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By Joseph P. Hodgson

JOSEPH P. HODGSON.-As to ".Safety First," I did not deal with that, but I will say that in 1913, the number of .men killed was nine, and that year the "safety first" program was commenced; the next year, 1914, the number was reduced to five. In 1915, I am pleased to say that the number was reduced to two in the Copper Queen's entire property, and we have none so far this year. THE CHAIRMAN (G. F. G. HERMAN, Bisbee, Ariz.).-I can say, further, that we have from eight to nineteen men per month incapacitated for more than 14 days. Those are classified in this State as serious accidents. The number of men injured has apparently increased, -but that is clue to our improved method of recording and classifying accidents, no matter how slight. C. W. GOODALE, Butte, Mont.-The importance of cooperation among the members of the staff of an industrial organization is well brought out by Mr..Hodgson's paper.

Jan 12, 1916

By N. Mao

This paper presents the results of a research effort to develop a technique for estimating in situ stresses by measuring stress-induced velocity anisotropy around a borehole. Relevant parameters required to make the estimate were identified and measured in the laboratory on a 35.6 cm (14.0 in) cube of Nugget sandstone with a 10.2 cm (4.0 in) diameter hole, under biaxial loading. Two pairs each of radially and tangentially polarized transducers were placed inside the hole with displacement directions either parallel or perpendicular to the principal stress directions. With this configuration, relative travel times were measured by both a pulsed phaselocked loop technique and a cross correlation of digitized waveforms. The biaxial velocity data were used to back-calculate the applied stress. The standard deviation of the differences between the calculated and applied stresses is 0.62 MPa (90 psi) for a stress level up to 8.63 MPa (1250 psi).

Jan 1, 1984

During the latter part of the Civil War an income tax was placed upon coal and the production of coal in the various collection districts during the life of this tax is shown in the tabulation that follows. The counties that include coal deposits are shown for each district. No record exists now of the production by counties, but it is calculated here on the basis of tax collected. The table, of course, does not include any territory not under the control of the northern armies, and not all of that in some cases.

Jan 1, 1942

By Barton K. Cross, Ronald J. Sparks

The question of' providing training for miners in the United States has been answered by Public Law 95-164, which stipulates not only that the mine operator will provide training, but generally how much and what material will be covered. At an operation employing 500 miners, the eight-hour refresher training requirement will necessitate an annual investment of over two man-years in labor, a figure that does not include the time spent in training new hires and job transfers. Such an investment can and should be cost effective, while catering to miners' needs.

Jan 6, 1979

By L. E. Garner

Aggregate resources should be considered primary factors in urban development. Resource planning is essential because land values, previous construction, or zoning restrictions often preempt exploitation of aggregate deposits within an urban area. Sequential land use practices allow development of area prior to urban development population and aggregate consumption trends in Austin, TX, indicate that substantial quantities of reserves will be lost to future development. Aggregate materials are present in nearby areas. However, a comparison of transportation and reclamation costs indicates that reclaiming land after mineral extraction can be significantly cheaper than transporting materials longer distances.

Jan 1, 1982

By Dennis V. D’Andrea, Joseph L. Condon

Laboratory crater tests were performed in two-dimensional plates and three-dimensional .blocks of Holston Limestone. Fractures in the blocks were exposed by sawing through the crater region, and a technique employing dye penetrants is described for examining the resulting fracture systems in both blocks and plates. Crater fracture patterns changed from extension of radial fractures to spa11 fractures parallel to the free face as charge depths were increased. The amount of fractured rock was found to be greater than the amount determined from standard crater volume measurements. Graphs and photographs illustrate these findings.

Jan 1, 1971

By Phillip E. Butler

The Pennsylvania Dept. of Transportation is actively engaged in the utilization of coal mine refuse in the construction of highway embankments. Long-held objections for utilization are invalidated by research, laboratory data, and construction performance. Laboratory, engineering, construction, and environmental impact analyses are presented from the construction of the Cross Valley Expressway embankment at Forty Fort, Luzerne County, Pa. Comparisons of laboratory data and construction measurements are given. The refuse embankment is the first of its size in Pennsylvania, and possibly the U. S., covering 1.5 million cu yd.

Jan 1, 1977

SAFETY in open-pit mining receives a minimum of notoriety, but its significance in profitable and efficient operation is as great as in other methods of mining. Open-pit operators from Arizona, New Mexico, and Mexico, exchanging ideas on safety at a symposium during the recent annual meeting of the Arizona Section, AIME, discovered unanimity of opinion that the successful safety program depended on managerial and supervisory support with the vigorous cooperation of the employes. Safety Organization The operation of a safety program is dependent on the ability of management to project enthusiasm through the company organization. This requires a suitable method of communication to the supervisory and employee levels so as to approach them in a manner that will inspire cooperation. Safety committees comprised of various combinations of managerial, supervisory, and employe personnel are frequently used for this purpose. The open-cut mine of the Chino Division of the Kennecott Copper Corp., Santa Rita, N. Mex., operates its program through management, a full-time safety engineer, and two safety committees, the Departmental and Shops Safety Committee and the Central Safety Committee. The Departmental and Shops Committee consists of one employe from each department and shop and the safety engineer. The selected members serve for a period of one month. This committee meets once a month, the only representative of management being the safety engineer because it is believed that employes will speak more freely when their supervisors are not present. The duties of the committee are to report hazardous conditions and make recommendations, but the committee has no authority to approve or disapprove. The Central Safety Committee is comprised of supervisory personnel from all departments and shops, and one member of each of the five bargaining organizations. Members of this committee are permanent and hold meetings soon after the departmental group to take action on recommendations of the latter. The committee system insures participation of all levels of personnel in the accident-prevention program, thereby stimulating safety consciousness. The New Cornelia branch of the Phelps-Dodge Corp., uses three groups to conduct the safety program; the General Safety Committee, Division Safety Committee, and the operating-unit safety meetings. The unit safety meetings provide for active participation of employes in discussion of various phases of the program. Within these groups vigorous opinions regarding safety are often expressed and the workmen are critical if they recognize a digression from the adopted safe-practice procedures. These monthly discussions give the program vitality. The Division Safety Committee is composed of the division superintendent and the division department heads, and follows an established pattern in discussion of all phases of the safety program. The General Safety Committee, made up of the manager, division, and department heads, makes decisions on policies and procedures; it renders decisions on recommendations and suggestions. Management at the New Cornelia branch feels this system of safety committees and meetings has been active in advancing the program. Safety Codes In the open-pit mines of the Southwest, a safety code for operation or service of equipment, methods

Jan 1, 1952

By R. J. Wasilewski

The variation of elastic modulus with terrzperature between -150° and 600°C has been investigated. Compounds close to equiatomic cornposition exhibit very low modulus values and very high damping near room temperature. The modulus rises steeply with both decreasing and increasing temperature. The modulus minimum and the high damping are believed to be connected with a martensitic transformation of TiNi. In bulk material this transition takes place slightly above room temperature quite rapidly and reversibly, with a hysteresis of some 30°C, depending on the Ti:Ni ratio. Oxygen impurity affects the reaction time. THE inter metallic compound TiNi has been reported to exist in an ordered CsCl (B2) structure,l-6 the homogeneity range varying rather markedly with temperature.7 Some investigators reported this compound to decompose peritectoidally to Ti2Ni and TiNi, between 650° and 800°C,3 or below,',g the latter workers also suggesting the decomposition to be strain-sensitive. Purdy and Parr were the first to report a structural transformation to a new, possibly hexagonal, structure at 36°C,7 a temperature indubitably too low to allow the significant amount of diffusion necessary for a peritectoid reaction. Mechanical-property investigations of compositions close to TiNi just above room temperature have qualitatively shown remarkable anomalies, particularly with respect to clamping,' and to a "memory" effect, that is, plastically deformed material returning to its original shape on being heated through the "decomposition" temperature.9 The original explanation of this effect—based on the apparent experimental observation of Ti2Ni or TiNi, formation in the material subjected to tensile or compressive stresses, respectivelp—does not appear plausible. It appeared more probable that the change involved may be a martensitic, diffusionless structural transformation, similar to those observed in other B2 structures such as 0 brass,'' AuCd,ll and AuZn.12 The existence of such transformation has been established recently,', although the symmetry of the low-temperature structure is not yet known. The present paper describes the elastic modulus and damping behavior of polycrystalline TiNi in both structures. EXPERIMENTAL METHOD 1) Specimens. All the specimens to be investigated were prepared from premelted electrolytic nickel and iodide titanium by nonconsumable arc melting and casting in a water-cooled copper mold. The cast billets were canned in mild steel, extruded, and swaged to approximately 0.375 in. diam. The cans were then removed by pickling or machining. Preliminary material balance checks indicated no significant weight losses on melting (less than 5 mg on a 50-g button); hence nominal compositions have been used throughout. In agreement with previously reported observations7 no single-phase material was obtained at—or close to—Ti 1.00 Ni1.00 a small amount of finely divided second phase always being present. It was found that most of the oxygen impurity was associated with this minor phase, which was identified as oxygen-bearing Ti2Ni—possibly Ti,Ni,O. The amount of this phase—at any given Ti:Ni composi-

Jan 1, 1965

By A. S. Malicsi, I. Iwasaki

The removal of hydrophobic coatings of flotation collectors from iron ores becomes of interest when a duplex flotation process is considered for upgrading, when a pelletizing process is considered for a concentrate floated with a fatty acid or a soap collector, or when a disposal of froth products from cationic silica flotation is of environmental concern. Ozone can oxidize organic compounds rapidly, thereby removing the hydrophobic coatings of flotation collectors. Ozone is widely used for treating and purifying drinking water, waste water treatment, and for chemicals processing (Murphy and Orr, 1975; Rice et al., 1980). Its uses in metallurgical operations, however, are very sparse (Allegrini et al., 1970; Chernobrov and Rozinoyer, 1975; Ishii et al., 1970; Iwasaki and Malicsi, 1985; Matsubara et al., 1978). Yet, its high reactivity and the absence of potentially hazardous byproducts become of interest in destroying flotation reagents adsorbed on mineral surfaces or remaining in mill water for recycle or for discharge. Duplex Flotation A duplex flotation process, as applied to oxidized iron ores, would involve a fatty acid flotation of iron minerals followed by an amine flotation of the siliceous gangue from the rougher iron concentrate. Such a process has been used in the Florida phosphate fields. Fatty acid coatings cannot be removed as readily with a simple acid or alkali treatment from iron oxide surfaces as from Florida phosphates. A combination of reagents, such as lime and quebracho, lime and alkali phosphate, or sulfuric acid and oxalic acid, has therefore been proposed. In a previous article (Iwasaki et al., 1967) , the use of activated carbon was found to be effective in removing fatty acid coatings both in the duplex flotation and the pelletizing processes. The use of ozone offers another approach to the removal of fatty acid coatings from iron oxide surfaces. To investigate the possible application of the duplex flotation process, a specularite ore from Michigan analyzing 36.5% iron was used. A 600-g (1.3-1b) sample was ground in a laboratory rod mill together with 250 g/t (0.5 lb per st) of sodium silicate to -150 µm (-100 mesh). This was transferred to a Fagergren laboratory flotation cell, and deslimed four times at 20 µm (quartz equivalent). The deslimed pulp was transferred to a laboratory conditioner, diluted to 40% solids, and conditioned with 250 g/t (0.5 lb per st) of soda ash and 250 g/t (0.5 lb per st) of oleic acid. The conditioned pulp was then transferred back to the Fagergren cell, floated until barren of froth, and the rougher froth product was returned to the cell and cleaned. The results are presented in Table 1. The cleaner concentrate at this point analyzed 45.3% Fe. The cleaner concentrate coated with fatty acid was transferred to a 2-L (0.53-gal) beaker. While the pulp was agitated with a glass T-stirrer, ozone was bubbled into the agitated pulp for 15 minutes at a rate of 10 mg/min (0.00035 oz per min) ozone (250 g/t or 0.5 lb per st 03 feed). It was observed that the pulp ceased to froth after about 10 minutes. The amine flotation of siliceous gangue from the ozonated pulp was carried out first by conditioning with a dextrin, a commonly used starch depressant for iron oxides. This was followed by flotation with a stage addition of an ether amine at increments of 100 g/t (0.2 lb per st). Three stages were required to float the siliceous gangue to near completion. The three froth products were combined and cleaned twice. When the cationic flotation Rougher, Cleaner 1 and Cleaner 2 cell products were combined, an iron concentrate analyzing 64.5% iron was obtained at an overall iron recovery of 72.8%. Pelletizing Fatty acid flotation concentrates have been pelletized successfully in northern Michigan mills. But at other locations, fatty acid coatings on iron flotation concentrates proved so undesirable in agglomeration that other methods of concentration had to be sought. For example, a sinter mix containing iron ore concentrates upgraded by fatty acid flotation resulted in decreased productivity. This occurred because the micropellets of particles with the hydrophobic coating are less tolerant of moisture. Thus, the bed permeability is lost (Beebe, 1965). The agglomeration of concentrates obtained by the fatty acid flotation alone, and the hydrophobic coatings destroyed by ozonation or by the duplex flotation process, is not expected to cause any difficulty since the surfaces of the concentrates would be hydrophilic. Removal of the fatty acid coating with activated carbon, indicated by the loss of floatability, was shown to restore the decrepitation temperature of wet balls during drying cycle (Iwasaki et al., 1967). Disposal of Cationic Silica Flotation Froths Recent demands of iron blast furnaces place the silica content of the magnetic taconite pellets at about 5%. Conventional process for magnetic taconite involving fine grinding and magnetic separation often produces magnetic concentrates analyzing in excess of 5% silica. This is due to the presence of the middling grains of siliceous gangue and magnetite. Cationic silica flotation of magnetic taconite concentrates (DeVaney, 1949) may be used to reduce the silica content. But the amine coating on siliceous gangue becomes of environmental concern when the flotation tailings are discarded in tailing ponds.

Jan 1, 1986

By W. L. Fink, H. C. Stumpf, L. A. Willey

VERY few studies of the aluminum-sodium system have been reported. Heycock and Neville1 were unable to detect any solubility of sodium in liquid aluminum. Mathewson2 prepared an equilibrium diagram consisting of two horizontal lines at the melting points of the two metals. He was unable to detect solubility at either end of the system. Scheuer3 first developed a method for the chemical determination of sodium in aluminum and then studied the solubility of sodium in liquid aluminum. His method involved heating aluminum with excess sodium under hydrogen until the two liquids were at equilibrium, chilling the melt, and analyzing the aluminum-rich layer. He reported the solubility to be 0.l0 pct sodium at 700°C, 0.115 pct at 750°C, and 0.128 pct at 800°C. This paper on the aluminum-sodium system is the twenty-first in a series from the Aluminum Research Laboratories describing equilibrium relations in aluminum alloy systems. MATERIALS The alloys for this investigation were prepared from electrolytically refined aluminum (99.99 pct) and sodium of reagent quality conforming to specifications of the American Chemical Society.4 Chemical analysis of the aluminum showed the following impurities: 0.002 pct Si, 0.002 pct Fe, 0.002 pct Cu, <0.001 pct Mn, 0.002 pct Mg and 0.001 pct Na. The sodium content of all alloys was determined by either chemical or spectrographic standard procedures5 No. IIA and No. SII respectively. DETERMINATION OF LIQUIDUS The boundary of the miscibility gap was determined by a modification of the method of Scheuer3 and of that used by Henry and Cordiano.6 As finally developed the method was as follows: A fused alumina crucible 25 mm id with 75 mm depth and 1.5 mm wall thickness was placed on the bottom of a container made from a 4-in. length of 1 ½ -in. iron pipe and pipe caps. The space around the sides of the crucible was filled with 40 mesh crystalline alundum. A charge of 60 g of aluminum and 10 to 15 g of sodium was placed in the crucible, the sodium first being trimmed and rinsed with iso-pentane to remove kerosene. The container was tightly capped, filled with argon (99.6 pct pure), and heated to the desired temperature (between 670 and 800°C). A positive pressure of argon (about 8 mm of mercury) was maintained through a small pipe to the upper cap. After soaking for from 2 to 5 hr, the container was quenched by directing a stream of water onto the base, so that solidification of the aluminum-rich layer was rapid and essentially unidirectional. Analyses were made on drillings taken from the bottom of the ingots after machining off the surface layers. Pre-

Jan 1, 1948

By E. C. Tveter, R. B. Tippin

The separation of minerals by heavy liquids is a standard laboratory technique which goes back at least 50 years, but commercially economic application of this principal to ore concenfration has been prevented by problems of liquid recovery. Recent investigations which demonstrate the different types of liquid-solid systems on various ores using the two major heavy liquids - methylene bromide and tetra-bromoelhane — are discussed in this article. As a result of these studies, over-all liquid recovery, and especially desorption of heavy liquid from the mineral surface, was improved by the use of surfactants as well as both thermal and mechanical energy. The examples given in the article indicate that heavy liquids which should find major commercial use in ore concentration are now available. The sink-float process using a heavy liquid medium has long been recognized as a nearly perfect method of mineral separation based on the differences in density of the minerals. The use of heavy liquid separation (HLS) offers the potential of an ore dressing method that provides the maximum possible natural concentrate grade and high recovery. In addi- tion the process requires no water, causes no waste disposal problems, and is extremely simple — the crushed ore is merely introduced into a liquid whose density is between that of the minerals to be separated. The minerals lighter than the liquid float while those heavier than the liquid sink. As such, HLS has become a standard laboratory tool but major commercial application has been prevented by the requirement for complete recovery of the heavy liquid, imposed by economics and materials handling hazards. Investigations, both in the United States and abroad, have shown that improved liquid stability as well as improved chemical engineering techniques brighten the future for HLS in ore dressing. The available commercial heavy liquids suitable for mineral separation cost from $2.00 to $20.00 per gallon1 and on this basis alone their complete recovery is economically desirable. Possibly of even greater importance is the fact that any residual heavy liquid remaining on the products could create a serious toxicity hazard. Consequently, to make the process commercially attractive requires (1) essentially complete removal of the liquid from the ore products and (2) optimum recovery of the liquid for reuse. Until recently these two stipulations have remained as obstacles but laboratory and pilot plant research now in progress suggests that these may soon be overcome. Various heavy liquid recovery systems have been proposed to meet these requirements and might be classified into one or more of the following three general classes: Solvent Wash (Fig. 1): The dissolution of the heavy liquid from the ore with a less expensive solvent,

Jan 1, 1969

By L. C. Noakes

Despite a long history of mining. Australia had no Commonwealth organization dealing specifically with the mineral industry until 1946, when the Australian Bureau of Mineral Resources, Geology and Geophysics was established. Once formed, this new organization quickly expanded to include a Mineral Economics Section, which was established in 1947. Australia&apos;s slowness to develop a Federal bureau for the mineral industry can best be explained in the light of her political history. From a group of British crown colonies, Australia evolved into six sovereign states which had no unified federal government until they formed the present Commonwealth in 1901.

Jan 1, 1970

By Charles L. Elliot

Data on worldwide geophysical activity in application to mining exploration programs is collected annually as part of a continuing service to the mining and petroleum industries. The data for mining geophysical activity in 1966 has been made available to SME by the Society of Exploration Geophysicists. This data is based on 288 voluntary responses for a return of 59% to 487 questionnaires that were circulated among worldwide mining companies, universities, and government institutions that conduct or utilize mining geophysical surveys or carry out allied research efforts. This data is herein published in condensed form with the permission of the SEG.

Jan 11, 1967

By Victor deWolfe

INTRODUCTION The Henderson Mine, located near Empire, Colorado, utilizes a continuous panel caving system to extract ore as one of the world&apos;s major producers of molybdenum. Any mine using a caving-by-gravity technique of mining must rely on closely controlled draw of the caved ore. This control is essential to insure proper caving action, to avoid damaging load concentrations of weight and to minimize the dilution of ore with waste material. Henderson is no exception. Draw control is a major factor in all production planning, from long- range plans to short-range and day-to-day ore scheduling. Draw control is reviewed constantly and administered daily in an effort to optimize production efficiency, ore recovery, and cave management. MINING METHOD The cave at Henderson is massive, moving slowly through large panels that are 244 m (800 ft.) wide by 610 m (2,000 ft.) long. Generally two cave areas are drawn at one time. The areas under active draw vary in size but can be as large as 244 m (800 ft.) by 244 m (800 ft. ) containing 400 draw points. Each draw point contains 45,360 mt (50,000 st) on the average and takes about two and one half years to exhaust. A complete panel is worked for seven to ten years. No pillar exists between panels, but rather a buffer zone of broken ore, or "static face," is left in each panel to be drawn with the adjacent, yet-to-be-caved panel in efforts of minimizing dilution of a working area from an exhausted one. (Figure 1) Production drifts are driven on 24.4 m (80 ft.) centers through the ore body. Between the production drifts are funnel-shaped draw bells on 12.2 m (40 ft.) x 24.4 m (80 ft.) centers to receive ore from the cave. Each bell is accessed by two draw points, one from the production drift on either side, thus forming a 12.2 m (40 ft.) x 12.2 m (40 ft.) draw pattern. Extraction of the ore is by rubber-tired, 3.8 m3 (5 yd3) load-haul-dump equipment. The LHDs then tram the ore a maximum of 49 m (160 ft.) to ore passes. Cave initiation and bell development are done from the undercut drifts which are parallel to and 17 m (55 ft.) directly above the production drifts. Longhole rings are drilled and blasted from the undercut drifts to define the bells and establish the undercut for caving. (Figure 2) DRAW CONTROL Since the cave line at Henderson is constantly advancing, it is necessary to be continually initiating new cave at one end while exhausting it at the opposite end. There must exist, therefore, an angle on the ore-waste contact in the broken rock from initiation to exhaustion. The basic concept of draw control is to keep this angle as smooth and even as possible, particularly at the time of exhaustion. If this is achieved, draw points are exhausted more or less in a line, avoiding pockets of remaining ore surrounded by exhausted areas. These pockets would cause spotty ore extraction at the time of exhaustion, increasing the amount of dilution occurring while introducing the potential for significant weight problems in the production area. To arrive at the desired angle on the ore- waste contact, maximum tonnage percentages are assigned to each row of draw points increasing at 10% or 15% increments (depending on cave size and velocity of draw) working away from the cave line. The available tonnage indicated by these percentages is the maximum allowable tonnage to be extracted from each draw point until the available tonnage percent- age is increased. As the cave moves, these percentages increase for each draw point regularly. However, in general the tonnage drawn from each draw point is kept at about 50% of this allowable maximum in order to maintain adequate available tonnage in the cave to sustain production for seven months if cave initiation were to cease. This available tonnage cushion is a safeguard built into the draw control program at Henderson to accommodate fluctuations in the rate of cave advance. When draw points move past the row of 100% tonnage availability, they are drawn past the desired 50% at the same increments per row until exhausted. (Figure 3) To achieve proper draw control, the number of LHD buckets to be taken from each draw point is assigned daily. The actual buckets taken, which may at times deviate from the

Jan 1, 1981

By D. E. Spiller

Introduction Precious metals recovery from ore can generally be accomplished using gravity concentration, flotation, and/or hydrometallurgical (leaching) techniques. The objective of this paper is to show why gravity concentration can be an important part of recovery systems that employ leaching as the primary unit operation. A brief discussion of modem gravity concentration equipment is also presented. Discussion Gravity concentration of ores has generated increasing interest in recent years. Reasons for this interest include: • Gravity concentration is environmentally attractive. There is little or no use of reagents. Hence, it is relatively nonpolluting. • The cost of cyanide has continued to increase. Therefore, cost savings may be realized whenever leaching feed tonnage can be reduced by preconcentration. • Compared to flotation and leaching, gravity equipment costs are low per processed ton. Field installation costs for gravity circuits usually are less because many&apos; units are supplied as self-contained modules. Also, the cost required to supply services, particularly power, to a gravity plant site are also less. In situations where preconcentration at coarse particle size is applicable, significant grinding equipment savings may be possible. • Gravity circuit operating costs are also relatively low compared to typical flotation and leaching circuits. Reagents, power, maintenance, and manpower savings in a well-engineered gravity plant may be realized. Again, if grinding is reduced, significant power and steel (media and liners) savings are possible. •In recent years, more efficient gravity concentrating devices have been developed. Benefits to Precious Metal Leaching Gravity beneficiation can complement precious metal leaching in two ways. First, the recovery of coarse liberated values before leaching may reduce leach time requirements and may reduce reagent consumption. Second, gravity preconcentration can reduce the size of a leach plant by decreasing the quantity of material to be leached. Coarse gold and silver have been shown to leach rather slowly. Kameda (1949) and Habashi (1967) have investigated the kinetics of cyanide leaching systems. They concur that in a heterogeneous reaction, the rate of gold and silver dissolution is directly proportional to the surface area. Thus, the instantaneous rate of dissolution for spherical 0.37 mm (400 mesh) gold is theoretically -25 times faster than for the same amount of gold at .841 mm (20 mesh), based on data from Fuerstenau, Chander, and Abouzeid (1979). Conversely, coarse liberated, +.841 mm (20 mesh), gold is more readily recovered by gravity concentration than is fine, -.037 mm (400 mesh) gold. Therefore, it is apparent that the two recovery systems complement one another. Figure 1 data demonstrates the potential synergism. A sample of - 3.327 mm (6 mesh) Nevada gold-bearing ore was cyanide leached using conventional bottle-roll test procedures. Gold extraction was determined as a function of leaching time. A second sample split from the same leach feed material was hand jigged to remove a coarse heavy mineral fraction, including virtually all of the +.210 mm (65 mesh) liberated free gold. This second sample, with the coarse gold and heavy minerals removed, was subjected to an identical cyanide leach procedure. Figure 1 presents the resulting comparative extraction data. Note that the percent gold extraction for the sample containing no +.210 mm (65 mesh) free gold includes the coarse gold recovered by gravity. The data show that the sample containing coarse gold required about 72 hours of leaching time to achieve 80% extraction. This compared to about 22 hours of leaching time for 80% gold recovery from the sample that contained only -.210 mm (65 mesh) free gold. Thus, there was a 69% reduction in leaching time. The improved extraction data is not wholly attributable to coarse gold removal, but rather it was the combination of gold removal and rejection of other heavy mineral cyanide consumers or leach retardants. Further investigation was not warranted at this time. Preconcentration is the second manner in which leaching systems can benefit from gravity concentration. The premise is that preconcentration can reduce the quantity of leach feed, which, in turn, may reduce leaching costs. Figure 2 presents preliminary data developed by CSMRI for US Minerals Exploration (USMX). Centennial Exploration Inc., in agreement with USMX, is proceeding with evaluations to determine the suitability of various processing schemes for recovery of gold values from the Montana Tunnels property. The data shows how the ore can be preconcentrated by gravity techniques to result in a reduced feed tonnage to secondary extraction techniques, presumably flotation or cyanide leaching. Testing has shown that Reichert cones, followed by treating the cone concentrate on spirals, can deliver about 88% gold recovery in about 13% weight, that is, 87% weight rejection. Consequently, fine grinding and reagent costs are attributable to only 13% of the plant feed rate. Cost data is not yet available, but the potential exists for significant cost savings.

Jan 1, 1985

By J. M. Lovejoy

Tests are described that show that back pressure on flowing wells is a waste rather than a conservation of natural forces. Stop-cocking, however, gives encouraging results. VARIOUS estimates have been made as to the percentage of oil left in a field after the wells have become so small that it is no longer practical to produce them. Engineers have given the matter much study and can substantiate their figures by data taken from fields from the Appalachians to California. The average oil sand has a porosity of about 20 per cent., which gives a capacity per acre-foot of approximately 1500 bbl. The sand thickness in Oklahoma pools averages 20 ft. With 20 per cent. porosity and a 20-ft. sand thickness, the capacity per acre is 30,000 bbl., yet these pools will seldom yield more than 5000 bbl. to the acre. These figures support the theory that at least 50 per cent. of the oil remains in a sand after all known methods to recover it have been used. Certainly we can never recover all the oil, unless in years to come oil becomes so rare and yet so necessary that the price will justify sinking shafts into the abandoned fields and regular mining methods of drifting and stoping. However, with improved methods of production and conservation of the natural agencies of expulsion, we should be able to increase recovery. It is with this in mind that this paper, relating to the effect of back pressure on wells, is presented. Except in the fields where production is caused by hydrostatic pressure, the chief agency for getting, oil out of the sand into the hole is the gas that occurs with the oil. After the pressure has declined, probably gravity also is a factor, but not an important one. With a given amount of oil in a sand and a given amount of gas occurring with the oil, the ratio of gas volume per barrel of oil is the key to the ultimate recovery from that sand. Our efforts should be along lines that will, lower this ratio and make the gas, the agency given us by nature, do its work better and yield more of the oil, of which it is a part.

Jan 1, 1924

By Charles Clayton

IN connection with certain coöperative work carried on between the Ordnance Department of the U. S. Army, the U. S. Bureau of Mines, and the U. S. Geological Survey during the year 1918, it was the writers&apos; privilege to make metallographic examinations of a great many test bars and other pieces of nickel-steel gun forgings. The material studied came from many different steel plants throughout the country and consisted both of forgings that had been accepted by the Ordnance Department and those that had failed to pass the specified physical tests. In general, these tests require a minimum tensile strength of 95,000 lb. (42,091 kg.), an elastic limit of 65,000 lb. (29,483 kg.), an 18 per cent. elongation, and a 30 per cent. reduction in area. An average chemical composition of the steel studied is 0.38 per cent. carbon and 2.9 per cent. nickel. The steel was of both basic electric and basic open-hearth manufacture. The object of the metallographic work was two-fold: to assist in detecting defects in the steel and, if possible, to suggest remedies for the same; and to obtain as much information as possible in regard to the relation of the microstructure of the metal to its physical properties. The greater part of the investigation was devoted to defective metal that failed to pass the ordnance specifications. Many types and kinds of more or less serious defects thus came under observation, but most of these were due to the attempts by the manufacturer to meet the enormous demand for ordnance steel with more or less inadequate equipment, and were soon remedied. Two defects, the so-called "flakes" or "snowflakes" and a certain "fibrous" or "woody" fracture, were encountered so frequently and were so disastrous to the steel that the greater part of the study was devoted to them. As a result much information as to their nature and their effects

Jan 2, 1919

By T. T. Read

A SHORT time ago, one of the large steel companies courteously furnished the author with detailed records of the operations of a considerable number of iron blast furnaces over a period of two months. These will be made the basis of a subsequent study, but in beginning this study one of -the first points requiring consideration was the probable error of the records, for what otherwise might be regarded as significant differences of performance under different conditions were perhaps really due to errors of observation, or failure to observe and record variations. If it is assumed, in blast-furnace records and calculations, that known quantities or factors are constant, when as a matter of fact they are variable, the effects of these variations may erroneously be ascribed to some other cause. This is quite clearly expressed in the following quotation from E. Buckingham:1 Theory always operates on an ideally simplified picture of reality because real phenomena are unmanageably complicated. The results obtained are not exactly true for any, real phenomenon, though they may be for an ideal one; and the approxi¬mation with which a theoretical equation, however obtained, represents the actual facts, always depends on the approximation in essentials between the ideal picture and its real prototype. The purpose of dimensional reasoning is to find out how some quantity which is involved in the phenomenon under consideration is related to certain others; or to find the relation connecting two or more quantities which vary, or may vary, simulta-neously during the course of the phenomenon. Since we know that we must have a complete equation to start with, we begin by thinking the matter over, to see whether the quantities we have in mind are the only ones involved. Usually it is evident that they are not; so we next make a list of all the quantities we can think of which might under any circumstances be important. Upon considering this list, which is often a long one, it is usually evident that under the actual circumstances a number of these quantities may safely be ignored; so we cross them off the list and thus pass from our most general conception of the phenomenon to an ideal simpler one, in which these quantities are not involved at all.

Jan 3, 1925

By P. K. Lucke

ONE of the interesting features connected with the great continental uplift, which formed the table land of Mexico, is the great depth to which oxidation and secondary enrichment of orebOdies occurred. This table land lies at altitudes ranging from 4000 to. 7000 ft. (1219 to 2133 m.) above sea level. Several cases could be cited of oxidized zones in copper and lead mines which reached depths of upward of 1500 ft. (457 m.) below the surface, and it is universally admitted that the oxidized zone in a mine is limited by the permanent water level. FIG. 1. Mexico, however, like many other countries in which deserts occur, is visited at rare intervals by tremendous floods, which sometimes cause variable water levels over long periods. The following cases bear out the idea, and lead to interesting conclusions of commercial value. My paper demonstrates that carbonates of ores may occur below water level and sulphides above, and I will endeavor to explain the causes of the phenomena in certain specific instances.

Jan 6, 1918