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By Nathaniel Arbiter, Robert R. Beebe

The U.S. mining industry - once world dominant and largely independent of foreign sources for most major commodities - has today lost this position, with little likelihood of regaining it. But even though Business foretold The Death of Mining four years ago, the current outlook is less pessimistic because of drastic restructuring of the industry; a major reduction in its labor force, with resulting large increases in productivity; and introduction of improved technology. In the face of this clearly changed climate for domestic mining, it is important to examine the role of the development and research establishment in the industry's future, which raises a complex set of questions: 1. Where and to what extent can new technology assist in the continued survival of the industry ? 2. What is the past record of the research establishment in developing new technology; and what is the industry's record in supporting and adopting it ? 3. In the light of these records, and of the drastic cuts in both research staffs and expenditures, what areas should be the most important targets for new technology? New Technology vs. Old A1 though mineral processing's origins can be traced back through five or six millennia, its modern structure is about 200 years old.

Jan 1, 1989

By Paul C. Schmall

In recent years, several geotechnical methods have been successfully applied to ash pond closure. These include modified dewatering techniques with wellpoints and wells to physically drain the material, soil mixing to create a soil-concrete material for improved accessways as well as improved ground structures within ponds to act as barriers or gravity dams, and geotechnical instrumentation to serve as an early warning of the potential for liquefaction. Given the parallels between fly ash and sluiced in place mine tailings, the application of fly ash remediation method to mine tailings shows significant potential, as illustrated in this paper.

Feb 1, 2024

By E. Watkins, S. J. Schatzel, K. Ajayi, V. Gangrade

The National Institute for Occupational Safety and Health (NIOSH) has been conducting research related to unconventional gas wells positioned in abutment pillars associated with active longwall mining in Southwestern PA. The Ventilation and Explosion Prevention Team at NIOSH is working toward predicting and characterizing a hypothetical casing breach and its effect on the safety and health of mine workers, gas well workers, and the general public. An important aspect of this research is to distinguish coalbed gas from shale gas in the mine environment as an indicator of a casing breach. Another noncoal gas source exists in the region in the form of underground gas storage fields which also have the potential to migrate towards mine workings. During a potential gas breach, the addition of excess natural gas to coal mine environments will exceed the anticipated ventilation loading and can create a mine explosion hazard. A methodology has been developed to distinguish these gas sources. Gas samples from the three sources were analyzed by gas chromatography, and interpretive methods were applied to distinguish the populations graphically and statistically. Bivariate plots using the hydrocarbon index and the normalized molecular ratios showed that the sources occupy different regions of the graphs. All samples of non-coal gas contained under 2% CO2, and all coal samples included greater than this amount. A multivariate T-squared (T2) test analysis relating comparing data in pairs showed a low p-value (dimensionality parameter) and a high T2 test statistic for all comparisons, indicating that these sources are different and distinct gas populations. Ultimately, stakeholders from the mining and gas production industries can utilize this method to identify gas sources and a potential casing breach.

Mar 2, 2022

By Bob Snashall

"We could blow that damned fort out of existence if we could run a mine shaft under it!," claimed one Union soldier eyeing Bobby Lee's Petersburg defense line protecting Richmond and Jeff Davis. Puncture the line, take the hill behind, and the Confederate capital would celebrate Yankee jubilee, 1864. That was June. At 4:15 am, July 30, mine boss Henry Reese, sergeant, 48th Regiment, Pennsylvania Volunteers, gingerly entered the mine, dark and muddy. Something had gone wrong. The mine, set to go off at 3:30, hadn't exploded. It had to be checked. The main mine gallery, 1.5 m high x 1.4 m wide (5 x 4 ft) narrowing at the top to 0.6 m (2 ft), forced Reese into a tuck run. Some¬where up there in that 155 m (511 ft) stretch running 6 m (20 ft) under the rebel works, moisture may have doused the charge. Or maybe a fuse splice went awry. Or maybe there was live fire steaming like a locomo¬tive towards 3.6 t (4 st) of black powder. Perhaps now Reese wished this sprint was shorter. For the mine boasted a record break¬ing distance in military annals. Lieut. Col. Henry Pleasants Credit goes to Lieut. Col. Henry Pleas¬ants, commander of the 48th, who overheard the soldier's claim for what a mine could do. In peacetime, Pleasants had been a railroad engineer working on the Sand Patch Tunnel. He later switched to mining, working for the Philadelphia & Reading Coal & Iron Co. He even married the daughter of the editor of The Miners Journal. Grieving over his wife's early death, Pleasants plunged into the ven¬ture of war. This Union officer took the mine idea to the brass. Headquarters was less than enthu¬siastic. "The chief engineer of the Army and the rest of the regular Army wiseacres said it was not feasible," Pleasants wrote to his uncle. The pros scoffed, ticking off a list of horribles: No mine has ever gone that far, men will suffocate and be crushed, the rebs will countermine, the men will be trapped. Still, Pleasants eked out a signal to proceed. In the meantime, the miners of the 48th had commenced to dig anyway. It was like the 48th to proceed against the odds. The 48th Pennsylvania Volunteers The 48th had been recruited from Penn¬sylvania's Schuylkill County, anthracite coal mining country. The mines had been torn with strife over wages. Strikes and rioting were prevalent. Murder was not unknown. This "Buckshot" resistance centered in Pottsville, PA. It forced the governor to call in troops. Coal was in high demand to run the war machine. Pennsylvania operators worried that a military draft would stir up violence, closing the mines altogether and further shift¬ing business to emerging Illinois mines. Men lit out for the western gold fields in the tens of thousands. Secret societies bred in the mining camps invited draft agents to dance at the end of a rope. Jobs threatened by cheap, victorious South¬ern labor were one warning intended to get immigrant miners interested in the war. Operators gave miners time off to help de¬fend Philadelphia during Lee's thrust to Gettysburg. Even so, the draft spawned new violence and troop intercession. Soldiers were not welcome. The dust had barely settled when into this Pennsylvania "minefield" marched the 48th on a recruitment drive. The 48th had been in Lexington, KY on guard duty. When it came time for the regiment to pull out, the townspeople had prevailed upon authorities to extend the regiment's stay. Finally ordered to move out, the 48th was feted by a hanky-waving, tear-dropping send-off to the lyrical lament Auld Lang Syne. And Kentucky fancied itself to be rebel. Striding smartly into the Pennsylvania coalfields, the 48th attracted 400 recruits and marched off under regimental colors specially made by the proud citizens of none other than Pottsville, PA.

Jan 1, 1989

By Elza Burch

Introduction The feeding to zero concept involves feeding 600 µm x 0 (28 mesh x 0) size raw coal to heavy media (magnetite) cyclones along with the +600 µm (+28 mesh) size coal. Traditional circuits employ desliming or removing the 600 µm x 0 (28 mesh x 0) size fraction and feeding the cyclones +600 µm (+28 mesh) size coal. The feeding to zero concept recirculates 600 µm x 0 (28 mesh x 0) fines in the circuit. At the same time, a portion of the fine material is continuously withdrawn and recovered. This, in turn, prevents a fines buildup. This concept eliminates desliming screens and secondary fines circuitry for recovery of 600 x 150 µm (28 mesh x 100 mesh) coal. The result is a very simple circuit. Feeding to zero at Island Creek Island Creek Corp. was the first involved with the new concept in 1976. The company needed a temporary plant for the 9.5 mm x 0 (0.4 in. x 0) raw coal at its Pond Fork mine, near Madison, WV, while a full-scale plant was being designed and built. At that time, the Childress Corp., of Beckley, WV, became interested in the feeding to zero concept. Island Creek awarded a contract to Childress to build a single cyclone modular plant, incorporating this feeding to zero concept. The plant was erected in three months. The Pond Fork modular plant proved successful in attaining the desired feed rate of about 63.5 t/h (70 stph), while maintaining good separating efficiencies and low magnetite consumption rates. The 9.5 mm x 150 µm (0.4 in. x 100 mesh) clean coal was recovered and the 150 µm x 0 (100 mesh x 0) size was disposed of to waste. The full-scale plant was completed about two years later and the Pond Fork modular plant was moved to Holden, WV. There, it was incorporated into the Holden 29 preparation plant as a separate circuit for cleaning -25 mm (-1 in.) coal. In 1976, a similar plant was installed in Virginia by another company. These two plants are believed to be the first two operational plants in the United States incorporating the feeding to zero concept. Island Creek subsequently contracted with Childress for an identical plant at the Coal Mountain operation in West Virginia. The plant operated for three years before the mine was closed. The unit was then moved to the Spurlock mine near Martin, KY where it continues to operate. The successful operation of the Pond Fork and Coal Mountain plants before and after relocation proved both the performance and moveability of this type of circuit when constructed in a modular fashion. Since the first two plants were built, Island Creek has incorporated the feeding to zero circuit in nine additional plants. A grand total of 33 cyclones have been installed using this concept. One is the Providence mine, near Providence. Providence preparation plant Island Creek contracted with J.O. Lively Corp. of Glen White, WV in July 1978 for the construction of the Providence preparation plant. The plant began operation in February 1979. The construction period was about halved by building the plant with modular design concepts. Prefabricated sections, floors, and sides were brought in as units and then bolted in place. The Providence plant has a good track record of processing coal at a feed rate of 454 kt/h (500 stph). Feed coal to the plant has an average ash content of about 18% and sulfur content of about 4.5%. It contains about 22% refuse. The coal product has an average ash content of about 8% and sulfur content is about 3%. Raw West Kentucky No. 9 seam coal is conveyed from a box cut in the Providence mine to a rotary breaker. The breaker is fitted with 74 mm-diam (3 in.-diam) opening breaker plates. Therefore, it is well suited for removing trash, roof bolts, wood, and pyritic balls that are common in Illinois Basin coal. The -75 mm (-3 in.) coal is conveyed from the rotary breaker to

Jan 8, 1987

Power cable costs are only a small part of total mining costs. So many mine operators consider power cable failure and resultant downtime as part of the cost of doing business. But, viewed in terms of lost production, these costs can be quite significant. Now one company, Cablec, seeks to cut cable costs by upgrading the polymer compounding process used to make cable insulating and semiconducting materials. Cablec is the leading manufacturer of electrical power cables in North America. And with about a third of the market, Cablec is the largest supplier of power cable to the mining industry in the United States. To improve its products, Cable has entered the polymer compounding business. In July, it began producing insulator and semiconductor polymer compounds at its plant in Indianapolis, IN. "This new facility provides a quantum leap over conventional compounding methods," said Harry C. Schell, Cablec's president and chief executive officer. "The Cablec polymers plant is producing a dramatically higher standard of polymer compounds that provide significantly higher levels of performance and improved life cycle costs for power cable." Cablec faces tough foreign competition in the wire and cable business. Competing on price alone is difficult, particularly when foreign producers are state subsidized. So Cablec feels the best way to compete is to establish new quality production standards. The company's new polymers plant is one way to do this. By increasing purity control and uniformity in polymer compounding, Cablec says its power cables will last longer and fail less often. A typical medium voltage cable consists of a conductor, conductor shield, insulation, insulation shield, metal shield, and jacket. The conductor shield and the insulation shield are conducting polymers. Contaminants and imperfections can occur within the insulation, at the conductor shield/insulation interface, or at the insulation shield/ insulation interface. Over time, these contaminants and imperfections can decrease the electrical strength of the cable or cause premature cable failure. The effort to minimize the number and size of any possible contaminants begins with pure polymer compounds mixed in a clean facility. However, most power cable manufacturers manually handle raw materials, use ethylene/propylene (EP) in bulk bales, and mix polymercompounds in open Banbury mixers. The quality and uniformity of polymer compounds is also impacted by temperature variations in the mixing process. This results in wide gradations of product consistency from batch to batch and ultimately contributes to power cable failure. Cablec says the improved polymer compounds from its state-of-the-art plant will be the purest and most consistent insulating and semiconducting materials available. The plant itself RCA spent $18 million to build Cablec's Indianapolis plant. RCA used the facility to mix specialty polymer compounds used to make video disks. RCA had two considerations in mind for the plant, cleanliness and uniformity of the compounds. However, when the video disk market failed to materialize, RCA sold the 46.5 dam 2 (50,000 sq ft) plant to Cablec for $3.1 million. Cablec invested an additional $3 million for modifications and increased production capabilities. Today's replacement cost for such a facility is estimated at $30 million. Cablec says the plant will set a new standard for performance and be economically difficult to duplicate anywhere. One of the essential elements of the plant's clean process environment is the air intake system. It filters contaminants greater than 2 um, less than one-fiftieth the current industry standard. All material handling and conveying areas in the facility are air-locked. This keeps out contaminants such as smoke, dust, and pollen. Banks of pneumatic pumps move polymer components through the system and continually filter the air. The plant also has a backup air intake system. No process downtime due to pump failure here. From the time raw material enters the plant, it is stored, transported, and processed in filtered air by an airtight stainless steel system. The stainless steel resists rust and corrosion. This further eliminates the danger of contamination from paint or rust particles in the conveyance network. A computer system allows a single operator in a central control room to monitor every aspect of the compounding process from air quality to line speed. The computer

Jan 12, 1988

Introduction The art of ripping has been around since the Romans built the Appian Way, but only in the last decade has ripping emerged as a close competitor with drilling and blasting in breaking up heavy rock formations. That is not to say ripping is as effective in every case, or even applicable in all cases, but it does provide a cost efficient option in many situations-if the optimum ripping techniques are used. Originally designed only to loosen hard pan, the ripper has evolved into a powerful tool, its capability increasing as the weight and horsepower of tractors increased. The tractor-mounted ripper extended the range of ripping because of the additional weight it brought to bear on the ripper tooth. Tandem ripping added the weight and power of a second tractor, extending the rippability range even further. Of course, the use of ripping depends on the type of material to be ripped. Not all rock can be ripped. Determination of rippability-an art in itself-is often inconclusive but there are certain universal principles that can be considered in the first step toward optimizing ripping techniques. Each of the three basic rock types, igneous, sedimentary, and metamorphic, exhibits characteristics that bear on rippability. Igneous rocks are the most difficult to rip because they lack the stratification and cleavage planes essential to ripping hard rock. Sedimentary rocks are generally the most easily ripped since they are built of layers differing in material, texture, color, and thickness. Metamorphic rocks vary in rippability with degree of lamination or cleavage caused by transformation from pre-existing rock. The condition of rock also affects its rippability. Although sedimentary rocks offer the best opportunity to rip and igneous and metamorphic the least, decomposed granite and other weathered igneous and metamorphic rock often can be ripped economically. Little or no trouble is encountered with hardpan, clays, shales, or cemented gravel. Likewise, highly stratified or laminated rocks offer good possibilities for ripping. However, thicklybedded rock formations generally must be drilled and blasted. The physical characteristics that favor ripping may be summarized as follows: • Fractures, faults, and planes of weakness of any kind; • Weathering, resulting from temperature and moisture changes; • Brittleness and crystalline nature; • High degree of stratification or lamination; • Large grain size; • Moisture permeated clay, shale, and rock formations; and • Low compressive strength. Ripper Types A second fundamental to optimal ripping is choosing the right equipment for a particular job. There are three types of rippers-the hinge, the parallelogram, and the adjustable parallelogram. In the hinge ripper the linkage carrying the beam and shank pivots on a fixed point at the rear of the tractor. The resulting arc from up and down movement causes up to 30° differential in the tip-ground angle. Thus, the tooth angle changes as it enters and proceeds to ripping depth. The parallelogram ripper's linkage maintains the same tip-ground angle regardless of tooth depth. This type of ripper has advantages over the hinge-type when ripping above maximum depth, but does not provide the aggressive tooth angle necessary for hard-to-penetrate materials. The adjustable parallelogram combines the benefits of both the hinge and the parallelogram rippers. It has the additional benefit of being able to vary the tip angle to the optimum angle of penetration. It can also be adjusted while moving, for the optimum ripping angle in any material. Several other factors that are important in selecting the right ripping equipment are down pressure available at the tip of the ripper, tractor flywheel horsepower, and the tractor's gross weight. Down pressure at the tip determines whether ripper penetration can be obtained and maintained. Flywheel horsepower rating determines whether the tractor has the power to advance the tip. The tractor's gross weight determines whether the tractor will have sufficient- to use the horsepower. The best ripping procedure depends on the job's actual conditions. While conditions, vary, it is still possible to apply certain proven techniques to a typical ripping situation. Caterpillar Tractor Co.'s Handbook on Ripping says optimum ripping efficiency can be realized if-before actual ripping begins-these questions and answers are considered.

Jan 1, 1983

By Waclaw Trutwin

The idea of automatic or remote control of the mine ventilation process generally, and methane concentration particularly, attracts the attention of mining engineers more and more. The advantages of introducing mine ventilation control systems are breaking traditional reluctance. The change of attitude is not only because of the requirements of modern exploitation technology, but it is also due to the recent progress in development and successful introduction of reliable monitoring systems and actuators in the form of controlled ventilators and doors [1]; [2], [3], [4], [5], [6]. Many 'years of theoretical and experimental studies of the dynamics of mine ventilation processes created the needed base for a proper design of an automatic control system [7],[8],[9], [10]. From these studies must, however, be drawn a fundamental conclusion, which may be regarded as the motto of this paper: An automatic control system for mine ventilation ill-conditioned or improperly designed is capable of creating hazard situations in response to random disturbances, much more, severe in consequence than a traditional ventilation system without any automatic or remote control! This statement is easy to prove if the dynamic properties of the ventilation process are taken into consideration. The ventilation process, as a matter of fact, is described by non-linear equations, and it must be expected that the process has more than one state of equilibrium. In other words, in the ventilation process may exist not only one but also more than one steady-states of flow, of which some are stable and others unstable. In certain circumstances, there may be no steady-state at all, and the process will oscillate [8], [11] , [12] . The state of flow in a network tends towards a steady-state and the actual steady-state established will depend on the initial conditions or disturbances in flow (fire,. etc.), which steady-state from the total number that will be . We frequently observe jumps from one steady-state to another. Disturbances in flow conditions which may cause such transitions are events of random character, occurring very rarely. Concluding, it must be stressed that there has to be a control system adjusted to the ventilation process in order to avoid situations mentioned above. There is only one alternative available and suitable for examination or study of the dynamics of a given mine ventilation problem: either by continuous monitoring of the real process, or numerical simulation of the process using a mathematical model. The advantages of the second method are obvious. This method allows consideration of every possible case very quickly and cheaply in relation to the first method. The aim of the paper is to show again that the simulation of the mine ventilation process and particularly a methane concentration process, separately or combined together with a control system, are real possibilities. A simulation method requires precise specification of the problem under consideration. For example, if we intend to examine a methane-concentration control system, the following items have to be specified: - expected target function of the control system. - structure of the control system. - mathematical model of control system, including sensor system, data preparation system, controllers, decision routine, regulators, etc. - structure of mine ventilation network. - mathematical model of ventilation process, including air flow and methane concentration processes. - pattern of disturbances which may occur in the controlled process as well as initial conditions on a 'start-up' of the system. Using typical computer programs for numerical solution of equations in the mathematical model of the problem involved, we are able, within the adequacy of the model, to simulate every case specified by the disturbances and initial conditions. As a result of simulation, it is expected that the following parameters could be defined: - transient flow in the network. - transient state of methane concentration in working areas. - stability of flow and methane concent¬ration. - stability of the control system. - range of control. - efficiency of control, etc. It is obvious that simulation methods readily allow for modifications to existing systems such that desired results will be obtained. Also optimisation problems could be solved by use of the simulation methods. In order to illustrate these general thoughts, a brief presentation of a mathematical model of methane concentration and

Jan 1, 1980

By Dwane K. Johnson

INTRODUCTION The small mining company faces the dilemma of how to finance the development of its properties because it doesn't have the financial resources to pay for the development costs from its own funds and doesn't have the financial strength which will enable it to borrow the money needed for development. However, with a little imagination and the right set of circumstances, a financing can be arranged for the small mining company. This paper is an attempt to describe the fundamental steps on how such financing can be obtained --in other words--what is required to provide ad- equate capital at a low cost for developing and bringing a mining property into production. Innovative financing follows from this. The basic keys for building and maintaining a successful mining company are "the four M's": Management, Market, Mine, and Money. The writer believes their order of importance is as presented and each is a building block required in order to obtain the type of financing that best suits the borrower and lender. MANAGEMENT Management must be experienced, highly organized, realistic, able to take advantage of opportunities and fulfill promises. Mine financing requires a variety of skills and concentration on the part of management. Management should have a thorough understanding derived from experience in geology, construction, mining and metallurgy and environment, marketing, and financial matters. Management must realistically visualize what will be produced at what cost and truly evaluate the risks associated so the proper security can be correctly designed. Since risk cannot be eliminated, the objective of management is to identify the risk which will be present in a given venture and assess the level of that risk which will be acceptable to the firm. Generally the risk which is present is subject to little or no modification. After management identifies and measures the acceptable risk the firm will proceed in a way that it will be least exposed. Management must determine its long-term objectives and strategies within the context of a constantly changing world. This question must be addressed before examining the sorts of risk which affect the development and operation of a specific property. The compilation and interpretation of "hard data" by competent people is good to a point but the experience and instinct of seasoned individuals are the important factors in management choosing a long-term direction for the corporation. The attitude towards risk greatly affects the goals of the corporation. Major risks in a mining project can be placed in four categories: 1. Market - Price, demand, substitution. 2. Costs - Capital, operating, financial. 3. Regulation. 4. Taxation. The feasibility study for a project is the fundamental tool used in the management of risk. Management should employ an experienced team with an established and well-understood set of ground rules for the preparation and assembly of a feasibility study. The necessary degree of realism is built into all levels of the study and if the project is technically feasible, the company's hurdle rate is then used to discount the base-case project cash flow in order to determine its financial viability. Understanding the geological significance of the mineral deposit is critical. There should be strong interplay among the pure regional geologist, the detailed mine geologist, the mining engineer, and the metallurgical engineer. The collaboration among these different disciplines will yield a feasibility study that is more reliable. The essence of this work must be appreciated and understood by the executives planning the financing. It is vital that they thoroughly understand the risks and make the proper risk assessments to cope with our rapidly changing market environment. Projecting future revenue values is paramount and is subject to much estimation. One way to mitigate fluctuations and the risk of falling prices is by selling production for future delivery. MARKET Since mining is a worldwide industry it is mandatory that bankers be aware of what is occurring in the marketplace, both foreign and domestic. The marketing of the production is a very important facet when considering financing for the small

Jan 1, 1987

By R. Edward Williams

The Spa1lation Process Because of the low thermal conductivity of many hard rocks, rapid heating of these rocks produces a thin surface layer in which the temperatures attain high values. Thermal expansion of this surface layer is constrained by the reminder of the still cool rock, and when stresses within the surface rock become high enough, the surface rock breaks away from the cooler rock behind it and flies or falls off as a thin flake called a spall. Then the next, newly exposed surface is heated, and the process continues. This process is the basis of spallation drilling. The hot gases from a jet burner provide the heat for spallation to occur, and their high velocity provides a scouring action that transfers heat to the rock and removes the spalls as rapidly as they form. Spallation is a process which works in very hard rock. It is dependent upon the thermal expansion coefficient and the thermal diffusivity of the rock but is also affected by any discontinuities in the rock. To date the efforts which have been made to evaluate the various rock according to their spallability has been minimal. As the success of this process is dependent upon the characteristics of rock it is expected that the study of rock mechanics will prove to be of greater value to this program than to the other mechanism for drilling and excavating rock. Commercial Uses of SPALLATION In the 19408s, the Linde Air Products Division of Union Carbide (UC 1 began developing spallation for use in mining taconite ore, which is presently the chief source of iron in the United States. In this work UC developed a jet-piercing tool that burned fuel oil with oxygen to produce spallation and contained mechanical cutters to remove rock that was not amenable to spallation. The UC jet-piercing machines have since produced about 40 million feet of shallow blast holes used for emplacing explosives in the taconite mines. During this work it was found that hole diameters could be increased by merely reducing the advance rate of the burners and that existing holes could be enlarged by making another pass through the hole with the same burner. The Browning Engineering Go. of Hanover, N.H., has developed a hand-held spallation burner to cut slots in granite. It has been used for a quarter of a century and is now standard equipment for quarrying granite throughout the world. This burner, which resembles a small jet engine oriented with its exhaust pointed downward, is the forerunner of a flame jet burner used to spall experimental holes in granite at maximum rates in excess of 100 ft/hr when operating in hard, competent granite. It uses No. 2 fuel oil, which is burned with compressed air. The system uses water to cool the burner and the exhaust gases. These gases, along with the steam produced from the cooling water, blow the spalls from the hole. Experimental Work Theoretical and experimental work has been accomplished by the Massachusetts Institute of Technology and the Los Alamos National Laboratory. This work is reported in Refs. (3) and (4). To verify the experimental results of this work laboratory scaled down field tests were conducted using two we1 1 characterized granites from quarries in Barre, VT and Westerby, RI, under defined heating conditions. In the laboratory tests a propane - oxygen heating torch was used to direct a flame at the granite surface and the spal 1 ing process was examined at various heating rates with a high-speed video taping system operated at 200 frame per second. This produced a time-lapse sequence where the onset of the spallation process was easily distinguished. Also the heat flux from the torch to a flat surface at various stand off distances and flows was measured. A similar set of tests was conducted using the more easily quantified and uniform heat source of a 1.5 kw GO2 laser. This allowed accurate

Jan 1, 1986

By Robert W. Miller, Rhoda S. Kriesel

INTRODUCTION With each passing year, the general public becomes increasingly aware of the potential hazards associated with many products and services previously considered unharmful. This new consciousness is especially evident in the workplace, both as the result of advances in scientific understanding and the influx of concerned, responsible professionals into the labor movement. For example, data produced by epidemiologists doing retrospective mortality studies which related mortality to toxic substance exposures has often been the prime motivating factor behind reductions in exposure standards. This certainly has been the historical case with many radiation exposure standards, including the limitations set for "Working Level" exposure. The concept of "Working Level" was introduced in PHS Publication No. 494 as a result of both the difficulty in relating radiation units to biological effects and the complexity of interaction of radon daughters with the physical environment. WORKING LEVEL One working level is defined as any combination of radon daughters in one liter of air that will alternately release 1.7 x 105 MEV of alpha energy during decay of 210pb (RaD). Its usefulness is that working level can be readily measured in both the field and the laboratory. Because of this, existing occupational exposure standards for radon daughters in both the U. S. and other countries use working level as their basis. In the U. S., the last twenty years has witnessed a significant reduction in radiation exposure standards from 120 working level months to 4 working level months per year. New international standards are being proposed that would be only slightly higher than these current U. S. standards. POTENTIAL EXPOSURES TO RADON DAUGHTERS The universal use of nuclear power, along with the transfer of nuclear technology to developing nations, has significantly increased the demand for fuels. As a result, uranium mining operations have expanded and increased numbers of miners are regularly exposed to radon daughters. In addition to potential exposures in uranium mines, other deep mines such as tin, gold, platinum and tungsten, also have potential exposures due to the distribution of uranium and its accompanying radium throughout much of the earth's crust. The energy crisis has also prompted interest in general population exposure to radon daughters in buildings. Many homeowners, using additional insulation and caulking to seal cracks and conserve energy, have reduced indoor ventilation to the point where working level exposure could become significant, especially in homes with unventilated crawl spaces. In fact, several serious exposure situations have occurred where mine tailings were used for building materials. MEASUREMENT TECHNIQUES Historically, rapid and convenient measurement techniques have been actively sought to improve the ability of ventilation engineers to limit mining exposures through ventilation control. A number of measurement techniques have been developed, including the Kusnetz method (further elaborated by Shalaynev) and techniques developed by Rolle and Tsivoglou. All, however, suffer from similar handicaps that prevent their usefulness in the mining environment. Their main limitation, in terms of minimizing employee exposure, is the elapsed time from the start of sampling until the results can be calculated. It is usually 40 to 90 minutes. This time delay presents both serious economic limitations as well as the potential for exposing workers to high radon daughter concentrations until calculations are completed and corrections made. Further reducing the utility of the Kusnetz and Tsivoglou techniques is the need for tedious calculations that increase the possibility of human error along with the need for cumbersome sampling equipment that is not ideally suited to the mining environment. INSTANT WORKING LEVEL METER To alleviate these problems, the idea of an Instant Working Level Meter was conceived by several groups. Several such instruments were proposed, built, and tested with disappointing results. Although the concepts behind them were sound, the instruments simply failed to measure working level without excessive distortion. In addition, their cost was relatively high, gamma background presented a problem, and sophisticated computations were necessary to determine working level. In short, these systems lacked the design engineering that would make them suitable for the demands of a mining environment.

Jan 1, 1981

Two years ago, about the only US mining ventures that could turn even marginal profits were precious metals. In 1986, the price of gold dramatically increased, as did earnings from precious metals, mostly gold mines. First quarter results for 1987 show that gold will again be the most profitable type of mining in the US. All this money being made in gold resulted in many new mines being opened and hundreds of old workings being re-evaluated. The Society of Mining Engineers has put together a technical program to aid in developing precious metals mines. Small Mines Development in Precious Metals will take place Aug. 30-Sept. 2, 1987 at Bally's Hotel, Reno, NV. For registration and hotel information see page 426 or contact Meetings Department, SME, P.O. Box 625002, Littleton, CO 80162-5002, 303-973-9550. Topics at this meeting will include Problems Facing Small Miners; Understanding Geology and Ore Reserves; Mining-Production Options to Maintain Competitiveness; Processing for Precious Metals; Financing and Economics that Work for Small Miners; and Understanding the Environmental and Permit Maze. Problems facing the small miner The small miner has found and started most of the productive mines in the world, says Alfred G. Hoyl, of Coal Fuels Corp., Rollinsville, CO. Hoyl will state his case for small mines development in the opening session of the program. He will make the point that government regulations must never restrict the small mine or miner. And the mining industry must be careful not to crowd out the independent miner. "To restrict small mining is to say to our young men and women entering the mineral industry `There is no place for you as an independent.' Small is a relative term and we must never feel that starting small will restrict our possibilities of becoming big." In the same session, David J. Starbuck, Starbuck Minerals Management, Mayer, AZ will take a look at why smaller precious metals mines are profitable. Mining companies have changed their perspective on reworking previously mined properties. They are looking at abandoned properties in older mining districts. The attitude is that smaller operations in these districts can be profitable with the present economic climate, more advanced technology, and better management. Small, independent mining companies may not have the same resources to draw from as the majors. But there are affordable technologies available that fit into the independents' budget. Betty Gibbs, of Gibbs Associates, Boulder, Co will explain how a personal computer can help a small operation to analyze data more effectively and respond quickly to changing conditions. Scenarios will be developed that will describe how PCs can be incorporated into all phases of an operation. Financial institutions are becoming more willing to consider venture capital for small precious metals mines, according to Stanley Dempsey, Denver Mining Finance Co., Denver, CO. This may be in the form of equity or debt, he says, but they want to receive assurances of the quality of the resource. And they will ask about the ore reserves. Next, the potential lender will want to know about the mining method and en¬vironmental regulations. Dempsey will discuss these requirements. Once a high grade gold mineralization is located and reserves proven, it must be evaluated to determine which recovery process is best suited for this deposit. Robert Martinez, Amselco Minerals, Ely, NV will evaluate processes for such a deposit located in Mexico. Heap leaching, vat leaching, and a carbon-in-pulp circuit are all viable options, he says. He will explain which recovery process proved to be the most economic. Other papers in the introductory session include How to Best Use Consulting Firms to Accelerate the Development Process, by H. Schreiber and R.P. Fronk, both of Behre Dolbear-Riverside, New York, NY. And W. Mayrsohn of Asamera Minerals, Denver, CO will present Marketing Fundamentals that Work for Small Mines. Understanding geology and ore reserves In the current scheme of fast-track exploration and mine development, insufficient time and thought is given to understanding the geology of a deposit as it relates to controls of mineralization and subsequent mine development, say D.S. Bolin and J.W. Rozelle, of Pincock, Allen & Holt,

Jan 6, 1987

By Jon P. Thorson, James W. Babcock

The second Belt Symposium in Missoula, MT, Oct. 9-16, 1983, was designed by its conveners as a field conference on progress in the Proterozoic Belt Supergroup rocks. The symposium was sponsored by the University of Montana with corporate support by Anaconda Minerals. Asarco Inc., Exxon Minerals Co.. Hecla Mining Co., Molycorp Inc., Noranda Exploration Co., and Utah International Inc. Introduction The Belt Supergroup is a thick middle Proterozoic sedimentary sequence deposited in a basin along the edge of the North American craton. Subsequent geological events have modified the basin so its remaining rocks outcrop in Montana, Idaho, northeast Washington, southeast British Columbia, and southwest Alberta. The Canadian equivalent of the Belt is called the Purcell Supergroup. Several similar age basins occurred along the western edge of the Proterozoic North American craton. The sedimentary sequence in the Belt basin aggregates is more than 18 km (60,000 ft) in the thickest section, where the bottom is still unexposed and the top removed by erosion. The limits of the Belt basin are poorly known because of later sedimentation and tectonic disruption. Knowledge about the Belt terrane has been difficult and tedious to obtain because the rocks are generally monotonous and fine-grained. During the last decade, new 1:250,000 scale reconnaissance maps of most of the Proterozoic Belt terrane were published, open-filed or otherwise brought close to completion. These maps provided the opportunity and challenge to consider the stratigraphy and tectonics of the Belt Supergroup as a complete basin. This broader view has contributed greatly to the solution of numerous geological problems created by previous, more focused work. Along with the availability of better geological base, there have been new exploration developments. The 1973-1983 decade saw proof of the commercial viability of the Cu-Ag mineralization at the Troy mine. The conclusion was also accepted that the Sullivan mine was a sedimentary massive sulfide. So was the conclusion that the Cu-Co mineralization at the Blackbird mine at Cobalt, ID, was the result of similar submarine exhalative processes. And there were discoveries of another Cu-Ag deposit "larger and higher grade" than the Troy mine and of two new massive sulfide districts. The Belt-Purcell terrane has long intrigued exploration and mining geologists as the host of Pb-Zn-Ag veins of Idaho's Coeur d'Alene district and of the Sullivan mine at Kimberley, British Columbia, one of the largest Pb-Zn-Ag deposits in North America. In the mid-1960s, Bear Creek Mining Co.'s discoveries of the stratabound copper-silver occurrences in the Revett Formation brought new interest in the basin's geology. The recognition of the Sullivan ore body as a sediment-hosted massive sulfide has also generated minerals exploration activity in the US portion of the Belt basin. This activity increased again after the Australian discovery of the world class Olympic Dam deposit at Roxby Downs, South Australia. This occurred because of some suggested similarities in the regional geologic setting of the Belt basin with the Adelaide geosyncline, the host for the Olympic Dam deposit. Technical Sessions Technical sessions topics included Lower Belt and Ravalli Group stratigraphy, Middle Belt carbonate and Missoula Group stratigraphy, Belt-Purcell stratigraphy, structure and tectonics, strata-related mineral resources, geochronology and geophysics, geochemistry and isotope geology, and a summary of significant problems. Evening poster sessions dealt with the topics and papers presented each day. Since the first Belt Symposium, mapping of the stratigraphic units and their tectonic setting has enabled geologists to progress in conducting a basin analysis. Belt Symposium II revealed no concensus on many aspects of belt geology. However, basic stratigraphic studies have reached the point that models can be constructed. The Prichard Formation is equivalent to the Aldridge Formation, but detailed correlations were not previously available. Studies were made by Finch and Baldwin (Cominco American) and by Cressman (US Geological Survey). These studies established important correlations of key markers within the Aldridge Formation to units in the Prichard Formation. This will help eliminate geologic interpretation problems at the US-Canada frontier and make Belt basin reconstruction more accurate. Also, the Sullivan ore deposit was shown to be formed in the deeper distal part of the basin. This contrasts to the more proximal depositional position of the Prichard Formation. Hamilton

Jan 6, 1984

By C. Allen Born

There is a saying, the greater the gratitude, the shorter the speech. That puts me in a quandary, because while I am grateful for the honor you do me today, I also want to talk about a subject very important to all of us. A few months ago, when I was asked to submit a title for my talk, I chose "Quality where it counts." I did so because perhaps the most important thing we at Amax did in our restructuring was to commit to quality in all our businesses - quality people, assets, and operations. Together, they create quality performance, which produces quality growth and increased shareholder value - our prime corporate mission. We are dedicated to quality. At Amax, quality means everything. Today, I want to discuss quality in another context - the need for it in American education and what we have to do to achieve it. If we ignore this need, we will all suffer the consequences. And they will be upon us far sooner than the next century, barely a decade away. I think almost everyone in this room works for an organization that had to restructure in the 1980s. Mining companies that did not, just didn't make it. Well, we did. As an industry, we are seeing improved profits, positive cash flow, and a more competitive position in the world marketplace. But all the restructuring we have done, all the hard work, and all the pain that was part of it, will be in vain if, down the road, we do not have the educated people we need as workers. We need them to design, manufacture, and market the products that use the basic materials produced by the mining industry. Simply put, we are at a crisis point in education today. That is no new revelation. Nearly six years ago, the National Committee on Excellence in Education published its report, "A Nation At Risk." Since then, things have not gotten better. The Wall Street Journal, Business Week, US News & World Report, and Fortune have all recently published stories about the sorry state of education in the United States. The subject has become the topic of TV talk shows as well as national commissions, and for good reason. The pool of young people entering the US labor force is dwindling. In the next decade, 10 million fewer will enter the job market compared to the 1970s. But fewer of them will be qualified to do the jobs available to them. Here's why: • 20% of today's high school students drop out before they are 18 and another 20% graduate functionally illiterate; 40% of today's fourth graders think the world is flat; • 20% of today's sixth graders cannot locate the United States on a world map; • one of every three ninth graders cannot figure the change for a two-item meal - and I'm talking about a Big Mac and fries. These are not just problems of minority kids in big city ghettos. Fully 40% of all young working Americans cannot add up their lunch bill. Three out of every five college freshmen need some kind of remedial work. These statistics are frightening. Particularly so because, while the work force is becoming less qualified and our public education system is getting worse, our products and technologies are becoming more complex. We are approaching a time when we will have machines that recognize handwriting but people who cannot write. We will have technology to take voice commands, but people who do not know what they are talking about. The accelerating pace of technology is changing the nature of the job market. The fastest growing jobs now, and in the years ahead, are in professional, technical, and sales positions requiring the highest education and skill levels. Most of these jobs will require the use of a computer. Yet, 95% of today's graduating high school students have never laid a finger on a computer keyboard. While our education system is failing the students it is supposed to serve, other countries are doing far better. In a

Jan 1, 1990

By J. W. Markham, M. Atiemo, K. Yoshida

INTRODUCTION In open-pit mining of high grade ores, operators of mining equipment in a pressurized cab are protected from inhaling radioactive aerosols by the use of filtered air through a pre-impactor and a HEPA (high efficiency particulate air) filter. At present, a limited amount of scientific information is available on the worker exposure to airborne alpha emitters in the mining of high-grade ores. (Miller, 1977, Nielson, et al, 1979). Factors affecting the worker exposure to alpha radiation are: the grade of uranium containing ores, the environmental conditions inside the mine pit, the protection factor of the cab filter system, the workers' attitude towards the hazards of radioactivity and the adopted contamination control in the cabs. Four series of field samplings were conducted during the fall of 1980 and the summer of 1981 at a mining site in northern Saskatchewan to investigate the extent of protection an equipment operator has against cab internal exposure to airborne alpha emitters with short and long half-life radionuclides. The variation in the breathing zone concentrationnof aerosols as a result of resuspension is discussed, however no attempt was made to isolate and investigate the factors and the associated parameters of the aerosol resuspension phenomenon. AEROSOL RESUSPENSION IN A CAB Mechanisms of Resuspension [Sources]: The presence of radioactive aerosols inside a cab and in the breathing zone of an operator may be due to initial concentration inside [(CO)], concentration in the filtered (HEPA) air (CC, at the flow rate [QC]), the concentration of the aerosol through resuspension ([CR]), the aerosol concentration in the air outside the cab ([CE]) and the concentration of the aerosol in the leaking air [(CL)], at the flow rate of [(QL)]. [Volumetric Equilibrium]: Assuming a uniform concentration of the aerosol in the cab, the aerosol concentration in the breathing zone[ (CBZ)] may be expressed as, [CBZ+CO+CC+CR-CL] If [QC] remains greater than [QL], there is pressurization in the cab. There is particle loss as a result of the leaking air from the tight cab, however the condition of adequate pressurization ensures that [CL ~ CC]. Also for dust particles filtered through HEPA [CC – 0] thus [CBZ + CO + CR] For the radioactive gases, [CBZ - CO + CC] Over long periods,[ CBZ – CE] but if [CBZ] exceeds the value of CE then the aerosol build-up in the cab has, occurred. [Resuspension]: Turbulent bursts are generated within the core region of a cab confinement, such that the down sweeps and the up sweeps of the air bursts cause resuspension of dust from the surfaces. Particles initiate resuspension by four major modes: namely, sliding, rocking, rolling, bouncing or sudden ejection or by combinations of these. Nevertheless the adhesion force between the particles and the surface must be overcome by ejection forces before ejection takes place (Kline et al, 1967, Corino and Bradkey 1969, Kim et al 1971). The phenomenon of particle resuspension from a surface may be expressed in a generalized form as: CR =f (Uz . R . D . H) where CR = Rate of particle resuspension UZ = Surface shear velocity R = Surface roughness parameter D = Aerodynamic diameter of particle H = Frequency of surface airburst A detailed discussion of the individual parameters was attempted for the case of farm tractor cabs by Atiemo (1981), such a treatment is out of the scope of this study. Only the effects of the properties of surface roughness will be discussed. Effect of Surface Properties The roughness of the surface from which dust particles are ejected plays a significant role in the deposition and resuspension of the dust. The roughness or the amount of protrusions greatly influence the intensity of air turbulence in the immediate vicinity of the surface (Schmel 1976). For most roughened surfaces therefore, larger particles may be dislodged in greater quantities than smaller particles. The deposition and the resuspension of dust particles from different surface materials of glass, cloth, metal and rubber have been studied in a wind tunnel (Atiemo, 1981). Field Conditions Mining equipments used are: a backhoe, a bulldozer with ripper attachment and two dumptrucks for high-grade ore operations; a payloader and two dumptrucks for waste handling; and a jumbo drill which drills dynamite holes for ore blasting. All this

Jan 1, 1981

By P. R. Taylor

The ability to process complex precious metal sulfide concentrates containing antimony and arsenic in an economic and environmentally safe manner is an important problem facing the mining industry in the United States. Preliminary experiments have been performed to evaluate the possibility of treating a complex sulfide concentrate to recover antimony and precious metals while fixing the arsenic and sulfur as calcium arsenate and calcium sulfate. The process involves: • the oxidation of the antimony and arsenic minerals in the presence of soda ash (to collect the sulfur dioxide and to keep the antimony and arsenic from forming volatile compounds), •the selective leaching of the arsenic from the roast calcine with a heated caustic solution, •the precipitation of the arsenic and sulfate with calcium hydroxide and/or calcium chloride, •the reduction of the residual antimony (collecting the precious metals), •fuming of the metal to form a marketable antimony oxide, and •crucible reduction of the residue to obtain a gold and silver dore. Batch experiments, using synthetic concentrates of stibnite and arsenopyrite, were performed to evaluate the first stage of this process. Variables studied included mixture weight fractions, temperature, gas flow rate, oxygen partial pressure, and time. All of the off gas dust and sulfur containing gases were collected and analyzed. The roast solid residues were subjected to arsenic leach tests, and the leach solutions were subjected to arsenic and sulfate precipitation tests. Introduction The ability to treat a precious metal ore containing appreciable amounts of antimony and arsenic is dependent upon the characteristics of the ore. For some ores, selective flotation can be used to separate the antimony and arsenic minerals, which are then treated separately. Unfortunately, many ores are not treatable in this way, and the concentrates obtained may contain appreciable contamination (which limits the ability to make acceptable grade antimony products). If we consider the added value of the precious metals in the concentrate, then it is worthwhile to evaluate the potential for new methods to treat these types of concentrates. In addition, the environmental difficulties associated with arsenic disposal must be taken into account in any proposed process. If there is appreciable arsenic present, then it must be either recovered as a marketable product or placed in an insoluble form. The problem addressed in this research program has become increasingly important as the precious metals mining industry is encountering more ores of this type. A large deposit of this type exists in Idaho where the antimony occurs primarily as stibnite, with minor amounts of kermesite and tetrahedrite (Cooper, 1951). The gold appears in association with arsenopyrite with minor amounts of other arsenic minerals. Flotation has been used to obtain an antimony concentrate. The concentrate, however, is contaminated with appreciable amounts of arsenic and also contains precious metals. If the ore or concentrate contains appreciable amounts of base metals, then the recovery process becomes much more difficult. There are a number of other known deposits of this type in Alaska, Canada, Montana and elsewhere. The objective of this research program was to develop a fundamental understanding of the oxidation and leaching chemistry of the arsenic and antimony minerals so that technology may be developed to treat these ores in an economic and environmentally safe manner. Wendt (1949) and Koster and Royer (1940) discussed the possibilities of oxidative roasting of stibnite concentrates containing arsenic. Wendt describes a process used at a smelter in Czechoslovakia (in the 1930s) where roasting was carried out in rotary furnaces at 723° K (500° C). Most of the arsenic was removed as fume, as well as 25% of the antimony. No sulfur dioxide containment was described. The oxide fume from the furnace went to a heated caustic water leach, where the arsenic was preferentially dissolved and then precipitated with calcium chloride as calcium arsenate following solid-liquid separation. The antimony oxide was sold, and the solid oxide discharge from the rotary furnace was sent to a blast furnace where the antimony metal, containing the precious metals, was recovered. The metal was fumed to obtain antimony oxide, and the precious metals were left in the residue and collected with copper. The copper

Jan 1, 1994

By R. E. Utting

INTRODUCTION Until recently, gamma doses had been largely ignored in Ontario uranium mines. This has been due to the assumption that these doses are small and have been more or less unchanged with time and hence their effects have been included automatically in the epidemiological studies that led to the establishment of radon daughter exposure limits. This assumption had to be challenged for two basic reasons. The first was that radon daughter exposures to miners have been progressively reduced over the years due to improved ventilation and ever more stringent regulations, while gamma exposures have presumably remained relatively unchanged. Therefore it must be assumed that the ratio of gamma to radon daughter exposure has gone up. The second reason is more philosophical. It is clearly inappropriate to make judgements on the significance of a potential industrial hazard when the magnitude of that hazard has not been fully assessed. Having decided that some sort of assessment of gamma exposures to uranium miners must be made, it was than necessary to determine how this should be done. Several options were available, for instance: (i) Wholesale personal gamma dosimetry for all mine and mill workers, (ii) Personal gamma dosimetry only for those workers suspected of receiving the higher doses, coupled with area monitoring to estimate the exposures of other workers, (iii) Area monitoring coupled with dose rate times time calculations for all. This would correspond to the generally prevalent method of assessing radon daughter exposures. It was argued that since radon daughter exposures are the major radiological hazard in uranium mines, to invest resources for assessing a lesser hazard to a greater degree of precision was not cost effective. (iv) Since gamma dose rate is related to ore grade, individual doses could be assigned from knowledge of work location and ore grade. Before deciding which of these options would be most appropriate, it was necessary to have some idea of the magnitude of the problem. Very few data were available in the literature and with the exception of a few spot dose rate measurements, and the results of a few gamma dosimeters issued to selected individuals by some of the mining companies, nothing was available. A rule of thumb of obscure origin is often quoted within the industry indicating that gamma dose rates underground will be about 0.25 mR/h per lb/ton or 5 mR/h per % U. This had been used by some to justify neglecting gamma radiation at least for ore grades of the order of 0.1% or 2 lb/ton, on the grounds that gamma dose rates would be of the order of 0.5 mR/h and therefore give rise to annual doses of only about 10 mSv (lrem). That is, it was assumed that gamma radiation was of limited concern compared to the hazard associated with the inhalation of radon daughters. We were thus faced with the situation of just assuming that no regulatory limits were being breached. This situation could not be allowed to continue. A program was initiated to investigate the gamma doses absorbed by uranium miners in three mines in Ontario, and extensive gamma surveys were conducted in the Quirke 2 mine of Rio Algom Ltd, Elliot Lake; Denison Mine, Elliot Lake; and Agnew Lake Mine, Espanola. Negative reaction was received from several mine company officials to the possibility of all miners being required to wear personal gamma dosimeters due to the logistical difficulties involved, and therefore part of the project was aimed at determining if a reliable correlation between gamma dose rate and ore grade in the work location could be deduced, in order that dose rate times time calculations might be used for gamma dose assessments. The results of these programs provided evidence that the gamma dose for some employees in the three mines investigated may be a significant fraction of the current maximum permissible annual dose of 50mSv (5 rem). When combined with radon daughter exposures in the manner recommended by the ICRP at their 1980 Brighton meeting (ICRP 80) the results indicated that some individuals will come close to the resulting limit and may even exceed it. The results also indicate that is probably not feasible to develop a reliable formula for

Jan 1, 1981

By Harold Stocker

INTRODUCTION This presentation is concerned with the administrative and technical capability of the Atomic Energy Control Board (AECB) to assure compliance with the individual exposure limit for radon daughters (currently 4 WLM per year). It is not concerned with the epidemiological bases for setting the exposure limit. Moreover, the intent is to show how sophisticated methodologies and advanced technologies, applied to radon daughter concentration measurements in uranium mines, convey the spirit of compliance by providing better estimates than do the historical methods. These better estimates mean that more accurate and more precise estimates of each worker's exposure are determined using these more modern methods and devices. The estimates so derived should provide more convincing evidence to an individual worker that his assigned exposure is a valid indicator of his true exposure. In addition, a perspective on the exposure estimate in relation to the exposure limit is given as further evidence that an exposure limit is not the dividing line between "safe" and "unsafe" exposures. A brief description is given of the compliance aspects of the Atomic Energy Control Regulations and of the limitations of purely statistical non-compliance procedures. Most of the emphasis of the paper will be placed on the uncertainties associated with conventional radon daughter exposure determination and the means being employed (and anticipated) to reduce these uncertainties. NON-COMPLIANCE Under current Atomic Energy Control Regulations (1978), the annual individual exposure limit for radon daughters is given without reference to the possible methods of sampling and calculation of radon daughter exposure and without any reference to possible uncertainties or their magnitudes. This is common in such statutes, the details of sampling, calculation, error analysis, and so on, being left for licence conditions or provided as a specific guideline to the licensee on the matter of compliance with the Regulation. Since the exposure limit is contained in the Regulations, compliance with it is absolute, as with any other law. In Canada, a state of non-compliance with the radon daughter exposure limit exists when an exposure (attributed to an employee) is reported by the licensee to exceed the limit. No uncertainty in the measurements or in the overall determination of exposure is reported nor is any requested. Removal of the worker and the loss of his services are the immediate and direct penalty suffered by the licensee for failure to maintain the exposure at, or below, the limit. A worker may be re-instated to employment for the balance of the reporting period only if the licensee can assure the AECB that further significant exposure to the worker will not ensue. In other jurisdictions, such as the United States, non-compliance is defined on a statistical basis. For example, NIOSH, the National Institute for Occupational Safety and Health presents procedures for calculating the 95% Lower Confidence Limit (LCL) in order to "compare the results of occupational environmental sampling to an occupational health standard and make a decision with a known chance of making an incorrect decision that a state of non-compliance exists" (Leidel and Busch, 1975). (In the nomenclature of this presentation, exposure limit would be used in place of "standard", in the NIOSH sense). Furthermore, it is emphasized in the NIOSH document that such numerical comparisons "are necessary only if the sample mean is greater than the standard". The NIOSH document points out, quite correctly, that the "statistical procedures presented below will not detect and do not allow for analysis of highly inaccurate results, i.e., systematic (non-random) errors or mistakes ... If a systematic error is known to exist in an instrument or analytical procedure then correct the sample mean of the data before analyzing for non-compliance". It is certainly not the purpose of this paper to criticize the sophisticated statistical approach to non-compliance as given in the NIOSH document or in similar approaches used in other jurisdictions. Rather, the purpose is to approach, with some introspection, the question of the determination of exposure by the employer for his employee and especially the employee's understanding of, and confidence in, the accuracy of the exposure determination and its relation to the exposure limit. DETERMINATION OF EXPOSURE Historically, in uranium mines, exposure to radon daughters for an individual miner

Jan 1, 1981

By C. J. Johnson, C. J. Bise

Coal extraction by continuous miners (CM) is currently the most common underground method in the US industry and accounts for slightly more than two-thirds of the nation's deep mining production (National Coal Association, 1987). Even if longwall mining should become more commonplace, it can proceed only after ventilation and access entries have been driven by CMs. Since an area of concern continues to be the effects of the dust generated on the health of mineworkers, this paper discusses the relationship between the elemental compositon of mining-generated airborne dust sampled from the immediate ventilation return of a CM and laboratory-generated dust derived from channel samples taken from the mines. There are several potential contributions of this type of study to the coal mining industry. First, after more fundamental knowledge of the causes of Coal Worker's Pneumoconiosis (CWP) is learned, the laboratory-generated respirable dust could be used to identify a potentially hazardous coal seam. Also, this study could possibly aid in understanding the fundamental causes of CWP by producing mining-simulated samples of coal dust that could be used in epidemiological studies. Further. assuming that there is no difference in the elemental composition of a drill-core sample and a channel sample from the same location, a mining company could predict anew mine's respirable dust elemental composition in the immediate ventilation return by using exploratory drill core samples of the roof, coal, and floor rock to prepare the laboratory dust. Ventilation engineers could then use engineering design and control measures during premine planning to reduce the incidence and severity of CWP by better ventilating the potentially hazardous coal seam. If this proper planning prevented any future changes to the ventilation equipment and mine design. much time and money could be saved. Scope of work To investigate the variability of the chemical characteris tics of respirable dust, airborne dust samples from eight underground coal mines located in the eastern and midwestern United States were collected with eight-stage Sierra Model 298 Marple cascade impactors, as well as 25 channel samples of mined material. Each channel sample was removed from the middle of the coal face before mining occurred. Sampling of the mining-generated dust was conducted by Lee (1986) by sampling the entire working sections, primarily for characterization purposes, to obtain information on the locational variability of dust characteristics. Research performed for this study used the elemental analyses of the mining-generated dusts he sampled in the immediate ventilation returns of CMs. The procedure that was used to produce the laboratory generated respirable dust was based on the Hardgrove grindability test since it reflects the pulverizing characteristics of coal. This test was chosen for several additional reasons. First, it is repeatable and reproducible. A consistent amount of input energy is used as well as a specified size range of feed material to be crushed (the channel samples). Second, it is thought to generate secondary dust in a way similar to that of the crushing and grinding of the coal and rock as they pass through the arc-shaped cutting path of the CM's cutter head. The potential effect on dust generation by this secondary grinding mechanism may be at least as much as that produced by primary fragmentation, which is dust produced by the cutting action of the bit against the coal or rock (Roepke. 1984). Finally, the Hardgrove grindability test is well known and is used in the coal industry to guide mineral processing engineers in estimating the capacity of mills used to grind coal. One hypothesis of dust researchers in the Generic Technology Center for Respirable Dust is that the elemental as well as the physical characteristics of coal mine dust will make a difference in the incidence and severity of CWP. Coal mine dust is generated not only from coal, but also from any rock partings contained within the seam or any roof or floor material mined with the coal. Thus, coal mine dust may not have the same elemental characteristics as the coal being mined. Given that hypothesis, mixtures proportional to each thick¬ness mined of roof, coal, and floor rock derived from the channel samples of the face areas from which the respirable dusts were generated by the CMs were used to produce the feed material that was pulverized in the Hardgrove machine.

Jan 1, 1990

By William A. Rhoades

INTRODUCTION Ever since miners were faced with the task of moving ore, some form of scraper has been in use. At first, men and beasts of burden supplied the power to move the scraper, and later, machines were developed for this purpose. In the early 1900s, a few mining properties used small pneumatic single-drum winches to pull loaded scrapers to a raise or ore pocket, and the empty scraper then was dragged back to the muck pile by a miner, as shown in Fig. 1. Just prior to 1920, an improvement was introduced, using two single-drum hoists. As illustrated by Fig. 2, the second hoist was used to return the empty scraper to the muck pile. However, this arrangement still re¬quired two men, one operating each hoist. The next developmental step was to eliminate the second man by locating the hoists side-by-side and using one man to control both hoists. As illustrated by Fig. 3, this involved the use of a tail rope over a sheave to pull the scraper back. The greatest progress in development and the great¬est increase in the use of slusher haulers occurred between 1920 and 1930. In 1921, the Sullivan Machinery Co. designed and built the first two-drum scraper oper¬ating on the principle illustrated by Fig. 4. It was powered by a 4.5-kW (6-hp) Turbinair(r) motor and would pull a 450-kg (1000-1b) load at 0.61 m/s (120 fpm). In 1922, this unit was shipped to the Verona Mining Co. of Caspian, MI, and it experienced immediate success in the Lake Superior iron ore district. Since the two-drum slusher was much less expensive and more efficient than hand mucking, the Lake Superior mines were saved from financial disaster when iron-ore prices fell 25% between 1923 and 1925. Immediately there¬after, a demand developed for slushers that would oper¬ate with electric power, which was considerably cheaper than compressed-air power. In 1923, the Sullivan Machinery Co. responded with the first electric-powered double-drum hoist. During subsequent years, design improvements in¬cluded separate tail-drum gearing to increase the tail¬drum speed, as welt as a number of safety features such as rope guides. One result of these improvements and their utilization by the Michigan iron mines was an in¬crease of 100% in the tons of ore per miner per day in those mines between 1924 and 1929. Since the two-drum slusher was capable only of straight-line mucking, it was not a practical machine for use in open stopes. In 1929, the Sullivan Machinery Co. introduced the first three-drum slusher. As illustrated by Fig. 5, two tail drums and one hauling drum were provided. A tail sheave could be placed at each side of the stope, and the ore then could be loaded and hauled to a central point from the entire width of the stope. During the 1930s, progressively larger slushers were demanded. By 1940, two- and three-drum units were available with motor power as high as 45 kW (60 hp), and slusher power continued to increase after 1940. Be¬tween 1951 and 1952, Joy Manufacturing Co. designed a 112-kW (150-hp) two-drum Blusher for the Climax Molydenum Co. Although there has been no demand for a slusher more powerful than this, 150- to 225-kW (200- to 300-hp) slushers are quite feasible at the pres¬ent time. During the last 30 years, many slusher improvements have been made to the operating life, operational safety, and ease of operation and maintenance. Increased tail-drum speeds have decreased overall scraping times. Rope guards, totally enclosed drums, and operator shields have reduced the hazards of injury due to wire¬rope breakage. Improvements in lubrication have made the slushers relatively maintenance-free, with long operating lives. The introduction of spring-actuated drag brakes prevented uncontrolled unreeling of dis¬engaged drums, allowing the development of practical remote-control slusher operation. Remote control now is available in a choice of all-air, all-electric, or air¬electric slushers. APPLICATIONS Quite simply, slushers are used to load and transport material (ore), generally over a short distance of from

Jan 1, 1982