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By R. J. Smith

Tar sand deposits in Utah represent more than 25 billion barrels of in-place bitumen, 96% of the known U.S. reserves. The technological development of a hot water processing strategy for the Canadian tar sands and the commercial Production of synthetic crude oil from this natural resource suggest that similar success may be possible for the utilization of Utah tar sands to provide our country with a new domestic source of synthetic crude oil. In this regard an active research program to develop a hot water processing strategy for Utah tar sands has been sustained at the University of Utah for the past five years. The hot water process for Utah tar sands differs significantly from that used for Canadian, tar sands due to the inherent differences in the respective bitumen viscosities and the nature of the bitumen-sand association. These differences have led to the identification of certain design criteria necessary to achieve satisfactory phase disengagement during digestion. (8,9) Subsequent flotation of the digested tar sand is a subject of current attention and is considered in this contribution for the Asphalt Ridge tar sand deposit. Although contact angle measurements of pure Asphalt Ridge bitumen indicate moderate hydrophobicity, air bubble attachment to the bitumen concentrate taken from a 37 liter flotation cell is not possible. This surprising result suggests that the flotation separation is dependent on air bubble entrapment rather than on attachment due to surface hydrophobicity. The occlusion of air bubbles in the bitumen is apparent from visual examination of the concentrate, especially at lower temperatures. It seems that a bitumen bubble agglomerate forms in the impeller region of the flotation cell, the effective density of which is such to allow the agglomerate to float to the surface. A factorial design of the major operating variables in the flotation separation indicated that the quality of separation was significantly dependent on the flotation temperature and to a lesser extent on the degree of agitation. For a flotation temperature of 77C recovery of 96.7 percent was realized at a grade of 61.0 percent bitumen. The improved separation at higher flotation temperatures was found to be due to the decrease in bitumen viscosity resulting in more effective rejection of coarse sand from the concentrate.

Jan 1, 1980

By S. G. Britton

"INTRODUCTION Strategies to improve performance in mining production, maintenance and mineral processing operations has evolved over time from straightforward time studies of individual machine performance in the 1960’s-1970’s to more sophisticated and better focused statistical and financial tools that have emerged in the 1980’s-1990’s . These newer strategies fall under the broad descriptive category of “Operational Excellence”. Operational Excellence is the term that embodies a number of performance evaluating strategies, the most popular and relevant to mining includes the following: • Six sigma for addressing errors within mining or milling process and work standards • Lean manufacturing to optimize organizational and operational performance through continuous improvement. There are other strategies that can be applied as well and this paper is not focused on a review of the pros and cons of any one strategy. However, the common threads in these strategies embody the expectation that expected results from these efforts can be quantified and that value will positively benefit the targeted work process, organizational unit, equipment deployment or time schedule when applied to it as part of continuous improvement efforts. The priority for proceeding with any one or more of these efforts is based on the calculated net value of the required resources to affect the change versus the estimated value gained. The practice is identified as the value mapping process. This discussion summarizes the value mapping process methodology used at a mature uranium mine and mill in northern Canada to work through the myriad of potential improvement efforts available through the operational excellence process for continuous improvement.THE RABBIT LAKE MINE AND MILL The Rabbit Lake Mine and Mill are located in northern Saskatchewan, Canada on the shores of Wollaston Lake (Figure 1).The mine was discovered in the mid-1960s and was built by Gulf Minerals, Ltd. The mill began processing uranium ore into yellowcake in 1975 from an adjacent open pit mine. A whole ore acid leaching followed by solvent extraction and precipitation forms the basic milling processes. The operation continued to produce yellowcake through a succession of open pits and finally opened an underground mine in the early 1990s. Since beginning production, over 200 million pounds of yellowcake has been packaged from the site, making it the sixth highest producing uranium mine of all time."

Jan 1, 2016

By R. A. Young

Effective October 1, 1958, President Eisenhower imposed, by Presidential Order, fixed import quotas on lead metal and lead concentrates and also on zinc metal and zinc concentrates. He did this in response to an escape clause case under the Trade Agreements Act filed by many of the large and small lead and zinc mining companies in September 1957. It is the purpose of this paper to give the experience of the lead-zinc industry under this order during its first two years and some observations on what might have come about under different conditions of U.S.A. consumption. Before considering these various experiences and observations, it seems wise to first outline briefly the basic problem which prompted such an order. Under the Tariff Act of 1930, the duty on pig lead was fixed at 2-l/8¢ per pound and slab zinc at 1-3/4e; per pound. The ores, concentrates and closely related products of these metals carried closely related duty rates. The Trade Agreements Act of 1934 and succeeding similar acts had given the, President the power to lower these and other duties. As a result-of various negotiations with foreign nations, by 1951 these duties had been lowered to their present day rates of 1-1/16 per pound on' pig lead and 7/10o, per pound on slab zinc. Related ores, concentrates and semi-fabricated forms of the metals received commensurate treatment. Under these acts, the Tariff Commission has the right to recommend to the President duty increases, specific quotas or both if find that imports have injured the affected industry. He has the right to accept or reject such recommendations, in whole or in part, in the light of the national interest,

Jan 1, 1961

By John P. Hunt

Early disseminated copper mineralization at Quebrada Blanca, Chile, accompanied 38 million year old quartz monzonite stocks and dikes. Early mineralization was followed by hydrothermal mineralization, mineralized breccias, and post-mineral faulting. After supergene leaching and enrichment, Quebrada Blanca was buried under late Tertiary gravels, and then partially re- exposed in Quaternary time. Based on work to date the orebody appears to contain more than one hundred million tons of plus one percent secondarily enriched copper ore, overlying several hundred million tons of protore averaging more than 0.5% copper. A source of industrial water appears to have been located. Pilot scale metallurgical tests, mine planning, and feasibility studies are in progress.

Jan 1, 1981

By Jack V. Hill

A unique type of grinding mill liner has been developed by Waagner-Biro A.G. of Austria. This liner system, termed an angular spiral lining (.ESL), was, successfully tested by Magma Copper Company's San Manuel concentrator in both their 10 Ft x 10 Ft (3.05m x 3.05m) and 12 Ft, 6 In x 14 Ft (3.81m x 4.27m) grate discharge wet grinding ball mills. Magma Copper's savings in power and grinding balls justified conversion of more than half of their twenty-six ball mills to the ESL system.1,2 The ASL system was provided by the Capitol Castings Division of Midland Ross Corporation, the exclusive licensee for such conversions. Capitol Castings proposed the following as expected benefits from installation of an ASL system in an 18 Ft x 21 Ft (5.49m x 6.40m) overflow discharge wet grinding ball mill at the Pinto Valley Operation of Cities Service Company: "(a) Reduction of power consumption per ton of ore milled = 12%, (b) Increased throughput = 6%, (c) Reduction of ball consumption = 6%."

Jan 1, 1982

By James A. L. White

Although the quantity of native sulphur associated with volcanism is large, production from such deposits represents only a very small percentage of total world production, This paper describes the location of the known deposits, the typical geology of these deposits and the nature of the sulphur mineralization. Methods of treatment are discussed and some comments are offered on the economic position of these occurences in world sulphur markets.

Jan 1, 1967

By Dapeng JIANG, Haixiao YANG, Shanxi WU, Jinwang Zhang, Guoliang XU, Yipeng ZHANG

The subsidence and deformation characteristics of strata with thick unconsolidated overlying strata is quite different from that with consolidated overlying strata. In this paper, based on the engineering background of large mining height mining in No.21 coal seam of Zhaogu No.2 Coal Mine in Henan Province, China, the PFC 2D numerical simulation and surface subsidence field observation was conducted to study the subsidence characteristics of strata with thick unconsolidated overlying strata (680m). Numerical model with dimensions of 2000m 750m was established and mining process with different working face lengths (150 m,200 m,250 m,300 m) and different working face advancing distances (40 m,80 m,120 m,160 m,200 m,240 m,280 m,320 m,360 m,400 m,440 m,480 m) were simulated. The surface subsidence characteristics and displacement field variation of rock strata under different working face advancing distances were observed and analyzed. The results show that with the advance of working face, the structure of subsidence-like basin will be formed in each rock stratum, and the influence range of mining on rock stratum is “bowl-shaped”, and the influence increases with the increase of advancing distance. The unconsolidated layer has a buffer effect on the mining-induced subsidence, and the buffer effect is realized by the expansion of the subsidence area. The movement law of strata under different working face lengths is similar to that under general mining, and the structure of subsidence-like basins will also be formed in each stratum. The influence range of strata after mining is ' bowl-shaped ', and the influence extent increases with the increasing of working face length; the internal subsidence of the unconsolidated layer is more intense and complex than that of the rock layer; the surface subsidence movement shows an obvious 'basin' phenomenon. The field observation data of surface subsidence also confirm the above results.

By H. Lyn Bourne

Daily everyone depends on the great variety of glass products, so much so that glass is often taken for granted. In fact most people do not realize how versatile glass has become. Consider the various uses and then try to imagine a day in which we are not influenced by glass. Common uses include container ware, table ware, window glass, lead crystal, automobile glass, and fiber glass. Several less common, but important, uses include laboratory ware, pharmaceutical, TV bulbs, light bulbs, glass ceramics, optical glass, fiber optics, and laser glass. Corning, Inc., a leader in specialty products, uses nearly 1 000 different compositions to manufacture about 60 000 different products (Edwards and Copley, 1977). Glass is such a complex product that-definitions vary and exceptions can be found for most definitions. Glass is an inorganic amorphous (non-crystalline) solid. Most glasses are produced by melting of a mixture of oxide raw materials, and then cooled to room temperature. Soda-lime-silica composition.s account for about 90% of all glasses melted (Anon, 1973). The properties of the glass product come mainly from its chemical composition. All of the different glasses require melting a combination of raw materials and forming the molten material into the desired shape. Both the melting and the forming processes use sophisticated technology and these technologies require experts to manage these production systems. The manufacturing process is continuous and takes place in tonnage quantities, so adjustments in the batch to achieve the desired finished product requires a great deal of expertise. Raw materials are fed to the batch mixing area in very large quantities (tons in most cases). As a result, impurities in the range of 0.1% result in addition of that impurity within the molten glass in kilogram amounts. More than twenty different industrial minerals are consumed in the manufacture of various kinds of glass (O'Driscoll, 1990). This chapter describes the major and minor ingredients of the various glass batches. It discusses the roles of the various oxides in the glass batch and most importantly considers the mineral raw materials which supply the glass industry. Each of the raw materials is described in detail in other chapters so the geology and mineralogy sections are kept brief here. Container glass, by far, accounts for the most production; followed by flat glass, fiber glass, and specialty glass of which table ware accounts for the greatest tonnage. [Table 1] shows the general production data for 1987 through 1990. Statistics for many of the uses do not appear because production volumes are small compared to the major uses. The glass industry is organized in four categories: containers, flat glass, fiber glass and specialty glass. The US Department of Commerce, Bureau of Census, publishes production data about the glass industry in three different categories: 1) glass containers, 2) consumer, scientific, technical and industrial glassware, and 3) flat glass. The Bureau has very complete statistics about the glass industry in these three categories but they report production data in different units according to industry standards. Therefore, [Table 1] gives the production data in dissimilar units. The production of most glass articles follows similar steps. The raw materials are mixed and the resulting batch is fed into the furnace. In soda-lime-silica glasses melting begins between 600 and 900°C. At these temperatures CO, and other gasses are released which create bubbles in the molten glass. To remove the bubbles and insure complete melting the temperature is raised to between 1 500 and 1 600°C. This is the melting-refining stage during which the refining agents in the glass batch serve to aid in the release of gas bubbles, homogenize the melt, and prevent the formation of scum on the surface of the molten liquid. At the conclusion of the melting-refining stage the glass is too fluid for working and the melt is cooled to about 1 100°C to attain the proper viscosity for working and forming to begin. After the glass article has been made, it must undergo annealing (slowly and uniformly reheated and cooled) to remove thermal stresses that were created during the forming process.

Jan 1, 1994

By Paul Saxton

INTRODUCTION AND SUMMARY This short paper deals with the history of the Nickel Plate Mine from 1898 to 1987, with the major focus on the time period 1984 to 1987. Operations from 1898 to 1955 are reviewed in a cursory manner with the emphasis on some interesting points in the early 20th century. A review of the major exploration work carried out in 1984 to 1987 is presented and the process of moving the project through feasibility, government permitting and construction is described. The Hedley gold camp is situated approximately 40 kilometres east-southeast of Princeton, in southern British Columbia. The area has had a long history of gold mining and between 1902 and 1955 approximately 1.5 million ounces of gold were won from several mineralized skarn orebodies. Most production came from the Nickel Plate and Hedley Mascot mines. AN EARLY TO MID 20TH CENTURY SUMMARY The Nickel Plate is one of the oldest mines in the province. The mine had operated from 1904 to 1930 and, after a period of shutdown, from 1934 to 1955. The mine was revived in 1967 for exploration and. will resume production during the first half of 1987. In the first period the Yale Mining Company and Daly Reduction Company operated for six years and Hedley Gold Mining Company for twenty-one. In the second period the mine was operated by the Kelowna Exploration Company Limited and Kelowna Mines Hedley Limited for twenty-one years. In the present period the mine has been revived by Mascot Gold Mines Limited.

Jan 1, 1987

By D. R. Hembree

This paper discusses the mineralogy and metallurgy of the Kuroko-style volcanogenic massive sulfide ores mined at Formosa Exploration Inc. I s Silver Butte Mine located in southwest Oregon. The principle ore type is pyrite + chalcopyrite + sphalerite, while subsidiary pyrite + bornite + sphalerite ore is also mined. The two styles of mineralization have distinctly different metallurgical behaviors and must be treated separately. Selective flotation methods are used to produce copper and zinc concentrates from the first ore type, while bulk concentrates are produced from the bornite ore. Careful mine planning and coordination between the mine and mill are utilized to achieve optimal recoveries.

Jan 1, 1993

By W. J. Mahoney

A good quality assurance program at a coal mine or power plant is accurate, timely, and consistent. The quality analysis is accurate due to a complete sampling of the coal stream loaded into the unit train. The sample analysis is accurate because standardized testing procedures are applied. Once the analysis is completed, the results must be communicated to the electric utility in a timely manner. The electric utility needs the results to better manage the coal inventory at the power plants. A quality assurance program can also help the mine deliver a consistent product to the electric utility's power plants. The importance of the three major tenants for a good quality assurance program: accuracy, timeliness and consistency, are based on major changes that are taking place in the electric utility industry.

Jan 1, 1994

By Richard H. Follis

Activated carbons used for precious metal recovery from carbon-in­pulp (CIP), carbon-In-leach (CIL) and carbon-in-column (CIC) circuits often become fouled with Inorganic precipitates (i.e., CaCOa, CaS04, SI02), and adsorbed organics (I.e., petroleum distillates from lubricants, dissolution products from wood) (1,2). These contaminants occlude carbon pore structure, such that diffusion controlled adsorption and desorption (elution) of Au(CN)2-and Ag(CN)2· are lmpeded(3,4l. Thus, after only short exposure to the hydrometallurgical environment, the characteristic carbon activity is reduced(5l. By frequently regenerating circuit carbons, activity losses associated with fouling can usually be minimized. Regeneration typically Involves (not necessarily Is this order): 1) elution to remove adsorbed ionic species (e.g., until they are removed from the carbon pore structure, Au(CN)2" and Ag(CN)2-are contaminants )(6l; 2) acid washing with 3-15% (wt/wt) hydro­chloric acid, to dissolve soluble Inorganic precipitates such as CaC03; 3) neutralization and/or water rinsing to remove residual acid; 4) thermal regeneration at 350-650 °C to remove adsorbed and/or pyrolyzed organ­ics; and 5) screening to remove fine carbon particles produced by Impaction and abrasion in the various stages of the hydrometallurgical circuit (i.e., adsorption, elution, and regeneration). The overall process Is purposed to reopen the circuit carbon pore structure, and to maintain the activity and qualities of virgin (new) carbon(7,8l. While most CIP, CIL and CIC operators regenerate circuit carbons, few evaluate the effect of the process on carbon quality in a systematic way. Process control usually Involves Iterating critical regeneration parameters such as: acid concentration; time frames for acid washing and water rinsing; and temperature and residence time for thermal regeneration. Carbon quality control frequently comprises only sporadic activity tests performed on non-representative grab samples dipped from conve­niently accessed carbon storage bins. So, regenerated carbon quality may vary and (at times) be unacceptable for effective hydrometallurglcal circuit operation. To verify the efficacy of the regeneration process and assure accept­able adsorption and elution rates for Au(CN)2· and Ag(CN)2·, operators should systematically assess regenerated circuit carbon quality. This paper, then, suggests methods for effectively assessing circuit carbon quality, Including: carbon sampling procedures; standardized physical and chemical test methods; and criteria and general procedures for comparative adsorption rate (activity) testing. Also, a two tier quality control testing scheme is proposed, and techniques for presenting and interpreting test data are demonstrated.

Jan 1, 1992

By Frank Florez

INTRODUCTION The Magma mine, located at Superior, AZ, some 100 km (60 miles) east of Phoenix, was discovered in 1875, and was originally mined for silver. The property was acquired by W. B. Thompson who formed the Magma Copper Co. in 1910. Prior to 1965, square-set mining and cut-and-fill methods were used to mine steeply dipping veins at Magma. Ore was moved to the chute by gravity in the square-set stopes and by slusher in the cut-and-fill stopes. The stopes were gobbed with waste from the surface and development headings in the mine. The original square-set method of mining was modi¬fied in order to extract replacement ore bodies in limestones that dipped at 0.52 rad (30°) and varied in thick¬ness from 4.6 to 38 m (15 to 125 ft). Mining of these ore bodies was done by starting at the upper lateral ex¬tremities and retreating to the access openings. The timber square sets and cribbing were used as temporary support, the ground being allowed to cave as the mining retreated. It was found, however, that mining in such a manner was very inefficient; safety hazards increased be¬cause of the blocky and fractured characteristics of the ore and the weakness of the hanging wall. GENERAL DESCRIPTION The undercut-and-fill method of mining is now the primary stoping method used at Magma, and no pillars are left for later extraction. Four ore bodies are pres¬ently being mined. These ore bodies (see Fig. 1) dip conformably at 0.52 rad (30°) to the east with the host limestone and shale sediments, their average strike length is about 122 m (400 ft), and their thickness varies from 9 to 38 m (30 to 125 ft). The ore bodies are separated by varying thicknesses of barren rock. Three of them come together at various elevations primarily in the area of the main feeder fault, but there is an apparent horizontal displacement of about 23 m (75 ft) on the main feeder fault which complicates stope layout. The method as originally applied has been modified in order to mine more efficiently ore bodies that are 9 m (30 ft) or more in thickness. The modification was made to mine all the ore at a given elevation before dropping down for the cut below. The advantage of this is that mining is done on a horizontal plane and there¬fore muck removal can be mechanized. Also multiple faces are made available, thereby improving overall min¬ing efficiency. Another major change made was to carry a 6.7-m (22-ft) face when mining the top sill and a 6.4-m (21-ft) face when mining under fill rather than the former standard 3.35-m (11-ft) high face. This change lengthens the mining cycle and cuts mat prepa¬ration costs in half. Stopes are laid out (see Fig. 2) so that maximum mining panels are available and yet one stope crew will not tie up too much of the ore reserve. One stope might consist of as many as 20 panels, and these panels are divided evenly on opposite sides of the extraction cross¬cut. The stope layout is planned so that the maximum tramming distance to the orepass will not exceed 30.5 m (100 ft) on the crosscut and 38 m (125 ft) in the panel. Stope panel distances are sometimes exceeded when min¬ing to the extremity of the ore body. Tramming distance limits are determined by the efficiency of the com

Jan 1, 1982

By Fan Yang, Meigui Yi, Lan Xiang, Zhuoying Yang

This paper reported the calculation of the reaction heat in titanium-bearing blast furnace slag (abbreviated as TBFS)-H2SO4 system based on estimation of thermodynamic data of CaTiAl2O6 by mechanical mixture model. The thermodynamic analysis indicated that CaTiAl2O6 was more stable than the corresponding ternary oxides as CaO·TiO2, CaO·Al2O3, and Al2O3·TiO2. TBFS-H2SO4 system was highly exothermic since all of the reactions involved in the system were exothermic. For example, the increase of temperature from 25 to 200°C led to the increase of the release of the reaction heat in TBFS (100.0g)-H2SO4 system from 181.28 to 189.73kJ. Eighty-five to 90%of the reaction heat in TBFS-H2SO4 system was contributed by the reactions among CaTiAl2O6, CaTiO3, CaMgSi2O6, and H2SO4.

Feb 1, 2021

By R. J. Sweigard

Stabilization of slopes on abandoned mine land (AML) invariably involves some channelization to control surface drainage. Typically, these channels are lined with riprap or gabions, depending on gradient and expected flow rate. The conventional treatments pose a problem since they require frequent maintenance and replacement at times. Soil bioengineering presents an alternative to the conventional measures. In soil bioengineering, living woody plant material forms the main structural component in systems designed to prevent shallow mass slope failures, surface erosion, and channel bank erosion. An experimental application of soil bioengineering techniques was designed and implemented at an AML site in eastern Kentucky. The design consisted of live fascines and live stakes using willow branches installed with coir geotextile. Channel gradient, length, and cross-section were considered in the design.

Jan 1, 1998

"There is no fast way to add more than 6.4 km (4 miles) of track, most of it underground, to the Sound Transit Link Light Rail system in the Seattle, WA area. But increasingly busy traffic through a number of neighborhoods made it necessary. That’s the challenge a joint venture of several tunneling contractors faced when starting its portion of the $1.9-billion Northgate Link Extension in 2013, and the project is still keeping it busy three years later. Completion will mean just a 14-minute ride from Northgate to downtown for an extra 60,000 passengers by 2030. The transit lines will join 57 km (36 miles) of new and under-construction light rail lines running north, east and south from Seattle.Voters tired of long commute times approved the project, a key part of Seattle’s regional mass transit system, in 2008. The extension is part of a plan adopted by the Sound Transit board in 1996 to connect the region’s major activity centers. The Northgate project is meant to relieve one of the area’s most congested traffic areas by connecting the Northgate, Roosevelt and University District neighborhoods to existing routes including downtown Seattle and SeaTac airport.Sound Transit dedicated more than a quarter of the project’s estimated expense — $440 million -— to general contractor JCM Northlink LLC to handle the tunnel segment. JCM is a joint venture formed in 2013 between Jay Dee Contractors of Livonia, MI, Frank Coluccio Construction of Seattle, WA and Michels Corp. of Brownsville, WI. The contractors regularly collaborate on large-scale projects and brought experience working on a different section of the extension they completed in 2013.JCM project leaders knew immediately that the schedule would be demanding. They had to work around the clock six days a week to meet deadlines. The extension runs through highly populated areas, meaning contractors had to follow strict city noise restrictions between 10 p.m. and 7 a.m. Plus, supervisors needed to manage the scheduling of nearly 250 employees to keep work running smoothly.“We had eight different operations and four different subcontractors trying to work in one tunnel,” said Chris Wood, JCM site project superintendent. “It’s important that we keep work linear. That’s the biggest challenge: keeping a path in and out so you can pour concrete, keep electricians and carpenters going, and accomplish everything else.”"

Dec 1, 2016

By Juan L. Yarmuch, Hyam Rubinstein

Open-pit mine planning is a key process in the definition of mineral reserves and their efficient removal. Ore in open-pit mines is extracted bymining sequential pits called pushbacks. Pushbacks need to satisfy complex geometrical constraints to ensure the efficient usage of equipment. To date, no algorithms that both account for those constraints and maximizeNPV, are commercially available. Therefore, pushback design is done by engineers relying on their expertise to incorporate practical features intothe output of commercial tools. In this article, we present a new algorithm that produces close to optimal practical pushbacks and we illustrate its performance in real mines.

Jun 25, 2023

By C. C. Graham

In this paper, the problems which occurred on introduction of dense medium cyclone coal washing in Australia are reviewed together with the subsequent measures taken to obtain a satisfactory magnetite. Specification and testing of magnetite for exploration, quality control and in-plant check purposes is discussed. The disposition of Australian magnetite deposits in relation to the main coal producing areas is examined, and comments are made on the economics of mining, milling and beneficiating magnetite for coal preparation uses alone. Future trends in the Australian Coal Industry are predicted, and the need for a planned approach for the future is demons traced.

Jan 1, 1982

By AMIR EHSAN KIANFAR, Moritz Ziegler, TOBIAS HARTMANN, ELISABETH CLAUSEN

Accurate positioning information is a central step in the process of increasing automation and digitalization in the mining industry. Underground mine environments are prime examples for GNSS-denied environments that pose several additional challenges for the employed indoor positioning system (IPS). This work presents an ultra-wideband (UWB)- based system specifically tailored for underground mine environments. Initially, the UWB technology is reviewed and basic concepts of IPS are explained. Both software and hardware are described in detail, including positioning algorithms, electrical circuitry and the mining-compatible enclosure. Finally, the results from field tests in an underground mine are presented and positioning performance and the influence of dilution of precision (DOP) on positioning results are empirically evaluated. The results show that the localization errors are mostly below 1 m. Individual zones of high localization errors are suspected to be caused by noise interference. Significant DOP influence due to the mining drift geometry can be confirmed. The system shows promising localization results for different mining applications, while influences negatively impacting the accuracy are presented as items for future research.

Jan 1, 2024

By Bradley V. Nelson

A number of improvements in the ventilation system at the Henderson Mine have been made during the past several years. A key to these improvements was the installation of a computerized monitoring and control system. Employees can determine the status of any of hundreds of devices in seconds through the use of this monitoring system. Mine dewatering pumps, the water treatment system, the ore haulage system, electric power usage as well as other devices are controlled and/or monitored. Ventilation parameters which are monitored include air flow volumes and carbon monoxide as well as the status of main surface fans, mine doors, and ventilation raise doors. Main underground fans are also constantly monitored and can be controlled remotely. These underground fans can be programmed to automatically start and stop to follow work schedules, optimizing air quality while reducing power cost. Mine air heating costs have been reduced by replacing an electrical resistance heating system with a direct fired natural gas heating system at the portal of the ore haulage tunnel.

Jan 1, 1995