable advantage to the drilling and blasting department, sometimes works to the disadvantage of the track department. Drilling and blasting of the porphyry apparently becomes easier as depth is attained, but highly siliceous streaks throughout the pit still present difficulties. At no place is the rock of such character as to allow excavation without preparations. On certain higher elevations or levels, old fill material can be dug by the shovels, but only at some risk to equipment because of bank slippage. PAST METHODS At the outset of shovel operations in 1906, drilling was done with the Ingersoll-Rand F-24 piston drill mounted on an un-

Jan 1, 1947

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 John P. Bryant, Samuel Kincheloe

Plant nutrients are obtained by plants from both the air and the soil. Carbon dioxide, a gaseous form of carbon and oxygen, supplies the carbon which usually makes up 50% or more of plant structure. Plants are capable of absorbing carbon dioxide directly from the air. Most of this nutrient comes from the air, although some carbon dioxide is absorbed by soil water and taken into the plant as liquid. Growing plants use more water than any other substance. Plant nutrients in the soil dissolve in soil water; soil water entering the plant carries nutrients into the plant structure. Water is then given off into the air as the plant breathes or transpires. Inorganic or mineral nutrients essential to plants are supplied through the soil. Inorganic or mineral composition of plants usually is less than 5% of their weight. In most crop-growing areas, much of this nutrition is inadequately supplied from the soil and must be provided by fertilizers. Proper fertilization of a soil to improve its mineral level and productivity is the role of the fertilizer industry. The 13 essential plant nutrients which often must be added to soils or plants and are generally considered fertilizer minerals are categorized as primary minerals (nitrogen, phosphorus, and potassium); secondary minerals (calcium, magnesium, and sulfur); and micronutrients, or trace minerals (boron, iron, manganese, copper, zinc, molybdenum, and chlorine). These classifications are based on quantity of nutrients utilized by plants and have no implication as to function or degree of necessity. One point of major significance to the fertilizer industry is that not all nutrient carriers supply that nutrient in a chemical form which will be available to the plant. Thus, a product which may be considered a fertilizer mineral may require substantial processing or may require substantial processing or treatment before it becomes useful as a nutrient carrier. Fertilizer Industry Scope Until about 1900 demands for higher crop yields were met simply by bringing new agricultural lands into cultivation. Since that time great progress has been made in increasing yields through fertilization and improved varieties and cultural practices. The importance of the fertilizer industry to the economy and welfare of the US is widely recognized. Fig. 1 illustrates the relationship of lime and fertilizers to crop yields. In terms of tons of fertilizer products, the growth of the industry has been phenomenal, particularly between 1950 and 1980. Table 1 shows the increase in fertilizer usage. Primary Plant Nutrients in Fertilizers Primary minerals (nitrogen, phosphorus, and potassium) are designated as such because [ ]

Jan 1, 1983

By Leland L. Mink, Roy E. Williams, C. Daniel Kealy

The passage by Congress of Public Law 92-500 (the 1972 Water Pollution Control Act amendment) has placed a new emphasis on the need for methods to minimize the escape of water from tailings ponds. Presumably tailings pond seepage will fall under the jurisdiction of the act. This paper is designed to explain some of the problems caused by leakage from tailings ponds and to present a method by which leakage can be minimized if not eliminated. Ground-water pollution in one case studied by the authors results from leaching of old jig tailings that are intermixed with the upper part of a sand and gravel aquifer. Anomalous concentrations of zinc, lead, and cadmium occur in ground water and soil samples taken from the portion of the sand and gravel aquifer containing the old jig tailings. Analysis of water samples from a tailings pond located in the area indicates that the pond water is not the source of the heavy metal concentrations found in the down-gradient ground water. However, leakage from the pond and water from the pond's decant system combine with natural ground water to compound the problem by raising the water table into old jig tailings which cover the upper portion of the aquifer. The ground water which discharges from this aquifer causes a statistically significant decrease in the quality of the receiving stream water. The shallow ground water contains oxygen; therefore, corrective measures include lowering the water table to depths below the old jig tailings which cover the aquifer at ground level. This, in turn, implies elimination of ground-water recharge by subsurface leakage from the pond. A laboratory study and a mathematical model study have revealed that seepage from such ponds can be minimized by a carefully designed and managed peripheral tailings discharge system. The process involves maintaining the slime zone and the free-water surface near the center of the pond so that the slimes retard downward and semilateral movement of pond water. The free-water surface must not be allowed to extend out over the nonslime portion of the tailings during or after pond operation. Mathematical model output reveals that this method can hold the subsurface discharge rate to as little as 110 gpd per acre of pond water area. The technique should be applicable to any tailings pond system in which the loss via seepage of poor quality water must be minimized or where recharge of ground water by losses from a tailings pond must be minimized, regardless of the reason.

Jan 1, 1974

By John R. Anderson, Michael B. Bever

CARBON may affect the alloys of copper in several ways. Provided an alloying element does not oxidize preferentially, even minute quantities of carbon dissolved in liquid alloys of high copper content will form insoluble carbon oxide gases under oxidizing conditions. Gas porosity in castings of high-copper alloys can originate in this way. In certain alloys of low copper content carbon may have considerable solubility and may be an effective deoxidizer. In these alloys carbon will tend to suppress porosity. Carbon that is dissolved in the molten alloy is likely to be retained in the solidified metal. The effects of carbon on the physical properties of the metal are determined by the form in which it is present. Depending on the alloying elements and also on the cooling rate, carbon may precipitate as graphite, may form a carbide or may remain in solid solution on cooling. As solid copper does not dissolve a measurable amount of carbon, a carbon-bearing solid solution can occur only if the copper is alloyed with a large excess of an element that forms such a solution. The effect of graphite or of carbides on the properties of an alloy will depend on the size and distribution of the precipitated particles. As Hensel pointed out in discussion,1 the formation of an alloy carbide may interfere with the age-hardening properties of a copper alloy. One would also expect that the presence of carbides or of graphite will under certain conditions promote corrosion or high-temperature failure in nonferrous alloys. Since carbon has an effect on the soundness of castings and affects the structure and behavior of solid copper alloys, its solubility in the liquid alloys is of practical importance. The isothermal change of carbon solubility with alloy composition is of interest from a theoretical standpoint. The binary alloys of copper with manganese and with nickel are used industrially and serve as a base for more complex alloys. The solubility of carbon in each of these metals is known, but its solubility in their alloys has not been determined. EARLIER DATA The solubility of carbon in molten pure copper is about 0.0001 pct at 1100 °C and 0.0004 pct at 1475°C.1 Various authors2-6 have published data on the solubility of carbon in molten manganese. There is agreement on a value of about 6.7 pct in the temperature range 1217° to 1245°C. This solubility value corresponds to the formula Mn3C. It increases very little with increasing temperature. Corson7 states that carbon is soluble in copper-manganese alloys and that the

Jan 1, 1947