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By S. Cotten

The regulations for current coal mine refuge alternative (RAs) require that there be sufficient life support to sustain survival of the maximum number of RA occupants for a minimum of 96 hours. This minimum duration assumes that rescuers can access the RA through the mine within four (4) days of the onset of the mine emergency. But what if mine rescue teams cannot reach the miners within the RA before the RA air supply is severely limited or exhausted? What could be done to extend the duration of the RA life support capability into the 97th hour and beyond? In addition, what might be done to enable more rapid rescue of the miners if they could not be reached through the mine in a timely manner? This paper describes a systems approach to address the potential need to extend the duration of RA life support beyond current design limits. New capabilities that need to be developed in advance of such a dire need include more reliable emergency communications, highly mobile and efficient drilling rigs, enhanced directional drilling techniques, and a standardized RA external life support interface along with RA internal design modifications to accommodate use of an external supply. The following sections describe the need, intent, and technology gaps associated with each of these capabilities. Government agencies and the mining industry should consider initiation of a program to investigate, design, and develop effective means of extending the current 96-hour time restriction on coal mine refuge alternative life support. The development process should begin with a focused study to refine the overall operational concept, investigate the current state-of-the art of critical technologies, identify specific technology development priorities, and make recommendations for a development program to provide in an organized, efficient, and timely manner a solution for extended refuge alternative life support and occupant rescue.

Jan 1, 2011

By M. A. Trevits

Mine seals are used in underground coal mines throughout the United States to isolate abandoned mining areas from the active mine workings. The objective of adding inert gas to a sealed mine area is to quickly reduce the oxygen concentration in the sealed area to a level that will not support combustion. In the US, this process is typically conducted from the ground surface and because surface-to-mine access is constrained by so many variables, it is very difficult to provide a source of inert gas to the place where it is needed underground. To alleviate this problem On Site Gas Systems, under a NIOSH contract, built a nitrogen (N2) generation plant that operates in the underground environment. The purpose of this work was to create a reliable in-mine mobile plant that would extract N2 gas from the mine atmosphere and use the gas to create and maintain a safe sealed mine area. The plant is based on a novel design using pressure swing adsorption technology and is sized to fit on a standard shield car or mine dolly for easy transport in and around an underground coal mine. The plant is capable of producing about 900 scf/min of N2 gas at ambient mine conditions. After an exhaustive series of laboratory tests, the plant was field tested at the NIOSH Safety Research Coal Mine (SRCM) where a portion of the mine was sealed and rendered inert. During the test, the movement of the injected gas was monitored at several underground points to observe the progress and extent of the inerting process. This paper describes the plant design and the results of the field test at the SRCM. Disclaimer: The findings and conclusions in this report are those of the authors and do not necessarily represent the views of NIOSH.

Jan 1, 2009

By J. M. Somers

Methane is a powerful greenhouse gas and the principal component of natural gas. Coal seams often contain significant quantities of methane, and underground coal mines must ensure that methane released into the mine during coal extraction does not build to dangerous levels. This is accomplished in part through the use of large-volume ventilation systems that remove methane from the mine and release it to the atmosphere. Although the methane concentration exhausted is quite low (typically <1 percent), the volume of air that ventilation systems move is so great that they actually constitute the largest source of methane emissions from underground coal mines. Each year underground coal mines throughout the world emit more than 500 billion cubic feet of methane from their ventilation systems. An air pollution control technology, thermal oxidation using a flow-reversal reactor (TFRR), has emerged as a potential solution to ventilation air methane (VAM) emission mitigation. One manufacturer of TFRRs has demonstrated its oxidizer design (the VOCSIDIZER?) at coal mines in the United Kingdom and Australia. These demonstrations tested the technology?s effectiveness at oxidizing low-concentration methane, as well as its ability to tap the excess heat to produce steam for electric power generation. In early 2007, the first U.S. demonstration of this technology began operation at an abandoned mine in West Virginia. The project is designed to prove the technology?s operational robustness, its ability to reliably oxidize methane at concentrations typical of mine exhausts, and its safety. If the technology is employed at active underground coal mines, it offers the potential to mitigate substantial quantities of global methane emissions. This paper discusses global VAM emissions, options for the recovery and utilization of VAM, and existing projects such as WestVAMP in Australia and the ongoing U.S. VAM demonstration project.

Jan 1, 2009

By Micheal Mitariten

The recovery of natural gas (primarily methane) from coalbed deposits is an established and growing source for the U.S and accounts for over 7% of its total production. The gas from on-purpose degassing is often pipeline quality and requires only compression and dehydration before being admitted to the pipeline grid. However, the raw gas is sometimes contaminated with nitrogen (N2) and/or carbon dioxide (CO2) that can make the gas inadmissible to the transportation pipeline. In general, the N2 must be 4% or less and the CO2 should be 2% or less. In treating gas produced from virgin coal bed or from degassing operations where CO2 is present, amine systems are traditionally used for its removal. Engelhard’s Molecular Gate® adsorption system offers an attractive alternative due to its skid-mounted equipment, low pressure and unattended operation, cost and high on-stream reliability. Natural gas can also be sourced from the residual coal, after the coal has been mined. This source can be the GOB or gas remaining in abandoned coal mines. This source is almost universally contaminated, often with a combination of CO2 and N2. If the source is from air intrusion, the source can also be contaminated with O2. In treating CO2 / N2 contaminated feeds, the Molecular Gate adsorption system removes both of the impurities along with a level of O2 within a single processing step. To date, ten Molecular Gate units are underway with three systems having been installed for upgrading gas from abandoned coal mines. The technology consists of specialty adsorbents and advanced PSA processes that are jointly offered as fabricated packaged equipment by Engelhard Corporation and Guild Associates, Inc. Upgrading requirements, project challenges and project arrangements are discussed in this paper.

Jan 1, 2005

The extraction and processing of minerals from surface mines and quarries can produce significant fugitive emissions as a result of site activities such as blasting, unpaved road haulage, loading, primary crushing and stockpiling. Uncontrolled fugitive dust emissions can present serious environmental, health, safety and operational issues impacting both site personnel and the wider community. Simulation technology is finding increasing use for the purposes of advanced warning of potential problem emissions in addition to providing a basis for future planning applications where demonstrable compliance with regulatory requirements are necessary. The initial re-entrainment and subsequent dispersion of fugitive dust presents a process complicated by the combination of the in pit topography, the surrounding natural topography and the dynamic nature of emissions from these sites. These factors impact upon the accuracy and reliability of the conventional Gaussian plume based computational prediction methods employed for regulatory compliance and IPPC applications. This paper proposes that optimal modelling of open pit emissions may be more accurately achieved by the use of a multi-scale predictive modelling approach utilizing computational fluid dynamics (CFD) methods for high resolution near source dispersion and conventional Gaussian based methods for far field dispersion modeling. This paper presents a numerical based flow and dispersion analysis of a typical UK based open pit utilizing CFD in conjunction with a conventional Gaussian plume based methods. Typical operating emissions and meteorological conditions are obtained from long term data records collected at a large operating quarry extraction operation in the UK. Emissions are modelled using a Lagrangian framework within conventional atmospheric boundary layer (ABL) profiles expressed as functions of turbulence and velocity parameters under assumed neutral conditions. Results are presented in terms of the impact of site topography on in pit retention as compared to the Gaussian based method.

Jan 1, 2009

By Thomas V. Falkie

In bygone days, the mineral engineer spent most of his/her time on the technical aspects of designing, building, and managing mineral operations. In the last few decades, mostly since the 1960s, increasing amounts of engineering man power and time are spent dealing with federal, state, and local government agencies, laws, and regulations. It is not possible to plan, design, build, or operate mines, mills, smelters, or even office buildings without detailed consideration of government regulations and permitting requirements dealing with environment, worker health and safety, etc. The four chapters of this section provide an overview of the social, legal, political, environmental, and health and safety aspects of mineral enterprise. The section sets the stage for the working engineer/geologist in industry, government, consulting, or academe and points the direction for more detailed study in each of these areas.

Jan 1, 2008

By W. G. Pariseau

This paper is concerned with the problem of calculating ground movement that occurs as a consequence of mining. Solution to this problem is straight forward in principle but rather complex in practice. The complexities of the problem arise from the ever present constraints of technology and economics. However, the identity of these constraints is often obscured by a faulty understanding of the fundamentals of the problem and a lack of appreciation of current computational capabilities and costs. As a consequence much time and effort are sometimes wasted on gathering information that is of little direct value to ground control programs while important input data are neglected. This paper therefore reviews the essentials of the process of formulating the ground movement problem. Emphasis is upon application to longwall mining, although much of the discussion applies to other mining methods as well. In many instances, the hardrock term stope could be substituted for longwall panel because the fundamentals of the problem are independent of the excavation dimensions and numerical values of rock properties. The latter affect the outcome of a calculation, of course, but not the underlying formulation. Only the major features of the problem are discussed in this brief review. The objective of the discussion is the generation of a reliable basis for developing a technically sound ground control strategy. Any rational ground control strategy must be consistent with basic physical laws. For example, under static conditions, equations of equilibrium must be satisfied. Operational, that is, tactical, details may vary widely with local conditions, but the basic approach must be based on physical laws. The alternative empirical route is intrinsically unreliable and unnecessary as well. Empirical descriptions of ground motion caused by mining including recent statistical curve fits to specialized data sets (for example, surface subsidence data) as well as antiquated "theories" of ground pressure and so forth, however dignified with mathematical formulas, are hardly worthy of the label "theory". Extrapolation of statistical data is always accompanied by rapidly diverging confidence limits, while empirical equations simply lack predictive capability under circumstances even slightly different from those of the original development. For example, applications of surface subsidence curves generated empirically in the United Kingdom to coal mines in the United States generally result in considerably greater subsidence and wider subsidence troughs than actually observed. Locally, empirical criteria for specific ground control actions may be made practical through trial and error, but such tactics are no basis for the development of a general ground control strategy. In this regard, it is worthwhile to recall the traditional scientific distinctions between speculation, hypothesis and law, between the empirical and the proven.

Jan 1, 1982

By Robert B. Crookston

This paper presents a resume and status report: of mining technology appropriate to commercial oil shale development in the United States today. The nature and extent of oil shale in Colorado and Utah are discussed. A look at most of the prominent activities now underway emphasizes the questions "where are we now?" and "where are we going?" Various alternate mining methods are reviewed and a bibliography of oil shale mining is included.

Jan 1, 1980

By Lawrence A. Engel

The idea of using apron feeders for in directly feeding crushers is not a new one, it has been used for many years. To the best of our knowledge, the first apron feeder was used for bulk material handling in 1893. Since that time many design innovations have been introduced, with perhaps the most effective one being, the application of crawler tractor components in the feeder construction. This design was first introduced to the mining industry in 1954 by, national iron company. These feeders were used in the handling of run-of-mine iron ore on the Minnesota iron range, at the rate of 1,000 ltph. The material was fed over a grizzly to a jaw type of crusher. The crusher output was also handled by an apron feeder. Today, the apron feeder is being used in hard rock surface mine operations feeding crushers at rates of 5,500 ltph and over.

Jan 1, 1983

By John A. Cherry

Groundwater monitoring conducted at an acidic uranium-tailings impoundment in Wyoming and at one in Ontario indicate that at these unlined impoundments situated on unconfined sand aquifers, there exist well-developed plumes of contaminated groundwater. In the literature similar results are reported for three other unconfined aquifers at mill sites in Wyoming. The plumes extend for distances of several hundred to a few thousand metres from the impoundments. They have a zone of low pH beneath and adjacent to the impoundments and a larger zone of near-neutral pH that extends to the outer limits of the plume. The low-pH zones are generally small compared to the neutral-pH zones because the advance rate of the fronts of acidic tailings-derived water is slower than the average velocity of groundwater which controls the advance of the high pH zone. The retardation of the acid fronts is a result of geochemical reactions that cause hydrogen-ion consumption in a slowly-advancing neutralization zone. Within each of the plumes, the low-pH zone has high concentrations of transition metals, heavy metals and radionuclides. In the neutral-pH zones all of the metals and radionuclides occur at significantly lower concentrations and are generally at levels much below those specified in national drinking water standards. An exception at two of the Wyoming sites is uranium, which exists at levels close to the drinking water limit. A second exception is ferrous iron at the Ontario site, which occurs at high concentrations throughout most of the plume. The carbonate alkalinity at this site is much lower than at the Wyoming sites. The field data indicate that even in very permeable aquifers with a low content of silt and clay, there is normally a considerable capacity for acid-front retardation. To take advantage of this capacity to control contaminant movement in the design, operation and decommissioning of tailings impoundments, it is necessary to be able to predict the rate of acid-front movement. This predictive task has two primary components. The first is to identify the important flow paths from the tailings through the groundwater system and the second is to determine the degree of acid-front retardation that will occur as the tailings-derived water moves along these flow paths. Relative to these two tasks determination of the effects of hydrodynamic dispersion is generally of lesser importance and simple approximations that do not involve field or laboratory tests are probably adequate. In order to effectively apply the acid neutralization concept to the prediction of subsurface contaminant migration at uranium-mill tailings, improved representations of hydrogeochemical behavior of iron and aluminum in the neutralization zone and of the consumption of hydrogen ions through reactions with carbonate and aluminum-silicate minerals are required. Along with the hydrogen ions contained in the tailings water the ferrous iron represents a large potential source of acid but the extent to which ferrous iron converts to ferric iron to produce acid is problematic. Detailed hydrogeochemical evaluation of existing contaminant plumes and the performance of laboratory experiments under appropriate redox conditions ore prerequisites for the improvement of predictive models. However, the existing field and laboratory data provide a basis for preliminary predictions of contaminant migration from acidic uranium tailings impoundments.

Jan 1, 1982

By James J. Anthony

High-angle in-pit conveying is an idea that is being seriously studied by a number of important mining companies in North America. Its application in coal, iron ore, and copper mining is expected to reduce mine operating costs considerably compared with the conventional discontinuous methods, such as truck haulage, which are fuel, labor, and maintenance intensive. Therefore, a mine introducing high-angle in-pit conveyors will gain a competitive advantage over mines which continue to use conventional haulage. While the capital outlay for in-pit conveyor systems obviously varies depending on pit depth, the angle of incline, and the capacity required, it generally will not exceed that of truck haulage when all Factors are taken into consideration. These factors include the cost of excavation, as measured in relation to the volume of minerals mined, the cost of building truck ramps and roads, and the capital outlay for the trucks and maintenance facilities. In many cases, in Fact, the capital outlay will be less than for truck haulage. This cost advantage will best be realized by new mines, whereas in existing producing mines the switch from truck haulage may involve the abandonment of a substantial capital asset.

Jan 1, 1983

By L. Maday, V. J. Perrone, M. B. Gilbert

INTRODUCTION The Brightwater conveyance system project in King County, Washington consists of constructing about 21 km (13 mile) of effluent and combined influent and effluent tunnels. The influent tunnels will collect sanitary sewerage from the service area and convey it to the new wastewater treatment plant (WTP) site at Route 9 in Woodinville, Snohomish County, Washington. The effluent tunnels will convey treated effluent from the WTP to an outfall at Puget Sound. The tunnels will be constructed primarily under residential areas, busy suburban roads and state highways. This 21 Km (13-mile) soft ground tunneling project faced many design challenges because of the geologic and hydrostatic conditions, long tunnel reaches, and large overburden depths. Boring depths typically varied from 45.7 M (150 ft) to 114 M(375 ft) with a maximum depth of 185.6 M (566 ft) deep. Conventional empirical methods for non-mountain terrain based on a 1:1 ratio of drilled footage per foot of tunnel bore (National Research Council, 1984) for evenly spaced borings provided a good start for the Phase 1 explorations. However a continuation of this approach with closer spacing was deemed inadequate by the geotechnical team for efficiently allocating geotechnical exploration resources for the final design. This paper presents a rational approach that was used to evaluate the value of a particular geotechnical exploration boring, pump test or other in-situ test. The method of analysis, which was initially developed for rock tunnels (Einstein et al., 1978), was modified for the Brightwater project. The method considers the potential construction cost risks resulting from variable geologic conditions, the cost and reliability of the geotechnical exploration and the earned value (cost risk reduction) of performing the geotechnical exploration. PROJECT DESCRIPTION In total the land portion of the conveyance system will include approximately21 Km (13 miles) of tunnels ranging in size from 3.0 to 5.2 M (10 to 17 ft) minimum inside diameter with overburden depths of 30.0 to 168 M (100 to 550 ft) and

Jan 1, 2005

By Robert J. F. Goodfellow, William A. T. Meakin, Kevin Clarke, Eric Cole

INTRODUCTION New York City Water Supply—Background and History Since the 1840s New York City has been developing the largest, and perhaps, the best urban water supply and distribution system in the world in terms of quality and reliability. One of the engineering marvels of the modern world, the New York City water supply and distribution system serves a population of almost nine million, including residents in nearby Westchester, Putnam, Ulster, and Orange Counties of New York State. The average daily supplies presently are running at almost 60 m3/sec(about 1,350 million gallons per day). The watershed for the New York City water supply system, under the jurisdiction of the New York City Department of Environmental Protection (NYCDEP), consists of19 reservoirs and three controlled lakes, and covers an area of approximately5,000 square kilometers (about 2,000 square miles) north of New York City. The watershed total available storage capacity is about two billion cubic meters (560 billion gallons) and is divided into three reservoir systems: the Catskill and Delaware watersheds located West of the Hudson River; and the Croton watershed, located east of the Hudson River, see Figure 1. The first element of the system, the Croton Aqueduct, started operations in 1842.Through several updates and expansions, the present-day Croton system was in place by 1911, providing a potential maximum supply of about 15 m3/s (about 340 mgd). The system includes ten reservoirs and three controlled lakes. In the late 1800s, with the tremendous growth of the City, New York City water engineers began to look for additional sources. Between 1907 and 1929, the City acquired water rights and constructed a new 148 km (92 miles) Catskill aqueduct to convey water from the Ashokan Reservoir to the City. The Catskill system also included the construction of the Kensico and Hillview Reservoirs just north of the City,

Jan 1, 2005

By Jayant Khambekar

In a typical mining operation, run-of-mine (ROM) ore is crushed and screened to achieve a desired particle size distribution and then fed to processing equipment for further refining. For an operation to be profit-able, the ore handling system from the mine face to final product loadout must be reliable. However, the handling system components do not always perform as expected and production problems related to material flow occur. These problems include flow stoppages, limited live storage capacities, uncontrollable flow, segregation, spillage and dust emissions. These result in lost production, excessive maintenance, safety and environmental issues, and even structural failures, in some cases. Flow patterns Flow problems and their solutions can be best understood by first learning how bulk solids flow. As bulk solids discharge by gravity from a storage system, primarily two types of flow patterns can develop: funnel flow and mass flow. Figure 1 shows a schematic of these two flow patterns. In funnel flow, only a portion of the material is in motion during discharge, which flows toward the outlet through a channel that forms within the stagnant material. Funnel flow occurs when a hopper is not sufficiently steep and smooth to ensure flow along the walls or when the outlet is not fully effective. In mass flow, all of the material is in motion whenever any is discharged ? there is no stagnant material. Mass flow occurs when the hopper is sufficiently steep and smooth to ensure flow along the hopper walls. Shallow valleys are not permitted and the outlet must be fully effective. A third type of flow pattern, called expanded flow, can develop when a mass flow hopper (or hoppers) is placed beneath a funnel flow hopper or bin. The mass flow hopper is designed to activate a sufficiently large flow channel in the funnel flow hopper to avoid flow problems due to a rathole. The flow pattern in which a bulk solid discharges from a storage system strongly influences the flow problems that may occur in that facility.

Jan 1, 2009

By J. D. Patsey, G. T. Lineberry

Introduction A computer program was developed for analyzing drum-type hoists before modifying an existing system or designing a new one. Its use permits the preliminary evaluation of a system before seeking technical assistance from mine hoist designers and manufacturers. The user-friendly program accepts a variety of data and analyzes hoisting systems for any balance state. The program can be used to calculate skip capacity to yield a desired production rate, solve for drum face width, select a permissible wire rope, estimate total horsepower of the hoist plant, and estimate annual power cost. Whether required for a large mining complex or for a relatively small operation, a hoisting system must be carefully designed to ensure the efficient, reliable, and safe flow of material. Because shaft sinking and hoist installation can total between 2.5% and 3% of the cost of opening a deep mine, proper hoist selection is critical. Today&apos;s mining engineer has at his disposal the most powerful design and analytical aid ever, the microcomputer. There is, unfortunately, limited software for the study of hoisting systems, unlike that for other materials handling equipment (Manula and Albert, 1980; Prelaz et al., 1964; Bucklen, 1969; Thompson, 1985). HOIST reduces a time-consuming set of calculations to a concise package of interrelated subprograms. A literature review revealed no common-user programs to analyze hoisting systems, although at least four major hoist designers/manufacturers/installers have their own in-house programs. To provide a tool with which the mining engineer could preliminarily analyze a materials hoisting system or could check the calculations of a hoist contractor, a computer program was developed. HOIST was written in BASIC for the IBM-PC for ease of program adaptation and to encourage field use on compatible systems. Details of program development are omitted, since the basic principles of hoisting analyses are relatively straightforward, simple, and readily accepted (Harmon, 1973; Nordberg, n.d.; Adler, 1957). Program features, intended usage, and benefits of the computerized solution are emphasized over theory development and mathematical rigor. Background The mine hoist system that is selected and installed at a mine is the "lifeline" of that mine, with installations lasting 20 years or more. Thorough study is warranted to ensure that productivity demands are met at a minimum cost per ton. The increased cost of a large, powerful, high-speed hoist must be offset by increased production to justify its selection. To optimize this tradeoff, an extensive hoisting analysis should be performed. The analyses to properly size the skip (or cage), the drum, and the hoist drive are conducive to computerization, permitting rapid evaluation of changeable operating and design parameters, such as velocity, acceleration, state of balance, and productivity demand. The program is particularly useful in conducting sensitivity trials, such as investigation into the effect of change of productivity on skip capacity and on horsepower of the hoist drive. HOIST is currently limited to the study of drum-type hoists with cylindrical drum(s). However, only minor changes to the program would permit analysis of friction hoists and conical drum configurations. Model development and testing The program is based on accepted equations and physical relationships. Examples of manual calculations formed the basis for decision points and program branching. Data is input in the order that it would be needed if the problems were solved manually. The choices are arranged likewise. HOIST was developed in sections, with manual solutions performed to check program logic. The testing became more rigorous as sections were completed. Output from one section becomes input for following sections, as appropriate. The simplified flow diagram of HOIST is given in Fig. 1.

Jan 2, 1987

By Xiaodong Ma, Marek Pawlik

Introduction Guar gum is a natural nonionic polysaccharide. Guaran, the functional polysaccharide in guar gum is a chain of (1?4)-linked ß-D-mannopyranose units with a-D-galactopyranose units connected to the mannose backbone through (1?6) glycosidic linkages. The poly-mannose chain is randomly substituted with galactose units at a mannose-to-galactose ratio of 1.8-1.0 [1]. Cheng et al. reported that the molecular weight of natural guar gum is 1.935 • 106 [2]. Robinson et al. found that the molecular weights of five grades of commercial guar gums ranged from 4.4 • 105 to 16.5 • 105 [3]. Guar gum is frequently used as a depressant of naturally hydrophobic gangue minerals (e.g. talc) in the flotation of PGM-bearing ores, and according to Harris et al. a wide variety of modified guar gums are available for this application [4]. The Canadian potash industry utilizes natural guar gum as a blinder of water-insoluble slimes in the flotation of potash (sylvinite) ores, a process that is carried out in a saturated KCl/NaCl brine. The role of the polysaccharide is to prevent a primary amine collector from adsorbing onto these unwanted fine minerals. Hydrophobic interactions between basal cleavage planes and the polymer appear to be involved in guar gum adsorption on talc [5,6]. Liu and Laskowski stressed the importance of metal impurities on mineral surfaces that can serve as the active sites for polysaccharide adsorption [7]. This postulate was consistent with the later observation that, in addition to hydrogen bonding, the adsorption of guar gum onto talc was strongly dependent on the surface concentration of magnesium sites [8,9]. In this case the adsorption process was visualized as metal-polysaccharide complexation facilitated by the cis-configuration of the hydroxyl groups of mannose monomers. Hydrogen bonding and chemical interactions with metal sites were also suggested by Rath and Subramanian in their studies on guar gum adsorption on biotite mica [10]. In a recent study, Wang et al. demonstrated that hydrogen bonding is the main driving force behind guar gum adsorption onto talc [11]. Their data also show that the adsorption density of guar gum on alumina is about five times lower than the adsorption density on talc. Pawlik and Laskowski reported that the amount of guar gum adsorbed on illite was independent of ionic strength over a wide range of salt concentrations [12]. The adsorption density of the polysaccharide on illite did not measurably change with ionic strength from that of distilled water to about 50%-saturated KCl/NaCl potash brine (~3N KCl/NaCl). However, guar gum adsorption on dolomite dramatically decreased over the same range of ionic strengths, and no reliable explanation of that phenomenon was proposed [12]. The water structure-breaking or structure-making capabilities of background ions are usually ignored when discussing adsorption mechanisms of polymers. The concept of “structure breakers” and “structure makers” was first introduced by Gurney, and Frank and Wen who postulated that each ion is surrounded by three distinct regions of water structure [13,14]. In the first layer, water molecules are tightly bound to the ion. The second region extends farther away from the ion and is referred to as the region of structure breaking. Only at larger distances, where the ionic field is weak, water molecules form the “normal” ice-like structure. It is also now recognized that the first layer can be further divided into the primary and secondary hydration shells. Small, strongly hydrated ions reinforce the “normal” structure of water and the region of structure breaking disappears. In contrast, large less-hydrated ions disturb the ice-like structure and generate an extensive region of structure breaking. Experimentally, these two effects can readily be distinguished through viscosity measurements. Increasing concentrations of structure-breaking ions result in a decrease of the viscosity of aqueous solutions. On the other hand, increasing amounts of structure-makers significantly reduce the viscosity of water. Alkali metal cations form a sequence known as the Hofmeister series. The Hofmeister series - Cs+, K+, Na+, Li+ - orders the cations from the least hydrated to the most hydrated. Based on literature viscosity data for aqueous

Jan 1, 2005

By David H. Carlson

The paper presents background information including data on diesel exhaust pollutants, amounts produced and standards, a desscription of mine characteristics that affect air monitoring techniques and a review of portable measurement techniques and instruments for gas and particulate matter sampling. A general review of the laboratory and field research being conducted at Michigan Technological University in the area of diesel emissions is presented. The facilities involved with this research are also described. These are the Experimental Mine, the diesel dilution tunnel test cell, the Mine Air Monitoring Lab (MAML), the portable instruments and the production mines used.

Jan 1, 1979

By H. Walqui

This paper describes operating gains resulting from standardizing operating and maintenance practices to improve the performance of the pebble crushing circuit at he Empire Concentrator. Past efforts have led to improved operating time but no additional throughput gains. New Operating standards and checklists were developed and introduced. The content and frequency of the review process indicators was then changed to take advantage of the new practices resulting in an increased feed rate and availability through the pebble crushing circuit.

Feb 27, 2013

By J. R. Riding

In situ solution mining of uranium has environ- mental advantages over conventional mining. The leaching of uranium, however, alters the groundwater quality in the aquifer where the mining occurs. Currently, regulations require that the groundwater be restored to approximately its original condition. This paper, a summary of a study funded by the Bureau of Mines, reviews the state of the art in restoring groundwater quality after in situ uranium leaching. Current restoration practices discussed include disposal of liquid wastes in deep disposal wells and evaporation ponds, producing from all wells during restoration, and recirculating water purified in surface plants. Methods for predicting the effectiveness and cost of current techniques are presented. Possible alternative techniques are also described. Two restoration operations are discussed.

Jan 1, 1979

By M. J. Sapko

The design requirements for mine ventilation seals have undergone a radical transformation in recent years, and these revisions have greatly increased the cost of the seal designs and their construction. For example, in the past two years, new federal regulations have increased the minimum design requirement to withstand explosion pressures from 138 kPa (20 psig) to 345 kPa (50 psig) or 827 kPa (120 psig), depending on whether the sealed mine volume is monitored continuously or not. Moreover, there is still a possibility that under certain conditions (such as detonations) even higher pressure requirements may be necessary. The ability of a monolithic, stand-alone mine seal to reliably withstand the full range of current and future pressures is becoming increasingly difficult and expensive. In an effort to develop a practical alternative, the West Virginia Office of Miners? Health Safety and Training (WVOMHST) and the National Institute for Occupational Safety and Health (NIOSH) have collaborated in a research effort to develop a practical, economic, and safe mine sealing technique that can enable mines seals to meet the full range of new explosion pressure design requirements. The basic idea is to use a barrier of common mine gob and rubble in combination with a conventional mine seal so that the pressure resulting from a gas explosion is reflected, absorbed, and attenuated, so the pressure on the ventilation seal is reduced. This paper discusses some concepts and preliminary test results for a ?Passive Mine Blast Attenuator? (PMBA) that can provide a useful alternative to increasingly larger and stronger stand-alone mine seals. Numerical models and full-scale experiments conducted at the NIOSH Lake Lynn Laboratory Experimental Mine (LLEM) show that the use of a PMBA can signficantly reduce the blast pressure and impulse on conventional ventilation seals. DISCLAIMER: This information is distributed solely for the purpose of pre-dissemination peer review under applicable information quality guidelines. It has not been formally disseminated by the National Institute for Occupational Safety and Health (NIOSH). It does not represent and should not be construed to represent any agency determination or policy.

Jan 1, 2009