Search Documents

By Pablo Velasco

After becoming an oil exporter in 1972, Ecuador experienced a remarkable economic upswing until 1978. However, in 1979 and 1980, a combination of only slight improvements in agricultural output and stagnating petroleum output lowered rates of real gross domestic product (GDP) expansion to 5.3% and 4.3%, respectively. In current prices, the GDP in 1980 was estimated at $11.3 billion.2 The slowdown in growth is attributable in large part to the low level of petroleum exploration in the 1972-79 period. Ecuador's production of petroleum declined 4.5% in 1980 compared with that of 1979. The consortium of Corporación Eatetal Petrolera Ecuatoriana (CEPS), the State oil company, and Texaco, Inc., operating in Ecuador's eastern jungle region was responsible for nearly all of 1980 production, The volume of total crude petroleum exports from Ecuador dropped from 44.7 million barrels in 1979 to 40 million barrels in 1980, although crude petroleum export revenues rose from $1,032 million in 1979 to $1,294 million in 1980- The latter figure represented 54% of Ecuador's 1980 merchandise exports. Crude oil exports accounted far about 46% of the value of 1979 merchandise exports. The total nonfuel mineral exports in 1980 were valued at less than $1 million.

Jan 1, 1982

By Nikolaos Petropoulos, Daniel Johansson, Håkan Schunnesson

The optimization of the explosive performance is of great importance since blasting is extensively used in civil projects and mining industry. An improved understanding of how the energy from the detonation of an explosive is transmitted to the surrounding rock mass is critical information for the proper selection of an explosive. Therefore, an extensive study was done on the influence of the blasthole diameter on the Velocity of Detonation (VoD) and the detonation front curvature. In this study, pure emulsion was tested in confined conditions to better simulate the behavior of the detonation of the explosive in rock mass conditions. A methodology is proposed for carrying out the evaluation of the explosive performance. The blasthole diameter of the samples varied from Ø 25 mm (0.98 in) up to Ø 65 mm (2.55 in). The confiner was a magnetic mortar cylindrical sample with mechanical properties similar to a rock mass. A streak camera and VoD probes were used to measure detonation front curvature and VoD. The data was analyzed and correlated with the physical parameters of the explosive. Additional to this, the data was further used to calibrate numerical models in LS-Dyna.

By A. M. Khan, I. Fatt

A new type of steady-state thermal conductivity apparatus is presented and described. Results of measurements on Berea sandstone, Solehofen limestone, Bandera sandstone and Bakelite are presented. The measurements were made in a two-hour period at temperatures up to 80C and axial stress to 125 bars. Results show that the thermoelectric heat pump and tbermistor-controlled electric heater can be used in a steady-state apparatus for measuring thermal conductivity of mck. Steady-state is reached within two hours. With Urethane foam insulation, not more than 3 per cent of heat is lost by radial flaw. The apparatus can be used as a comparator or for absolute thermal conductivity determinations. INTRODUCTION Calculations of heat flow through the earth, both for scientific purposes and engineering design, require thermal-conductivity data on rock. However, the high variability of rock often makes handbook thermal-conductivity data of little value. For engineering purposes in particular, it is desirable to have measured data on samples of the rock mass under study. Laboratory methods for measuring thermal conductivity of rock use either time-dependent1-3 or steady - state temperatures .4,5 The time -dependent (transient) techniques give thermal diffusivity from which thermal conductivity can be calculated if the specific heat and density of the sample is known. Most transient techniques require lengthy and involved sample preparation. Also, the accuracy tends to be low. Steady-state techniques are more exact and are the most commonly used. The steady-state methods, however, also have their difficulties. A long time is often required to reach steady state, troubles have been encountered in establishing good thermal contact between the heating and cooling elements and the rock sample, and unaccounted-for heat losses are difficult to prevent. In this paper we present a description of a new type of steady-state thermal conductivity apparatus. With it, measurements have been made in two hours at temperatures up to 80C and axial stress to 125 bars. Sample preparation is simple and no thermal conductivity standard is needed.6 APPARATUS Our steady-state apparatus is an improvement . over the conduction pile used by Zierfuss and van der vliet4 and Somerton.5 A linear heat flux is established through a conduction pile consisting of the rock sample and a standard material of known thermal conductivity in series. The thermal conductivity of the sample is calculated from the ratio of the temperature drop across the sample

Jan 1, 1966

By J. de Villiers

Through the application of the Rietveld method, X-ray diffraction analysis has developed into a fully quantitative method for phase analysis. The Rietveld method itself has progressed from an unstable method that required constant operator intervention to a stable and robust method suited even for routine analysis. The method, however, requires a good initial calibration and periodic checks on its precision and accuracy. Several examples from minerals processing, pyrometallurgy, and materials science show the strengths and applicability of the method. The limitations and sources of error are also discussed.

Jan 1, 2007

THOSE who dabble in mining are no exception to the general rule that persons who know a little of a subject are apt to think that they know all about it. Some colour may be lent to the belief that anyone may succeed in mining when we hear stories ofsuccessful diggers and others who have accidentally dropped on to something good through no foresight on their part; but we seldom hear the contra account of the far larger number who fail. Those who go more thoroughly into the subject, and take up mining as a profession, know very well that it is not one that can be carried out successfully by rule of thumb, but when they point this out to the public they are credited with trying to make it a closed profession; and outsiders are for ever trying to find out the "trade secrets," at the same time spurning the main one, which is hard study, experience, and the ability to apply their knowledge. The average mining man is not rich. When he amasses a fortune it is generally through speculation rather than professional work, thought no doubt his technical knowledge may have assisted...

Jan 1, 1910

By Gabriel S. Esterhuizen

A survey of pillar conditions was carried out at 21 operating limestone mines that use the room-and-pillar method. The surveyed pillars were all located in rock that was classified as ?Good? to ?Very Good? using the RMR rock mass classification system. The average pillar width was 14.5m and the width to height ratio varied between 0.53 and 3.52. Pillar conditions were rated using a visual rating system that accounts for the effect of geological structures as well as stress related fracturing. Nine cases of pillar instability were recorded. Pillar instabilities were typically limited to individual pillars or small groups of pillars within an otherwise stable layout. Local geological structures such as inclined discontinuities or weak bedding planes contributed to instability in seven of nine cases that were observed. Elevated pillar stresses contributed to instability in the two remaining cases. All the unstable pillars observed had width-to-height ratios of less than 1.5. It is concluded that pillar instability is most likely to be caused by unfavorable geological structures in pillars with width to height ratios of less than 1.5. Stress related instability such as rib spalling becomes more prevalent when the average pillar stress approaches approximately 20 MPa.

Jan 1, 2006

By R R. Large

"An EXTENDED ABSTRACT is available for download. A full-length paper was not prepared for this presentation. The formation of orogenic gold deposits is commonly a two-stage process. The first stage involves preconcentration of gold in a suitable source rock, such as carbonaceous black shales, felsic volcaniclastics or komatiitic volcanics. The second stage involves the release of gold (commonly with arsenic and/or tellurium) from the source rocks during metamorphism or magmatic intrusive events, accompanied by the replacement of source rock pyrite by pyrrhotite. Recent research suggests that synsedimentary gold trapped in the structure of diagenetic arsenian pyrite in reduced carbonaceous facies of continental margin sedimentary basins is an ideal source of gold for many orogenic and Carlin-style gold deposits. The conversion of diagenetic arsenian pyrite to pyrrhotite during basin closure and greenschist facies metamorphism is the key process releasing gold and arsenic to the metamorphic fluid, with metal deposition in structural trap sites to form the deposits.CITATION:Large, R R, 2015. Orogenic gold deposits – a two-stage process of gold enrichment, in Proceedings PACRIM 2015 Congress, pp 335–340 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Mar 18, 2015

By R. Barbaroux, B. Dubreuil, D. Lemieux, K. Nasrallah, G. Garcia Curiel

"As conventional, base or precious metal ores are being depleted worldwide, the mining industry is being faced with an increasing exploitation of arsenical sulfidic ores. Enargite is a common example. The downstream processing of these ores poses several problems ranging from high smelter penalties to finding economical and environmentally acceptable means of disposing of the arsenic. Several hydrometallurgical processes have been proposed for extracting arsenic from enargite concentrates, notably using alkaline sulphide leaching (ASL). However, few alternatives exist for removing the arsenic from solution and stabilizing it in an approach that is both economical and environmentally acceptable.Dundee Sustainable Technologies (“DST”) has developed a technology for the sequestration of arsenic by vitrification. Originally optimized to cost-effectively stabilize arsenic trioxide (As2O3), DST has shown that vitrification could be successfully extended to the stabilization of other arsenical products, such as calcium arsenate (Ca3(AsO4)2) and sodium arsenate (Na3AsO4). In this paper, the vitrification of arsenic in the context of an enargite leaching was newly explored and demonstrated successfully.Two processing routes are being proposed by DST for the removal of arsenic from enargite concentrates and their stabilization by vitrification. In one route, arsenic is extracted by ASL and is then precipitated under atmospheric conditions and with minimal use of reagents and time. The resulting precipitate is an unstable hydrated ferric arsenate (FeAsO4•2H2O) that is then vitrified in a non-dilutive method, significantly increasing the stability of arsenic and meeting the TCLP guideline. Unlike is required for scorodite, stabilization was successfully achieved without concern for the crystallinity of the precipitate."

Jan 1, 2017

By Onur Kacar

A numerical approach has been implemented to assess the settlements of pile groups bearing in clay. A three-dimensional finite element program has been utilized to model a vertically loaded spread footing, a single pile and a pile group. The soil-pile interaction was modeled with elastoplastic soil and interface elements. The model has been calibrated using a full scale load test results of the spread footing. The load transfer between the pile cap and the piles was assessed and compared with the measured load transfer. The predicted settlements for the pile group were compared with the settlements measured in the load test. Two analytical methods were adopted to predict the pile group settlements and they were compared with the measured values. The group effect determined from the analysis and load tests was compared with the interaction factor approach. It was shown that three-dimensional finite element approach can be successfully utilized to assess the soil-structure interaction and hence the load transfer between the piles and pile cap. Also, it was shown that the interaction factors method and equivalent footing method work well in the working stress range, however they underpredict the settlements at larger loads. INTRODUCTION The settlement and stress distribution of pile groups have been traditionally estimated based on load test results of single piles and accounting for the group effect using various approaches. The most widely used approaches are group interaction factor method, equivalent footing method and equivalent pier method. These approximate solutions are used very commonly in practice as full-scale load tests on pile groups are very expensive and time consuming. These approximate solutions are also adopted in design guidelines such as AASHTO LRFD Bridge Design Guidelines (2014) and FHWA Design and Construction of Driven Pile Foundations (2016). In this paper, a full-scale foundation load tests performed by Nagai et al. (1961) and reported by Koizumi and Ito (1967) is modelled using a three-dimensional numerical method. The settlements and the stress distribution on foundations were then compared with the traditional equivalent footing method and interaction factor methods.

Jan 1, 2018

The final criterion of the payability of gold production must necessarily be the cost per ounce. While the cost per ton or per cubic yard is a convenient comparative measure, and a necessary one for operating control, it is obviously insufficient as a means of judging whether or not, in a given set of circumstances, a gold production operation can be carried out economically.. An alternative method, which brings in the ton or cubic yard as a unit of measurement, is to determine what the cost per ton or per cubic yard is equivalent to, in terms of pennyweights or grains at the ruling price of gold. But this is really only another way of stating that the cost per unit of product is the ultimate measure of profit or loss, and not the cost per unit of input material.It is accordingly of interest to determine the cost per ounce of Australian gold production and to compare it with the cost of gold production in some other parts of the world which are important gold producers. The attached Appendix, which it, is believed comprises practically all Australian mines of any consequence (except Big Bell and the de Bernales group, for which. insufficient information is available), shows this information for financial years ending between 31st December, 1938, and 31st December, 1939, the production of gold dealt with being some 1,084,000 oz., or 66% of the total Australian production for 1939. For comparative purposes the cost of producing gold at representative mines in New Guinea, New Zealand, Fiji, Canada, India, West Africa and Rhodesia, and the average cost of production at the Rand mines, has also been included.It is obviously necessary in a study of this sort that the basis of the cost should be clearly set out and should be as nearly as possible uniform. The "cost" as calculated herein comprises the following items:

Jan 1, 1940

In a previous paper of this series (1) ,t it was shown that during the creep of single crystals of lead, deformation proceeded to some extent by means of slip on the octahedral plane. This was demonstrated by the formation of slip lines on the surface of the specimens. At first, a few broad bands appeared after a small amount of extension had taken place; then, as extension proceeded, more and finer lines made their appearance until the whole of the specimen surface was covered.The present paper deals with the process of deformation of the creep specimens as revealed by X-ray examination..By the use of this method, the changes taking place in the crystal structure can be observed.EXPERIMENTAL METHODThe method of preparation of the specimens, their size and the purity of the lead employed have been reported elsewhere (1). At frequent intervals, the specimens were dismantled from the test rack and X-ray back reflection photographs were taken. The specimens were so marked and set up on the X-ray machine that the same spot on the specimen surface was always examined.Copper radiation was employed, the tube voltage and current being 45 kV. and 25 m.a. respectively and the exposure time 2 hrs. The slit diameter was 1 mm.

Jan 1, 1941

By Dennis R. Dolinar

Controlling coal mine roofs subjected to high-horizontal stress conditions has always been difficult and uncertain. Traditional supports such as wooden cribs and posts, concrete donut cribs, and standing supports collapse and fail when the roof and floor move horizontally as mining progresses. The former U.S. Bureau of Mines (USBM) (currently the U.S. Department of Energy (DOE)), in cooperation with Western Fuels-Utah, Incorporated, conducted research to provide an alternative for traditional secondary support systems in a 3-entry gate road system subjected to high horizontal movements. The support system used in several other coal mine operations, consisted of internal high-strength galvanized resin-grouted cable supports. The system virtually eliminates the necessity for external crib, timber, or concrete supports. The support system consisted of 2.4 m (8 ft) full-column resin grouted bolts and 4.8 m (16 ft) long cable supports installed in conjunction with wire mesh and "Monster-Mats." Cable loading and roof deformations were monitored to evaluate the behavior of the immediate and main roofs during first and second panel extractions. Additionally, cable trusses were installed on the longwall headgate to protect the coal conveyance system from roof and pillar falls created by the formation of cutters and gutters. The test results indicated that the designed support system successfully maintained the roof during the extraction of two longwall panels and dramatically reduced the cost of secondary support. This paper describes the theory of high-horizontal roof movements, the advantages of vertical cable supports and cable trusses, and presents the roof and cable measurements made to assess the support performance during longwall retreat mining.

Jan 1, 1996

By Christopher J. Bise

INTRODUCTION Traditionally, the purpose of mine ventilation has been defined as diluting, rendering harmless, and sweeping away hazardous gases which may accumulate in the mine workings. In addition, the ventilation sys- tem must keep the mine air as close as possible to the composition of the outside air in order to provide a healthy environment for mineworkers and to prolong the useful life of mine equipment. When plans are being developed for opening a new mine, sound engineering dictates that a calculated estimate be made of both the maximum air quantity requirement and the accept- able mine pressure (head) loss related to the under- ground layout and projections. For efficient mine ventilation, the two problem areas that demand constant attention are (1) the control of fugitive air (through short circuiting or leakage) and (2) the provision of ample airway capacity. Fugitive air is that quantity of air which enters the mine but short-circuits the mine ventilation system through improperly constructed stoppings, doors, overcasts, etc.; it represents an investment in power consumption (to move the air by the fan) while providing little, if any, return on the investment. Typically, efficient mines are those that can deliver 50% of the air handled by their fans to the working faces. Paying close attention to a new mine's ventilation system at the outset can provide numerous benefits during the later stages of mine development. It has often been noted that the main cause of trouble in many mine ventilation systems can be found within a radius of 2500 ft from the fan installation. Ample airway capacity is essential to maintain the mine's head loss within manageable limits. Similar to an industrial ventilation setting, the mine's headings are the ductwork for the handling of air. The difference between normal industrial ventilation and mine ventilation is apparent when more capacity is needed; a plant can install new or larger ductwork, but it is difficult for a mine to install new entries or enlarge existing entries. Air velocities are usually maintained under 800 fpm to limit the stirring up of dust in the intakes. With this in mind, it is usually simple to estimate the number of entries required in the ventilation system. The problem that often arises, though, is when the system's capacity is overestimated and the required head loss to achieve the desired quantity is excessive. Experience has shown that problems are often encountered when the mine's condition requires much more than three inches of water gage to ventilate the mine on a day-to-day basis. Thus, if proper attention is not paid to these parameters in the preplanning and initial development of a new mine, it may take an extremely high water gage to achieve the ultimate quantity. This chapter will not delve deeply into the theory of mine ventilation; many other books have been writ- ten on that subject. Instead, the following sections will point out the important considerations that should take place before the development of a mine ventilation system, as they relate to normal industry practice. FUNDAMENTALS OF AIRFLOW Volume Versus Water Gage The common method of measuring ventilation pressure is in terms of inches of water gage (wg); one inch of water corresponds to a pressure of 5.2 psf (pounds per square foot). Further, airflow follows a square-law relationship between volumes and pressures where, as shown in the previous section, a doubling of quantity requires a quadrupling of pressure. Table 1 shows how heads and quantities are related. Using Table 1, it is readily apparent that, for the same mining conditions, an increase in pressure from one inch wg to three inches wg results in only 34,640 cfm when the initial quantity is 20,000 cfm. Pressure Losses Pressure losses in mine ventilation systems are caused by friction between the heading walls and the air stream and by shock losses due to abrupt changes of the air stream caused by obstructions. The common pressure-loss formula for mine ventilation is: [ ] where H is the pressure loss in inches of water gage, K is the coefficient of friction, L is the length of the airway in feet, 0 is the airway perimeter in feet, V is the air velocity in feet per minute, and A is the cross-sectional area of the airway in square feet. Table 2 lists the US Bureau of Mines schedule of friction factors (K) for mine airways. Recent research indicates that the values of K are slightly less for coal mines than those presented in the USBM Schedule, and Table 3 should be consulted in this instance. It should be

Jan 1, 1986

By Carlos G. C

Geological mapping in various parts of Australia is revealing a number of regional associations of mineralizations with particular persistent stratigraphic horizons. Such associations have been found in the Broken Hill district of New South Wales, in the Nairne district of South Australia, in the Mount Isa district of Queensland, and in the Cloncurry district of Queensland. The question has arisen as to whether such mineralizations are of sedimentary origin, syngenetic with the enclosing formation, or whether they are of epigenetic origin, and reflect the selective hydrothermal mineralization of favourable beds on a regional scale.The purpose of this investigation is to test the use of sulphur-selenium ratios as an objective method of distinguishing between suphide deposits or sedimentary or supergene origin, and sulphide deposits of hydrothermal origin, with particular application to the districts mentioned. At the same time much data has been accumulated as to the selenium content of Australian sulphide ores. The possible use of sulphur-selenium ratios for this purpose was first demonstrated by Goldschmidt and his coworkers (1933; 1935), and later by Carstens (1941). Goldschmidt and Strock (1935) found that pyrite of sedimentary origin had a sulphur : selenium ratio of the order of 200,000 or more, whereas pyrite of hydrothermal origin had a sulphur: selenium ratio of 10,000 to 20,000. Carstensfound that sedimentary pyrite deposits in Norway contained less than 1 g. Se per ton, whereas hydrothermal pyrite deposits contained 20 to 30 g. per ton.

Jan 1, 1954

By Glenn Lyle

CEMI Introduction ? Introduction of the CAF VOD Project ? Who are our partners ? What were the Outcomes What is next

May 1, 2011

By Christopher J. Bise

INTRODUCTION The level of analysis for the selection of extraction practices and equipment for surface mining differs significantly from that found in underground mining. In underground mines, manufacturers establish the operating ranges of the equipment, such as continuous miners, loaders, and boring machines, based upon required clearances. Even in conventional mining, the drilling and blasting operation is usually based upon experience. The selection of extraction methods and equipment for surface mining is more complex. Since the equipment is not as confined as in deep mining, and the topography tends to be highly variable over the limits of the mining property, primary extraction equipment must be sized according to reach and capacity. Further, due to the depositional characteristics of the over- burden, drilling and blasting patterns must be well- designed to provide proper fragmentation. The following sections provide a general outline for the selection of surface excavation practices and equipment. The topics covered include drilling and blasting and the selection of both primary extraction equipment (draglines and stripping shovels) and secondary extraction equipment (bulldozers, scrapers, and front-end loaders). EXCAVATION FUNDAMENTALS Before discussing surface excavation, a few concepts need to be defined. A bank cubic yard of material is equal to one cubic yard as it lies in its natural or undisturbed state. A loose cubic yard is that portion of a bank yard that after being disturbed has expanded to measure one cubic yard. Swell is the volume increase of a material when it is removed from the natural state. Finally, load factor is the percentage decrease in the density of a material from its natural state to a loose state. Table 1 summarizes these concepts for various materials. When the amount of overburden to be removed to meet a predetermined production rate is established, a suitable bucket size for the primary stripping equipment must be calculated. The following equation can be used: NBSIZE = TBC/[(OF) (BF)] (1) where NBSIZE is the actual bucket capacity after considering the swell percentage, the bucket fill factor, and the operating efficiency; TBC is the theoretical bucket capacity (TBC = COB/NTCPM); OF is the operating efficiency; BF is the bucket fill factor; COB is the cubic yards of overburden to be removed; NTCPM is the number of theoretical cycles per month [NTCPM = (SMH) (3600)/(CT)]; SMH is the scheduled monthly hours; and CT is the cycle time in seconds. FRAGMENTATION PROCEDURES Drilling Where the strength of overburden material is such that blasting is required prior to excavation, drilling is conducted before using explosives. Two methods of drilling are found in mining-percussion and rotary. Penetration of overburden by percussion drilling is carried out by successive impacts of the bit into the rock. Emphasis in this section will be placed on rotary drilling, because for surface mining it is the most versatile form of overburden penetration. Rotary drilling is conducted by rotating a rigid string of tubular rods to which a rock-cutting bit is attached. The drilling energy is supplied by rotation as well as thrust and the cutting action consists of abrasion, scraping, spalling, or chipping. The size range for rotary-drilled blastholes is from 4 in. to 15 in., with 9 in. and 12 in. being common sizes. There are two types of bits used for rotary drilling-drag and rolling cutter. Drag bits are applicable in soft to medium-soft shales or their equivalents. Since their effectiveness is dependent upon a successful scraping action, drilling thrust is lower for drag bits than for rolling cutter bits; the opposite is true for torque requirements. The bits are usually fitted with replace- able tungsten-carbide cutters. Rolling cutter bits, though available in a two-cone configuration, are usually of the three-cone style (Table 2). The rolling cones have either large steel teeth for drilling through soft formations or small tungsten-car- bide inserts for drilling through hard formations. Table 3 relates bit size and weight to the strength of the rock. The major advantage of rotary drilling is its rapid drilling rate. This is due to the fact that, unlike percussion drilling, rock penetration is uniform with depth. Blasting Effective bank preparation deals not only with the fragmentation of overburden but also with the mini-

Jan 1, 1986

Many mining operations generate seismic signals. They come from blasting and from underground mine failures such as rockbursts, longwall first caves, coal bumps and pillar collapses. Traditionally, seismic motions have been of concern on a local scale because they could create damage to structures at the mine site or at neighboring locations. Recently, there has been much interest in the Comprehensive Nu-clear-Test-Ban Treaty (CTBT). It prohibits any kind of nuclear explosion, anywhere. So distinguishing signals of mining operations from a possible nuclear explosion becomes important. An international monitoring system (IMS) is being implemented to verify the Treaty when it is ratified. It includes a large number of seismic stations capable of recording signals at regional (200 to 2,000 km or 124 to 1,240miles) and teleseismic (more than 2,000 km or 1,240 miles) distances.

Jan 1, 2000

By S. C. Richardson

The electrowinning (EW) circuit involves using DC current in the electro-deposition of metal ions dissolved in an electrolyte solution. This paper reviews a method to generate an electrical model of an EW circuit and the resulting DC current flow and voltage distribution within the circuit. Topics include: calculations of electrolyte conductivity and the equivalent electrical resistance of the electrolyte within the electrolyte pipelines connected to the EW circuit; the breakdown and classification of the EW cell voltage into thermodynamic and ohmic potentials; methods to calculate current flow and voltage distribution throughout the EW circuit; and the development of an interactive visual computer model to simulate an actual EW circuit. Stray current can pose an electrical shock hazard to employees in electrolytic tankhouses. Increased awareness and understanding of the voltage distribution and potential stray current in an EW circuit can help mitigate the risk of tankhouse workers becoming exposed to electrical shock hazards. INTRODUCTION Copper electrowinning tankhouses include rectifiers that deliver large amount of DC current to a series of electrolytic cells consisting of electrolyte solution and electrode pairs. The design also includes metallic conductors in the circuit which include the rectifier’s positive and negative busbar, inter-cell busbar, and the anode and cathode frames. The electrolyte acts as an electrical conductor that transfers current via ions between the anodic and cathodic electrodes within the cell. Electrolyte, as a conductor, is also capable of conducting DC current outside the main EW circuit through the connected electrolyte feed and return pipe systems. DC current that follows a path other than the intended path through the series of electrolytic cells is referred to as “stray current”. Developing an electrical model enhances understanding of voltage distribution and current flow in an EW circuit. This paper discusses the derivation of the various representative components of the model as well as the output from the model. An interactive visual computer simulation incorporates the model output in order to aid in the understanding of voltage distribution and current flow throughout the EW circuit.

Jan 1, 2019

By Chris Breeds, Ken Johnston, Omar Fulton

This paper describes the process typically involved with obtaining permits for quarry blasting in urban, suburban, and rural areas, including: content of SEPA (State Environmental Policy Act) driven documentation; regulatory review and comment; public interaction; public hearing(s); periodic regulatory review(s); negotiated solutions; monitoring; and the role of public interaction, demonstration(s), and education in the process. Case studies will be presented of projects undertaken in the State of Washington that illustrate the approach and regulatory interaction as well as the importance of public interaction throughout the process. Reference will be made to the role of national and state regulations and local codes and ordinances that provide additional limitations on blasting practice. The paper will also present regional blast monitoring data that are used in the permitting process and the layout and design of blast monitoring systems typically required as conditions of the quarrying permit. Our presentation will include tools, which we have developed to assist with explaining blast related phenomenon to both regulators and the public.

Jan 1, 2003

Your assessment of the minerals business. The global mining industry is strong. I am convinced it will continue meeting the mineral needs of the world population for the foreseeable future. Two developments have greatly enhanced the industry’s ability to produce low-cost minerals. Technological innovations have resulted in increased productivity. And there has also been a dramatic expansion of the number of favorable countries in which to explore and mine. Capital expenditures have flowed to the low-production-cost mines and development projects as well as to the exploration projects and stable countries having the best potential for valuable mineral discoveries.

Jan 1, 2000