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By Eugene F. Poncelet

GLASS squares compressed on edge by steel jaws in poor contact with them developed jagged "partial-contact" cracks caused by the formation of local tensile stresses. Compressed by steel jaws in perfect contact, they developed smooth "release cracks" during release of the pressure, caused by the formation of local tensile stresses during the release of the pressure. All these cracks were parallel to the pressure. A Microflash photograph of a disintegrating specimen under sufficient pressure reveals a network of fractures roughly normal to each other, also "release cracks" and a disintegration cloud. The Griffith theory is amended to account for the formation of a first crack. A new theory, based on the theory of thermal agitation and wave propagation, is proposed to account for the progress, velocity and forking of cracks. The network of fractures is shown to have been caused by reflection at a free boundary of pressure pulses emanating from a first crack. Postulating equal distribution of energy in the pulses emitted on either side by a crack in particles, the smaller fragments are shown to continue fracturing preferentially while some of the coarser fragments remain as residual pieces. As comminution of the smaller fragments proceeds, the solid is reduced to a collection of residual particles of smaller and smaller sizes, accounting for the disintegration cloud. EXPERIMENTAL PROCEDURE Apparatus A crushing device was built, consisting of a steel frame and two polished, hardened steel jaws to crush plate-glass samples. This device was placed between the plates of a hydraulic press (Fig. 1). For centering the pressure, the upper jaw was guided vertically and provided with a rounded top. Later it was replaced by a steel jaw ground out of a 2 1/2-in. steel ball, the lower surface being ground to an optical flat (Fig. 2). Four rollers were used to eliminate side thrust. Screws and centering blocks served to center the specimen in the crushing device. Hard steel shims placed under the shoulder of the upper jaw limited the "follow through" of the jaw after failure of the specimen. Stress distribution was first examined with a polariscope mounted on an optical bench, using monochromatic light. It was found convenient to mount polaroid, quarter wave and monochromatic plates directly on the device itself. A Microflash unit* was used as light source for the photographs. A gold-paint pattern was fused on the plate-glass specimens and connected into the spark-

Jan 1, 1944

By T. E. Leonitis, F. N. Rhines

The oxide scale that forms upon magnesium at elevated temperatures is non-protective in the sense that the rate of oxidation is constant and thus does not decrease with the growth of the scale as it does with other common metals. This generality was first stated by Pilling and Bedworth' and has been verified by Suzuki2 and by ScheiL3 Pilling and Bcdworth argued that the nonprotective characteristic could be predicted from the fact that magnesium oxide occupies less space than does the metal from which it springs, wherefore the scale is not expected to cover the metal so completely as to exclude all direct access to the atmosphere. Linear oxidation is thought by Scheil to be common to all cases where the reaction occurs at the oxide-metal interface. According to Suzuki," he product of high-temperature oxidation in the air is MgO contaminated with no more than traces of a nitride. When the metal is allowed to burn freely a fume composed of cubic crystals of Mg04 is given off. Delavaults observed that excrescences form upon liquid magnesium and alloys in thc course of oxidation. Beyond this it is known that the use of atmospheres containing small quantities of sulphur dioxide6 or carbon dioxide serve to retard the oxidation of solid magnesium and that beryllium' and calciums minimize the oxidation of molten magnesium exposed to the air. Many studies have been devoted to the nature and rates of "protective" oxidation, as exemplified by the cases of copper and iron, wherein the rate is characteristically parabolic. The theory of parabolic oxidation is well developed.~ Linear oxidation, on the other hand, has received but little attention and the theories proposed by Pilling and Bedworth and by Schcil have yet to withstand careful examination. There exists no comprehensive survey of the high-temperature oxidation of magnesium (or of any other metal that exhibits linear oxidation) over a broad range of temperature, under a variety of atmospheric conditions and covering a large group of alloys. It has been the purpose of the present investigation to provide such a survey, with the object of gaining a fuller understanding of linear oxidation, especially as applied to magnesium. Experimental Procedure and Results The experimental studies undertaken were of two kinds: (I) measurement of the rate of oxidation as influenced by the 6c

Jan 1, 1946

By Charles H. Fulton

Experimental work on the process was begun on a laboratory scale at Cleveland, Ohio, in 1914, and transferred to East St. Louis, Ill. in 1916, where a commercial sized furnace was in technical operation until January, 1918. The essential steps of the process are as follows: Oxidized zinc ore or roasted zinc concentrate is mixed with crushed coke and coal-tar pitch and formed into briquets 9.25 in. (23.5 cm.) in diameter and 21 in. (53 cm.) long in a manner similar to that used in the manufacture of graphite or carbon electrodes, except that much less care and time are required. The composition varies with the nature of the ore; in one case, the composition was 100 parts ore, 70 parts coke, and 18 to 20 parts pitch. These briquets maintain their form and volume during and after the distillation of the zinc. This object is gained by using coke as the matrix and coal-tar pitch as the binder; this pitch becomes coke on heating and unites the ore particle and the original coke particle into a continuous mass. The briquet is an electrical conductor but only to such a degree that it can be used as a resistor. The size of the ore, as in the present practice of zinc metallurgy, ranges from fine material, such as flotation concentrate, to coarse table and fine jig concentrate. The process is not restricted to any particular kind of ore, but is applicable to pure high-grade ores, ores high in iron, and complex zinc-lead ores, since the residue from the distillation IS held immovably in place within the briquet. In the case of complex zinc-lead ores, the distillation is conducted with a carefully regulated temperature so that a high percentage of lead is retained in the distilled briquet, which may then be smelted for its lead content. In present zinc smelting practice, the reduction fuel is from 40 to 50 per cent. of the weight of the ore. In the briquet, it is from 60 to 85 per cent. of the ore; but unless the ore is exceptionally high in residue, the distilled briquet may be used over again as coke, or as a high ash coke for fuel purposes in power production. Baked briquets are hard and tough and may be handled roughly

Jan 1, 1921

By Gordon E. Agar

"The effectiveness of two this compounds for copper sulphide depression in molybdenite flotation has been examined with batch bench scale tests.Both thioglycollate and trithiocarbonate proved to be good depressants producing results similar to those f or cyanide and hydrosulphide. Ferrocyanide was ineffective on the samplematerial tested.The depressing effect of thioglycollate was readily reversed by an addition of xanthate.IntroductionThe separation of molybdenite from copper sulphide minerals has been practised for a long time(1,2). The predominant scheme is to depress the copper sulphides while capitalizing on the ready floatability of molybdenite to effect the separation. There are many recipes for depressing the copper minerals(3), among them the addition of sodium cyanide, particularly if the copper is present primarily as chalcopyrite; some form of sulphide whether as sodium sulphide, sodium hydrosulphide or phosphorous pentasulphide when there is a mixture of copper sulphides such as bornite, chalcocite and chalcopyrite; potassium ferro cyanide may be used when chalcocite is the principal copper sulphide; peroxide or hypochlorite; as well as heating in the presence of lime. There are different mechanisms by which these reagent s depress the copper sulphides. Sodium cyanide and the sulphide-producing compounds apparently reduce the potential of the copper sulphides to the point where they ar e rendered unfloatable by xanthates . Another study(4) showed that after cyanide addition, chalcopyrite was rendered un floatabl e until the REDOX potential (measured with a gold electrode with a saturated calomel reference) had become more positive than -250 mY. If the chalcopyrite pulp was conditioned after the cyanide addition the REDOX potential increased slowly, whereas if air was introduced during the conditioning the potential r ose much more rapidly. The difference in time to reach the desired level of - 250 mV was as much as 20 minutes' 25 minutes being required when air was absent but less than 5 minutes when air was admitted . These observation s demonstrate some of the problems encountered when using cyanide to depress sulphides such as chalcopyrite . The system adjusts to counteract the depressing effect of the additive. To overcome this tendency it is necessary to either add an overdose at the beginning or to more or less continuously add the regulating reagent. In this study cyanide solution was added under automatic control to maintain a specified REDOX potential."

Jan 1, 1984

By Eugene F. Poncelet

GLASS squares compressed on edge by steel jaws in poor contact with them developed jagged "partial-contact" cracks caused by the formation of local tensile stresses. Compressed by steel jaws in perfect contact, they developed smooth "release cracks" during release of the pressure, caused by the formation of local tensile stresses during the release of the pressure. All these cracks were parallel to the pressure. A Microflash photograph of a disintegrating specimen under sufficient pressure reveals a network of fractures roughly normal to each other, also "release cracks" and a disintegration cloud. The Griffith theory is amended to account for the formation of a first crack. A new theory, based on the theory of thermal agitation and wave propagation, is proposed to account for the progress, velocity and forking of cracks. The network of fractures is shown to have been caused by reflection at a free boundary of pressure pulses emanating from a first crack. Postulating equal distribution of energy in the pulses emitted on either side by a crack in particles, the smaller fragments are shown to continue fracturing preferentially while some of the coarser fragments remain as residual pieces. As comminution of the smaller fragments proceeds, the solid is reduced to a collection of residual particles of smaller and smaller sizes, accounting for the disintegration cloud. EXPERIMENTAL PROCEDURE Apparatus A crushing device was built, consisting of a steel frame and two polished, hardened steel jaws to crush plate-glass samples. This device was placed between the plates of a hydraulic press (Fig. I). For centering the pressure, the upper jaw was guided vertically and provided with a rounded top. Later it was replaced by a steel jaw ground out of a 2 1/2-in. steel ball, the lower surface being ground to an optical flat (Fig. 2). Four rollers were used to eliminate side thrust. Screws and centering blocks served to center the specimen in the crushing device. Hard steel shims placed under the shoulder of the upper jaw limited the "follow through" of the jaw after failure of the specimen. Stress distribution was first examined with a polariscope mounted on an optical bench, using monochromatic light. It was found convenient to mount polaroid, quarter wave and monochromatic plates directly on the device itself. A Microflash unit* was used as light source for the photographs. A gold-paint pattern was fused on the plate-glass specimens and connected into the spark-

Jan 1, 1944

By Francis Foley

Small cracks in a plane normal to the axis of steels are found to be prevalent around the water hole of drill steels that have been in service for an unknown period of time. Cracks are not found on the exterior surface of the steels nor to any noteworthy extent even in the water holes of new drill steels. Undulations and serrations are found in the surface of the water hole in new steels; it is in the bottom of such defects that cracks appear to develop in used steels. The occurrence of the more serious of these cracks at the point in the steel to which it has been heated for forging or hardening is thought to account to a great extent for failures in that region. DURING the past, year, we have investigated the cause of 88 per cent, of the failures by breakage near the bit end of some 1-in. hollow, hexagon, drill steel used in a metal mine. This breakage in the region of the bit end of the steel was peculiarly interesting because we have been led to believe that breakage is most prevalent at the shank end of drill steels near the place where the steel is gripped in the chuck of the machine. Records from the mine showed, also, that the breakage was much more frequent in the short lengths of large gage than in the longer steels; this is shown graphically in Fig. 1. Because the shorter lengths of steels are usually those that have been made from broken pieces of long steels or from steels that have worn down in service, rarely being made from new stock, those who are wont to suspect steel of tiring and crystallizing in its old age through fatigue are quite apt to see in this instance an argument in favor of their idea. As usual, a careful examination of these steels reveals defects of a character hardly to be overlooked in favor of such a cause. The steel in question was rolled with a sand core and was of the usual drill steel composition: Carbon 0.85 per cent., manganese 0.60 per cent., silicon 0.25 per cent., phosphorus 0.02 per cent., and sulfur 0.02 per cent., with the exception that in some specimens there was vanadium, not exceeding 0.20 per cent.

Jan 3, 1924

By M B. Lovitt

At present there exist few engineering calculations and principles to determine exclusion zones for personnel in the vicinity of underground explosive storage areas. This paper is written to encourage discussion and proposes a method to determine the impact of a blast wave through scientific principles. Then, criteria to determine an ‘exclusion zone’ need to be identified.There has been much work on identifying the pressure rise and the energy of explosive detonations underground but little to develop standards in criteria that can be used to identify exclusion zones for useful design of underground infrastructure. The author presents the newest ways of calculating the pressure rise and energy (impulse) of detonations and puts forward some criteria to be used in the identification of personnel exclusion zones to spark peer review and discussion of these criteria towards developing standards for engineering design.Smith and Sapko (2005) used Sach’s scaling laws to develop a relationship for both pressure rise and impulse, which incorporated factors for intersections. Intersections will have a large bearing on the peak pressure and the energy of the blast wave and needs to be incorporated. Silvestrini, Genova and Leon Trujillo (2009) developed the scaling laws with an energy concentration factor to predict blast propagation in partially confined geometries and together the author has proposed equations for determining the peak side pressure and the impulse in a tunnel. Given the weight and type of the explosive, the average cross-section of the tunnel complex and the number and type of intersections encountered, the pressure and impulse can be calculated.The military has carried out many investigations as to the influence of blast pressure and impulse on the human body. Lung damage, ear damage and velocity knock down are three of the criteria standards that are presented as possible methods of determining exclusion zones around an underground explosives magazine. The author suggests that the velocity curves presented by Bowen, Fletcher and Richmond (1968) could be an acceptable method of determining an exclusion zone as it incorporates the side pressure and the energy of a blast wave to determine the ability of the wave to knock down personnel.CITATION:Lovitt, M B, 2014. Exclusion zone calculation for underground explosives storage, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 319–326 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

By George S. Rice

[This paper gives the details of tests to determine the design of stoppings capable of withstanding a pressure, applied to either side, of 50 pounds per square inch,6 as required by section 104 (a) of the Operating Regulations to Govern Coal-Mining Methods on Leased Lands on the Public Domain, issued in 1921 and quoted oil page 3. The regulations stipulated that the design for such stoppings must lie approved by the United States Bureau of Standards. The engineers of that organization felt that knowledge of the strength of such structures with respect to the effect of explosion pressure in comparison with static pressure and the strength of the walls which furnish the abutments was insufficient to justify drafting specifications; therefore, tests were needed to give information upon which a rational system of design could be based. A study was then begun; and several series of tests, made by the United States Bureau of Mines and the United States Bureau of Standards in cooperation, grave information upon which are based the detailed specifications offered in a later section. Slabs representing concrete stoppings, with and without reinforcement, were tested at the Bureau of Standards in Washington D. C.; afterwards, full-scale stoppings were tested at the Bureau of Mines Experimental mine near Pittsburgh. It was assumed that in most coal mines the strength of the floor and roof could not be relied on, but dependence must be placed upon the strength of the coal ribs. The tests developed that the maximum strength of an unreinforced concrete stopping was obtained when it acted as a flat arch buttressed against the coil ribs; but its ultimate strength depended upon the resistance to movement offered by the coal strata against which its ends were restrained or buttressed.]

Jan 1, 1931

By Dario Libano, Luis Carlos Tenuta, Juliano Penteado

"Subway Line Number 4 is one of the newest additions to the Rio de Janeiro subway network. Built for the 2016 Olympic Games, the 16 km (10 mi) subway line connects the south of the city, with the famous beaches Copacabana and Ipanema, to Barra da Tijuca, west of the city, where the Olympic Village and Park are located. This new line, which is estimated to cost USD 2.2 billion, is composed of six new stations: Jardim Oceanico, São Conrado, Gavea, Antero de Quintal, Jardim de Alah (JA) and Nossa Senhora da Paz Station.Brasfond participated in the construction of foundations and retaining structures for all the stations, and each presented unique challenges. This is not surprising considering the difficulties present in a highly-dense urban environment including utilities, nearby buildings, traffic limitations and noise.Of all the stations, JA Station was the most challenging to construct. It was located in the middle of an avenue between residential high rises supported on shallow foundations as well as below sea level in sands with high groundwater. The other five stations were located in small parks and squares, with no buildings in the immediate proximity.The interaction between designer, owner, and the general and foundation contractors during execution was key in safely reducing the construction schedule. This was significant given the unforeseen construction delays the client, Linha 4 Sul JV, faced due to permitting, utility and traffic relocation so that the owner could fulfill their final goal: making the line operational by the Olympic Games. The GeologyGeologically, the state of Rio de Janeiro is formed by Pre- Cambrian crystalline basement rock. Its geological formation is very old and comprises two major orogenic cycles, a Mid-Precambrian formation, called “Transamazonian” (Proterozoic Era), and an Upper-Precambrian to Cambrian called “Brasiliano” (Proterozoic-Paleozoic). The soil structure of the city of Rio de Janeiro is represented by a core of crystalline rocks, usually granite-gneiss, covered with a layer of residual soil and/or partially weathered rock.JA Station is located in a Rio de Janeiro neighborhood called Ipanema, between the Atlantic Ocean and Lagoon Rodrigo de Freitas. Local geology is composed of a superficial soil, consisting of medium to coarse grained sands. Top of rock varied significantly from 20 m (66 ft) deep on one side of the station to 33 m (108 ft) on the other. Water is at a depth of 3 m (10 ft)."

Jan 1, 2016

By Isaac A. Zlochower, Michael J. Sapko, Eric S. Weiss, Kenneth L. Cashdollar

This paper describes dust explosion research conducted in an experimental mine and in a 20-L laboratory chamber at the Pittsburgh Research Laboratory (PRL) of the National Institute for Occupational Safety and Health (NIOSH). The primary purpose of this research is to improve safety in mining, but the data are also useful to other industries that manufacture, process, or use combustible dusts. Explosion characteristics such as the minimum explosible concentration and the rock dust inerting requirements were measured for various combustible dusts from the mining industries. These dusts included bituminous coats, gilsonite, oil shales, and sulfide ores. The full-scale tests were conducted in the Lake Lynn experimental mine of NIOSH. The mine tests were initiated by a methane-air explosion at the face (closed end) that both entrained and ignited the dust. The laboratory-scale tests were conducted in the 20-L chamber using ignitors of various energies. One purpose of the laboratory and mine comparison is to determine the conditions under which the laboratory tests best simulate the full-scale tests. The results of this research showed relatively good agreement between the laboratory and the large-scale tests in determining explosion limits. Full-scale experiments in the experimental mine were also conducted to evaluate the explosion resistance characteristics of seals that are used to separate non-ventilated, inactive workings from active workings of a mine. Results of these explosion tests show significant increases in explosion overpressure due to added coal dust and indications of pressure piling.

By Alfred H. Bell, Paul F. Simpson

The year 1929 was one of continued activity in the petroleum industry of Indiana and Illinois but the new production obtained has not been sufficient to offset the decline in the production of the old wells. (Table 1). Drilling costs have not changed materially during the year and the prices of crude oil have increased slightly so that the prospects for further testing are good. The total number of completions in Illinois in 1929 (Table 2) was greater than for any year since 1917. More than half of these were in the Dupo field, St. Clair County, and the development of this field was largely responsible for the increased ratio of producers to dry holes in the state as a whole. In the southeastern Illinois field which has produced about 97 per cent. of the state's petroleum the number of completions was greater than in 1928 and the percentage of producers was a little higher. Although a number of scattered wildcat tests were drilled most of them were not located on favorable structure and the only noteworthy discovery was that of a shale-gas horizon at shallow depth in Morgan County in western Illinois where gas has been produced for a number of years from sandstone and limestone. In Indiana the number of completions (Table 2) was only slightly less than in 1928. Most of the new wells were drilled in the southwestern part of the state, which now produces more than 90 per cent. of Indiana's petroleum.

Jan 1, 1930

Date of Congress: Sept. 20-25, 1915 The Committee on Local Affairs is fortunate in having secured for a period expiring on Oct. 17, an option. on 100 rooms at the Palace, St. Francis. and Fairmont hotels, the three leading hotels in San Francisco, with the distinct understanding that, on account of previous engagements whereby the hotels are reserved to their capacity, the rooms must be vacated on Sept. 25, 1915. The rates applying will be $7 per day for the court rooms occupied by two people, or $10 per clay for outside rooms occupied by two people. The outside rooms will face the main thoroughfares at the Palace and will face the city and bay at the Fairmont, which latter is situated oil a high hill. The St. Francis faces an open square and is flanked by main thoroughfares on two sides. If any members desire their reservation to extend either before the 20th or beyond the 25th, the Committee will use its best endeavors to secure such extension in individual cases, in accordance with the indicated wish. The Hotel Bureau in San Francisco is co-operating actively and closely with the. Committee on Local Affairs of the International Engineering Congress and offers the choicest rooms in any of 350 hotels other than the three mentioned above. This Hotel Bureau controls 50,000 rooms in San Francisco, which are all provided with hot and cold water, stationary basins, baths, house telephones, elevators, and bell boy service.

Jan 9, 1914

By Anders Lagerkvist, Agne Rustan, Tommy Edeskar

Landfills containing house hold waste produces methane gas during its decomposition. This gas can be recovered and used for heating purposes or as fuel for vehicles. Gas drainage pipes are installed in the landfill and connected to a central suction point. Methane gas is produced over several years until the slots in the drainage pipes are clogged by black mud. And the bottom of the vertical drainage pipe are filled with mud and water. The hypothesis tested in this paper is if blasting in the drainage pipes could activate the production and flow of methane. In the first part of the paper we have studied what maximum linear charge concentration could be used in different size of PE-pipes place vertical on surface and surrounded by air. The results show that the maximum linear charge concentration for single blasts varied between 5-10 g/m, depending on the diameter of the pipe (90, 100 and 160 mm). Maximum charge concentration for the repeated blast varied between 2-5 g/m. In the second part of the paper we study what maximum linear charge concentration could be used in situ for full scale blasting tests in a special prepared test area which later on was dug out after the blastings. Up to 40 g/m could be used without damaging the top of the 160 mm diameter pipes. The increase in methane flow was so low so it could not be measured c4.5 m3/h. The conclusion therefore is that the stimulation of methane flow by explosives is not enough.

Jan 1, 2000

By E. V. Given

In speaking of the concentration of oxidized lead ores at St. Anthony, I do not wish to infer that we believe all lead ores of a similar grade can be mined and treated at a profit. Our oxidized ore is low grade, and certainly the mine under present economic conditions could not operate on this ore by itself. Because of favorable conditions we can operate the oxide circuit at a small profit although we do not allocate overhead, pumping, and certain other general charges to it. Operations in the oxidized zone were started as an essential part of the development necessary to outline and mine the sulphide ores. Nearly all of our oxidized lead ore comes from the Collins section of the mine, with the transition from oxide to sulphide ore lying between the 600 and 700-ft levels. Production from sulphide ore is much larger than from the oxide, both in tonnage handled and in grade of ore. The oxide and sulphide ores are never mixed either in the mine or the mill. The oxide mill sections have their own ore bins, grinding, and flotation departments. The lead minerals contained in the ore, in the order of their importance are cerussite, wulfenite, galena, anglesite, and vanadinite. The ratios of the different lead minerals change as the ore is pulled from different parts of the mine-total lead being approximately 50% to 60% cerussite, 20% wulfenite, 15% galena, zero to 10% anglesite, and 2% vanadinite. About 3% of the lead was unidentified.

Jan 5, 1950

By Westcott PC

The development of a coal project must pass through a number of phases, taking it from a preliminary evaluation through to detailed mine planning. For each phase the cost of data acquisition and associated mining studies increases exponentially for linear improvement in confidence levels. Within each mine planning phase there are four main activities. These are the economic ordering of reserves, the development of optimum mining strategies, mine design and scheduling and finally the estimation and analysis of costs. The development of the optimum mining strategy is one of the most critical aspects of a mining evaluation and the application of economic criteria at this stage of the project is vital to the successful generation of an optimised mine design. A recommended method is the concept of Economic Ratio (ER) which expresses the likely cost of mining from a given location in a deposit in terms of the value of the product likely to be recovered. The most economic mining strategy will be to mine areas from low economic ratio to high economic ratio within the constraints of reasonable equipment application. When selecting equipment a particular machine should not be selected solely on comparison with other machines in isolation, but the total application must be analysed. In the final selection the decision should be based on whether or not it has a beneficial effect on overall project economics. The economic assessment must take into account the time value of money, so the real cost can be determined with adequate provision for capital and operating costs over different time spans.

Jan 1, 1983

By G. W. Holl, E. G. Fairon

A review of some aspects of shaft design by G. W. HOLL, Pr. Eng., Associate of the Mine Managers Association of South Africa (Fellow)- and E. G. FAIRON, Pr. Eng., F.S.A.I.E.E., M.I. Cert. M.E.E., M.I.Mech. E., (Visitor) The reasons for the choice of vertical shafts in preference to inclined shafts of comparable capacity are set out, and the conclusion that the former are superior in virtually all circumstances is substantiated. The shapes of vertical shafts sunk during the past sixty years are recorded, and the shapes of those sunk during the past decade are analysed. From this analysis it is clear that circular and elliptical (and quasi-elliptical) shafts are the shafts of the future and that the rectangular shape will conceivably become obsolete. The determination of the size of shaft is discussed and it is forecast that shafts of a diameter of 13 metres and over will probably be sunk without difficulty in the years to come. The depth of the shaft is determined intrinsically by the Mines and Works Regulations applicable to factors pf safety of winding ropes and the design of the winding system. The Blair winder is the most suitable winder for hoisting rock, men, and material from a depth of, say, 3000 metres on an economic basis. Tertiary shafts will, therefore, be unnecessary in future deep-level mining projects at the limit of depth now envisaged. A few other factors relating to shaft design are described.

By K. Chandrasekaran

Carbon-in-leach (CIL) is conventionally used in gold processing operations when preg robbing components are present in the ore. The CIL process is becoming attractive even in other situations when preg robbing is not a factor. This article discusses the situations in which the CIL circuit can be applied. The results of leach-carbon-in-pulp (CIP) and CIL circuits at Newmont Gold's Mill No. 2, No. 4 and No. 5 are presented for comparison: Mill No. 4 was initially designed for a throughput of 4.1 kt/d (4500 stpd). It is presently operating at an average throughput of 6.6 kt/d (7300 stpd) without significant changes to the original design. The design considerations of the Mill No. 4 CIL circuit are discussed. The design of any processing facility represents a series of decisions. The final design options should ultimately be resolved based on economic as well as technical considerations. Every design decision represents some sort of compromise. From the designer's view point, these decisions are made with the conviction that the final product represents the most prudent long-term expenditure. The design throughput rate is a decision reflecting a maximum return for the capital and operating costs that the plant will incur over the life of the mineral resource. The result is a plant that is not operator proof or maintenance free because it would be too costly to build. At some point, the incremental increase in the level of capital expenditure would not adequately compensate for the lower

Jan 1, 1991

By A. T. Schwennesen

Every engineer has occasion to trace or copy a map, plan, or other drawing on paper too thick for the ordinary way of using tracing-cloth or tracing-paper. When the figure is small and simple a copy may be made by holding the original against a window-pane, coverin& it with the paper, and tracing direct by the aid of the strong sunlight from outside. The need of utilizing this principle on a larger scale and in a more convenient position led Dr. J. C. Branner to plan the table of which the following is a description : This table was first made in the form of an adjustable glass-top table with a mirror beneath, in 1887, while Dr. Branner was State Geologist of Arkansas. Later it was modified as experience suggested until the form as here described was evolved. The device consists essentially of a drafting-table with a plate-glass top, upon which the original drawing and the paper are laid, and a mirror mounted underneath to reflect the light of the sky up through the drawing. The glass top is hinged and fitted with two arms and thumb-screws, so that it can be raised and fixed to any position, either inclined or horizontal. The mirror is pivoted and revolves about a horizontal axis, so that it may be tilted to any angle. The hood of cardboard or black cloth prevents the reflection of light from the tracing, and may or may not be attached to the table.

Jan 1, 1912

By Oscar Lachmund

DURING the driving of the main haulage level at the Copper Mountain mines of the Canada Copper Corpn., Ltd., near Princeton, B. C., some very rapid driving was clone, though no claim for a world's record is made. Conditions, however, were unfavorable for economical operations. The cost of power was high, for the fuel was of poor grade; besides, during the time the work was in progress, very little other power was needed so that most of the power cost was charged against the footage. The FIG. 1.-DRIFT ROUND USED IN MAIN HAULAGE LEVEL, COPPER MOUNTAIN, B. C. transmission line consisted of No. 4 galvanized iron wire with the result that the line loss was considerable. The voltage transmitted was about 30,000. The plant was operated under a lease, which was due to expire about the same time this work was supposed to be completed; an extension was refused, therefore speed was most important. The plans called for a straight adit 2900 ft. (884 m.) in length. At a point 2800 ft. (853 m.) from the portal, two raises were to be put up to the next nearest workings, a difference in elevation of about 800 ft. (243 m.). One of these was to be a two-compartment hoistway and the other a zigzag ore pass, or muck run. A location for these raises had been determined by a number of diamond-drill holes, but the material to be penetrated by the alit was not known. It seemed imperative to get the tunnel work completed as rapidly as possible, in

Jan 3, 1919

GENERAL GEOLOGIC DESCRIPTION The Escalante silver deposit located in southwestern Utah is an epithermal vein system consisting of a hanging wall and footwall vein, straight along strike and dipping about 1.2 rad (70°). The veins consist of banded quartz and calcite fissure fillings between volcanic wall rock that are silicified for some distance outward from the vein. The walls of the veins are sharp and the silver values are irratically distributed through the various bands within the vein. Only very low grade values are to be found in the wall rock on either side. The portion of the veins above the water table have been somewhat leached, with some secondary enrichment below. The mineralogical zonation is variable due to a fluctuating water table through geologic time. SAMPLE MANAGEMENT AND COMPOSITE ASSAYS Sample data for the deposit was obtained from over 50 diamond drill holes, from channel samples cut at intervals across a drift, and from scattered surface trenches along the out- crop of the hanging wall vein. Fig. 13-1 is a surface plan map showing the vein outcrop and some diamond drill locations. Two cross sections through the vein are shown in Figs. 1 3-2 and 13-3. Nearly all of the diamond drill core of vein material was assayed in 0.3-m (I-ft) lengths, from which weighted composite assays were manually calculated for the horizontal thickness of the vein. The centers of the composited samples were plotted in north and east coordinates and by elevation, and locations and assay

Jan 1, 1980