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By M. F. Dibble

Borehole mining, as a viable concept for recovering deep sedimentary mineral resources, historically has been long on promise but short on results. Application of the technology seems best suited for recovery of Miocene phosphate resources on the East Coast of the US. These deposits are potentially enormous in scope and underlie sections of the coastal plain and the near offshore continental shelf. The landside portion of these deposits is generally too deep to mine using conventional open pit methods. Recovery of these minerals requires development of alternative mining systems. A significant proportion of the offshore resources appear to occur at depths below the practical limits of cutter suction dredges and may even exceed the limits of bucket ladder dredges. In most cases, these deposits are covered by overburden that must he removed and redeposited before the mineral section can be mined. The depth, thickness, composition of overburden and environmental sensitivity of a particular location would greatly impact mining system selection. It is probable that dredging would not be an environmentally acceptable mining method in many of the areas underlain by this phosphate rock. In cases such as these, new processes such as borehole mining must be evolved or the resource will remain undeveloped. A borehole mining test program, using an existing prototype borehole system, successfully extracted Miocene phosphate in northeast Florida in 1984-1985. The top of the main mineral zone was 80 m (261 ft) below the surface. Based on the results of this test program. plus subsequent laboratory testing and engineering studies, a new prototype borehole mining system was designed in 1986. This design incorporated several innovations of offering potentially higher productivity and lower power consumption than was achieved with the previous test unit. Economic restraints at that tine prevented the actual construction and testing of this new design.

Jan 1, 1992

By J. Le Gac, P. Zettwoog, J. Brenot, J. W. Piechowski, J. C. Nenot

INTRODUCTION The components of the radiological hazard in uranium mines are: 1. exposure to radon and its short-lived daughters 2. exposure to long-lived ore dust 3. external irradiation With usual ventilation rates, the dosimetry relative to the first component only requires to know each miner&apos;s exposure to a potential energy of short-lived radon daughters, (EXP[a]). Presently in France, (EXP [a]) is determined monthly from radon concentration measurements. This is an indirect method based on ambient monitoring. After technical improvements achieved during the last years it is now possible to measure (EXP [a] directly with an individual portable monitor developed by the CEA. The purpose of this study is to analyze and compare monitoring results obtained by these methods. DETERMINATION OF INDIVIDUAL MONTHLY EXPOSURES TO[ a] POTENTIAL ENERGY OF SHORT-LIVED RADON DAUGHTERS [(EXP a) amb, by ambient monitoring] In this case, it is necessary to know individual monthly exposures to radon, (EXP Rn). As mentioned by Pradel, J.,et al., 1976, (EXP Rn) is estimated for each worker from informations both on radon concentrations (CRn) measured as described by Pradel, J., 1962 and times (t) elapsed in the various workplaces. (EXP Rn) is obtained by summing up the elementary radon exposures CRn x t relative to all the workplaces where the miner was successively engaged during the month. (EXP Rn) values are then corrected by an average equilibrium factor of 0.17 deduced from many measurements and taken as representative for all the French mines. Finally: [ ] (EXP [a])ind are obtained directly with an individual portable instrument described by Chapuis, A.M., et al., 1972. Samples of the air breathed by the miners are pumped continuously and radioactive aerosols containing the short-lived radon daughters are trapped on a filter. The principle of the measurement lies on track-etch detection of the a particles emitted by the deposited material. The technical aspects have been described by Duport, P., et al., 1980a, Mouden, A., et al., 1980 and Chapuis, A.M., 1980. As mentioned by Duport, P., et al., 1980, fiability is good. Indications on measurement reproducibility are given by Pradel, J., et al., 1977. PRESENTATION OF THE STUDY The underground miner population of a French mine was followed during a period of 8 months (October 1978 - May 1979). Monthly exposure values for each miner determined either by ambient or individual monitoring, (EXP [a])amb or (EXP[ a])ind, had been previously collected by Le Gac, J., 1980, who had also defined miner categories according to the characteristics of the work performed. Four groups are considered here, the "whole mine" and three categories as described in table 1. Data on the individual monthly exposures,(EXP [a])ind and (EXP [a])amb, were taken again to analyze the individual cumulative exposures instead of the "month by month" values. The influence of the mean habits of each individual on its surveillance results is thus taken into account. Each miner has a cumulative exposure as an individual of the "whole mine" and as an individual of each category in which he has worked. The results obtained by ambient or individual monitoring were analyzed (characteristics and adjustements of the distributions) and then compared (description of the discrepancies observed). Obviously, such a study is possible only if all the exposures are cumulated on a same duration. As miners have not worked during the same number of months in the course of the data collecting period, one must considerer the individual average monthly exposures, [(EXP a)]amb and[ (EXP a)]ind, instead of the individual cumulative exposures, which is typically the same thing. To be sure that the exposure values of a given group (A, B, C or D) are representative, only those miners really present for at least 3 months in this group were taken into account. Finally, each individual of the "whole mine" and of the various work categories has a pair of average monthly exposure values : [(EXP a)amb] determined by ambient monitoring and[ (EXP a)ind] deduced from the track-etch recording of the individual portable monitor. Let a miner be engaged during n months[ (3 < n < 8)] in a group, the following relations then apply : [ ]

Jan 1, 1981

By Gunter Zimmermeyer, Hartmut Eicker

Radon in the Environment Radon, as a natural nobel gas, can be detected nearly everywhere in the environment as a decay product of ubiquitous uranium. As it is emanated from soil and rocks measurable concentrations have been found at the surface of soils and in even higher concentrations in enclosed spaces like, for example, mines and buildings. While above soil surface activities caused by radon have been found in an order of magnitude of up to 1 pCi/l (Weigel, F. 1978), concentrations in enclosed spaces and mines are higher because of the lack of atmospheric circulation. Beside air circulation the relevant figure depends on the Ra226-concentration in the surrounding rocks or building material, as well as on emanation coefficient and the diffusion coefficient. While representative Rn222concentrations in well ventilated buildings are reported to be in an order of magnitude of 1 pCi/l maximum values up to one order of magnitude higher have been found in badly ventilated brick buildings (Ettenhuber, E., Lehmann, R., Clajus, P., 1978) (Aitken, J.H., et al., 1977). Just now it was stated that the reduced air circulation due to German legal regulations on energy conservation will increase radon exposure of the public considerably (Jacobi, W., 1979). Radon in Mines Radon exposure of workers is, of course, a matter of concern in uranium ore mines where relatively high concentrations of the uranium to be mined are present. Measures to protect workers&apos; health have been implemented, based on experience on dose-effect relationship. They serve to meet exposure standards by limiting inhalation of radioactive particles, in reducing radon concentrations or in limiting working hours. Both improved measuring devices and capacity as well as the lower discrimination threshold enable to measure radon concentrations in other mines, e.g. in coal mines. It is known that radioactivity in coal is small compared with that in other minerals and even soil, rocks. Nevertheless, radioactive elements were identified in coal and so the question was whether the concentrations of radon in coal mines might be a subject of concern. The problems encountered when measuring radon in coal mines are described below, as the measuring device has to be flame proofed which is an important additional requirement. Measured radon concentrations in British coal mines have already been published (Duggan, M.J., Howell, D.M., Soilleux, P.J., 1968 (Dungey, C.J., Hore, J., Walter, M.D., 1978). The authors found concentrations of up to 14 pCi/l in Cornish mines. In most cases the values were in the order of 2 pCi/l. These results were consistent with measurements reported from U.S.-coal mines (Lucas, H.F., Gabrush, A.F., 1966). Such concentrations of radon were not considered to represent a hazard for British miners (Ogden, T.L., 1974). In Germany, too, first measurements have been carried out in five coal mines in the Saarland in the 60&apos;s. Air samples were taken at different places in the coal mines, dried, fed to an ionisation cell and measured by a device including reference cells. Samples taken at ventilated places showed radon concentrations consistent with the lower British results. They all kept within the standards of the first German regulation on protection against radiation. Measuring the radon daughters was renounced because of the relatively low radon concentrations and the requirements for flame proofness in coal mines. Moreover, it can be ascertained that because of the effective ventilation the disequilibrium factor between the decay products and the radon concentration remains far below the value of one (Muth, H., 1978) (Keller, G.). In 1979 the committee on mine safety and health protection in coal and other mines of the EEC proposed to have measured and evaluated radon concentrations in European coal mines to find out whether they complied with international standards. Great Britain and Germany agreed to this proposal and by commissioning such measurements to scientific institutes complied with the request to harmonize the methods used. In the Federal Republic of Germany, e.g. Westfälische Berggewerkschaftskasse (WBK) and Staatliches Materialprüfungsamt; Dortmund (MPA) were requested to carry out the measurements in coal mines of the Ruhr coalfield whereas Saarberg Interplan was responsible for the Saar coalfield. The WBK measurements are reported in later paragraphs.

Jan 1, 1981

By S. Bhattacharyya

"Coal bearing rocks are low in strength and have geological discontinuities and moisture sensitivity. Roof and rib falls have been a major cause of fatalities in underground coal mines. Mine engineers design ground control plans for regulatory approvals. A number of tools are available from the National Institute for Occupational Safety and Health to design and assess these plans. Quite often there is not enough geotechnical input information available and databases are used. The Appalachian coalfields have extensive scientific studies to provide reliable regional inputs. Coal mines in the western United States and Canada have different geological and geotechnical conditions, and less available data. Use of inappropriate inputs in the tools like Analysis of Retreat Mining Pillar Stability, Analysis of Longwall Pillar Stability, and Analysis of Roof Bolt Systems may result in unsafe designs or over conservative designs. The tools are convenient but thorough understanding of the inputs, calibration with local test results, and operating experience are important for effective designs. The current paper contains examples of difficulties faced during mine planning using the above tools. INTRODUCTION Sedimentary rocks hosting coal are low in strength. Scientific studies have placed the uniaxial compressive strength (UCS), triaxial strength, tensile strength, and shear strength of sandstone, shale, siltstone, and claystone in lower range of strength among engineering materials (1). Coal bearing rocks are most often weakened by the presence of faults, folds, slickensides, bedding planes, fractures, and other geological discontinuities. The host rocks and coal are often moisture sensitive. The geological discontinuities allow water to enter the strata from surface and underground sources. This combination of low strength, structural weakness, and presence of water results in difficult ground control situations for the coal mines. Roof (2) and rib falls have been a major cause of fatalities in the U.S. coal mines. Modern underground coal mines necessitate wide entries to accommodate large production equipment. The gate roads, section entries, and the mains entries of a highly productive coal mine are as wide as twenty feet (6.1 m). The actual intersection width is nearly thirty feet (9.1 m). Adequate ground control is a key to maintain safe production in such mines. Mine engineers and their consultants routinely submit ground control plans to the regulatory authorities. While working out the necessary roof control system the mine engineers are required to demonstrate the adequacy of the proposed system. There are multiple ways to demonstrate this. Some engineers use traditional load calculation approach of roof bolts (3). Some engineers use numerical methods. Yet a large number of professionals resort to a suite of ground control design software offered by National Institute for Occupational Safety and Health (NIOSH) commonly referred to as “NIOSH Ground Control Toolbar.” These software tools are free and easy to download from NIOSH website. The software included in this package are: Analysis of Longwall Pillar Stability (ALPS), Analysis of Retreat Mining Pillar Stability (ARMPS), Coal Mine Roof Rating (CMRR), Analysis of Roof Bolt Systems (ARBS), Analysis of Horizontal Stress Effects in Mining (AHSM), Analysis of Multiple Seam Stability (AMSS), Analysis of Retreat Mining Pillar Stability - Highwall Mining (ARMPS-HWM). These tools are relatively simple to install and use in planning and designing coal mines and prospective coal mines. They accept a number of geotechnical inputs which are collected from either an existing mining operation or from exploratory boreholes. In the absence of necessary input parameters, the programs use default values which are drawn from databases of geotechnical properties built by NIOSH."

Jan 1, 2015

By Q. Feng

Study has been carried out to apply fly ash as a high strength, water-resistant precast construction material through geopolymerization. Experiment results show that the working conditions such as water content, the concentration of NaOH, curing temperature and curing time, significantly affect the mechanical property of geopolymer matrix. Through optimization, an above 100 MPa compressive strength has been achieved with the geopolymerization products. Water soaking tests show that the geopolymerization product has a very high water resistance without losing noticeable compressive strength even after a one month soaking time. To elucidate the geopolymerization mechanism, microscopic techniques such as SEM/EDS and XRD are also applied to investigate the microstructure, the elemental and phase composition of geopolymerization products. The findings of the present work provide a novel method to apply fly ash as a high strength water-resistant precast construction material. INTRODUCTION Geopolymerization is a process of obtaining a polymeric structure from aluminosilicate by dissolving aluminosilicate sources in a strong alkali (NaOH) solution and condensation of free [SiO4] and [AlO4] tetrahedral units (Davidovits, 1985 and 1988; Duxson et al, 2005; Xu et al, 2000, 2002). After being cured at an elevated temperature for a specific time, the geopolymerization product shows a greatly increased compressive strength. Therefore, the product can be used as construction material such as bricks and road pavement. It provides a very promising method to apply aluminosilicate sources such as fly ash as a construction material for a sustainable development. Fly ash is a main industrial waste which needs to be carefully impounded because of its potential impact on environment. Many efforts have been tried to utilize fly ash as construction materials substituting Portland cement because of the following reasons. First, fly ash is a silica/alumina rich aluminosilicate source, which makes the specific material an ideal source for the geopolymerization process. Secondly, the particle size of fly ash is usually very small (< 50 microns) and it saves a lot of energy consumed during the production process of cement for grinding. Thirdly, fly ash is an ideal source for geopolymerization because it is reactive to alkali even at a mild temperature. In the past twenty years, many studies have been carried out on the geopolymerization of fly ash to a make construction material. However, there are some disadvantages associated with the conventional methods of treating fly ash (Palomo et al, 1999; Perera et al, 2007; Temuujin, et al, 2009; Hanjitsuwan et al, 2014). First, in most of the proposed geopolymerization methods, a salient amount of sodium silicate has to be added. It therefore makes it complex whether the achieved mechanical strength of geopolymerization products is due to a real geopolymerization process between fly ash and alkali or just largely the solidification of metal silicate in air at an elevated temperature. Secondly, the authors have noticed that when the geopolymerization product is soaked in water, the solution pH increases sharply and the products will lose its compressive strength gradually in water. Thirdly, the compressive strength of the geopolymerization product as reported is usually less than 50 MPa.

Jan 1, 2019

By J. L. Thompson

Lime made in a coal-fired, rotary kiln has two distinct parts: the interior of the pebble which constitutes about 90% to 95% of the mass, and the surface dust layer which is the remaining 5% to 10%. Uncalcined carbonate and impurities in the product may be present and will not contribute to the hydration reaction. Sulfur, in the form of calcium sulfate, inhibits the hydration reaction out of all proportion to its actual weigh-percent concentration (perhaps more than 100 times). If the stone contains dolomite, this will also inhibit the reaction as its rate of hydration may be nearly 3 orders of magnitude slower than that of calcium oxide. Early in the hydration reaction dolomite appears practically inert, but given sufficient time, it will hydrate. The heat of reaction is nearly the same as for calcium oxide, but the time necessary for complete hydration may be hours; whereas for calcium oxide it is only fractions of a second. The dust layer on the surface also inhibits hydration as it contains a collection of compounds which have low heats of hydration. This layer can tie up additional calcium and delay or prevent its reaction. In addition to these chemical processes, an important physical characteristic of the lime is its specific surface area. This is effected by the thermodynamics of calcination and the original stone grain chemistry. All of these factors are agglomerated in the gross slaking rate of lime. The total hydration of the chemical complex that is called commercial lime can be more easily understood by considering the individual reactions and combining them to form a more comprehensive picture of the situation. Calcium oxide and magnesium oxide hydrate similarly, differing only in their reaction kinetics. Dolomite, however, reacts differently than a simple mixture of calcium oxide and magnesium oxide as its hydration is strongly influenced by calcination thermo-dynamics and stone grain structure. Finally, the dust hydration and composition should be examined Understanding these reactions enables one to examine hydration reaction graphs with more insight into the factors which determine their shape.

Jan 1, 1978

By Trevor Riters

Advancements in Emerson’s instrumentation technology have brought value and efficiencies to the mining industry. Some think of instrumentation as basic sensors and valves, but with our solutions, we equip operators with real time, accurate, and reliable data that they can use to operate more effectively. Working with Caltrol, an Emerson Impact Partner, we successfully engineered wireless skids for a leach pad to provide precise measurement and control of raffinate solution to many of the large mine sites in Arizona.

Feb 1, 2025

By W. H. Jang

Recent research has demonstrated the utility of the FTIRIIRS adsorption density equation for in-situ study of surfactant adsorption phenomena at the surfaces of reactive internal reflection elements (IREs). The FTIR/IRS adsorption density equation is given by (formula) where r is the adsorption density, A/N is the absorbance per reflection, e is the integrated molar absorptivity, 4 is the bulk concentration of the surfactant, d. is the effective thickness, and dP is the depth of penetration. Previously, the validity of the adsorption density equation had been confirmed only from solution depletion measurements with high surface area powders. Now, FTIR/IRS studies of Langmuir-Blodgett (LB) monolayer films of fatty acid transferred to the surface of a calcium fluorite IRE .clearly validate the reliability of the FTIR/IRS adsorption density equation for quantitative analysis of surfactant adsorption. In addition, it was confirmed from the observed absorption band at 1550 cm-1 that the first monolayer of the transferred fatty acid chemisorbs at the surface of fluorite.

Jan 1, 1994

By V. Nasri, S. Rashidi

The Baltimore Red Line Project is a proposed 22.7-km (14.1-mile) long east-west light rail transit (LRT) line envisioned to connect the areas of Woodlawn, Edmondson Village, West Baltimore, downtown Baltimore, Inner Harbor East, Fells Point, Canton and the Johns Hopkins Bayview Medical Center Campus. The Red Line LRT System has two tunnel segments - the Cooks Lane Tunnel (CLT) and the Downtown Tunnel (DTT). The CLT segment is roughly 1,920-m (6,300-ft) long. It commences at the west portal located at the highway ramp for I 70 (to be removed) and terminates at the east portal that is at the intersection of Edmondson Avenue (U.S. Route 40) and Glen Allen Drive. This segment of the project consists of the following construction components: approximately 1,458 m (4,786 ft) of tunnels, 143 m (469 ft) of cut-and-cover tunnel, and 318 m (1,045 ft) of retained cut (U) section. The approximate horizontal alignment for the Red Line LRT Project is shown in Fig. 1.

Jan 1, 2014

By Fengmei Chen, Na Liu, Bingrui Li

A quantitative analysis method based on the airflow influence factors is proposed to evaluate airflow stability and discriminate airflow directions in complex ventilation networks. The law governing the changes in airflow caused by changes in branch resistance is revealed based on simulations of the airflow state in a ventilation network and a statistical analysis of the results. Then, three evaluation indices (airflow sensitivity, airflow influence factor, and critical airflow) are determined, matrices of the airflow influence factors and critical airflows are constructed, and a quantitative evaluation method for ventilation system stability and an airflow directional discriminant are proposed. The results show that the stability of a ventilation system can be reliably evaluated based on the airflow influence factors. The sensitive branches and dominant branches in the ventilation network can be determined based on the matrix of airflow influence factors, and the reversed branches, the trigger branches, and the reversal conditions can be determined based on the critical airflow matrix and the matrix of airflow influence factors. The proposed airflow directional discriminant is suitable for evaluating arbitrarily complex networks.

Nov 15, 2023

By Gerd L. Klinkenborg

Throughout the history of the development of mechanical conveyor systems, it has always been an established condition that loads are received at one end of a single-belt conveyor and discharged at the other end, in a point-to-point, manner. In systems designed for the continuous transport of granular materials, vertical or horizontal changes in; direction have always necessitated two or mope single-belt conveyor units, with a corresponding number of motors, assemblies and attendant mechanical and human or electronic controls. Besides these high-cost factors, every change in the system&apos;s direction also calls for a transfer point that involves a drop-back of the material from one level to another with some loss of the material and, in many cases, unwanted additional granulation. Where changes in the location of the discharge point are wanted, the conventional single-belt conveyor must either be replaced by a shorter or longer unit or additional units purchased and installed, or existing units rearranged. This can take days or weeks, and involve special expenditures as well as loss of valuable production time.

Jan 1, 1962

By H. M. Breza

Recovery of phosphate concentrate at Texasgulf’s Aurora Mine has received increased attention as existing mine and plant production capacity limits are reached. Improved mining techniques and mineral processing of waste streams offer the potential to increase phosphate production without major expansion. Innovative processes are being identified to maximize pit mining recovery and to minimize waste stream losses. This results in more effective use of the phosphate. Texasgulf Inc., is a wholly owned subsidiary of Elf Aquitaine. The Aurora phosphate rock mine and refining facilities are located in North Carolina on the south shore of the Pamlico Sound. Phosphate ore is recovered by open- cast mining techniques. Draglines and other earthmoving equipment are used to uncover and mine the phosphate ore. The mined ore then is slurried and pumped about 8 km (5 miles) to the processing plant.

Jan 1, 1994

By R. O. Burt

The paper examines the role of gravity concentration in both an historical contest and within the modern mineral processing industry. It discusses the principles of the process on both a theoretical and practical level. The importance of ore mineralogy and its effect on the choice of the wide range of equipment available is discussed. The paper concludes with a discussion on the relative merits of gravity concentration to other forms of processing and seeks to answer the classical question - "Quo Vadis?"

Jan 1, 1982

By Yoginder P. Chugh

About 65% of the U.S. coal production (835 n tons was surface mined. A similar proportion of the Illinois basin&apos;s production (112 m tons) is also surface mined. The state of Illinois is the third largest producer of surface mined coal in the United States, after Wyoming and Kentucky. The Illinois coal basin encompasses Illinois, southwestern Indiana, and western Kentucky, as shown in Figure 1 (dotted lines). The estimated surface mineable coal resources in the basin amount to about 27 billion tons, and about 69 m. tons of coal is surface mined annually (Tab1e I).

Jan 1, 1981

By M. J. Friedel

The U.S. Bureau of Mines (USBM) investigated the use of seismic refraction tomography (SRT) for identifying changes in pillar loading at the Foidel Creek coal mine near Steamboat Springs, CO. The inherent three-dimensional problem of directly imaging coal pillar conditions is reduced to a single plane by considering only the load transmitted from the pillar to the adjacent floor rock. Floor strata velocity tomograms (at two pillar locations) illustrate that loading conditions are non-uniform and non-unique. Also, time sequence monitoring, as the longwall face advanced, demonstrated that SRT can be used to image spatial changes in loading conditions. SRT offers the mining engineer a simple and flexible technique for evaluating pillar conditions, and thus, both present and alternative mine designs.

Jan 1, 1994

By H. G. Kok

The concept of mobile crushers is not new. Since their introduction about twenty-five years ago, mobile crushers have continued to gain widespread attention. Mobile crushers reduce minerals in the open pit mine to sizes which can be transported by belt conveyors thereby reducing or eliminating the need for truck haulage. Truck haulage is costly, often representing as much as 50% of the total mining cost. More than 80% of the cost of truck haulage is for fuel and lubricants, tire replacements, and operators&apos; wages, and it can be anticipated that the cost of these items will continue to increase at a steady pace. Compared to truck haulage, transportation by belt conveyor is more energy-efficient, less labor-intensive, and less subject to inflation. Therefore, mobile crushers, used in conjunction with belt conveyor systems, offer one of the most effective available methods of reducing energy consumption and mining costs,

Jan 1, 1980

By R. L. Loofbourow

R. L. LOOFBOUROW 26.1-WATER FLOW, USE, EFFECTS, CONTROL Sec. 4 discusses ground water as a mineral resource. This section deals wits its control. 26.1.1-IMPORTANCE Flows of water have an important effect on the cost and progress of many deep excavations. The existence of water limits the methods which can be used in some and presents hazards in others. Lowering the water table may reduce the flow of springs and wells, with legal liability. Some mine water is used directly, some must he treated before discharge, and some contains dissolved solids which can be recovered profitably. Loss of life, property and production from flooding is grim, yet scores of mines have worked safely for generations under ground-water reservoirs, lakes and even the sea. The need for better control of mine water is measured by the severity of recent floodings, concern with the quality of the effluent, and direct and indirect costs of working wet ground. There are improvements in freezing in shaft sinking. pumps and their controls, treatment of mine water before it is pumped, bulkhead construction and certain efforts to reduce inflow. 26.1.2-PUMPING COST VS. COST OF WET WORK Direct operating and capital costs of pumping are more evident than indirect items, such as provision of extra standby power and maintenance facilities, The total of all these items is the cost of pumping water from an excavation. This is only a part of the extra cost of working wet as compared to dry ground. Less productive methods, less efficient equipment and more expensive explosives may have to he used in wet work. Wet rock plugs certain equipment. Maintenance is higher and labor less efficient. The cost of any means of keeping work drier should he weighed against the higher cost of working wet. The true difference in the cost of working wet ground, compared with dry, probably is known at only a few places (see checklist of cost items. Sec. 26.4.13). An extreme comparison of mining wet and dry ore is available from Ambrosia Lake, N.M. (Fig. 26-1). Between 1963 and 1967, a wet and a dry mine were worked by the same company under conditions otherwise comparable. Average operating cost at the wet mine, pumping 1,600 to 1,800 gpm. was $16.06 per ton of ore. The comparable figure for the dry mine was $7.50. Principal differences included the fact that the wet mine was worked from drainage levels below the, ore with rail haulage, the dry mine in the ore with off-track equipment loading and hauling to the shaft. In the wet mine, ore packed in chutes, cars and skips and froze in surface bins, stockpiles and trucks. Wet-mine costs included drilling drain holes, building and maintaining ditches and pumping water containing abrasive solids.

Jan 1, 1973

By Paul I. Eimon

Mineral exploration involves the selective collection, organization, evaluation, and interpretation of geologic and economic data. The sources for such data are original field observations and existing documented information of various types (publications, reports, maps, company files, etc.). As exploration data are accumulated, their storage, organization, indexing, and J retrieval is key to exploration progress. Initial use of computers by mineral exploration groups indicates that data can be meaningfully handled by machines. A computer system makes possible: 1) the storage, manipulation, collation, and selective retrieval of data; 2) the utilization of data banks developed by geologic service groups and available at relatively low cost; and 3) the retrieval of stored data more rapidly, completely, and effectively than by manual systems. Profitable utilization of these capabilities is conditioned by several factors, particularly by the information selected for the system data base and by the design of the computer system itself.

Jan 1, 1968

By Karoly C. Kocsis, Kyle A. Scalise, Marcelo Bandeira Teixeira

Understanding the effects of various ventilation and climatic parameters on the work conditions and comfort in underground mines is critical for efficient ventilation system design and cost savings and to ensure the health and safety of mine workers. To understand the effects that the thermal damping effect (TDE) and the thermal flywheel effect (TFE) have on the ventilation system design and potential cooling system, this paper compares and analyzes the wet and dry-bulb temperatures at the bottom of an intake shaft using two modeling software packages, Ventsim™ and Climsim™. The comparison shows the consequences and importance of taking into account the TDE/TFE when predicting the climatic conditions in future underground mines, especially when deciding on whether a cooling system should be employed in order to provide adequate climatic conditions in the production stopes, dead-end development headings, and throughout the mine. Ventsim™is able to account for the TFE, while Climsim™does not. Both software packages have their uses and are used within the mining industry, but it is useful to understand their limitations and where the future of underground climatic modeling will lead.

Oct 9, 2020

By J. D. Vincent

The Henderson mine site is located at Red Mountain on the eastern slope of the Rocky Mountains, approximately 50 miles west of Denver, Colorado. The mill site is on the west slope of the Rockies and mine and mill will be connected by an underground railroad. All designs were for a 30,000 ton per day molybdenite processing plant with provisions for future expansion up to 50,000 ton per day if required. This plan did not just happen, but slowly evolved over the course of about three and one-half years. Many plants cannot afford such a delay, but the Henderson project will take about 10 years from the time it was first discovered until production starts, and these 3 !1 years were well spent in determining the best possible plant site and flowsheet. This paper describes some of the problems and the decisions made in arriving at the final location and flowsheet. The paper does not go into the specific details of how protection of the environment went hand-in-hand with all other project considerations. This particular aspect of the planning has been well-reported on other occasions and considerable information on it is readily available.

Jan 1, 1972