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By Marshall D. Himes, Erich U. Petersen

The Sheep Creek deposit near White Sulfur Springs, Montana, is a middle Proterozoic stratabound and strataform sediment- hosted massive sulfide containing 4.5 million tonnes (5.0 million tons) averaging 2.5% copper, and 0.1% cobalt in the upper sulfide zone (USZ) and 1.8 million tonnes (2.0 million tons) averaging 6% copper in the lower sulfide zone (LSZ). Mineralization is hosted in lower Newland shales and debris flow conglomerates in the graben controlled Helena Embayment of the Belt Basin. USZ mineralization is characteristically high in Cu, Co, Ni, Ba, As, Pb, and Ag. Mineralogically it consists of barite, pyrite, chalcopyrite, and tennantite with minor amounts of marcasite and various cobalt minerals. By contrast, the LSZ is higher in Cu, Ni, and Au and lower in Co, As, Pb, and Ag. Mineralogy of the LSZ consists of dolomite, pyrite, and chalcopyrite. In both zones significant amounts of cobalt, nickel, arsenic and copper are incorporated in pyrite rims. Tennantite controls the abundance of numerous trace elements in the ores. Fluid incluison data from hydrothermal USZ barite suggests that mineralization took place at temperatures in excess of 250°C, under relatively oxidizing conditions (f02 = 10-32+1) from moderately saline (1.1-3.5m), metal rich brines. The absence of barite in the LSZ suggests mineralization under more reduced conditions (f02 = 10.38+2). Mineralization was initiated following basin subsidence in excess of 400m. Contemporaneous with movement along a growth fault(s) was upward migration of hot metal-bearing fluids which permeated the loosely consolidated sediments and may have vented on the sea floor. The differences in the two zones may be explained by the effects of variations in local basin morphology and solution chemistry.

Jan 1, 1990

By Leo J. Miller

Sulphur deposits occur within the upper Salado and Rustler gypsum beds. The formation of a sulphur ore deposit requires large quantities of oil and /or gas, which must find a permeable pathway into the gypsum beds. The Permian beds weathered for 90 million years in an arid environment prior to the advent of a more humid climate at the beginning of the Cretaceous Period. The first Cretaceous rivers enlarged the sinkholes and caverns of the Karstified Permian terrain and deposited coarse conglomerates and sands into the subsurface gypsum beds. The resulting network of caverns and sinkholes, filled with permeable sediments, permitted the free flow of hydrocarbons through gypsum. All sinkholes and caverns infilled with Cretaceous sediments that have penetrated beds of gypsum are associated with sulphur deposits. The Castile gypsum lies directly upon the oil bearing Bell Canyon formation. Sinkholes and caverns filled with Cretaceous sediments occur in the lower Castile. Biogenic calcite and sulphur are adjacent to these sediments. However, faults assume the major role in generating the initial permeability, allowing hydrocarbons to penetrate the lower Castile gypsum. The largest faults are associated with biogenic calcite and sulphur for a distance of 250 meters up from the bottom of the Castile gypsum. It is apparent that two dynamic geological processes are generating permeability within the gypsum beds: (1) the Cretaceous sands decending with difficulty into gypsum and (2) the faults creating permeability in the lower Castile but losing energy through soft bed deformation in the overlying gypsum. Both of these sulphur ore controls are mappable, but another large sulphur deposit of the magnitude of the Culberson Sulphur Mine has yet to be discovered. It appears that massive sulphur deposits are associated with the largest scale examples of each system, that is (1) high densities of deep sinkholes and separately

Jan 1, 1992

By H. C. Schultz

CERTAIN local conditions were known to govern in large measure the successful adaptation of liquid-oxygen explosives to the large-scale blasting at Chuquicamata. The wide variation in hardness of the rock at this mine makes necessary the use of several types of explosives. A locally manufactured, relatively cheap black powder is used in blasting the softer waste and capping material, where a thorough displacement breaks the ground sufficiently. Before the introduction of L. 0. X., an ammonia dynamite of about 65 per cent. strength was being used exclusively in the harder underlying orebody, where a strong disruptive as well as a displacement effect is required for breaking the ore. The L. 0. X., as developed for quarry work in the United States, has a practically pure carbon cartridge material, and the explosive has an extremely high rate of detonation, which should therefore cause it to usefully develop its full disruptive effect in hard ground. The individual blasts at Chuquicamata are large, black-powder shots which dislodge masses of rock averaging 250,000 tons each, and dynamite shots which dislodge masses averaging 80,000 tons each. It was desirable, therefore, to attempt to displace with L. 0. X. the faster, higher-cost ammonia dynamite used in the smaller blasts. It was not considered practical to attempt to displace the low-cost black powder with L. 0. X., as estimated costs showed no probable saving if this were accomplished. Since it was desirable to continue shooting relatively large blasts for effective operation, it was recognized that larger storage facilities than had been in use would be needed for the liquid oxygen. Because of the difficulty of transporting and discharging large storage containers of efficient design, the better plan appeared to be storage and discharge of the liquid at the place of manufacture, and transportation of the filled soaking boxes to the blast.

Jan 1, 1928

By William H. R. Gibney

Pine Point Mines is located in the Canadian Northwest Territories about 1,200 kilometers north of Edmonton, Alberta, close to the south shore of Great Slave Lake. Originally staked in 1898 by prospectors en route to the Dondike gold rush, it was not until the mid-1940's that large scale exploration was undertaken and Pine Point Mines Limited was formed. Work was suspended in 1954 while studies were made to resolve one of the main problems, transportation. In 1961, under the Canadian Government "Roads to Resources" program, agreement was reached for the construction of the Great Slave Lake Railroad and plans made to put the property into production. A townsite was established in 1963 and plant construction commenced. The first production was in the form of high grade ore which was shipped to Kimberley, B.C. for concentration starting in late 1964. The Pine Point concentrator started up in late 1965 when hydroelectric power became available. Initially designed for 4,540 metric ton per day, facilities were expanded to a designed 7,250 metric ton per day in 1969. Problems associated with mining operations at Pine Point are not unique but are augmented by the isolation, climatic conditions, topography, high water table and spread out nature of the operation. Maintenance of a reasonably stable work force is the number one problem but this is gradually being overcome as more local people are trained and oriented from a nomadic type life to an industrial society. The purpose of this paper is to give a general description of the mining operations of Pine Point Mines Limited and to review some of the problems encountered in the operation along with problems associated with northern operations.

Jan 1, 1970

By G. Landry

"IntroductionThe Cape Breton Development Corporation (CBDC) is constantly searching for new technologies that will enhance the marketability of its coal. One technology which shows promise in utilizing Cape Breton coal while at the same time addressing the traditional problem of high sulphur inherent in the Sydney coalfield, is coal-water fuel (CWF).The CWF process incorporates a physical change of coal by grinding it to a flour-like consistency and then mixing additives with it to make a fine coal slurry. The result is a liquid dispersion of fine coal particles that looks like oil, stores and flows like oil but burns like coal. The equipment used in the CWF process is standard mineral processing equipment but application of the equipment to manufacture a specific coal particle size and the specified additives used in the mixing make it a unique technology.The product, which CBDC calls Carbofuel, consists of 70% coal, 29.5% water and 0.5% chemical additives. The key properties of the product are: (I) high solids loading, (2) medium viscosity levels, (3) stability against sedimentation, and (4) good atomization behaviour. Carbofuel specifications are presented in Table I. Carbofuel ProductionThe production of Carbofuel is based on the controlled processing of coal. This involves wet grinding of the coal to a specified fine size (-250 um), cleaning this fine product to remove mineral impurities and formulating (with chemical additives) a finalproduct from the cleaned coal. The flow diagram in Figure 1 indicates the processing circuit for Carbofuel production.CBDC operates a 7-tonne-per-hour continuous production plant complete with state-of-the-art control, full laboratory services, product storage and loadout facilities. The plant has operated since 1983 and has produced over 12000 tonnes for several major demonstrations. The plant is shown in Figure 2."

Jan 1, 1987

By Tyler J. Irving

"In June, 2018 Lalor Mine experienced a catastrophic blade failure of its No. 2 main exhaust fan. Over the following weeks several key changes were made to the auxiliary ventilation system to allow for continued production while the exhaust fans were repaired. Issues encountered during the period of reduced ventilation, and how they were addressed, are analyzed. The suspected causes of the exhaust fan blade failure and how risk was mitigated during the reconstruction are discussed. Key findings and lessons learned are summarized.INTRODUCTIONLalor Mine, owned and operated by Hudbay is an underground zinc, gold, and copper mine located in northern Manitoba. The mine workings extend to a depth of 1090 metres below surface and is ventilated using a push-pull system. The mine exhausts through the main exhaust raise; equipped with two, 2,500 HP fans and upcasts through a portal located at the neighboring mine, Chisel North. Air is supplied to Lalor Mine through four intake fans: The Production Shaft Downcast (two, 250 HP), Lalor Downcast (one, 400 HP), and Chisel Downcast (one, 600 HP) fans (Figure 1). In June 2018, Lalor was scheduled to produce approximately 4,000 tpd.On June 20, 2018 at 2:52 pm Lalor Mine experienced a fault on the Variable Frequency Drive (VFD) for Exhaust Fan No. 1. Time would be needed to complete the necessary repairs. The No. 2 exhaust fan was ramped up from 525 rpm to 800 rpm, 90% of its capacity, in order to maintain underground ventilation. At 4:46 pm, approximately two hours after the rpm was increased, exhaust fan No. 2 experienced a catastrophic blade failure. The forces from the blade failure resulted in the top half of the fan housing being thrown 50 metres from where it was installed; with several large sections of fan blade being thrown further (Figure 2)."

Jan 1, 2019

By Tomio Tsuchiya

A construction method of bored steel H piles enclosed in soil-cement columns was developed by modifying steel H piles from the driven pile method and soil-cement walls from the earth retaining wall method, the latter of which produces cut-off walls and reduces noise and vibration during construction. In this method, as shown in Figure 1, rich mix cement paste is injected into the bottom of a borehole which is pre-excavated by a casing auger, then a steel H beam is inserted before it hardens. Since the high vertical resistance of piles created with this method has already been confirmed by loading tests, we have been investigating the applicability of this method for piles used to temporarily support building loads for top-down construction on raft foundations until the raft foundation is completed. However, in the loading tests, the bearing capacity at the lower end of the pile was measured by loading from ground surface on test piles which had friction cut to eliminate the shaft friction resistance. In other words, the loading tests neglected the effect of overburden pressure of soil above the lower end of the pile. To investigate the effect of the overburden pressure on the bearing capacity at the lower end of the pile, loading tests were carried out on a bored steel H pile of lm diameter (test pile) bored in a compacted sandy soil layer with N-value of 100 to 250 at each stage of a three-step excavation. The results show no significant change in the bearing capacity at the lower end of the pile when the depth ratio of the pile is in the range of 5.7 to 22.7.

Jan 1, 1996

By Gabriel Voicu, Marc Bardoux, Ross Stevenson, Nuno Machado, Jean Lafrance

"Abstract - Paleoproterozoic gold is found in two different volcano-sedimentary sequences and intrusive bodies separated by 600 km within the Guiana Shield. The 2148 Ma St-Élie granite intruded volcanic-arc sequences of the Paramaca Group, whereas the 2096 Ma Omai stock penetrated backarc/ island-arc sequences of the Barama-Mazaruni Group. Positive ?Nd values suggest that both suites lack Archean crustal contaminants, which is consistent with an origin in an oceanic arc setting.The St-Élie deposit is characterized by quartz-dominant vein systems with minor metallic minerals, represented chiefly by pyrrhotite and pyrite. Biotite, calcite, and chlorite are the main gangue minerals. The alteration haloes at St-Élie feature a strong potassic zone, which is absent at Omai. The ore-controlling structures are ductile and the vein emplacement is related mainly to contemporaneous, high-angle ductile shear zones. The St-Élie deposit formed at temperatures >350°C. The stable isotopes values are compatible with a magmatic or metamorphic source.The Omai deposit is characterized by quartz-dominant vein systems with minor metallic minerals, represented mainly by pyrite. Ankerite, scheelite, albite, and chlorite are common gangue minerals. The vein emplacement is syn- to post-deformation and it is related mainly to brittle shear zones with low-angle, stockwork networks and breccia zones. The Omai deposit formed in a temperature range between 250°C and 170°C, much lower than the St-Élie deposit. The stable isotope values are compatible with mixing between deep-seated fluids of metamorphic or magmatic origin with surface- derived fluids.These differences between St-Élie and Omai are interpreted in terms of their depth of formation. Alteration patterns, structural elements, and paragenetic assemblages indicate that the gold at St-Élie was most probably emplaced at a mesozonal depth, whereas the Omai setting yields evidence for epizonal conditions."

Jan 1, 1999

By G. L. Oldright

THE development of the reverberatory furnace for smelting copper ores up to 1912 was described by E. P. Mathewson1 with details concerning the great changes in dimensions of the furnace. Hayward2 tabulated similar data, giving developments up to about 1924, in his revision of Hofman's Metallurgy of Copper. DEVELOPMENT OF PRESENT SMELTING PRACTICE Dimensions of Furnace Very great changes were made in the dimensions of the smelting hearths of the furnaces in the period from about 1800 to 1906, the length increasing from about 11 to 116 ft., and the width from 8 to 19 ft. During this time the tonnage smelted increased from about 9 to 250 tons per 24 hr., and the tons of charge smelted per ton of coal increased from 0.77 to 4.3. Dr. Mathewson3 noted no great difference in the "fuel ratio" (i. e., the amount of charge smelted per ton of coal) as the furnace at Anaconda was increased in length by stages from 85 to 116 ft. (The percentage of copper in the slag, however, was lowered from 0.42 to 0.36 per cent. The other components of the slag were not given.) Lump coal fired on grates was the fuel used during this period. The practice in operating the furnaces over such a length of time naturally varied considerably, and there were differences in composition and nature of the charge, yet it may be considered that the improvements made were due largely to increases in the size of the furnace used. Doubt-less some credit should be given to improvements in operating technique, like changes in the methods of firing the lump coal. However, the greater part of the advance was probably due to factors influencing the

Jan 1, 1928

By D. E. Hilts, R. P. Miller

A large coal-fired power plant is to be built jointly by Pacific Power & Light Co. of Portland, Ore., and by The Washington Water Power Co. of Spokane, Wash., at a site about six miles northeast of Centralia, Wash. (Fig. 1 inset). Coal will be mined from nearby open-pit strip mines that will require cut slopes up to 250 ft in height. Coal will be removed from a seam approximately 50 ft thick, located at the bottom of the mine slopes. The coal fields are located in a geologically complex area that has been subjected to intense folding and faulting. As a result, the soils and rock comprising the overburden are highly fractured, jointed, and faulted. Classical approaches to slope stability based solely on the physical proper- ties of the overburden appeared to offer little assurance that theoretically stable slopes would remain stable. Further, the scope of the exploration program that would be required to provide a reasonable evaluation of the primary and secondary structural features from which a statistical evaluation could be developed was considered impractical. As part of the development of an overall mining plan, an experimental Pilot mine excavation, about 225 ft high and about 400 ft long at the base, was made within the coal field. This experimental cut was to provide firsthand knowledge of cut slope behavior. Our purpose at the Pilot mine was to observe the cut slopes as the excavation proceeded. To accomplish this purpose, the highwall of the experimental cut was instrumented to measure face movements and movement at depth behind the face. In addition, a detailed geologic map of the soil and rock conditions exposed in the highwall was prepared, and continual visual observations of the mine were made, noting significant events as they occurred.

Jan 1, 1970

By F. Clerf

Blast-furnace practice in France is determined more or less by the character of the ores used. Some French ores are siliceous and others are calcareous, therefore by proper burdening a self-fluxing mix can be used. This is not always done, as, for instance, at the Caen furnaces of the Société Métallurgique de Normandie, where the Soumont ore of Normandy is used, which is a carbonate. This ore is crushed and calcined at the mines, yielding a product that is porous and well sized for blast-furnace consumption. The lines of the Caen furnaces are given in Fig. 1. Originally these furnaces had 16 tuyeres; six are plucked and these tuyeres are set in the lower part of the bosh. Operating data on these furnaces are given in Table 1. Table 1.—Operating Data, Caen Furnaces Kg. Lb. Blast-furnace burden per ton of iron: Soumont ore....................................... 200&2100 (44004650) Limestone...................................... 900-1000 (2000-2200) Basic Bessemer slag (bearing phosphorus).............. 100 (220) Scale.............................................. 35 (77) Manganese ores.................................... Coke.............................................. 1100-1200 (2420-2640) Coke analysis: Ash, per cent............. 9.5 H2O, per cent............. 4.0-10.0 Flue dust (approx.), kg. per metric ton.... 200 (440 lb.) Blast volume, cu. m. per hr.............. 83,000 (49,000 cu. ft. per min.) Preesure, lb............................ 12 Blast heat, deg. C....................... 650 (1200" F.) Top temperature, deg. C................ 250 (445O F.) Top-gas analysis, per cent............... CO2 9.3 CO 30.3 O2 0.3 H2 3.0 Production per 24 hr. about 510 tons. MINETTE ORES The most important ore deposits in France are in the west of Lorraine. They are called the "minette" ores and are of two classes-—siliceous and

Jan 1, 1937

White Cliffs Opal Field is situated in the Parish of Kirk, County of YulguIgra, New South Wales, about sixty miles from Wilcannia, and the River Darling. It is near the southern edge of the cretaceous basin of the interior of Queensland, New South Wales, and South Australia. Twelve miles south-west of the field, at Tarella station, a well sunk at the homestead reached silurian slates at about 127 feet, getting no water but about three miles north-west from the originally discovered and no central portion of the field, a well was sunk by the Momba, Pastoral cmpany, to a depth of 270 feet all the way through the cretaceous beds. At 250 feet a small supply of water was struck, but so much mineralised as to be quite useless. The temperature also was so high that a man could not keep his hand in it and Mr. Hogarth, the Momba Pastoral Company's manager, informs me that so corrosive was the action of the water, that a shovel left down the well for a fortnight, at the time of sinking, was completely eaten away.The well having now fallen in, I have been unable to obtain any of the water for analysis.The western termination of this portion of the cretaceous basin is situated about twenty miles from the present opal field, at which distance the slates, with (probably) devonian sandstones and quartzites resting unconformably thereon, commence to appear. The opal bearing country, as far as at present known, follows a more or less well defined line running north and south towards the northern end of the field, and north-east and south-west towards the southern extremity.

WILLIAM B. PRICE* and PHILIP DAVIDSON, ? Waterbury, Conn. (written discussion?).-A few months ago we became interested in deter-mining the properties of. pure nickel as a metal for certain uses and arranged, in cooperation with the International Nickel Co., a series of tests on cold-rolled so-called pure nickel and both longitudinal and transverse specimens annealed at 100° intervals from 250° to 1050° C. Since our tests confirmed; in a general way, the observation put forth by Messrs. Browne and Thompson, they are given at this time. A bar of hot-rolled nickel about 9 in. (23 cm.) wide and 1/2 in. (12.7 mm.) thick made by the International Nickel Co. -of grade A stock, analyzing nickel plus cobalt 98.48 per cent., iron 0.67 per cent., manganese 0.27 per cent., carbon 0.07 per cent., silicon 0.162 per cent., sulfur 0.033 per cent., and copper 0.22 per cent., was cold-rolled down to a final thickness of 0.134 in., malting a 73.2 per cent. reduction by rolling. Test specimens, 9 by 1 in., cut parallel (marked longitudinal) and normal (marked transverse) to the direction of rolling were milled to give a test section 2 5/8 in. long by 0.5 in. wide. Three specimens each, longitudinal and transverse, were annealed at one time in a nichrome-ribbon-wound tube furnace. The temperatures were taken with a Pt-PtRh thermo-couple in conjunction with an Engelhard millivoltmeter. An annealing period of .) 2 hr. plus 25 min. preheating was used in all cases. Sclero-scopic and Brinell hardness tests and static tensile tests were made on the rolled and annealed specimens. The tensile tests are shown in graphic form in Fig. 7. The Brinell and scleroscopic values are given herewith.

Jan 11, 1919

By Rainer Schmid-Fetzer

"Isothermal solidification can be used to form strong bonds at low temperature which remain solid at much higher temperatures. The exploitation of this principally known technique for new high temperature processing steps and applications in electronic assemblies is the aim of this study. Three basically different variants of this bonding process are given. A comprehensive metallurgical screening is performed and the feasibility of binary material systems is given on a quantitative scale. Very attractive systems are spotted that have never been studied before. Detailed examples for the realization of bonding with complex ternary and multilayer systems are given. A highly innovative example is the chip stack or wafer to wafer bonding of unmetallized Si or SiC at 250°C. The proposed bond remains stable above the melting point of Si and enables entirely new semiconductor manufacturing steps. An improved die bonding process was proposed and verified experimentally.1. IntroductionIsothermal solidification can be applied to form strong bonds at relatively low temperature which remain solid at higher temperatures. The general principle of this type of joining processes is illustrated in Fig. 1. An interlayer of a low melting component, Lo, is provided as a foil or a thin film coating between the high melting joint components, Hi. The assembly is heated under moderate pressure to the bonding temperature, T 8 , and a liquid filler forms, either because the melting point of Lo is exceeded or due to a eutectic reaction between Lo and Hi. This molten filler initiates a rapid interdiffusion or reaction diffusion between Lo and Hi and the subsequent approach to an equilibrium state results in isothermal solidification. The solid phases formed at T 8 in the joint region exhibit a remelt temperature above T8 •"

Jan 1, 1996

By A. E. Harnsberger

IT is obvious that details such as the physical and chemical properties and methods of heat-treating of the materials mentioned must be omitted in a paper on the subject of materials used in oil-refinery pumps1. Only sufficient identification is given to enable the reader to find additional information from already published data. Also, the subject covers many operating conditions, but only the most important services can be discussed here. To the oil refiner, the pumping equipment is of vital interest. Practi-cally every refinery operation, from the charging of crude oil to a dis-tillation unit to the shipping of refined products, involves some form of pumping machinery. Intermediate refining stages involve a great variety of pumping machinery to carry out successfully the aims of the refinery technologist. Temperatures at which oils are handled vary from -50° F. to 900° F. Pressure may vary from subatmospheric to 2000 lb. per sq. inch. The selection of the proper materials involves a careful consideration of the conditions under which the particular pump must operate. The four most important considerations are temperature, pressure, corrosion, and abrasion or wear. Consideration of temperature "may be divided into three parts: (1) extremely low temperatures of the order of -50° F.; (2) temperatures at atmospheric or slightly above; (3) temperatures from 200° to 900° F. Pressures fall into two groups: (1) pressures reached by ordinary trade pumps of the order of 200 lb. per sq. in., (2) extreme pressure conditions involving pressures as high as 2000 lb. Noncorrosive and corrosive oils can be treated differently. Under the latter falls corrosion in the presence of water, which will occur up to 250° F. The second type of corrosion occurs in high-temperature pump-ing equipment, with temperatures from 450° to 900° F., largely caused by sulfur compounds. There are also other corrosive conditions met with

Jan 1, 1935

By H. V. Winchell

An examination of the Pachucal and El Oro districts in July, 1920, although cursory and incomplete, disclosed facts of more than passing interest to the student of ore deposits; and inasmuch as the literature on these districts is limited and some of their features are little understood, a brief account may prove valuable as reconnaissance observations to be amplified by more thorough and competent observers in the future. Where no complete monograph of important mining camps has been published a partial presentation of the data is often useful, as it sometimes leads to fuller and more accurate descriptions, by stimulating discussion. The following notes are presented with the consent of A. F. Main, managing director of the El Oro Mining and Railway Co., whose fund of information and wide acquaintance with both districts alone made it possible to cover so much ground in a short time. I have also to acknowledge gratefully the aid of C. A. Lantz and D. S. Calland of the Santa Gertrudis and Real del Monte companies, respectively, in connection with permission to inspect geological features of interest in their mines. Pachuca, Hidalgo, Mex. This mining camp has been described by Ezequiel Ordonez2 and by A. F. J. Bordeaux.3 Reported to have been discovered in 1522, it has produced silver valued at between $250,000,000 and $375,000,000. Subsequent to the discovery here, by B. de Medina, of the patio process. in 1557, its production was increased, and now, using cyanidation, the output of the camp is nearly one-sixth of the production of the world. Topography and Climate The Pachuca district consists of a series of mountains and valleys, the minor features of which extend in an easterly and westerly direction and reach elevations of 8000 to 10,000 ft. (2438 to 3048 m.) above the

Jan 1, 1922

By R. A. Keaens, L. C. IlsLey

For the purpose of this paper a junction box is defined as " a fused switch, inclosed in an explosion-proof compartment, provided with a ready means of connecting or disconnecting a trailing cable without exterior arcing." Junction boxes are intended for use 'at the end of trailing cables and are usually placed outby the last open break through in a coal mine. Where it is not safe to supply current through nips or a trolley, junction boxes suitably spaced along an entry permit tramming and movement of machines with safety, although not with any considerable speed. Junction boxes serve a threefold purpose: (1) They prevent dangerous arcing at the outby end of trailing cables; (2) they afford a safe, practical means of fusing or otherwise protecting the trailing cable; and (3) they provide an efficient means of transmitting power from the power circuit to the trailing cable. The first purpose is the most important from a safety standpoint. Dangerous arcing-that is, arcing capable of igniting gas-frequently takes place where trailing cab1c.s are attached to the power circuit. At either 250 or 500 volts only a very small arc is required to ignite mine gas. The usual method of connecting the trailing cable to the power circuit by cable nips or trolley taps does not make a good electrical connection; hence arcing may be expected at any time. Furthermore, junction boxes eliminate the personal hazard attendant upon the use of trolley taps by removing the possibility of accidental contact with trolley hooks or wire and the liability to injury due to manipulation of the device. In the general scheme of mine ventilation it is assumed that gas will not accumulate where junction boxes are placed, an assumption based on the normal operation of the ventilating system. In many

Jan 1, 1929

By Warren C. Bracewell

At your request, am submitting to you several suggestions for the amendment of our safety code which you felt should be kept up to date. These suggestions are the individual opinions of the several members of the Committee who have expressed (heir wishes to me in the form of letters. I am particularly indebted to Mr. Daniel Harrington, Mr. Wm. Littlejohn, Mr. Warren Bracewell, Mr. James Dalrymple and Mr. Hugh McLeod. Would recommend the following amendments Section 1. Wherever practical trolley wires should be securely hung to roof or timbers at a distance of not less than five feet and not more than six and one-half feet above rail and six inches to one side. Spacing of hangers should be from 20 to 25 feet apart. (Adopted.) Section 2. 500 volts D. C. should not be allowed in any installations. Trolley wires should not carry a voltage of more than 250 D. C. nominal voltage. (Referred back.) Pumps, mining machines and other small machinery units where A. C. power is used should not carry a voltage of more than 220. Large units, such as pumps, motor generator sets and hoists, may use a volt- age of 2,200. When power cables to operate same are of armored construction, cables should be securely buried at all crossings, on haulage-ways or where men travel. (Adopted.) Section. 4. Where power lines pass through door frames or stop- pings they should be properly insulated with porcelain compound or rubber tubing. Wherever men get on or off man-trips in vicinity of trolley wires same should be secure1y guarded and should also be protected with a cut- out switch, same to be pulled before men are allowed to get on or off trips. (Adopted.)

Jan 1, 1926

By Willard S. Morse

The purpose of this paper is to record the results obtained in the use of the Russell process at Aspen, Colo., covering a period of fourteen months, from November, 1891, to December, 1892, during which time over 30,000 tons of ore, purchased outright, on sample and assay, were treated. No explanation of results will be attempted, nor will the chemistry of the process be discussed. The plant was designed by C. A. Stetefeldt, of San Francisco, and built by the owners by day-labor. The entire plant is run by water-power, supplied by a flume 3600 feet long, and a pipe-line 2200 feet long, with a head of 170 feet. A Pelton water-wheel, 8 feet in diameter, is used, from which power is transmitted by wire-rope to the main counter-shaft in the mill, a distance of about 250 feet. From the mill to the samplingworks, about 150 feet, power is also transmitted by a wirerope. A Pelton water-wheel, 3 feet in diameter, supplies power to the dynamo for lighting the plant. The works are located adjoining the Santa Fe railroad tracks, and all ore and supplies are unloaded directly from the cars to the works. The sampling-works are provided with crushers and rolls, the latter being used only on samples, and not on the entire ore. The Cornish hand-quartering method is used in sampling ore for purchase. After crushing and sampling, the ore goes to the bedding-floor, where it is made up into mixtures suitable for the mill. The drying-plant consists of four Stetefeldt double shelfdryers, six shelves high, and is provided with dust-chambers and a stack 65 feet high for draft. These dryers are fired with gas

Jan 1, 1896

By William L. Fink and

Although age-hardening in an aluminum-base alloy containing magnesium was observed by Alfred Wilml as early as 1911, it was not until 1919 that a theory of the mechanism of age-hardening was proposed. Merica, Waltenberg and Scott12 studying the age-hardening characteristics of Duralumin at the National Bureau of Standards, came to the conclusion that age-hardening of this alloy is caused by the precipitation of CuAl2, which is less soluble at the aging temperature than at the solution heat-treating temperature. It was postulated that the initially precipitated particles were of fine colloidal dimensions but that they gradually grew in size. Maximum hardness was assumed to be attained when the particles reached a critical size. This theory was at first received skeptically, since it was a widely accepted view at that time that the solid solution state should represent the hardest state of an alloy. Shortly thereafter, Jeffries and Archer proposed their well-known slip interference theory.'13 They suggested that age-hardening was accomplished by the precipitated particles acting as keys interfering with slip along crystallographic glide planes. They showed that this mechanical theory led direct,ly to the assumption that maximum hardening would be produced by particles of a critical size, which they suggested would have a diameter approximately 10 times that of the aluminum atom. This contribution amplified and explained so rationally the age-hardening theory of Merica, Waltenberg and Scott that wide acceptance resulted. Objections to Simple Precipitation Theory From time to time certain experimental facts have been cited against the simple precipitation theory. The first of these was the discovery by Fraenkel4 that, although the electrical resistivity of Duralumin decreased during aging at higher temperatures (from 100' to 250' C.), it increased during aging at ordinary temperatures. This was not in accordance

Jan 1, 1940