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By B. Mather

I was very glad to see an article on alkali-silica reactivity of concrete aggregate in Mining Engineering. It is a topic that should be reexamined every 45 or 50 years. The last paper (that I am aware of) that examined this topic in Mining Engineering was Rexford (1950), who, at the time, was chief of the Petrography Section, South Pacific Division Laboratory, Corps of Engineers, Los Angeles, CA. The paper was presented at the AIME meeting in Los Angeles in October 1948. Professor Leming's discussion generally reflects the current state-of-the-art; but it provides no references for more detailed information. I wrote a paper titled "How to avoid excessive expansion of concrete due to alkali-aggregate reaction," which was included in the Proceedings of the Second International Conference on Alkali-Aggregate Reactions in Hydroelectric Plants and Dams (Mather, 1995). In it, I emphasized that it was the hydroxide ion (not the alkali ion) that caused the pore fluid in the concrete to dissolve reactive silica, producing alkali-silica gel capable of taking up water, thereby, causing swelling and rupturing of the concrete. Excessive expansion will not occur unless there is a sufficient amount of potentially expansive alkali-silica reaction product and water. However, this will not occur unless there is a sufficient quantity of reactive aggregate and unless there is a mechanism by which the pH of the pore fluid can get well above the normal value of about 12.6. In many cases it is sufficient that the cementitious medium not only contains portland cement but also contains ground granulated blast-furnace slag (GGBFS) or a pozzolan such as fly ash, silica fume, metakaolin, volcanic glass or calcined shale - any of which may have been included for economic reasons. If, however, the person selecting the ingredients for concrete knows that the aggregate is reactive (with no alternate nonreactive aggregate available) and if the person knows that the only available portland cement has a high alkali content (i.e., more than 0.6% Na2O.; Na2e = % Na20 + 0.658 x % K20), then it becomes irresponsible not to establish what reasonable precaution one should take. Professor Leming suggests using one of the following: 30% Class F fly ash, 50% GGBFS or 7% to 10% silica fume. My view is that one should prepare test specimens with the cement and aggregate to be used and then expose these both with and without at least two dosages of each of the various alternative materials (slag, pozzolan and lithium salts), so that the required dosage. to control the expansion is established. Then, economic considerations can dictate the final selection. I would point out to the author that GGBFS is not a pozzolan and is not an admixture, but, rather, it is a latent hydraulic cement. An admixture is, by definition, a material other than hydraulic cement. I commend him for suggesting that aggregate producers should become more active in knowing about the reactivity of their products and the precautions needed to permit those products to be safely used in concrete.

Jan 1, 1997

By George H. K. Schenck

INTRODUCTION Boom-type miners mechanize the cutting and loading of a solid face of rock having a compressive strength as high as 124 MPa (18,000 psi). These machines differ from continuous miners in that they can cut much harder rock for any given cutter-motor power. In addi¬tion, a boom miner is more flexible, selectively cutting any cross-sectional configuration from circular to square arched faces, 1.57-rad (90°) turns, and mixed faces. The basic operational concept of these machines, utilizing a rotary cutterhead mounted on a slewing boom, gained importance in Europe during the 1960s, where, due to their specialized use in roadway development in coal mines, they were called roadheaders. Along with continuous miners, boom miners share the ability to fully mechanize the cutting and loading of rock, while eliminating the need for explosives and the shattering and overbreaking of roof and rib rock that is normal with conventional mining methods. The machines are crawler-mounted and most have a gathering-arm load¬ing system that, moves the cut rock onto an internal chain conveyor, subsequently discharging the rock into a transport device such as a shuttle car or a bridge conveyor. Light- to medium-weight boom miners are cutting softer rocks and minerals successfully, including coal, shale, rock salt, potash, phosphate, weak sandstones (e.g., those associated with uranium ore), and other softer ores such as iron. Heavy third-generation ma¬chines have extended rock cutting to much harder ores such as borate, oil shale, and almost all rock associated with coal seams, except for the harder sandstones and limestones. The mention herein of specific manufacturers, brand names, or models is made by way of example only and does not constitute endorsement by the author or The Pennsylvania State University. All costs cited herein are in 1977 US dollars. TYPES OF BOOM MINERS All boom miners utilize drag bits fixed on a rotary cutterhead that is powered by an electric motor mounted in the slewing and elevating cutter boom. The cutter¬head is either a ripping or a milling type, and the machines have evolved through three generations (Boldt, 1977). As shown in Fig. 1, only one boom miner cur¬rently produced can use both ripping and milling cutter¬heads on one machine. Table 1 lists boom-miner models built in large num¬bers and/or successfully used or marketed in North America. Table I excludes machines of which only a few have been produced and none sold in North America, such as those manufactured by Demag, Eick¬hoff, Paurat, and Thyssen-Rheinstahl. The only Ameri¬can boom miners are the F-16 and Super ROC¬MINER"', manufactured by AEC, Inc., State College, PA. Machine Generations Three generations of boom miners can be distin¬guished according to their times of introduction, their rock-cutting abilities, and their weight classifications. Fig. 2 illustrates typical best-practice production rates possible from boom miners for selected cutter-motor power sizes (other power figures can be interpolated). Productivity is related directly to the energy applied to the cutterhead and indirectly to machine weight; higher

Jan 1, 1982

By Jonathan Ludlow, Robert A. Jankowksi

Introduction A longwall operator can make few changes to increase output, significantly reduce respirable dust, and decrease power consumption. Reducing drum speed, and thereby cutting with increased pick penetration, is one. This article defines the benefits of deep cutting in terms of reduced dust production and power consumption. It also identifies the practical aspects of high pick penetration in terms of shearer performance and coal loading. Before examining some practical aspects of reducing drum speed and looking at the theoretical background, it is worthwhile to summarize what is meant by high penetration and deep cutting, and what potential benefits and pitfalls may be expected. Deep cutting (in the sense of high penetration rather than wide web) can be defined in one or more of the following ways: • Cutting with an average pick penetration distance higher than that used in the past. • Cutting with a pick penetration higher than the longwall operator would have used if the advantages of deep and slow cutting were not considered. • Cutting with a well-designed shearer drum below 40 rpm. All these definitions are slightly arbitrary. They are given to provide a basis for discussion and to make the point that any move towards deeper, more efficient cutting can result in operational benefits. The benefits of deep cutting appear in many different areas. The most noticeable benefit, provided suitable instruments are available, is the reduction of airborne respirable dust. During an experiment on a longwall in the Pittsburgh seam, a nearly four to one reduction in dust levels was seen when drum speed was halved. Not all studies have shown such a big reduction, but it seems that some benefit is almost always obtained when drum speed is reduced. Production rate and specific power consumption are also affected (in a positive sense) by reducing drum speed or increasing pick penetration. Although these changes may not be as spectacular as those in dust level, they contribute to the economic return of the longwall operation. Similarly, improved washability through fines reduction may have a beneficial economic effect. Cutting with shearer drums operating at lower speeds does have some possible deleterious impacts that an operator should be aware of. For example, cutting reactions - loads imposed on the picks by the coal being cut - will be increased as a deeper cut is used. Steps must, therefore, be taken to ensure the stability of the shearer and provide an adequate haulage effort. These increased cutting reactions also result in higher loads on the power transmission system (gearboxes, ranging arms, pick boxes, etc.) from the shearer motor(s) to the pick tip. These higher loads must be anticipated and provided for with the necessary hardware. In particular, extra haulage power must be provided with low drum speeds, since haulage effort required increases roughly in proportion with pick penetration. Because the drum will be rotating more slowly or will have fewer picks, the load on shearer components will also be more variable. If suitable, robust equipment is not used, this increased vibration will decrease reliability. Benefits of Deep Cutting Lower dust levels, decreased specific power consumption, and improved product washability are the most noticeable benefits of reduced drum speeds. Although the benefits will vary greatly with mining conditions and the type of coal, some examples of what can be expected are described below. Reduced Dust Levels Figure 1 shows principal results of a study on the effects of reduced drum speed conducted on a longwall in the Pittsburgh seam (Ludlow, 1981). This figure shows that average dust production was reduced by about 70% when drum speed was halved. By making some assumptions about such quantities as coal density, it is possible to apply this proportional reduction to the quantity of respirable dust liberated per ton of coal mined. When this is done, two kinds of results are obtained: • At 70 rpm, about 1 g (15 gr) of airborne respirable dust is created for every ton mined (roughly one part per million). At 35 rpm, only 0.28-0.37 g/t (3.9-5.1 gr per st) of coal mined become airborne respirable dust. • At 35 rpm, nearly four times the amount of coal may be mined before the compliance level is exceeded, compared with 70 rpm.

Jan 3, 1984

By M. J. Laurila

On-line analyzers have made possible the use of fuzzy logic control (FLC) for local control applications in coal preparation plants. This includes FLC of heavy media vessels, cyclones and froth flotation. These circuits can be controlled in real time in a feedback control loop with an on-line analyzer. Control strategies can be developed using statistical process control (SPC) techniques that are first applied offline in manual mode through the use of control charts. This process is then brought on line using FLC in conjunction with an on-line analyzer to automate the circuit. The advent of programmable logic controllers into preparation plants has enabled the implementation of regulatory control. As shown in Fig. 1, regulatory control is the first level of process control as applied in a processing facility. This involves: ? The logic sequences for start-up and shutdown procedures and all the necessary interlocks for safety purposes. ? The systems to control individual parameters and individual units. Various sensors provide information that is useful to plant operators, while some can be used as signals to simple control loops (Couch, 1996). This is the foundation of process control. It must be performed well if there is to be any success in supervisory control (level 2). Supervisory control involves overseeing and directing the operation of unit operations, or of the whole plant. This includes more complex loops with cascade control. And it may involve feed forward and feedback systems. There have been some applications of this kind of control to thickeners and flotation circuits in coal preparation plants. In addition, there are supervisory controls on some dense medium circuits and, in a few cases, on jigs. These are implemented with proportional integral- derivative (PID) controllers that are available in the PLC or distributed control system (DCS) operating soft ware. In a coal preparation plant, there might exist 200 to 400 PID loops. In a coal-pro¬cessing plant, this number is increased to 2,000 loops or more. It is important to use the correct algorithm or function to control a process variable to the point where automatic control shows a markedly decreased variability over manual mode. This is where fuzzy logic and neural networks have an advantage over traditional model-based expert systems. Because dynamic models of coal preparation unit operations do not exist, model-based expert systems that are prevalent in minerals processing cannot work. Not enough is known about the dynamics of the gravity separation processes employed in coal preparation. Therefore, an artificial intelligence (AI) system consisting of a subset of fuzzy logic, genetic algorithms and neural networks must be applied in coal preparation plants if a supervisory control system is to be successful across a range of feed characteristics and variable set points.

Jan 1, 1997

By R. Svensson

Mill linings play an important role in all comminution processes with tumbling mills. The selection of the lining profile and the lining material has a great impact on overall grinding cost and mill throughput. The cost of the lining itself is minor in comparison with media cost, energy cost and the value of the material produced in the mill. The total scope must be considered when mill linings are selected. This paper presents case stories and explains the technology behind: •Different lining profiles in rubber and steel. •Orebed magnetic lining. •Steel-cap rubber lining. Introduction Grinding mill liner life has been a subject of extensive studies. Performance of mill liners in terms of wear life and how it relates to lining material and profile is abundant in the literature. The most economical liner material and design has been selected based on "cents per ton milled". The evaluation of the effect of liner design on power and grinding media consumption which is a far more significant part of grinding costs, has been less comprehensive (Fig. 1). Lining materials Steel. The most commonly used lining material is steel, calculated on a volume basis. Steel linings have a long tradition and a number of alloys is being developed. The main families are: •High chromium white iron. •Cr and Ni white iron (NiHard). •High chromium martensitic steel. •High carbon Cr-Mo perlitic steel. •High carbon Cr-Mor martensitic steel. •Manganese austenitic steel. Steel linings can be successfully used in most applications. Rubber. Rubber linings were introduced 25-30 years ago and are today commonly accepted to be the most economical linings for secondary and regrind applications. Rubber linings offer long service life in these applications and generally [ ] outlast steel in wet applications, where corrosion is a factor on the steel. The cost relation between rubber and steel has changed dramatically over the years. In the past rubber was more expensive than steel, but today, the cost of rubber is close to half the cost of steel, calculated on a volume basis (Fig. 2). The present cost situation has made rubber linings an economical alternative in several primary grinding applications as well. The wear rate of rubber in primary applications is higher than the wear rate of steel, thus giving a shorter wear life. This can be compensated for by making the rubber lining thicker. The cost advantage of the rubber allows for this, but care must be taken not to jeopardize the grinding performance. Rubber, as a lining material, offers some distinct advantages over steel lining, such as: •Ease of installation and removal. •Safer to handle, less risk of injuries. •Faster replacement. •Reduces noise.

Jan 1, 1996

By C. J. Pritchard, R. D. Jacksha, K. V. Raj, C. Sunderman

"In underground metal/nonmetal mines, repeated short-term exposure to high levels of airborne contaminants can become a serious health issue. Currently there are no common mechanisms to control or mitigate these localized short-term high exposures to contaminants. To improve miners’ health and safety, the National Institute for Occupational Safety and Health (NIOSH) Spokane Mining Research Division (SMRD) is developing a Smart Monitoring and Control (SMAC) system for real-time monitoring of mine air quality, with integrated countermeasures to reduce these high concentrations of airborne contaminants in localized sections of mines. To develop and test a SMAC system that could be implemented in an underground mine, researchers at SMRD built a test apparatus incorporating a fan, louver, ducting, and sensors combined with atmospheric monitoring and control software. This system will institute effective countermeasures to reduce contaminant levels, improving miner safety and health. INTRODUCTION Underground metal/nonmetal mines continue to expand production by employing large equipment in greater numbers and by mining into hotter and more challenging environments which are typically further away from fresh air sources. This challenges the ventilation system to provide adequate quantities of fresh air to the workplace and increases the potential for air contamination that could result in immediate and long-term health and safety issues to miners. Currently, airflow and contaminant concentration in mines are typically based on full-shift sampling which may not identify repeated short-term exposures to high levels of airborne contaminants such as NO2, CO, diesel particulate matter (DPM), and dust with high silica contents. Repeated short-term exposure to these high levels of airborne contaminants over long periods can lead to serious health issues which are documented in several studies [1-3]. Based on data from ventilation surveys and contaminant measurements, mines often attempt to minimize exposures by sending more fresh air to the working areas where contaminant concentrations are found to be high. Real-time monitoring and control allows mines the option to monitor contaminants and implement controls in real-time. Real-time monitoring and control of the underground mine atmosphere has been a subject of research since its beginning in Finland at Outokumpu’s Pyhäsalmi Mine in 1969 [4]. More recently, the real-time monitoring and control problem has been addressed using ventilation on demand (VOD) systems [5, 6]. A VOD system attempts to deal with this issue on a global level by controlling fans, louvers, and doors to increase airflow to active work areas and reduce contamination. However, a VOD system only controls the distribution of airflow and may not be helpful in mines where the airflow is at maximum capacity. Researchers at SMRD are developing a SMAC system for real-time monitoring of mine air quality with an integrated ability to trigger countermeasures to reduce high concentrations of airborne contaminants in localized sections of mines. The SMAC system will be designed, tested, and refined in the laboratory environment before being demonstrated in an underground mine. This system will not only try to minimize the gaseous contaminants by regulating the airflow locally, but will also potentially employ a filtration system to reduce the airborne contaminants in case the maximum airflow capacity is reached. This paper presents development of the test apparatus for the SMAC system."

Jan 1, 2017

By M. Kalin, H. D. Sharma

INTRODUCTION The inactive uranium mill tailings investigated in this study are located in two mining districts, Elliot Lake and Bancroft, Ontario, Canada. The sites exhibit a mixture of surface features consisting of dry areas, with or without vegetation and areas covered with water. On the edges of the water bodies, indigenous vegetation has invaded the tailings beaches [ Typha] spp. (Tourn.) L. a dominant plant on these tailings beaches, has been studied for the uptake of radium-226 and lead-210 (Kalin and Sharma, 1981b) from the tailings. It was found that most of the radium-226 remains in the roots of the plants, and that the solubility of radium-226 in control soil differs from that in tailings. The uptake of radium226 by vegetation and other biota is related to the solubility of the element in water. Factors controlling the solubility of radium-226 in uranium mill tailings are of interest in assessing the environments effects of these wastes. Rusanova (1962) found that the soluble or extractable amount of radium from the soil is inversely related to the total concentration of calcium and magnesium in the soil. Experimental work has clarified some aspects of the leachability of radium from uranium mill tailings (Levins, et al., 1978; Wiles, 1978, and others). Halvik, et al., (1967) studied the effects of pH and chemical composition of surface water on the liberation of radium from uranium mill tailings and uranium ore. He found that an increase in pH, up to a value of pH 9, decreases the amount of radium released from the tailings and the ore. A positive effect of calcium chloride was noted on the leachability of radium. Benes (1981) reviewed the physicochemical forms of radium and its migration in water. Based on experimental work, he identifies primary factors which determine leaching of radium from uranium mill tailings. The ratio of the volume of the leaching solution to the weight of the leached sample; the composition of the leached solids and the leaching solution, and finally the pH of the leaching mixture are of importance. He emphasized that a scarcity of field data exists, which would relate experimental work to the actual situation in the tailings ponds. Inactive tailings ponds in Ontario are 16 to 23 years old, and processes which are of importance in evaluating the long term effects of uranium mill tailings in the environment can be studied. The objective of this work was to investigate the leachability of radium-226 from the tailings under field conditions. MATERIALS AND METHODS Sample Collection A description of inactive tailings sites in Ontario, where the tailings and the water for this study were collected, can be found in a report by Kalin (1981). [Typha] spp. specimens were excavated from 16 different locations along with the tailings attached to the roots of the plants. A tailings sample from the surface area, around the plant, was secured before the excavation. Tailings samples were also collected from dry areas free of vegetation in the vicinity of the wetland stand. Surface samples (depth 0-5 cm) and samples at a depth of 20-25 cm were secured. Water was collected from the shallow tailings beaches around the vegetation stands. Sample Preparation The roots of the plants were washed free of tailings with distilled water. The resulting thick slurry of tailings was allowed to settle in the wash basin for five minutes and the supernatant water was decanted. The remaining saturated tailings were brought to 400 ml volume with distilled water in a beaker. The slurry was mixed with a magnetic stirrer for two 24-hour periods. After the first 24-hours, the slurry was allowed to settle for 20 minutes and the supernatant water was removed. The remaining sediments were suspended again in distilled water (400 ml total mixture) and leached for a second 24hour period. The tailings slurries from the different locations had a solid to liquid ratio, which ranged from 0.5 to 1.4 grams of dry tailings material per millilitre. The variations in the ratios are the result of differing fractions of coarse and fine tailings on the sites. In the final leachates, the solid to liquid ratios of the samples were 0.17 (±0.1) g/ml. The pH of the surface water and the leachate was determined with an I.L. Portomatic pH meter. All the tailings samples were homogenized in a mortar and brought to dryness at 75 to 85°C. Approximately 0.5 g of tailings were

Jan 1, 1981

By Chester A. Rowland, David M. Kjos

INTRODUCTION Mineral ore comminution is generally a feed preparation step for subsequent processing stages. Grinding, the fine product phase of comminution, requires a large capital investment and frequently is the area of maximum usage of power and wear resistant materials. Grinding is most frequently done in rotating drums utilizing loose grinding media, lifted by the rotation of the drum, to break the ores in various combinations of impact, attrition and abrasion to produce the specified product. Grinding media can be the ore itself (autogenous grinding - primary and secondary), natural or manufactured nonmetallic media (pebble milling) or manufactured metallic media - steel rods, steel or iron balls, or a combination of autogenous media and steel balls (SAG milling). This chapter covers rod and ball mills utilizing manufactured metallic grinding media. MILL DESIGN The interior surface of rod and ball mills exposed to pulp and/or grinding media are protected from wear and corrosion by rubber, metallic or a combination of rubber and metallic wear resistant materials. Rod and ball mills essentially draw constant power, thus are well suited for use of synchronous motors with power factor correction capabilities as drive motors. A net of approximately 120 to 130 percent of running torque is required to cascade the charge in these mills. The pull-in torque is about 130 to 140 percent with the pullout torque to keep the motor in-step (in-phase) generally in excess of 150 percent. When rod and ball mill are started across-the-line the starting and pull-in torques result in inrush currents exceeding 600 percent which possibly result in high voltage drops. To deliver 130 percent starting torque to the mill the motor design must take into account the maximum anticipated voltage drop. Motor torque decreases as the decimal fraction of the voltage available squared. E.g., a motor rated 160% starting torque with a 10% system voltage drop will deliver 160% x or 129.6% torque to its output shaft When it is not possible or practical to start a fully loaded synchronous motor across-the-line it is possible to utilize the motor's pull- out torque to start the mill. By using a clutch, normally an air clutch. between the motor and the mill, the motor is brought up to synchronous speed before the clutch is energized. If the motor has an adequate amount (175 or greater) of pull-out torque the pull-out torque starts the mill without major disruptions on the electrical system. Since the energy release at initial cascade of the mill charge is an inverse function of acceleration time, a minimum acceleration time of 6 to 10 seconds or more is recommended to prevent damage to the mill or the mill foundation. Economics at the time of plant design and mill purchase determine the drive to be used. The simpliest drive is the low speed synchronous motor with speeds in the range of 150 to 250 RPM connected to the mill pinionshaft by either an air clutch or flexible coupling. Using a speed reducer between the motor and pinionshaft permits using motors having speeds in the range of 600 to 1000 RPM. In this speed range, if power factor correction is not required induction motors can be used; squirrel cage where there is no restriction on inrush current; slip ring where a slow start and low inrush current is required. Air clutches can also be used to ease starting problems with squirrel cage motors. In some areas of the world induction motors and starters are less expensive than synchronous motors at a sacrifice of motor efficiency and power factor correction. Dual drives, that is two pinions driving one gear mounted on the mill, become economical for ball mills drawing more than 3500 to 4000 horsepower (2600 to 3000 kilowatts). Further developments of the low frequency, low speed synchronous motors with the rotor mounted on the mill shell or an extension of one of the mill trunnions could improve the cost picture for these "gearless drives", making them practical for large ball mills. The percent of critical speed, which is the speed at which the centrifugal force is sufficiently large to cause a small particle to ad- here to the shell liners for the full revolution of the mill is given in mill specifications. Critical speed is determined from the following: Where D is mill diameter inside liners (specified in meters). Cs is critical speed in RPM. When D is specified in feet

Jan 1, 1998

By John A. Lombardi

The solutions used in the heap biooxidative treatment of sulfidic refractory Au ores and concentrates generally load free mineral acid and metals and, in time, become bacteria activity inhibiting due to increased acid and/or metals concentrations. The metals­acid loaded, bacteria action inhibiting, solution must then be: (a) processed, prior to re-use, to reduce its metals and/or free acid mineral acid components to bacteria activity compatible levels; or (b) neutralized prior to discharge. The most common solution processing method to adjust acid-metals contents is liming to effect the precipitation of dissolved metals and neutralization of acid sufficient for solution discharge or re-use. The lime treated solution is high in suspended, colloidal precipitates and requires clarification before reuse. Most often the "bleedstream" of biooxidation solution that is lime treated is discharged and req­uisite volumes of fresh water added to the process inventory. Depending on the level of bacterial action in a heap and the acid consumption characteristics of the ore/concentrate being biooxi­dized, the bleedstream taken from the circulating, irrigation on­off, solution, can be large (e.g., 10%-20% and sometimes as much as 50% of the daily solution flow requires treatment-dis­charge). The lime charge for bleedstream treatment is commen­surately large and represents a major operating cost of the bioox­idation process. The patented HW Process Technologies (HWPT) Engineered Membrane Separation (EMS™) method is an alternative biooxi­dation bleedstream treatment method. EMS™ processing removes the multivalent base metals ions in an acid solution to a small volume EMS™ "concentrate" (typically 25% of the feed­stream volume). The EMS™ method coincidentally makes a "clean acid" EMS™ "permeate" (typically 75% of the feed­stream volume). In general, EMSTM processing enables 75% of the bleedstream volume to be recycled; allows up to 2/3 of the bleedstream acid to be recovered as a 10% solution; and delivers dissolved metals in a highly concentrated stream suited to a. metals recovery or b. reduced cost lime neutralization for a given metals precip­itation relative to full bleedstream neutralization (typically a 30% lime use savings). Implicit in the recovery of process water for re-use and reduced reagent addition requirement for metals precipitation is a significant reduction in wastewater-sludge dis­posal facility size.

Jan 1, 1998

By Raja V. Ramani

The importance of adequate mine ventilation for good health and safety conditions has been recognized since miners first ventured underground. All over the world, governmental actions on mine health and safety are directly connected with mine atmospheric environments, their assessment and control. These governmental actions have been, and continue to be, motivated by the repeated experience that potential threats from inadequate control can evolve, either suddenly or in the long-term, into unpleasant reality, with catastrophic health and safety consequences. These governmental actions are also an indication of the remarkable professional progress being made in the field of mine ventilation. Surely, no developed country allows underground mines to operate without strong measures for the control of ventilation under legislation specific to the mining industry. In the United States, the federal government's responsibility for mine safety has evolved over the years through a series of legislative actions beginning in 1865 and continuing through 1910, 1941, 1947 and 1952. The passage of the Federal Metal and Nonmetallic Mine Safety Act of 1966 (the 1966 Metal Act), the Federal Coal Mine Health and Safety Act of 1969 (the 1969 Coal Act) and the Federal Mine Safety and Health Amendments Act of 1977 (the 1977 Mine Act) are the most important milestones in this long legislative history. The 1977 Mine Act adopted many of the stronger features of the Occupational Health and Safety Act of 1970, perhaps the most significant piece of legislation on workers' occupational health and safety in US history. It is worthwhile to ask, "How far has the United States come in mine ventilation since the enactment of the 1969 Coal Mine Health and Safety Act?" This Act and its associated regulations defined in great specificity many standards for healthful and safe operation of a mine and, in some cases, what constitutes good ventilation practice. Since 1969, the frequency and severity of disaster-type events in US coal mines has decreased markedly. A primary reason for this achievement is the strength of the Coal Mine Health and Safety Act's regulatory features in such areas as methane concentration control, dust control, intrinsic safety and explosion-proof enclosures, minimum air quantity and quality standards and air distribution and escapeway provisions. These requirements greatly influenced mine ventilation planning, engineering and practice, leading to increasingly safe working conditions.

Jan 1, 1993

By Pierre Mousset-Jones

I wish to submit this discussion item on an article published in the August issue of MINING ENGINEERING by Wala and Altman on the use of canonical diagrams for mine ventilation networks. The advent of computer graphics and plotting has enabled complex underground mine layouts to be displayed and plotted in a variety of three-dimensional formats that very closely represent the actual mine layout. This enables the ventilation engineer, who is familiar with the mine, to display the network pressures, quantities, resistances, and fan locations in a way that can be understood by everyone working at the mine. This is important if the ventilation engineer needs to discuss changes in ventilation with production personnel, or escapeway procedures with union personnel. While I can see some of the potential advantages in the canonical ventilation network diagram described by Dr. Wala, it appears to me that the fact that the diagram has no visual equivalence with the actual mine layout means that most mine personnel will find such a diagram almost incomprehensible. The authors have alluded to certain advantages with the canonical diagram. However, I find it difficult to see that these outweigh the fact that almost everything listed can be duplicated in isometric diagrams, and in a manner that is easily understandable to everyone in the mine from the miner to the mine manager. Perhaps the authors could show a clear example comparing an isometric with a canonical portrayal of a mine network showing the information that is immediately available from the canonical network that is not obvious in the isometric diagram. There is in this country a severe problem with the need for ventilation engineering and expertise to be used and accepted by mine management as a necessary part of mine planning. It seems to me that to introduce the requirement for further explanation so that management can understand a canonical network display will hinder more than help the recognition that ventilation engineering needs at so many mines in this country. However, I can see that the canonical approach does have its place in assisting the practicing mine ventilation engineer, wherever he or she exists in the organization chart for a typical mine. In this case, could the authors please indicate the availability of the computer program to compute and plot the canonical diagram for a mine ventilation network.

Jan 1, 1990

By Pierre Mousset-Jones

I wish to submit this discussion item on an article published in the August issue of MINING ENGINEERING by Wala and Altman on the use of canonical diagrams for mine ventilation networks. The advent of computer graphics and plotting has enabled complex underground mine layouts to be displayed and plotted in a variety of three-dimensional formats that very closely represent the actual mine layout. This enables the ventilation engineer, who is familiar with the mine, to display the network pressures, quantities, resistances, and fan locations in a way that can be understood by everyone working at the mine. This is important if the ventilation engineer needs to discuss changes in ventilation with production personnel, or escapeway procedures with union personnel. While I can see some of the potential advantages in the canonical ventilation network diagram described by Dr. Wala, it appears to me that the fact that the diagram has no visual equivalence with the actual mine layout means that most mine personnel will find such a diagram almost incomprehensible. The authors have alluded to certain advantages with the canonical diagram. However, I find it difficult to see that these outweigh the fact that almost everything listed can be duplicated in isometric diagrams, and in a manner that is easily understandable to everyone in the mine from the miner to the mine manager. Perhaps the authors could show a clear example comparing an isometric with a canonical portrayal of a mine network showing the information that is immediately available from the canonical network that is not obvious in the isometric diagram. There is in this country a severe problem with the need for ventilation engineering and expertise to be used and accepted by mine management as a necessary part of mine planning. It seems to me that to introduce the requirement for further explanation so that management can understand a canonical network display will hinder more than help the recognition that ventilation engineering needs at so many mines in this country. However, I can see that the canonical approach does have its place in assisting the practicing mine ventilation engineer, wherever he or she exists in the organization chart for a typical mine. In this case, could the authors please indicate the availability of the computer program to compute and plot the canonical diagram for a mine ventilation network.

Jan 1, 1989

By Eung Ha Cho, Olufemi Olajide

A laboratory study of dense-medium cycloning with Redstone seam coal has been conducted. The investigation was performed using a 203 mm (8 inch) Heyl-Patterson dense-medium cyclone. Two underflow orifices of 31.8 mm (1.25 inch) and 44.5 mm (1.75 inch) in diam were used. The effects of feed pressure and particle size were determined on the cycloning performance. In addition, the cyclone performances with coal-rich and refuse-rich samples were also investigated. The results of the experiments were analyzed using basic hydrodynamic principles. The increase in the feed pressure increases the specific gravity of separation (defined as the specific gravity of a particle which has an equal chance of reporting to either the overflow or underfow) for both underfow orifices. The specific gravity of separation increases as the particle size decreases with the smaller underflow orifice. However, an opposite relationship has been observed with the larger under flow orifice. The separation becomes sharper with the smaller under flow orifice than with the larger underflow orifice especially for fine material. Also, the separation becomes sharper with an increased feed pressure for a coal-rich sample. However, this effect is less significant for a refuse-rich sample. This means that a high feed pressure is required for a coal-rich sample to maintain a good separation. It also means that applying a high feed pressure provides a good separation regardless of the type of feed coal in terms of ash content in the solid.

Jan 1, 1990

By Sol Ribakoff

In 1973, the Subcommittee on Contracting Practices, U. S. National Committee on Tunneling Technology, National Academy of Sciences, under the able chairmanship of A. A. Mathews, Engineering Consultant, now residing at Kaunakokai, Hawaii, undertook the task of developing recommendations for improving the performance of underground construction work, reducing the costs of such work and minimizing the adverse atmosphere in which owners and contractors were performing their respective functions. The author of this paper and C. F. Palmetier, a highly qualified engineer, were thereupon appointed as staff assistants to the Subcommittee, to conduct a study of existing practices and relationships in Western Europe and in the United States, to formulate recommendations, for consideration by the Subcommittee and the participants in a future workshop, as to desirable improvements in United States contracting practices and relationships. The study, conducted during 1973 and 1974, involved personal interviews of a total of 162 owners, engineers, contractors, lawyers, and educators throughout the United States and in 7 European countries, the analysis of 178 replies to questionnaires sent to individuals and firms involved directly or indirectly in construction, and a workshop, near Napa, California, attended by 53 representatives of all segments of the construction community. The resulting report, entitled "Better Contracting for Underground Construction", was submitted by the Subcommittee to the National Academy of Sciences, which approved its publication in 1974 by the National Technical Information Service. It may be obtained from that Service at Springfield, Virginia, 22161, as Accession No. PB-236 973. It was recognized-during the course of the study that while contracting practices and owner-contractor relationships were the

Jan 1, 1981

By Dewey W. Geary

Nast Tunnel is the first granitic rock tunnel to be excavated by a machine for the U. S. Bureau of Reclamation. The machine has excavated 7500 ft of the tunnel at this writing. There have been a few good performances; 31 ft in 1, 8 hr shift, and 75 ft in a 24 hr period, but generally, progress has been slower than that accomplished in the area with reasonable comparable conditions using conventional excavation techniques. Button cutters, now employed using compression as the means of fracturing the rock, gives an average cutting rate of 3 ft per hr. Disc cutters manufactured by the machine builders were tried briefly but the snap rings holding the discs to the hub failed. However, disc cutting rates of 6 ft per hr were recorded for short periods of time in a competent coarse grained porphyritic granite. The snuck was also considerably larger and more easily handled than that produced by the button cutters. Hughes Tool Co. is working on a disc type cutter and expects to have them on the cutterhead by mid June. LOCATION, DESCRIPTION AND BID OPENING Nast Tunnel is located 2 miles west of the Continental Divide and 39 miles east of the town of Basalt in Pitkin County, Colorado. The outlet portal is in the Fryingpan River Valley at an elevation of 10,000 ft. The tunnel is a 3 mile long, 10 ft diameter gravity flow water conveyance structure and a part of the west slope water collection for the Bureau of Reclamation, Fryingpan-Arkansas Project. The bids were opened July 9, 1970, with 8 contractors submitting proposals. The contractors were given the opportunity to bid

Jan 1, 1997

By Charles L. Kimbell

In 1989, worldwide total mineral industry activity registered advances from 1988. The advances may not have been as noteworthy as in 1988, but they can be regarded as a worthy theme for this article. The momentous political events throughout the centrally planned economy world had only a modest impact on actual 1989 industry performance. But these events have such potential for worldwide changes in economic activity in future years that they dwarf the actual happenings of 1989 in the minerals industry. Perhaps most significant factor here is the rapidity with which these changes have occurred. The end of World War II saw the almost immediate division of the world into major economic blocs. Since then, mineral industry activities and other industrial operations have been governed, to a considerable extent, by this partitioning. The developed market economy nations have operated as a group, the centrally planned economy states as another group. And these two groups essentially have vied for recognition in mineral industry operations in the developing economy states of Asia, Africa, and Latin America. The boundaries of these country blocs were fixed in the late 1940s and early 1950s. They were not markedly altered in the ensuing 40 years. We have seen the establishment of the Organization of Petroleum Exporting Countries, the continued Arab-Israeli confrontations in the Near East, the upheavals surrounding the rise of Moslem Fundamentalism in Iran, and the establishment of Soviet-oriented governments in Cuba and Nicaragua. These events have had notable effects on the world economy. But they have done little to change the pattern of market economy states functioning as a competitive bloc with the centrally-planned economy countries.

Jan 1, 1990

By David K. Parrish, Duane A. Labreche

Analyses of jointed, cracked, and fractured rock masses are needed in a variety of applications, including estimates of reduced elastic modulus for damaged rock, estimates of permeability changes as a function of stress changes, and estimates of relaxed regions due to excavation. A numerical model of a jointed rock mass, the compliant joint model, was used to study the response of a jointed rock mass near a circular excavation. The joint constitutive model was linear elastic in shear and nonlinear elastic in the relationship between normal stress and joint closure. Initial stresses were geostatic and representative of stresses in a granitic jointed rock mass approximately 300 m deep. The joint properties were based on values that are representative of joints in granite. The results from numerical simulations were compared with analytical results for stress near a circular opening in an elastic medium having reduced Young's Moduli to simulate jointing effects on the rock mass. The differences in response between the two models resulted from the differences in the symmetry of elastic properties, the shear modulus of the continua, and the constitutive relations in the two models. For example, the non- linear anisotropic elastic continuum described by the compliant joint model resulted in vertical stress concentrations near the side of the circular opening that were 25 percent higher than the vertical stress concentrations in an isotropic linear elastic medium. The equivalent Young's Moduli in the vertical and horizontal directions of the compliant joint continuum deviated by as much as 50 percent from the values assigned to the modeled region prior to excavations. The perturbations of the elastic Young's Moduli extended to distances of two to three radii from the opening.

Jan 1, 1986

By Gustavo G. Pilger, Jair C. Koppe, João Felipe C. L. Costa

Information for mine planning requires to be close spaced, if compared to the grid used for exploration and resource assessment. The additional samples collected during quasi-mining are frequently located in the same pattern of the original diamond drill holes net but closer spaced. This procedure is not the best in mathematical sense for selecting a location. The impact of an additional information to reduce the uncertainty about the parameter been modelled is not the same everywhere within the deposit. Some locations are more sensitive in reducing the local and global uncertainty than others. This study introduces a methodology to select additional sample locations based on stochastic simulation. The procedure takes into account data variability and their spatial location. Multiple equally probable models representing a geological attribute are generated via geostatistical simulation. These models share basically the same histogram and the same variogram obtained from the original dataset. At each block belonging to the model a value is obtained from the n simulations and their combination allows one to access local variability. Variability is measured using an uncertainty index proposed. This index was used to map zones of high variability. A value extracted from a given simulation is added to the original dataset from a zone identified as erratic in the previous maps. The process of adding samples and simulation is repeated and the benefit of the additional sample is evaluated. The contribution of an extra sample was evaluated locally, globally and in terms of variation in total measured resources. The procedure showed to be robust and theoretically sound, mapping zones where the additional information is most beneficial. A case study using coal seam thickness illustrates the method.

Jan 1, 2002

By G. E. McCormack

The Development of Water Gel Explosives Water gels are the most recent development in the evolution of commercial explosives. They are replacing dynamite, having proven themselves to be excellent alternatives to this standard explosive. Water gels have steadily displaced dynamite over the past two decades. In 1959, they accounted for only 1 percent of the market they share with dynamite. By 1965, the share taken by water gels had grown to 21 percent. In 1970, it was more than 40 percent, and today it is nearly 70 percent. Black powder, the first widely used explosive, revolutionized mining and quarrying in its day. However, when nitroglycerin dynamite became available about a century ago, it rapidly displaced black powder because it provided greater energy for a given weight, much higher velocity for better rock shattering performance, and water resistance for easy use in wet conditions. During the past 50 years, ammonium nitrate has played a dramatically increasing role in explosives. First, it was used as an ingredient in dynamite, and then as the main component in nitrocarbonitrate blasting agents. About a quarter of a century ago, it began to find use in a simple and inexpensive mixture with fuel oil that has revolutionized the explosives industry and today fills about 80 percent of the requirements for explosives in the U.S., and certainly the bulk of blasting requirements worldwide. Water gels and dynamites share the remaining 20 percent. Water gel explosives based on ammonium nitrate have also been developed during the past quarter century. These explosives contain

Jan 1, 1979

By R. V. Ramani, R. Bhaskar

The results of an experimental and theoretical research study into the spatial and temporal behavior of dust in mine airways is summarized. Experiments were performed in mine airways under both controlled and normal operating conditions. The controlled experiments included the documentation of the concentration and deposition patterns for two types of dust under three different velocities. Real-time dust concentration data were obtained using real-time aero- sol monitors (RAM-I) to determine the dispersion coefficient in a mine airway. On the theoretical font, a convection diffusion model of the dust flow phenomenon in mine airways was developed considering dominant mechanisms affecting dust transport and deposition in mine airways. Mechanisms modeled include turbulent and gravitational deposition, coagulation, and dispersion. In this paper, comparative analyses of results (8 the controlled experiments with the predictions of the mathematical model are presented. In addition to the overall validity, the adequacy of the individual components of the model, such as deposition in the total and respirable range, and dispersion are examined in light of the experimental observations. The experimental data from controlled experiments compared more favorably with model predicted results for similar conditions as opposed to the comparative analysis results for in-mine experiments. Specifically, the model predictions for an 800-m stretch of a mine airway reveal some significant differences from the behavior noticed in the mine. The deposition was lower than that predicted by the model. Consequently, concentrations were higher than those predicted by the model. The implications of these and other findings for further investigations are discussed in the paper.

Jan 1, 1987