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By M. H. MacLeod

"IntroductionTHE ORE PASS raise system at the H.B. mine as contracted by the Patrick Harrison Company consisted of two raises 6 ft. by 10 ft. in section •at 60 degrees to the horizontal. No. 88-88 raise, the longer of the two, was a straight raise 690 feet long, with access only from the bottom. In addition, this raise had two branch knuckle-backs only one of which was driven by. the Patrick Harrison Company. The other, known as 88-89 raise, was 540 feet long with knuckle-backs at 280 feet and 410 feet. Its access, also, was from the bottom (see Figure 1)."

Jan 1, 1954

By Edward H. Peplow

Mr. Chairman; honored guests; ladies and gentlemen: The subject assigned me today by your committee, according to the program, is "the mineral industry and the public land law I have preferred to title the talk, "you can't mine here", for that is the essence of the message we, as an industry, have been getting from so many agencies of the federal government for so long. And, unfortunately, we can expect to hear the same message even louder and clearer -- with increasing frequency in future, unless. Every segment of the industry devotes the necessary time, energy and creative thought to combat successfully the forces that are writing and broadcasting that message. I said that agencies of the federal government are sending the message --- and this, of course, is true, the forest service, for example, and some bureaus of the department of the interior as well as others have appeared in recent years to be abiding by a policy designed to block mineral entry to as much of the public lands as possible and to wipe out existing mining claims.

Jan 1, 1967

By T. A. O’Hara

SINCE May 1948, when tungsten carbide bits were introduced at the Flin Flon mine, they have been popular with the miners because of their fast drilling speed and low gage loss. The high cost of commercial carbide bits and tipped drill steel, however, prevented their use except for the hardest rock. In an effort to extend the use of tungsten carbide on a basis economically competitive with detachable steel bits, experimental work was begun in 1950 to test the feasibility of making tungsten carbide tipped drill steel in the mine drill steel shop. This work showed that tipped drill steel could be made locally at less than half the cost of the commercial product. The performance of the local tipped drill steel was comparable to that obtained with commercial carbide bits and tipped drill steel and the cost per foot drilled was much lower. Local tipped drill steel was adopted for all mine drilling in November 1951. Since then drilling costs per foot have been sharply reduced and footage drilled per manshift has increased markedly. Experience at Flin Flon has shown that production of satisfactory carbide tipped drill steel is not difficult and that highly skilled labor and costly equipment are not required. As long as wise selection of brazing materials is made and certain simple precautions are rigidly maintained, there is no reason why small mines with relatively unskilled labor cannot produce a satisfactory product. The following description outlines the technique used at Flin Flon for making carbide tipped drill steel and discusses characteristics of the brazing process that make special precautions necessary. Drill steel is forged to four-wing shape in a conventional steel sharpening forge. Standard steel dies are modified to minimize forging cracks around the central waterhole and to forge a blunt bithead on the steel. The steel is preheated to 1500°F and held at this temperature for at least 2 min. When the temperature has equalized throughout the steel section, the drill steel is transferred to the forging furnace and heated rapidly with a reducing flame up to 2000°F. This two-stage method of heating minimizes the grain growth and decarburization of the steel while ensuring that the steel temperature does not vary greatly throughout the forging zone. After forging the steel is allowed to cool in air to about 1600°F before being annealed in a bath of vermiculite. Despite the high hardenability of the 3 pct Ni-Cr-Mo drill steel used, this simple treatment anneals the drill steel sufficiently for milling. The forged and annealed drill steel is slotted on a plain horizontal milling machine that is equipped with a quick opening chuck and a slot depth stop. The full depth of the slot is milled in a single pass of the 3-in. milling cutter which is fed at 33/4 in. per min across the crown of each bit wing. The slots are cut to a width of 0.342 to 0.344 in. Maintenance of this slot width is necessary to ensure that the optimum brazing clearance of 0.002 in. will result after assembling of shims and carbide in the slot. Prior to March 1953, when the milling machine was installed, drill steel was slotted on a small manually fed ¾ hp milling attachment mounted on the bed of a lathe. Over 16,000 drill steels were slotted on this unit, and in view of its small size and low cost it gave excellent service. Brazing of Tipped Steel Drill steel that has been milled and cleaned in carbon tetrachloride is mounted in a rotating cradle holding six drill steels, the length of which may be from 2 to 12 ft. The slots in the drill steel, the shims, and the tungsten carbide inserts are thoroughly fluxed with a fluoride flux and assembled as shown in Fig. 1. Fig. 2 shows the brazing equipment in use. As the ring burner is lowered over the bithead a spring valve opens the gas lines, and the gas mixture, preset to give a slightly reducing flame, is fed to the ring burner where it is lit from a pilot flame. The ring burner heats the drill steel over a zone about 1 to 2 in. below the bithead, which becomes heated by conduction through the steel. By this means the bithead is heated rapidly and evenly, and contamination of the brazing joint with soot from the flame is avoided. The bithead is heated to the melting temperature of the brazing alloy within 1 min. This rapid heating minimizes the disadvantage of a non-eutectic brazing alloy. The brazing alloy, a nickel-bearing quaternary alloy, is placed at the bottom of the slot below the carbide insert, as shown in Fig. 1. As the brazing alloy melts it is drawn by displacement by the carbide and by capillary action into all parts of the joint to displace liquid flux from metal surfaces. As soon as the brazing alloy melts, each insert in turn is wiped by being moved back and forth along the slot. This action assists wetting of the carbide by the brazing alloy and assists in displacing molten flux from the joint. After continuous heating for about 75 sec, when the bithead has reached a temperature of about 1500°F, the ring burner is raised and the gas supply is shut off automatically by the spring valve. As soon as heating is stopped a hand press is placed on the bithead and the inserts are squeezed down firmly. This action minimizes the clearance between the bottom of the insert and the slot. Correctly brazed steel should maintain a clearance at the bottom of the slot of 0.001 to 0.002 in. After six steels have been brazed they are removed from the cradle and allowed to cool in air. As soon as each drill steel is cool it is dressed on a grinding wheel to remove excess flux and braze and is ground to the gage appropriate to the length of the drill steel.

Jan 1, 1955

By P. G. Claridge, G. P. Holmes, M. A. Redfearn

"IntroductionThe purpose of this chapter is to describe current practices in operation and maintenance of flotation circuits in Canadian mineral processing plants. The chapter consists of five sections covering description of equipment, typical flow diagrams, operating practice, instrumentation and maintenance. The approach to this subject matter is intended to be practical in nature, hence no deeply technical treatment of flotation theory will be included.Flotation is an extremely common method of beneficiation of ores. Reasons are the relatively low cost of flotation compared to many other processes, the wide range of minerals to which flotation may be applied and the high efficiency of separation which can usually be achieved. The flotation process as we know it today has been developed within this century and it has had a major impact on the mining industry. While flotation is commonly considered to be important because it provides a relatively simple method of separating minerals in complex ores, the availability of flotation has also made a multitude of mineral deposits in Canada into economically viable mining operations because it allowed production of a concentrate of valuable minerals from the ore. These concentrates could then be transported from geographically remote locations to centraHzed plants for further processing, although the existence of the remote and perhaps small ore body alone would not have justified installation of metallurgical processing plants on site and the grade of ore would not have justified direct shipment of ore.Some of the more common minerals concentrated by flotation are sulphides of copper, nickel, cobalt, iron, zinc, molybdenum, lead and arsenic, as well as non-sulphides such as sodium chloride, potassium chloride, phosphates, silicates, chromite, fluorite, limestone and coal. Precious metals are also recovered by flotation, generally in association with another sulphide mineral."

Jan 1, 1989

By C. O. Stee

Introduction The property of Siscoe Gold Mines, Limited, is located on an island of three hundred and forty acres in lake Kienawisik, about fifty-two miles east of Noranda and forty-three miles south of Amos, Quebec. Lake Kienawisik is a part of the Harricanaw drainage system and is crossed by the main line of the Canadian National railway at Amos. This waterway is navigable for boats with a draught of six to eight feet. Diesel engines have been the source of power at the property since milling operations were commenced in January, 1929. The power plant as first designed and built was for a milling capacity of 100 tons. Two Polar-Atlas units, each of 200 h. p., comprised the original installation. With the growth of the 'mine, more engines were added until, at the present time, there is installed at the Siscoe what is rated the largest Diesel plant at any mining property in Canada. Definition of a Diesel 'The Diesel engine is an internal combustion engine which works with liquid fuel introduced after, and independently of, the air charge at a time when communication between the cylinder and the outside atmosphere is cut off, and which ignites the fuel by compression-beat and stored combustion-heat unassisted by heat sources outside the cylinder itself while the engine is in stable operating condition"(1).

Jan 1, 1934

By G R. Ballantyne, A N. Mainza, M S. Powell, M Evertsson

Quantifying the performance of a comminution circuit relative to the potential that can be achieved is a standard requirement in justifying operating practice (the stick) and seeking operational upside (the carrot). Historic measures tend to focus on performance relative to laboratory design tests that provide a direct measure of specific grinding energy, such as the Bond Work index; or use dropweight breakage tests calibrated to a substantial database of operating sites, such as the SMC methodology. Although of value, such approaches are flawed in significant aspects. The assumption is that the laboratory prediction is correct and the single sample it is based on is representative. The database benchmarking methods are, by definition, locked into current average operating practice. These approaches do not illuminate upside; can ‘punish’ operations that are actually more efficient, such as grinding finer than standard operation; and reward poor operation that may comfortably produce a coarser grind than optimal for recovery or excess slimes that reduce recovery and increase water consumption and tailings volume. The authors present a set of established and easily applied measures, that draw together daily operation, highlight areas of opportunity and inefficiency, use simple samples and data, and can be used to provide the ‘size of the prize’ in pursuing specific areas of process improvement. These tools include size specific energy (SSE), overall operating effectiveness (OEE) which includes product quality, theory of constraints, and links this to the end objective – maximising the efficient production of valuable metals. Feeds into objective function driven by economics – financial return, energy, environment.

Aug 24, 2022

By J. E. Nesset

Significant progress has been made over the past few years on characterizing and understanding the hydrodynamic properties of frothers and collectors and their effect on flotation performance. This new understanding of the relationship between chemistry and hydrodynamics has dramatic implications for how flotation circuits should be controlled and optimized. Previous work has shown that choosing a frother of the right 'strength' is critical to the optimization process, and the work and discussion presented here serve to strengthen the argument. In addition, we know that collector chemistry may also influence hydrodynamic properties and needs to be taken into consideration. To achieve the desired outcome, the overall system of reagents and operating parameters must be considered. This paper describes and discusses some of the findings from recent research on frother and collector chemistry effects (on hydrodynamics), and how this knowledge can be used to improve flotation performance. Examples from plant experience are discussed.

Jan 1, 2009

By A. S. MacNeil

In the operation of any colliery, a factor of prime importance is power, of which there are three sources available--man, horse, and mechanical. It is obvious that man-power is by far the most expensive, .and that animal-power, also, is more costly than mechanical. In recognition of this fact, mine owners must, for economical ?reasons, reduce the use of man-and horse-power to the minimum by the increased use of machinery. In other words, they must aim at as complete mechanization as conditions will permit, taking advantage of existing physical factors that are favourable and making the necessary adjustments and modifications to meet the less favourable conditions. Intensive coal-mining, or the concentrated production of coal, has two distinct objectives: First, the maximum production from a limited area, in a given time, thereby reducing the pit-room to be maintained; and second, reduction in cost. To this end the coal should be mined, loaded and delivered to the surface by mechanical means wherever possible. To do this, however, requires the solution of many problems arising out of the varying conditions that obtain underground. The system of longwall mining lends itself to a more efficient scheme of mechanization than does pillar-and-room, as, in the former, the first step has already been taken by the removal of all the coal in one operation, and the opening-out of one face, the equivalent in mining area of fifteen or more rooms. Much has been clone of late towards mechanization in the longwall mines of the Sydney coal-field, and with gratifying results, but not without great difficulties and frequent disappointments. In 1923, an attempt was made with shaker conveyors. The introduction of conveying was a step to bridge the gap between the face, where the coal is produced, and the main transportation arteries leading to the slope or shaft. The receiving of the coal at the faces and its delivery to main loading points on main haulage-ways was the part allotted to the conveyors.

Jan 1, 1932

By Leland H. Johnson, D. O. Kennedy, John Hyvarenen

"SUMMARYA chronological list of accidents at metal and nonmetal mines and quarries (excluding coal mines) in the United States in which five or more lives were lost has been compiled. Many accidents were described in more detail because of their unusual features. Tables have been included to indicate the number of major disasters and fatalities by the principal product mined, by State, and by 10-year intervals prior to 1945. The number of men employed to indicate man-hours exposed is unavailable; therefore, a true picture of disaster data by principal product mined, by State, and by 10-year intervals is not possible. However, the data are presented to indicate the causes of major accidents and bring to the attention of those affiliated with metal and nonmetal mines and quarries the hazards involved and lessons that can be practiced to keep such accidents at a minimum.INTRODUCTIONThe Bureau of Mines was requested to compile a chronological list of accidents at metal and nonmetal mines and quarries (excluding coal mines) in the United States in which five or more lives were lost. Data have been assembled on accidents from all causes in these mines and quarries from the earliest available information (describ¬ing an underground mine fire in Nevada in 1869) to the last reported major accident (a fire at a company-owned bunkhouse in Arizona in 1945).The accounts of the accidents were taken from articles in the technical press and from published and unpublished reports of the Bureau of Mines. Many of the accidents have been selected for more detailed descriptions because of the unique circumstances and to show the variety of dangers involved.Tables are included to indicate the number of major accidents, with resultant fatalities, by the principal metal and nonmetal producing groups, by State, and by 10-year intervals previous to 1945. The interval 1869-1875 does not cover a 10-year span because of incomplete records previous to 1869.ACKNOWLEDGMENTSThe information contained in this report was obtained from the Engineering and Mining Journal from 1874 to 1947, the Mining and Scientific Press from 1908 through 1922, from various published and unpublished reports of the Bureau of Mines and from reports by State departments of mines. Daniel Harrington, retired chief of the Health and Safety Division; E. H. Denny, supervising engineer, District H; R. M. Loucks, mining engineer; and other engineers of the Division contributed much valuable information and assisted materially in the study."

Apr 1, 1949

By L. Yuan, A. Smith

"NIOSH’s mine fire simulation program, MFIRE, is widely accepted as a standard for assessing and predicting the impact of a fire on the mine ventilation system and the spread of fire contaminants in coal and metal/nonmetal mines, which has been used by U.S. and international companies to simulate fires for planning and response purposes. MFIRE is a dynamic, transient-state, mine ventilation network simulation program that performs normal planning calculations. It can also be used to analyze ventilation networks under thermal and mechanical influence such as changes in ventilation parameters, external influences such as changes in temperature, and internal influences such as a fire. The program output can be used to analyze the effects of these influences on the ventilation system. Since its original development by Michigan Technological University for the Bureau of Mines in the 1970s, several updates have been released over the years. In 2012, NIOSH completed a major redesign and restructuring of the program with the release of MFIRE 3.0. MFIRE’s outdated FORTRAN programming language was replaced with an object-oriented C++ language and packaged into a dynamic link library (DLL). However, the MFIRE 3.0 release made no attempt to change or improve the fire modeling algorithms inherited from its previous version, MFIRE 2.20. This paper reports on improvements that have been made to the fire modeling capabilities of MFIRE 3.0 since its release. These improvements include the addition of fire source models of the t-squared fire and heat release rate curve data file, the addition of a moving fire source for conveyor belt fire simulations, improvement of the fire location algorithm, and the identification and prediction of smoke rollback phenomena. All the improvements discussed in this paper will be termed as MFIRE 3.1 and released by NIOSH in the near future INTRODUCTION The MFIRE program, originally developed in the 1970s by the U.S. Bureau of Mines (now the Office of Mine Safety and Health Research (OMSHR) at the National Institute for Occupational Safety and Health (NIOSH)) and Michigan Technological University, has been a mainstay in modeling of the mutual influences between a fire and the mine ventilation system during an underground mine fire. Several update versions have been released over the years, including MFIRE 1.27, 1.29, 1.30, 2.0/2.01, 2.10, and 2.20. The MFIRE program logically comprises four parts: (1) a conventional network calculation, where it performs the basic network balancing without considering heat or mass transfer; (2) a temperature calculation to establish the reference temperature distribution before a non-steady state (transient-state) simulation; (3) a transient-state simulation that follows changes in ventilation step-by-step to produce a continuous description of the temperature distribution, smoke, and contaminant spread through the ventilation system during a fire event; and (4) a quasi-equilibrium simulation to predict the state of the ventilation system after a relatively long period of time (defaulted as 5 hours in MFIRE) as the fire reaches a quasi-steady state. In summary, MFIRE is a computer simulation program that performs normal ventilation network planning calculations and dynamic transient-state simulation of ventilation networks under a variety of conditions including the influence of natural ventilation, fans, fires, or any combination of these (Chang et al., 1990)."

Jan 1, 2016

By F. E. ROGERS

CUSHMAN TUNNEL No. 2 is adjacent to the Hood Canal, near potlatch, Wash. It is 17 ft. inside .diameter, about 13,000 ft., or two and one- half, miles in length, and is a part of the second unit of the Cushman power project. It connects Cushman ' Dam No. 2 and the penstocks that lead to the power house on the beach about 400 ft. vertically below the outlet of the tunnel. One shaft, 190 ft. deep, and two adits between 300 and 400 ft. long, were provided to speed up driving operations. his arrangement gave 6 working headings, which were operated 24 hr. daily. The shaft provided a manway, cageway, and skipway.

Jan 1, 1931

By P. G. Clites

The Bureau of Mines developed techniques for induction melting of titanium in a split, water-cooled copper crucible. Calcium fluoride, which was used as an inert slag cover during melting, formed an insulating layer of solid slag between the ingot and the crucible wall. This thin layer of slag effectively insulated the ingot from the crucible and prevented the molten metal from shorting across the slits between adjacent segments of the crucible. The technique provides a means of melting loose titanium sponge and scrap and eliminates the need for fabricating a consumable electrode. This report describes the equipment developed and compares ingots produced from two varieties of titanium sponge, from titanium scrap, and from previously melted titanium with ingots prepared by vacuum-arc melting and electroslag melting. A melting scheme was also proposed and evaluated whereby loose titanium sponge was consolidated into a first-melt ingot for vacuum-arc remelting.

Jan 1, 1969

By C. B. Solnordal, R. N. Taylor

"The heat transfer conditions on the outer side of a Sirosmelt lance have been determined using experimental measurements in combination with a mathematical model. By measuring the temperature distribution along the wall of an operating Sirosmelt lance, as well as the thickness of the slag layer on it, it was possible to use the model to determine both the heat transfer coefficient between the vessel contents and the lance, as well as the effective thermal conductivity of the slag layer. The effective thermal conductivity of the slag layer was found to be within the range 0.7-1.3 W.m-1. ºCl, which was in good agreement with measured values for magnetite. The outside heat transfer coefficient varied from 80 to ISO W.m-2. ºC1, which was smaller than quoted in the literature for metal/slag systems. The discrepancy was attributed to the high thermal resistance experienced between the lance wall and the slag layer, as well as the large quantities of combustion gases that envelop the lance tip and reduce convection and conduction from the melt to the lance. INTRODUCTIONThe Sirosmelt submerged combustion smelting process [1] utilizes top injection lances to deliver fuel and air into a metallurgical melt. To help prevent lance failure from attack by liquid slag, helical vane swirlers are used within the annular lance to impart swirl to the flowing air and enhance the heat transfer from the lance wall to the air. The swirlers thus cool the wall, and enable a frozen layer of slag to form on its outer surface. This solid layer ensures liquid slag cannot reach the lance surface, and the lance is protected.A mathematical model of the heat transfer processes associated with the lance has been developed, which combines heat balance equations with engineering correlations to predict the wall and cooling air temperature of the lance, as well as the air pressure distribution. An account of the model was given by Rankin et al. [2]. To further enhance the model and enable it to predict the thickness of slag accumulating on the lance, a detailed analysis of the heat flow through the slag layer commenced in 1992. By far the most complex aspect of the model, the flow of heat from the vessel contents to the outside of the lance was dependent on many factors, including:-"

Jan 1, 1996

By Scott W. Hazen

METHODS of determining the grade and tonnage in a mineral deposit have not kept pace with scientific developments in certain other fields. The most commonly used methods of computing ore reserves and grade based on diamond-core-drill sampling are the cross section, the polygon, and the triangle. Tonnage and grade may be determined by these methods with acceptable accuracy, providing a sufficient number of samples is obtained; however, there is no means of determining when an adequate number has been taken or conversely if the deposit has been sampled excessively. It is demonstrated that, by using the techniques of statistical analysis, a mining engineer will be able to determine when a sufficient number of samples has been taken to determine the grade of a mineral deposit within a specified degree of confidence. Using statistical analysis to compute the estimated average grade of the ore makes it possible to compute a fiducial interval, when a distribution of the sample means is normally distributed, so that there is 95-percent confidence in the statement that the true average grade of the deposit is covered by the interval. This report utilizes data obtained from the Maggie Canyon manganese deposit, Artillery Mountains region, Mohave County, Ariz. Four separate exploration projects were conducted at Maggie Canyon between 1936 and 1950. Present economic conditions have made it desirable to determine the amount of 5-percent manganese ore developed by these projects.

Jan 1, 1957

By Mikael Ramström, Mark Board, Clayr Alexander, Wes Johnson

"The Coeur d’Alene mining district in northern Idaho was first discovered more than 130 years ago and has been the most prolific silver producer in the United States. The Lucky Friday Mine of Hecla Mining Co. (Fig. 1) is the oldest (76 years) and deepest mine in the district, producing approximately 725 t/day (800 stpd) from narrow, vertical lead-zinc-silver veins. Cut-and-fill stoping is currently employed at depths to about 2.3 km (7,500 ft) below ground surface, making it one of the three deepest mines in the western hemisphere. The #4 Shaft, a winze collared at approximately 1.8 km (5,900 ft) below surface and extending to a depth of nearly 2.9 km (9,600 ft), was recently completed to exploit the high-grade 30 Vein, whose resource has been identified to at least 3 km (9,700 ft) in depth.As with other deep mines worldwide, a number of challenging engineering issues occur, including: high in situ rock stress, squeezing ground, seismicity, high in situ rock temperature that requires refrigeration, diesel equipment that generates diesel particulate matter (DPM) requiring high ventilation airflow rates and long travel times for worker transport to mining areas.This article describes the methods that Hecla Mining Co. has employed during the past 50 years to combat these issues, with emphasis on the current efforts to implement continuous, mechanical mining with automated or teleremote operation and to introduce electric and battery-powered equipment to minimize refrigeration and ventilation requirements for the removal of DPMs.Geology and in situ stressThe host rock mass is an extremely thick sedimentary sequence (the Belt Series) that has been tilted vertically by mountain building. The rock mass in the current section of the mine consists of weak, heavily foliated argillites and stronger, bedded silty-argillites termed siltite. The vertical veins (80° to 85° dip) are parallel to the foliation with an approximate width of around 3 m (10 ft). There are two regional, vertical faults that parallel the vein; one forming the hangingwall of the orebody and one approximately 90 m (300 ft) in the footwall. Additionally, there are several thin, rough northdipping faults that occur in the footwall of the vein which terminate on the major vertical faults.Both the argillite and vein package rock have moderate strength, with uniaxial compressive strength averaging about 100 MPa (14,700 psi with a range from about 70 to 125 MPa), and are moderately abrasive, with CERCHAR abrasivity index of 2 to 3. The major in situ stress is horizontal and oriented in a northwest direction with magnitude about 1.5 times the vertical stress. Consequently, the Lucky Friday stress magnitude at depth is analogous to significantly deeper South African gold operations."

Jan 12, 2017

By AIME AIME

The following changes of address of members have been received at the Secretary's office during the period of Jan. 1 to Mar. 15, 1907. This list therefore supplements the annual list of members corrected to Jan. 1, 1907, and brings it up to the date of Mar. 15, 1907. The names of members who have accepted election during January and February, 1907 (new members), are printed in italics. The large number of changes of address since Jan. 1, 1907, shows the importance of publishing these changes as frequently as possible, and the Bi-Monthly Bulletin has been selected as the means to present this information to the members of the Institute. By a simple method of cutting out these names and addresses and pasting them directly over the corresponding names in the annual list of members, the record can be kept practically up to date, and the value of the list correspondingly increased. For this purpose the changes of address have been printed only on one side of the page. The names of new members, being in italics, are readily distinguished from the others, and can be pasted in approximate alphabetical order on the margins of the pages.

Mar 1, 1907

By Packer M. L

Residual load development in driven precast piles has been widely researched in recent decades, however, cast-in-situ piles are often falsely assumed to be unaffected by this mechanism. In contrast to driven precast piles, driven cast-in-situ (DCIS) piles require adequate time prior to loading for the concrete hydration and curing processes to occur. The concrete-soil interaction is complex with several processes influencing the strains within the concrete pile after installation. DCIS piles are of particular interest given that residual loads can be developed during both the driving and casting phases. This paper present four instrumented end-bearing DCIS piles in which temperature and strains were monitored from installation, with the objective of identifying the processes contributing to the measured strains prior to testing. Strain gauges at all depths recorded a net compression and compressive strains continued to increase up to the commencement of load testing. Residual strains in the test piles were interpreted to be influenced by temperature (concrete and ambient soil and air temperature), consolidation of soils due to dissipation of excess pore water pressure, and swelling from absorption of groundwater. Several residual load interpretive methods are evaluated with the Siegel and McGillivray (2009) approach being consistent with the skin friction full mobilisation theory of the Unified Design method. The full-scale field tests presented herein provide a unique opportunity to examine DCIS pile behaviour, particularly given the lack of presently available published studies on these pile types.

May 1, 2022

By S. B. Hildebrand

One of the highest costs associated with insitu leaching of copper is electric power. The ReGeneration Project addressed this cost issue with a research and development project for BHP Copper. The project involved the installation of a reverse pump/turbine for the removal of excessive fluid pressure and the production of useful electric power. The motivation for the project was that a reverse pump/turbine unit could not only remove the excess hydraulic pressure, but when connected to an induction .generator, would produce electric power, which could be easily connected to the local power gird, thus reducing operating costs. The project was successful in removing excess fluid pressure and producing electricity. The unit is operating today, producing useful 2,300 volt power at a rate of approximately 128 kilowatts.

Jan 1, 1998

By B. Derin

In this study, the Reddy-Blander model was applied to predict the sulfide capacity (CS) values of FeOx-CaO-SiO2, FeOx-MnOx-SiO2 and FeO-MgO-SiO2 ternary slags, which were experimentally studied in the literature within the temperatures between 1,673 and 1,923 K. As expected, sulfide capacity values of the slags increased with increasing temperature and decreased with increasing acidic content. The model was found to be in agree-ment for all slag systems at the temperature range of 1,773-1,923 K. However, a disagreement occurred when the basic content was relatively high, at 1,673 K for FeOx-CaO-SiO2 and FeO-MnOx-SiO2 slags. The effect on the CS of high amounts of Fe+3 and Mn+3 in the slags is discussed.

Jan 1, 2011

By C. R. Hubbard

The Bureau of Mines investigated the mineral resources of the Upper Missouri River Basin as a part of a comprehensive water and related land resources study by six Bureaus of the Department of the Interior, coordinated by the Department?s Missouri River Basin Field Committee. The mineral resource study covers an area of about 18,000 square miles within the Great Plains physiographic province of central Montana. Exposed formations in the plains are chiefly flat-lying Cretaceous sediments. Plutonic rocks crop out in the central parts of the mountain uplifts, with steeply tilted Paleozoic sediments exposed around the margins. Oil and gas have been produced from several fields in the study area; total value exceeds $45 million. Coal-bearing formations of Upper Jurassic and Cretaceous age cover approximately 45 percent of the study area. Total production of coal since 1880 has been between 40 and 50 million tons, exceeding $75 million in value. Total metal production has been about $52.5 million. Industrial mineral production from 1925 through 1960, is valued at $24 million.

Jan 1, 1964