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By Frederic Benard

THIS paper constitutes a preliminary report on experimental work done to date on the anode-furnace treatment of blister copper containing relatively high percentages of nickel. The investigation has not been completed but results obtained to date are considered of sufficient interest to warrant publication. Blister copper treated by the Ontario Refining Co. originates from the Copper Cliff smelter of The International Nickel Company of Canada, of which the former is a wholly-owned subsidiary. The blister as received contains from 0.50 to 1.50 per cent Ni, and in this respect differs from that treated by other refineries on this continent. During the anode-furnace treatment, only a portion of the nickel content is removed, and of the amount remaining in the anodes a varying percentage is found in the anode slimes. This necessitates a more extensive treatment than ordi-nary, nickel-free material. Sometimes it is necessary to repeat roasting and leaching operations several times before the nickel content is low enough to permit the slimes to proceed to the doré furnace. This, besides increasing treatment charges, causes greater selenium stack losses and involves a longer tie-up of precious-metal values. Therefore this investi-gation was undertaken with the aim of determining the factors influencing the diversion of nickel into the anode sludge. From experience it was known that by keeping the nickel content of anodes at 0.55 to 0.60 per cent maximum a relatively satisfactory slime resulted. One avenue of attack therefore was developing improved elimination in the anode-furnace treatment so that the anode content mentioned above was not exceeded. Another angle was to improve the solubility, during electrolysis, of the nickel constituent in the anodes, thus decreasing the percentage diverted to slimes. It may he mentioned here that nickel dissolving electrochemically and entering the electolyte pre-sents no difficulty from a treatment standpoint. Both of these approaches were studied and are discussed sepa-rately below.

Jan 1, 1938

By H. W. Morse

MANY of our largest copper deposits contain both sulfide and oxidized copper minerals. The large porphyry mines, with a total copper content of from 1.3 to 2.0 per cent., send to their mills ore with an oxide content of from 0.2 to 0.4 per cent. While the recoveries of the sulfide minerals are considered satisfactory, using modern flotation practice, the recovery of the oxi- dized portion of the copper is small; better recoveries of oxidized copper were made before flotation methods were adopted. Beside the ores containing copper mainly as sulfide, with a small amount of oxidized copper,, many of the large porphyry mines have oxide orebodies, in which the propor-tion of non-sulfide minerals is so high that 2 present methods of treatment are ineffi-cient. In some of the well-known ore-bodies in the Miami-Inspiration district the copper occurs about equally in sulfide and oxidized forms; and the mixing of minerals is complete throughout the ore-body. There is no possibility of mining the sulfides and the oxides selectively. At the New Cornelia mine, however, the oxide, and sulfide ores are sharply separated, per-haps by an ancient water level, so that the oxide ore may be mined and treated in a plant built for that purpose, and the under-lying sulfide ore may be left for treatment in another plant. If copper is to be made from an ore having a total content of 1.5 per cent., it must be concentrated or removed from the ore by some. hydrometallurgical method. Flotation is generally successful for the sulfide minerals and flotation after filming with sulfide seems to work in many cases on the pure carbonate minerals. Callow has given an extensive report on the results of "sulfidizing" flotation on the ores of the. Magma mine.1 Where such ?a 'method can be applied, it ?may be, the best

Jan 9, 1922

By K. L. Meissner

IT is well known that the so-called pure aluminum contains noticeable amounts of impurities, chiefly iron and silicon, and many investigators have studied the forms in which these impurities exist. Hanson and Archbutt1 state that iron exists in aluminum as FeAl3 and silicon as elementary Si. The solubility of FeAl3 in solid aluminum is practically zero, or at, least so small that the smallest trace of this compound is always visible as a separate constituent. Under the microscope an untrained eye can hardly distinguish, without etching, the two constituents FeAl3 and Si, but they can be differentiated by suitable etching reagents. Anderson2 also found the same results. At the same time, Merica, Waltenberg, and Freeman3 found a constituent of unknown composition. They found a small thermal effect, at 610° C., on the cooling curves of aluminum with small amounts of iron and silicon. The intensity of this effect increased with increasing silicon content, until an effect occured at 576° C. due to the eutectic between aluminum and silicon. The authors ascribed the effect at 610° C. to a compound, which they call "X constituent." It was thought to be a ternary compound containing iron and silicon and could be observed under the microscope as dark particles. Dix4 has published unetched photomicrographs of an aluminum with 0.51 per cent. iron, 0.99 per cent. silicon, and 0.32 per cent. copper that shows three constituents: the dark Si (black in the figure), the lighter FeAl3, which appears in half tone and crystallizes in needles, and a constituent of the same color as FeAl3 but of different crystal form, often resembling Chinese script. This last constituent, Dix thought, is a compound of Fe and Si; not the compound FeSi but a compound containing a relatively small amount of silicon. Therefore an alloy of aluminum, iron, and silicon (the copper should not be regarded in the following considerations) must contain the following four constituents: Al, FeAl3,

Jan 1, 1926

By AIME AIME

SOMETIMES family backgrounds have a deep, if not direct, influence on a man's lifework. Dean Callen's paternal grandfather came from Wales and was a mine official in the Pennsylvania anthracite field: his maternal grandfather was a retail coal dealer. His father and uncle began their careers a. breaker boys but later completed their educational work and ended up in the professions. With that background, and with the slate region ( Pen Argyl, Pa.. July 1;. 1888) as his birthplace it is not strange that "Cope" Callen became a teacher of mining engineering. Graduating from Lehigh with an E.M. in 1909, lie taught there for two years while working for his master's degree. In 1911 he returned to teaching, going to the University of Illinois as instructor in mining engineering. In 1917, he was called to West Virginia University as head of the department of mining engineering and director of mining extension. The extension work was about dead. bus a new staff was built and the six-weeks' short course in coal mining was revamped into an attractive and efficient course which has grown in popularity with the years. Callen's work attracted the attention of the Federal Board for Vocational Education and in 1919 he was invited to serve that board as special agent for mining, taking a six-months' leave front the University.

Jan 1, 1944

By Robert Ammon

THE milling practice at Mascot, at present, consists of dry crushing to 5/8 in., jigging, fine grinding, and flotation. The ore arrives at the mill from two mines, No. 1 mine shaft being located in the mill proper, and No. 2 mine shaft being situated about 2000 ft. east of the mill. No. 2 mine is served by a coarse-crushing plant and aerial tramway to the mill bins. In jigging the No. 2 mine ore, 50 per cent. of the mill feed is rejected by the jigs as a jig tailing ranging from 5/8 in. to 10 mesh; the balance of this ore, after removing a jig concentrate, is ground to flotation feed. In jigging No. 1 mine ore, approximately 80 per. cent. of the mill feed is rejected as a coarse jig tailing, and the balance is ground to flotation feed without the removal of a jig concentrate. The difference in the treatment is due to the fact that No. 2 mine ore carries a higher content of zinc than that from No. 1 mine. Of the total concentrates produced, 30 per cent. are made on jigs and 70 per cent. by flotation. At present, 1350 tons of No. 2 ore and 800 tons of No. 1 ore are treated each 24 hours. The recoveries for a typical week are as follows: Heads 3.55 Jig concentrates 61.30 Jig tails 0.80 Flotation concentrates 60.60 Flotation tails 0.09 Extraction 86.50 The associated minerals are a straw-colored sphalerite, dolomite, calcite, pyrite, quartz, and a small amount of greenockite, which occurs as a stain and sometimes small crystals along with sphalerite, but is comparatively rare. Pyrite is present in small quantities, usually less than 0.5 per cent.; quartz occurs in bands a few inches thick. A complete analysis of the ore is as follows:

Jan 9, 1924

By David Kellogg

THE electrodeposition of iron has been practiced for many years. The earlier work along this line was directed toward the preparation of pure metal, but later applications were the production of thin adherent coatings as used in the electrotyping of Russian banknotes in 1860. A historical discussion of this subject, however, would require too much space and would be inconclusive. Much has been written on the subject but no two experimenters use the same methods or materials. During the war, the British army repair, shops developed a method for building up worn parts of automotive machinery, aero engines, etc., using the cold surfate . bath and low-current density method.1 The method has been used successfully in commercial work for the production of about 6000 repaired parts, therefore experiments along the same lines have since been made at the Westinghouse research laboratory. The apparatus used was essentially that described in the article mentioned except that the current -used for cleaning was obtained from a 3/4-kw. 60-volt, house-lighting generator, direct connected to a 2-h.p., 870-r.p.m., 440-volt, three-phase, type CS induction motor; while the plating current was taken from the storage batteries at any desired voltage. The anodes were made from 0.036 in. (0.9 mm.) Armco iron made into cylinders 5 in. (127 mm.) long by 8 in. (203 mm.) in diameter with a disk of 1/8 -in. (3.17 mm.) micarta, having a 2 in. hole in the center fitted at each end of the cylinder for stirring. The anodes were hung on a wooden rocker frame driven by a wooden connecting I rod directly connected to a small reducing gear, such as is sold by most of the apparatus supply firms; 3 gal. stoneware crocks were used as containers. The solution used by Thomas, 75 gm. of the crystallized ferrous ammonium sulfate per liter, was tried, using the current density recommended by him; namely, 0.33 amp. per sq. dm. Under these conditions the deposit on a ½ -in.

Jan 2, 1922

By Lucius W. Mayer

The iron-mines of the Ilsede Hutte Co. are at a town called Peine, about 20 miles east of the city of Hanover, on the railroad to Brunswick (Braunschweig). Hanover, the capital of the province, is a military city, of approximately 250,000 inhabitants, and Brunswick, 33 miles to the westward, has half this number. The mines here described are located in a hamlet called Bulten. The Ilsede Hutte Co. operates two mines, 1.5 miles apart, in this district, Gruber Bulten Adenstadt and Gruber Lengede Bodenstadt. It also operates an iron-mine north of Hanover, known as Eiserhaven, and one in the Harz mountains, the Georg Friedrich, at Grauhof. At Bulten a coarse conglomerate, a, Figs. 1 and 2, contains nodules of limonite, cemented together by calcium phosphate. The ore contains Fe, 25 to 45; Mn, 3 to 4; P, 0.6 to 1.4; '5, 0.02 to 0.09; and CaCO,, 40 per cent. Directly above the conglomerate is limestone, 6, followed by a mixture of clay and sand, c. Below the conglomerate is a hard clay floor, d. The ore-body, from 10 to 12 km. long, and from 3 to 4 km. wide, has a maximum thickness of 20 m. It dips approximately 10" S.E., and strikes K 45° E. Diamond-drill prospecting shows that the thickness diminishes to 1 m. as the end of the deposit is approached to the northward. At this point the deposit is 250 m. below the surface. In Fig. 2, the arrow, A, indicates the general direction of the work along the face, a, which is carried forward in a straight line for its entire length. At B, a train is approaching, which has come from the bench-track, shown at C, by a circuitous route, and is dumping the overburden behind the face, in the space left by the excavated rock. Formerly the work at this mine mas essentially open-cut, the deposit having a very strong outcrop. Figs. 1 and 2 show also

Jan 1, 1909

By W. A. Boyer

MINE hoisting ropes can be loaded to capacity only when the strength of each component is exactly known. Characteristic curves provide this information. When load and rate of acceleration are specified for an individual rope, the characteristic curve assumes a predetermined shape. The factor of safety, is the value drived by dividing the breaking strength of the rope by the total stress on the rope where it goes over the sheave wheel. Total stress is made up of the weight of skip and cage, weight of ore, weight of rope between sheave and rope clevise or attachment, friction load, and acceleration force. Bending stresses are neglected in this discussion, as sheaves and drum di- ameters are assumed to be large enough to reduce these stresses to a negligible value. Fig. 1 gives a family of characteristic curves for 13/8-in. improved plow steel hoisting ropes of 6 x 19 or 6x21 construction with hemp core. The curves will hold for all rope sizes of this specified quality when they are subjected to an acceleration rate of 1.608 ft per sec2. As acceleration rates of most deep shaft hoists are close to this value, this family of curves is presented as a standard. Deviation is slight for other rope sizes. Note that for 0, which is the most important, the deviation is within the width of the line on the graph. In a comparison of values for ropes of 5/8-in. and 21/4-in. diam the deviation ranges from a maximum 1.2 pct at 0 ft to 0.6 of 1 pct at 5000 ft. The value of 5.55 is the same for all ropes at 2500 ft. The tabulated columns to the left of the curves are values of total connected load for each size of rope from 5/8-in. to 21/4-in. diam inclusive. With each of these connected loads the safety factor characteristic curve for the rope will be as indicated.

Jan 1, 1957

By T. O. Massalski, J. D. Ayers

A massive transformation, 0 —am, takes place in a number of copper-based systems where the ß phase field at high temperatures extends to compositions that are slightly less solute-rich than the limits of the a phase field that exists at lower temperatures. Consequently a solid-state transformation becomes possible that does not involve a change of composition. Within this narrow range of composition the massive am phase can be usually produced in two ways: by a fast quench from the ß phase through the a phase field, or by pulse heating into the a phase field of an undecomposed ß phase retained at room temperature by a prior quench. In the Cu-Ga system the occurrence of a massive am phase has been also observed in alloys whose composition falls outside the equilibrium a phase field1 but not in the Cu-Zn system.2 We have recently observed the occurrence of the ß — am transformation also in two silver-based systems, Ag-Zn and Ag-Cd. The a and 0 phase fields in these systems again slightly "overlap" in a narrow composition range. In the Ag-Zn system this range falls between 36.7 and 40.2 at. pct Zn, and in the Ag-Cd system between approximately 41 and 42.5 at. pct Cd. Alloys were prepared from high-purity materials and cast in quartz tubes under inert atmosphere. No significant change of composition occurred during casting. After homogenizing in the 0 phase field, small pieces of each alloy were reheated at the homogenizing temperatures for a few minutes and quenched into various media. There were examined metallographic-ally after removal of surface layers that may have suffered from contamination or from the loss of the solute element. In the Ag-Zn system extensive areas of the am phase (identified by selective etching, observation of the slip trace variants, and occurrence of twinning) were observed in a comparatively low-zinc alloy (37.8 at. pct Zn) for all quenching media employed: iced

Jan 1, 1969

By R. S. Shoemaker, F. W. McQuiston

MAGMA COPPER COMPANY San Manuel Division The San Manuel Division of Magma Copper Company is located about 40 miles north of Tucson. Tons of ore milled in 1973 .were approximately 19 million at a rate of 63,000 tons per day. From this, about 42,600 tons of copper concentrate was produced as saying approximately 1.070 MoS2 and 28.5% Cu. As with many copper ores, the San Manuel ore contains a small amount of gold, but in this particular case, much of the gold follows the molybdenite. Separation of the molybdenite from the bulk sulfide concentrate by flotation produces a moly concentrate containing 9170 MoS2, just under 1.070 Cu, and 0.44 oz Au per ton. Cyanidation of this material for gold recovery is the only practice of its kind known to the authors of this Monograph. After thickening, the final moly concentrate is filtered for subsequent repulping with cyanide and lime. Gold dissolution takes place in three agitators in series with the pulp from the agitators being filtered in two stages in disc filters. Between filtration stages, the cake is repulped with barren solution. Filtrate from both filters is the pregnant solution. The Merrill-Crowe recovery unit is comprised of two solution tanks and the usual clarifying filter, vacuum tower, zinc dust feeder and Merrill precipitate press. Because of the presence of finely divided sooty chalcocite in the molybdenum concentrate, the cyanide consumption of 55 pounds per ton is extremely high. A compensating factor, however, is the low copper, moly concentrate produced. Precipitation of the gold from the highly copper-contaminated solution requires 0.56 pounds of zinc dust and 0.28 pounds of lead acetate per ton, high figures when compared to normal cyanidation practice. The Merrill- Crowe precipitate assays approximately 28.5% Cu, 8% Zn, 156 oz Au and 197 oz Ag per ton. The San Manuel cyanidation plant is an outstanding and innovative development.

Jan 1, 1975

By C. G. Dunn, M. Sharp

HE orientations of 137 crystals of copper pro-T duced by secondary recrystallization in a cube texture matrix have recently been obtained. Although the crystals were not prepared for the purpose of determining a secondary recrystallization texture, the orientations obtained proved to be rather interesting, as they confirmed, in part, published results and also disclosed additional preferred orientations in the texture. The investigation of Dahl and Pawlek,&apos; Bowles and Boas,&apos; and Kronberg and Wilson3 showed respectively that the orientations of secondary grains growing in a cube texture matrix were related to those of the primary grains by a rotation of approximately 30" about [001], the rolling direction, by 30" about a <111> direction, and by either 22" or 38" about a <111> direction. Kronberg and Wilson also obtained a component of the texture similar to that of Dahl and Pawlek except the angle of rotation about [001] was 19". When the cube texture matrix was deformed a small amount, Kronberg and Wilson also obtained large grains on subsequent annealing corresponding to a rotation of 15" about the normal to the sheet, but these grains should not be classified as secondary recrystallization grains since they were produced subsequent to mechanical straining. The particular texture obtained, of course, depends on the composition of the copper and the processing schedule employed, since these factors markedly affect the primary recrystallization texture and the sharpness of any cube texture produced. In the present investigation, pieces of OFHC copper in the form of M in. round bars were cold forged to a thickness of about 0.40 in., annealed 1 hr at 350°C in an atmosphere of hydrogen, and cold rolled crosswise to a thickness of 0.200 in. These pieces were given a penultimate anneal of 1 hr at 350°C in hydrogen and a 90 pct cold-rolling reduction lengthwise to a thickness of 0.020 in. Indentations generally were put in one end of the as-cold-rolled strips to promote earlier formation of secondary grains.

Jan 1, 1953

By Lloyd Pollish, Robert N. Breckenridge

SUCCESS in any underground mining operation is determined by accessibility of the orebody, which in turn is dependent upon maintenance of passageways to the mining zones and temporary support of the voids caused by extraction of ore. This is accomplished by one or a combination of the following methods: timbering, back-filling, pillaring, or, more recently, rock bolting. Timbering has usually been the principal means of maintaining these underground openings necessary for mining operations. Timber, however, does not prevent ground movement beyond the scope of localized sloughing, which is indicated by the gradual failing of the timber itself. Besides this, timbering has always been a costly process, and with the decline of available supplies of timber close to the mining areas, mining men have constantly sought other methods of controlling ground. Rock bolting is now replacing timbering at an ever increasing rate. Experience has proved that this form of ground support is just as applicable to blocky igneous rock as to stratified rock. Besides preventing sloughing of the walls and back of underground openings, Fig. 1, rock bolting has a stabilizing effect on the surrounding ground in much the same manner that steel reinforcing rods add to the strength of concrete structures. Further, rock bolting is flexible and may be applied to any shaped excavation, whereas timber sets are in a fixed pattern and the ground must often be changed to conform with this pattern. Rock-bolting installations were made in metal mines of the Northwest as early as 1939. An exhaust air crosscut was driven that year in one of the Butte mines of the Anaconda Copper Mining Co. The crosscut was rock-bolted and gunited at the time it was driven and is still being used to exhaust hot humid air from the 3400 level of the Belmont mine. It is interesting to note that no sloughing or caving has taken place in the 14 years it has been open. Even though these early installations of rock bolts were successful, few men recognized their potentiality until recent years, when the coal mines started their programs of mechanization and the great trend toward roof bolting began. In some areas of the Northwest stopes that previously required heavy timbering and close backfilling are now being mined by the more economical cut-and-fill and shrinkage methods. When used in conjunction with timbering, rock bolting increases the efficiency of the operation by decreasing hanging wall dilution and by making it possible to blast larger rounds. Most of the rock bolts installed to date in mines of the Northwest have been the 1-in. diam slot and wedge type, but there has been a recent trend to- ward using the 3/4-in. diam expansion shell bolt shown in Fig. 2. In addition to these commercially manufactured steel bolts, wooden bolts have been used with considerable success by the Day Mines of Wall&apos;ace, Idaho. Installation of the slot and wedge type requires three distinct operations, with tools for each operation: 1—drilling the hole to proper diameter and depth, 2—setting the bolt, and 3—tightening the nut. Holes are drilled and bolts set with pneumatic rock drills. A number of setting or driving tools have been used successfully, but most follow the same general pattern. Usually the driving tool is designed to accommodate a short length of drill steel on one end and the rock bolt on the other end. In this manner the hammering effect of the rock drill is transmitted through the steel and driving tool to the bolt. When machines not having stop rotation are used, slippage is allowed between the driving tool and bolt or between the drill steel and driving tool. The rock bolt nuts are tightened either with pneumatic impact wrenches or with hand wrenches. Impact wrenches are desirable because they are faster and assure adequate tightness. Expansion shell bolts have the following advantages over slot and wedge rock bolts: 1—No special equipment other than a wrench is needed for their installation. 2—Installation is faster. 3—They are removable. 4—Holes need not be drilled to a specific depth as the expansion shell will anchor anywhere along the length of the hole. These advantages are offset somewhat by the lesser strength of the bolt, since expansion shell bolts are generally made from 3/4-in. diam steel as compared to 1-in. diam steel for the slot and wedge type. One manufacturer, however, is now fabricating expansion shell rock bolts from steel of high tensile strength, which gives this ¾-in. bolt a much greater strength than that of the mild steel bolt. Table I illustrates tests made by the Anaconda Copper Mining Co. to determine the proper hole size to use with various types of bolts and to determine

Jan 1, 1955

By M. Safiuddin

Within the mining industry, open-pit mining has progressed to a point where a 200-yard walking dragline is as conceivable today as a 35-yard dragline was just a few years ago. This is possible due to the technological advances and efforts of the excavator manufacturers, and the manufacturers of electrical equipment have always stayed abreast with them in meeting their requirements and in making available their experience with the latest control techniques and hardware from various fields for application in the mining industry. An electrician in the coal mines of Pennsylvania who had not heard the word "thyristor" a few years ago now wonders how the old excavators with Rototrols® and magamps could ever work. The thyristors, transistors and operational amplifiers which were considered too fragile and delicate for a rugged and severe application like the mining industry, today constitute the brain of an excavator control taking commands from the operator and giving signals to the muscles of the excavator to move the bucket within a few seconds. A giant 200-yd bucket filled with overburden moves within seconds of the operator&apos;s command. Digging cycles for these excavators range from 40 to 60 sec depending on the swing angle. With these tremendous production capabilities of the excavators, the equipment failures and downtime become very expensive, and the loss of return on large capital investment for such machines becomes intolerable. Even though these failures and downtimes cannot be avoided completely, a good preventive maintenance program can help reduce them, thus increasing the machine availability. As we shall see below, the causes for breakdowns and failures of electrical equipment can be analyzed and a good preventive maintenance program can be drawn up to eliminate most of these failures.

Jan 9, 1967

By Carl Scholz

FROM the engineer&apos;s standpoint, labor organizations are of interest in so far as they &apos;affect efficiency, maximum production and unit cost, and in this respect the earlier labor organizations&apos; which existed in the age of the Guilds were ideal, inasmuch as they stood &apos; for competency of the apprentice and close cooperation &apos; between employers and employees. In later years the tendencies of the labor organizations have gone in the opposite direction and, instead of cooperating &apos;with the engineer, they usually oppose him on the mistaken theory, that labor-saving devices reduce forces, and costs; in this respect, the united Mine Workers of America have been particularly active and, to a very great extent, successful.

Jan 1, 1921

By J. S. Ll. Leach, J. B. Cohen, M. B. Bever

IN the work reported here, the heats of formation of a copper-rich Cu-Sn (a) solid solution and of the Cu.Sn (0 phase were measured by tin solution calorimetry. An approximate determination of the heat effect attending the precipitation of 6 from a supersaturated a solid solution was also made.

Jan 1, 1955

By AIME AIME

SCANDINAVIAN countries seem somehow to furnish an amount of talent and leadership far be- yond their population and among them Sweden is possibly at the top. The United States has many eminent citizens born in the latter, country or who are of Swedish descent and the engineers may claim their full share. In fact, the number is so large that a flourishing Swedish-American Engineering Society has been formed, including a number of A. I. M. E. members. Dr. Waldemar Lindgren is more scientist than engineer, though his work in economic geology has contributed directly to the development of many mines, and he is the valued- adviser of a number of our larger companies. Probably he, himself, values most highly the contributions he has been able to make to scientific theories of the occurrence and genesis of ore deposits. His fellow scientists have placed upon these the high mark of approval by electing him to the National Academy of Sciences, and Princeton University in 1916 recognized his position by conferring on him the degree of Doctor of Science.

Jan 1, 1931

By Reno H. Sales

Since leaving school my efforts have been geared to the task of making geology useful to the mining industry. The responsibility of the economic geologist or mining geologist of today has grown to be more, than that of merely finding a faulted segment of a vein. The very life of our mining industry is at stake. The tremendous progress in every branch of our modern civilization has been made possible through the developments in the use of metals. Many of our great industries, including that amazing field of aircraft production live and thrive only because of metals.

Jan 1, 1942

By Sales, Reno

Since leaving school my efforts have been geared to the task of making geology useful to the mining industry. The responsibility of the economic geologist or mining geologist of today has grown to be more, than that of merely finding a faulted segment of a vein. The very life of our mining industry is at stake. The tremendous progress in every branch of our modern civilization has been made possible through the developments in the use of metals. Many of our great industries, including that amazing field of aircraft production live and thrive only because of metals.

Jan 1, 1942

By Arthur F. Taggart

THIS paper proposes two units (believed to be new) for designating, respectively, capacity and efficiency for primary and intermediate crushers. CAPACITY Operators know that the tonnage of rock that can be put through a particular crusher in a given time depends primarily upon the crusher setting (open setting for jaw and primary gyratory types; closed setting for cones and rolls), and secondarily on the lithological character of the feed, its maximum size, the amount of fines that it carries, its moisture content, and the method of feeding. The proposed capacity formula states a relationship between all of these factors in one empirical equation, as follows: TR = TR80KK&apos;K" [I] in which TR, called "reduction tons per hour" is the new capacity unit; T = tons of feed per hour requiring crushing; R50 = "80-per cent reduction ratio"; and K, K&apos;, and K" are constants that approximate quantitatively lithological character, moisture content and method of feeding, respectively. T is the hourly weight of material in the feed that is coarser than the coarsest discharge of the crusher. (For marked departures from the normal specific-gravity range-2.5 to 2.7-T should be adjusted to corresponding volumes.) In the crushers under consideration, discharge from the crushing zone is limited by the minimum dimension of a particle. Passage of the particle through a square-mesh screen is limited by its intermediate dimension. Sheppard&apos; has shown that the ratio of maximum: intermediate: minimum average dimensions of broken rock fragments may be taken as about I.5: I: 0.6 for nonslabby rocks such as are normally crushed, while for slabby rocks, such as thin-bedded sedimentaries, foliated schists and gneisses, and the blocky metamorphics, the ratio is of the average order I.3:I:0.3, with the possibility of an even smaller relative value for the thickness. Gaudin2 finds the corresponding ratio for -150 + 200-mesh pyrite to be I.4: I : 0.6. The hourly tonnage of crusher-plant feed, therefore, should be diminished by the amount thereof that would pass a square-mesh screen of which the aperture is S times the crusher setting, where S is a shape factor with the value I.7 for rocks that break exceptionally cubic; 2 for ores and rocks that exhibit no appreciable tendency to slabbiness; and 2.5 to 3, or, in exceptional cases, even larger, when the tendency to slabbiness is marked. R80 is the quotient of the theoretical square-mesh aperture that would pass 8o per cent of the feed, divided by the aperture that would pass S0 per cent of the product. It is best determined from screen tests.

Jan 1, 1941

By L. D. Jaffe, H. C. Mann, E. L. Reed

The authors have recently published1 evidence that brittle transgranular fracture of polycrystalline metals does not originate at a point and propagate continuously across the material, but rather develops at numerous related points, leading to a series of micro-cracks which subsequently link up. Series of related microcracks were observed in steel specimens broken in fatigue—specifically, at the central portions which broke during the final sudden fracture. Microcracks of the type mentioned have also been reported by Baeyertz, Craig and Bumps&apos;2 in steel broken under impact. % Some micrographs published by Irwin3 may possibly be interpreted as showing discontinuous brittle cracks. This paper presents additional observations of the authors on discontinuous crack propagation under various conditions. Impact Test Specimens The impact test specimens were cut longitudinally from a steel forging containing: The heat treatment consisted of oil quenching in a heavy section from 1600°F and tempering at 1260°F. The resulting yield strength was 78,000 psi, tensile strength 100,000 psi, elongation 20.0 pet in 2 in., and reduction of area 55.0 pet. Standard V-notch Charpy specimens were broken in a 217 ft-lb machine at a striking velocity of 16.8 fps. Energy absorption and fracture appearance of this material in the V-notch Charpy test are plotted as functions of temperature in Fig 1. After test, fractured halves of specimens appearing 0, 40, 80, and 100 pet fibrous (to the unaided eye) were electroplated with nickel, sectioned longitudinally (perpendicular to the notch at its midpoint), polished and etched with Cohen-Hurlich-Jacobson (C-H-J) reagent (picric acid plus Zephiran chloride in ethyl ether)4 or with Vilella reagent (1 pet picric plus 5 pct hydrochloric acid in ethyl alcohol).5 Fig 2 and 3 show cracks near the path of final fracture in the specimen rated 0 pct fibrous. Many of the cracks terminate at the boundaries of ferrite grains, some at boundaries of pre- existing austenite gzains. * A crack often starts near the end of another. Cracks within the same ferrite grains are parallel or approximately parallel in several cases. A large percentage of the cracks are branched. Fig 4 illustrated a group of cracks in the 40 pct fibrous specimen. Branching and parallelism are well marked. No more than three directions of crack have been observed on the polished surface in one ferrite grain. This is consistent with the evidence6 that brittle fracture of ferrite takes place on the cube face crystallographic planes: the {100} planes. The appearance of some cracks suggests that they have passed around a projecting corner of a second ferrite grain by going outside the plane of polish. In Fig 5 cracks in the 80 pet fibrous specimen may be observed. Some of

Jan 1, 1950