Search Documents

By Robert W. Miller

There are many problems faced by ventilation engineers in deep underground mining operations, not the least of which is controlling miner exposure to radon gas and its daughter products. Radon gas is commonly found in uranium mining operations, but may also be present in other deep metal mines. For example, tin mines in England, iron ore mines in Sweden, gold mines in South Africa, and molybdenum mines in the U. S. have potential radon exposures. This is because uranium and accompanying radium ore are ubiquitous to the earth's crust albeit at low levels. The fact that the activity represented by one WL can be caused by a relatively low concentration of radon gas increases the difficulty of control. Since the source of the radon gas is usually widespread throughout a mine, local exhaust ventilation is not a viable control schema. The technique used to control exposure is then dilution ventilation and, in fact, huge amounts of air must be moved in order to reduce potential exposures to an acceptable level. An interesting comparison can be made of ventilation rates in different types of mines. It is estimated in modern coal mines, which are generally acknowledged to have high rates of ventilation, that about eleven tons of air are moved for each ton of ore mined. A typical operating uranium mine may have ventilation flows of 14-15 tons per ton of ore mined. This provides an idea of the scope and importance of ventilation in modern mining operations where radon is a hazard. Further pressure is put on ventilation engineers by the steady downward trend in exposure limits set by national and international standard setting agencies. Much of this tendency toward lowered standards is based upon longitudinal mortality studies of miner populations. Another important factor is the limited number of experienced miners available in the labor pool. For optimum production, it is important to have as many experienced miners underground in each shift as possible. However, the average daily exposure in a U. S. mine must be less than .3 WL to permit the miner to work underground for a full year. The ventilation system then must provide enough uncontaminated air to maintain the WL below the .3 TTL level to maximize production efficiency and minimize personnel turnover and the problems associated with it. Ultimately, the goal of the ventilation engineer and health physicist is to protect the working miner from harmful exposures based upon currently acceptable standards. U. S. Federal regulations require that in uranium mines all active work sites must be monitored every two weeks if they measure above .1 WL. Areas that have .3 WL ratios or higher must be monitored on a weekly basis until five consecutive weekly samples show the level has dropped below .3 WL. Also, exposure records must be kept for all individuals exposed to levels exceeding .3 WL. These requirements provide a strong economic incentive to have a ventilation system that minimizes exposure of any personnel. A good ventilation system requires careful planning, operation and backup in order to fulfill its mission of providing adequate clean air. Its proper operation also requires coordination with production personnel so it can be adapted as new areas in the mine open up and old areas are sealed off. The ultimate indicator of ventilation efficiency to control radon daughter exposure is, of course, monitoring working levels. Historically, this has been done using the Kusnetz, Tsivoglou, and Rolle's methods, among others. These methods all require cumbersome equipment and tedious calculations to obtain the measurements that results in WL. More important, however, they require a significant time lag between sampling and counting, typically 40-90 minutes. This time lag is, in fact, what can cause significant economic losses due to unnecessary downtime as well as high WL exposures. In a typical mining situation, a sampling technician using the Kusnetz method takes a sample, moves to the next location and takes another sample and so on. Forty to ninety minutes after the first sample, the technician will stop, run the activity count on the filter and calculate the WL. The technician may be one-half mile away or several levels removed from where the first sample was taken when it is counted. If the WL ratio is high the technician must then backtrack to the sample position. There are then two options. If the sample area is a working stage, it can be shut down or a second sample can be taken. If the first alternative is chosen; i.e., shutdown and correction of the ventilation, then another sample must be taken, followed by a forty minute wait for results. If the ventilation adjustment didn't correct the problem, then the whole process must be repeated with a minimum of forty-five minutes per sample cycle when using the Kusnetz method. It has been estimated from operating uranium mines that the cost per hour for downtime on a production slope is about $1,50O/hour. The time lag between sampling and resultant data can be very costly. If the second alternative is chosen to verify the first reading, the miners may be unnecessarily exposed to high levels while waiting for the result. Clearly, such a sampling system can be markedly improved by eliminating the excessive time lag between sampling and analysis.

Jan 1, 1981

By Steve Hansen

Heap leach mining has long been an economical method for mining oxide copper ore, and with the U.S. Bureau of Mines' discovery of a way to economically strip gold and silver from carbon, it has become the standard in the gold and silver indus­try as well. Sulfide leach in copper recovery has always been a byproduct of oxide leach mining, although the methods and recovery rates have increased with technology. SX-EW has kindled a new interest in large deposits and dormant dumps of low grade sul­fide ore. Gold ore bio-oxidation was developed in the 1980's, but not tried on a large scale until US Gold built the Tonkin Springs Mine. This mine was dedicated to the large scale production of low grade sulfide ore that could be recovered through the newly discovered bio-oxidation method. On a small scale the method worked, but as time went on, problems with the process were costly and recovery was poor. US Gold did not have the finan­cial backing to perfect the process, and the mine was closed in 1990. Since that time, Newmont Gold Company has taken the lead and has dedicated a lot of resources to the perfection and implementation of this method. Some small mines have discov­ered that if they can use this technology they may be able to dou­ble and triple current ore reserves, because they will be able to mine deposits previously considered not to be economical.

Jan 1, 1998

By Zachary D. Jones

"Construction projects involve a great many risks. The risk that material prices will increase during construction or that the work under construction will be destroyed by fire, storm or some other act of God are just a few examples. In tunneling or other underground projects, one of the most significant risks is that the subsurface conditions encountered during construction will differ from what is expected. Each of these risks must be accounted for, managed and, ultimately, the duty of one party or the other to bear. The primary tool used to allocate the risks among the parties is the construction contract.How construction contracts allocate the risk that subsurface conditions will differ have changed substantially over the past century. Today, construction contracts typically allocate the risk through a series of provisions. The first is the differing site condition or changed condition clause. That clause, when included, typically provides that the owner bears the risk that the subsurface conditions will differ from what is indicated in the contract documents or what is reasonably expected. The second clause is one that provides that the contractor must conduct a reasonable site investigation and bear the risk of any conditions that such an investigation would reveal. The third provision includes disclaimers that are often included in geotechnical data provided by the owner (or the owner’s engineer) to the contractor which attempts to limit the information on which the contractor can rely. Because these risk-allocating devices seem to be somewhat conflicting, a brief discussion of each provides critical insight in understanding how the modern construction contract allocates the risk of differing site conditions.At the start of the 20th century, construction contracts looked substantially different than they do today. For example, one contract for the construction of a public water tunnel in Chicago, IL, drafted around 1895 contained the following clause:The contractor must take the work entirely at his own risk. No extra allowance will be made for quicksand, hardpan or boulders."

Jan 1, 2014

The methods used by Newmont Gold Company for deposit modeling are illustrated by examples taken from current operations and advanced projects. These methods vary depending on a number of factors, including geology, mining and processing methods, economic feasibility, and availability of historical information. It is argued that successful modeling requires understanding these factors, knowing the strengths and weaknesses of the various modeling techniques, and being flexible in selecting the appropriate methods.

Jan 1, 1997

By J. R. Allan

Loki Gold Corporation of Vancouver, B.C. is developing the Brewery Creek Mine, an open pit heap leach gold mine, 55 km East of Dawson City in Yukon, Canada. Mineable reserves are 17.1 million tonnes of ore grading 1.45 grams/tonne, containing 24.8 tonnes (797 000 oz) of gold. Ore processing will employ a cyanide heap leach using run of mine ore. Permitting this mine has drawn significant attention because the heap leach technology, while proven in other jurisdictions, has not been used in Yukon. Engineering design of the heap leach was similar to that for heaps in more temperate climates, with the added issues of permafrost and cold weather operations. Solution balance with seasonal placement of ore and year -round leaching requires very large pond capacity for spring thaw conditions.

Jan 1, 1996

By Trenton Cohen, Joseph McDivitt, Shimi Tzobery, Joseph Sinacori, Scott Zylstra

The 2-mile Beach Interceptor Sewer Tunnel was constructed in 1954 to then-current mining and construction standards. Due to its lifespan and deteriorating condition it is now being stabilized and rehabilitated using conventional excavation and ground support. At its southern end, below the prestigious Three Arch Bay community, the original tunnel alignment passes under a historic landslide where the tunnel was sand-backfilled and capped by bulkheads in 1957 due to severe and caving ground conditions. Video footage of the existing 21- inch clay pipe showed that the pipe is severely damaged, which could potentially cause a catastrophic pipeline failure and sewage spill into the Pacific Ocean. This paper summarizes the thorough investigation and design process, tunnel realignment construction and monitoring, and the notable cooperation between all parties to overcome the project risks.

Dec 1, 2022

By Jennifer King, Julius Strid, Steve Brandon

The Upper Hidden Basin Diversion (UHBD) tunnel was driven approximately 1.2 miles by the drill-and-blast method primarily through massive granite bedrock geology for the Kodiak Electric Association, Inc. (KEA) in Kodiak, Alaska. The tunnel was driven from both portals with a constant 11-ft straight-leg horseshoe cross-section. The design and construction of the tunnel was driven by logistics and weather conditions due to its remote location and the rugged terrain of the project setting. This paper will discuss the successful completion of a challenging project over a ve-year period from preliminary planning through the design and construction phases.

Jun 23, 2024

By Steve Kral

The third edition of the Underground Construction Association of SME?s (UCA) Cut-ting Edge series of conferences focused on a challenge every underground construction and tunneling project faces: controlling ground water. The conference, held in New York City in November, is a partnership between the UCA and The North American Tunneling Journal. It featured more than 20 technical presentations by professionals from around the world. Each described specific challenges unique to their projects and the solutions employed. More than 175 professionals attended the two-day conference that included an exhibition featuring 17 companies. In addition to the technical program, the Cutting Edge conference featured a site visit to the $10.4 billion East Side Access Project. Sponsors of the meeting included Soil Freeze, Traylor Bros., Kiewit, Skanska, Moretrench, Aecom and The Robbins Co.

"2016 marks an amazing accomplishment in the mining industry as Hecla Mining Company celebrates its 125th anniversary. Mining Engineering sat down with Hecla chief executive officer Phil Baker to gain insight on this tremendous longevity and to get his view on how the company demonstrated persistence, gained perspective and created position for its future.How did Hecla Mining Co. get its start?In 1891, a group of investors formed Hecla Mining Co. to acquire and trade mining claims. No one envisioned at the time that the Coeur d’Alene Mining District would become known as the “Silver Valley” and produce more than 37.2 kt (1.2 billion oz) of silver. Or that Hecla would become the United States’ largest primary silver producer and the last of the pioneer mining companies from this illustrious mining district.What is the origin of the name “Hecla?”That will forever be a bit of a mystery. Some say it was named for an Icelandic volcano. Or that it was the foresight of James Toner, who staked the original claim, that it would go so deep, fulfilling the medieval folklore definition of the word ‘Hekla’ which is the ‘Gateway to Hell.’ I kind of like that one as the Lucky Friday Mine is almost 3.2 km (2 miles) below the surface. It has operated for more than 70 years and has all the markings of being one of only a handful of mines that will exceed 100 years of operations.Is there a common theme, or thread throughout the 125 years?Yes, Hecla has always had a solid core of high-quality and long-lived assets, positioning the company to survive times of low prices and reap the rewards when cycles turn and higher prices are realized. But it is not just assets. Hecla has extraordinary expertise that gives us the perspective of how to acquire, explore, develop and operate mining assets, particularly underground. Hecla is one of the few companies to acquire assets in the last five to six years and we did not have to write down the value of them, which means we paid a fair price for the asset despite the metals price reductions that followed the purchase. And our culture has given us the persistence to develop mines and complete projects despite lower prices. For almost a decade we have been advancing the #4 Shaft project at the Lucky Friday. We will complete this project this year, positioning the mine to operate for decades. We have elected to spend capital on San Sebastian (Durango, Mexico) and Casa Berardi (La Sarre, Quebec, Canada) to start a new mine and increase production, respectively. And we are spending money on new techniques, new equipment and innovations to our operations. Our insistence is on developing for the future, and our persistence in doing so is in contrast to many companies that are reducing investment and shrinking production"

Jun 1, 2016

By Philip-Mark Spanidis, Christos Roumpos, Francis Pavloudakis

This paper presents a quantitative research (QNR) methodology, collaboratively conducted with a group of mining experts, to assess the knowledge management (KM) landscape in the Greek mining industry. The methodology involves a framework for structured data collection using questionnaires, statistical analysis, discussion on critical methodological aspects, and interpretation of results focusing on key aspects of interest for designers of KM systems. The paper concludes by summarizing the methodology’s outcome and proposing perspectives for further research.

Aug 1, 2024

By David Green

Outokumpu introduced Flash Flotation in the early l 980's for the recovery of floatable material from grinding circuits. The devel­opment came as a result of surveys made in the company's own concentrators, initiated to find ways of improving recoveries with increasingly complex and lower grade feeds. Hydrocyclones make a separation on the basis of particle size and mass, resulting in the "over collection" of fine grained, high S.G. particles. Minerals such as pentlandite, galena and gold are easily caught in the circulating load around the mill until they are overground and rendered more difficult to collect. In the case of downstream flotation this may simply involve a reduction in the rate coefficient due to the finer sizes or, as is commonly experienced in gold operations, the mate­rial is slimed and coated onto non-floating gangue and is eventual­ly lost to tails. At the time of Outokumpu's plant surveys, it was a relatively common practice, particularly in Pb operations, to have a flotation section fed directly from the secondaiy mill discharge. This is called unit flotation and certainly provided an improvement but very high wear rates and frequent sanding out of the cells was com­mon. The development of the SkimAir® Flash Flotation cell by Outokumpu provided a subtle but quite dramatic departure from the unit cell approach. Now it was possible to feed cyclone under­flow into a special purpose-built flotation machine and recover a final grade concentrate from the grinding circuit (see Figure I). The cell was designed to handle the coarse circulating load, per­mitting the coarsest fractions to be short-circuited back to the mill and retain the floatable sizes for a nominal time of 30 seconds to several minutes. The short retention time was possible largely due to the relatively clean content of the cyclone underflow stream. This means that the fine gangue is not present except in small quan­tities and most of the fine to mid-size paiiicles are higher S.G. and more likely represent valuable mineral. This condition provides for very high rate coefficients. Typical operations would recover 40-60% of the valuable mineral in this fashion and improvements in recovery were typically from 2-5% for sulphide minerals to the 5-20% range for gold and silver.

Jan 1, 1998

By S. K. Kawatra

Three International Symposiums-the 14th International Minerals Processing Congress, the 17th APCOM, and the SME-AIME Fall Meeting special symposium on Design and Installation of Com¬minution Circuits-held in 1982 contained a number of papers relating to the fundamentals of processing coals and metallic minerals. In addition, many papers were published elsewhere in the literature. Some of the more significant papers are discussed below. Flavel reviewed key performance parameters, such as energy-size relationships for crushing processes and presented a fundamental basis for optimizing crusher productivity and possibly for development of improved crushers. Pascoe et al. reviewed the use of such control schemes. Magerowski and Karra presented a paper on a computer program developed to and in the design of crushing circuits. The problems associated with the scale-up to very large ball mills were discussed by Arbiter and Harris, and by Kavetsky and Whiten. Both groups concluded that limits on mill productivity arise due to problems associated with particle breakage and transport through mills. Austin et al., and Herbst and Rajamini reviewed the state of the art for the use of grinding circuit simulators for design and control. The effects of mill operating and design parameters on selection and breakage were discussed in detail. The philosophy of control and instrumentation for grinding circuits was extensively discussed by Kawatra and Seitz. Their paper presented for the first time an analytical basis for the inferential particle sizing technique and provided a fundamental basis for the implementation of this technique. The effect of reagent addition variation's on the performance of coal flotation circuits was reported by Kawatra and Waters, and on sulfide flotation circuits by Klimpel et al. Both groups analyzed circuit behavior from a kinetic point of view and reached similar conclusions on the effects of collectors and frothers. The computer simulation of an iron ore flotation circuit also was reported by Kawatra et al. In this work particular attention was paid to the phenomena of froth overloading. Klimpel et al., Kawatra and Waters, and Niemi et al. discussed the philosophy of control by flotation circuits from a fundamental kinetic point of view. An extension of the use of high intensity magnetic separators to the beneficiation of specular hematite was discussed by Thayer and Linkson.

Jan 5, 1983

"The use of gas, temperature and airflow sensors to monitor mine ventilation atmospheres offers numerous safety and operational bene-fits to mines. Detection of dangerous gases, changes to airflows and temperature variations are some of the obvious benefits of using at-mospheric sensors in mines. The prediction of atmospheric conditions between sensors however is potentially difficult, and gaining full cover-age a mine will require either many sensors or a method of extrapolat-ing and predicting atmospheric conditions between and downstream from a sparser array of sensors. A dynamic simulation computer model technique is demonstrated to predict time based changes to atmosphere conditions downstream from sensors. The dynamic model predicts the time based changes to conditions throughout a mine model by simulating the mixing, addition and dilution of gases and airflow changes in incremental time steps. The advantages of this method is that a real time overview of full mine atmospheric conditions can be viewed at all locations, even areas without sensors. In addition, with careful planning and consideration of potential contaminant sources, the number of required sensors can be minimized. INTRODUCTIONMine Ventilation Sensors The importance of continuously monitoring mine atmospheres has been recognised since mining underground began and the effects of harmful gases on human health were discovered. Through the 19th and early 20th centuries, coal miners would take canaries underground with them as an early detection system against life threatening gases such as carbon dioxide, carbon monoxide and methane (Figure 1). The canary, normally a very songful bird, would become agitated, stop singing and eventually die if not removed from these gases, signalling the miners to quickly exit the mine (Davalle, 2011)."

Jan 1, 2016

By Haydn H. Murray, Fred G. Heivilin

The "Hormite Group" was proposed for palygorskite (attapulgite) and sepiolite for their complex magnesium silicate composition and elongate crystals (Martin-Vivaldi and Robertson, 1971). These minerals occur in close association with each other and more complex structural variations may exist (Bailey, 1972). In 1862 Savchenkov used the name palygorskite to describe a mineral from the Palygorsk locality (Hay, 1975), near the Ural Mountains. Ovecharenko and Kukovsky (1984) mention that when mountain leather deposits were prospected in the Palygorsk Division mine it was assumed this unusual mineral was a variety of asbestos. Early mineralogists used the terms "mountain cork" or "mountain leather" when referring to palygorskite. Robertson (1986) mentions that it appears palygorskite was known since Theophrastus&apos; time, ca. 314 BC. J. de Lapparent used "attapulgite" for clays from Attapulgus, GA, and Mormoiron, France, because he thought them different from palygorskite, but the two types were proved to be the same (Bailey et al., 1971). The name attapulgite is still used for the Florida and Georgia deposits when the crystal length to diameter ratio does not exceed 10:1(Merkl, 1989). Georgia palygorskite clays are of much shorter length compared to classic palygorskite. In 1847 Glocker first used the name sepiolite which was called "Meerschaum" by Werner (1788) and Hauy (1801) namedit "Ecume de Mer." Brochant (1802) described low density and white magnesium silicates adding the name Talcum Plasticum and Ecume de Mer. In the Meigs-Attapulgus-Quincy district palygorskite (attapulgite) commonly occurs in two distinct forms referred to as short length palygorskite (Meigs Member) and long length palygorskite (Dogtown Member) (Merkl, 1989). Long length palygorskite crystals (> 10 pm) are rarely observed in the Meigs and Dogtown Members, but when present are in association with dolomite crystals. The short length form is usually less than 2 pm in length and has a low magnesium content whereas the long length form has a high magnesium content and a length greater than 2 pm. The distinctions in morphology are not only important because of the relationship to the origin of the deposits, but also in relation to activity in causing membranolytic activity related to data on palygorskite samples from 9 locations ranging from relatively inert to active in work reported by Nolan et al. (1989). The > 10 pm lengths amounted to only 51 of 17,401 fibers sized. The shortest lengths (< 0.5 pm) were relatively inert. This study pointed out that surface activity, morphology, and chemical differences may be distinctly different within the definition of palygorskite, or for that matter for any individual mineral so that health and other properties must be measured because the name alone does not necessarily indicate uniformity. Palygorskite (attapulgite) fuller&apos;s earth was first sold for drilling mud in 1941. The market for this use expanded slowly and has maintained a level of 7 to 10% of the total US production during the last few years. Most of the fuller&apos;s earth sold for drilling mud comes from the southern part of the Meigs-Attapulgus-Quincy district of Georgia and Florida. Palygorskite clays produced in this area are superior to most other fuller&apos;s earth for mud used in drilling salt formations, but because of high water loss, they are inferior to bentonite where the rocks drilled contain no saltwater. According to Oulton (1965), more than 90 different grades of fuller&apos;s earth are produced. Some of these grades are used for pharmaceuticals designed to absorb toxins, bacteria, and alkaloids; for treatment of dysentery; for purifying water and dry cleaning fluids, dry cleaning powders and granules; for the manufacture of NCR (no carbon required) multiple copy paper; for the manufacture of wallpaper; and as extenders or fillers for plastic, paint, and putty. Fuller&apos;s earth mined near Ellenton, FL, was used for making lightweight aggregates for the construction of concrete barges during World War I1 (Calver, 1957). Still other uses of fuller&apos;s earth and its suitability for uses in new products are outlined by Haden, Jr., and Schwint (1967), Haden, Jr., (1972), and Haas (1970). One special use of fuller&apos;s earth is as a carrier of platinum catalysts that are made in the United Kingdom from sepiolite clays mined in Spain. Other uses of sepiolite fuller&apos;s earth (Chambers, 1959) are similar to those of the palygorskite (attapulgite) type mined in the United States.

Jan 1, 1994

By W. M. Houston

Pebble phosphate deposits of Florida have been established from south of Hardee County to north of the Georgia state line or a distance of over 200 miles and mining has been carried out in an ever-increasing scale since 1885. Deposits that can be mined on the economical standards of today, however, are confined in a roughly circular area about 30 miles in diameter, generally within Polk and Hillsborough Counties, in the central part of the State. The phosphate deposit, which is in the Bone Valley formation of the Pliocene age, is covered by Pleistocene terrace deposits of unconsolidated quartz sand. The matrix is composed of approximately 1/3 clay slimes, 1/3 silica sand and 1/3 phosphate pebbles. The phosphate occurs usually as rounded pebbles, ranging in size from about 1 inch down to approximately 400 mesh. The phosphate may vary from well- defined hard pebbles to soft weathered rock. The color and character of the phosphate will range from smooth buff pebbles to black and white or "salt and pepper rock." The screen size distribution varies considerably between locations and the recoverable phosphate varies may extend down to the 400 mesh size in some areas, while in others the phosphate may cut off sharply at about 35 mesh.

Jan 1, 1962

By R. M. Bricka, T. J. Olin

Heavy metals contamination is an environmental problem at Army installations engaged in firearms training and munitions production. At these facilities, weathering and corrosion of expended munitions and leaching from wastewater lagoons, landfills and burn pits have resulted in heavy metals contamination of the soil. The principal metals encountered in firing-range soils are lead, copper and zinc. Cadmium and other metals, such as antimony, that are often incorporated in the munitions are sometimes seen in lesser concentrations. Mercury is associated with various propellants and, while present in much smaller concentrations, is of concern because of its acute toxicity. Chromium is primarily associated with plating operations. The transport of metals into groundwater has been confirmed at some locations, which has required treatment of the soil and groundwater at these sites. Certain treatment processes for contaminated soil produce metals-laden extracts, which also require treatment before reuse or disposal. Ion exchange is generally quite effective for removing metals from aqueous streams. However, resins are expensive and must be regenerated, and activated carbon is generally less effective for most metals and also requires regeneration. Therefore, alternative effective and economic sorbents are needed. Twelve sorbents were screened in initial batch testing. These included activated carbon, bark, chitosan, crown ether, corn cob, xanthate, clay (kaolinite and montmorillonite), peat moss, seaweed and reagent-grade zeolite (aluminosilicate, Sigma Product No. Z3125). Of these, zeolite demonstrated the highest capacity for Pb, Cr and Cd. For this reason, zeolite was selected for further testing in batch, kinetic and column studies. Materials and methods Zeolite. The zeolite used in the second-phase batch and column studies was obtained from a natural deposit of clinoptilolite-rich rock located in South Dakota (Rocky Ford SDH) (Desborough, 1996). Large blocks of the material were crushed and sieved into the following three particle size ranges: 0.5 to 1.0, 1.0 to 4.0 and 2.0 to 4.7 mm. This material demonstrates high structural stability in acidic solutions (pH 2.5) (Desborough 1996) and has a measured surface area of 30 m2/g. The measured total cation-exchange capacity (TCEC) was approximately 10 meq/100 g. This is well below what has been indicated for commercially available zeolite, which has been reported to be about 180 to 220 meq/ 100 g. The TCEC test was repeated (Method 9081, SW 846) using sodium acetate. The test resulted in a TCEC of 54.5 meq/100 g. The difference between values obtained for this material and published values for zeolites may be attributable to the greater heterogeneity in the material used in this study, compared to commercially available materials, or to the effect of the relatively large particle sizes utilized. Batch studies. Seven batch studies were conducted using synthetic metal solutions and soil extracts (Table 1). Extract composition: The P-extract was prepared by sequential surfactant extraction of organics from a burnpit soil followed by acid extraction of metals. The pH of this solution is approximately 1.1. A number of metals and organic compounds were present in the soil. Analysis of the extract was restricted to Pb, Zn and Cu concentrations for this study. The FBH extract was produced from a firing-range soil that was oxidized with a 0.01 M CaO solution and then extracted with 0.1 M acetic acid. This was filtered through 0.5-µm Whatman No. 5 filter paper and stored at room temperature. The pH of the FBH extract was approximately 4.5. pH Control: Calcium carbonate (CaCO3) may be present in the zeolite horizon or bed. Calcium ion (Ca") is released from the exchange sites when in contact with solutions containing ions for which it is more selective, such as lead. This results in a rise in solution pH over time. Acid washing removes most of the carbonates, eliminating the need for a buffer. Batch studies were conducted using both acid-washed zeolite (AW) and unwashed zeolite (UW) for performance comparison. The UW zeolite was rinsed with distilled deionized (DDI) water to remove fine soil particles. Both materials were dried at 105°C (220°F) overnight, so that the dry mass could be determined. Column studies. Ten column studies were conducted. Because it was expected to have the best hydraulic properties, the largest particle

Jan 1, 1999

By J. R. Sturgul

Computer simulation models of working mines can be of great benefit to the mining engineer. Such models have been used in mine planning (how many trucks/shovels to have, when to add new equipment, etc.), mine design (should a dispatch system be purchased, where is the optimum location for mine equipment, etc.). A simulation model can be thought of as a dynamic representation of the mine and can easily be changed to reflect a wide variety of different situation. Often a major problem for the engineer is how to present the results of the simulation study. Using PROOF(R) animation it is now possible to interface the results of the simulation and project them onto a computer screen in cartoon fashion.

Jan 1, 1993

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 &apos;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 &apos;start-up&apos; 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

"In this review, the sintering process is first described and analyzed to identify the key factors affecting the structure and quality of iron ore sinter. Under fixed sintering conditions, the granulation efficiency, nucleus stability and primary sinter melt volume and properties of the iron ore are considered to have significant influence on the sinter structure and quality. Discussion is then focused on the effects of the chemical, physical and mineralogical characteristics of the iron ore on these key factors, and the consequences for sinter quality.IntroductionSintering is the most economic and widely used agglomeration process to prepare iron ore fines for blastfurnace use. In modern blast furnaces, particularly in East Asia, iron ore sinter typically constitutes more than 70% of blast-furnace ferrous feed. High sinter quality and efficient sinter production are thus required both to withstand extreme operating conditions and maintain the productivity of modern blast furnaces. At the same time, due to the increasing depletion of high-grade iron ore resources from large deposits, iron ore producers are moving toward developing an increasing number of deposits of much smaller sizes to maintain current production levels. It is therefore imperative for iron ore producers to develop long-term strategies to maintain the quality and consistency of the iron ore fines they are producing while exploiting these smaller, complex deposits. The iron grade, gangue minerals present, overall mineralogy as well as physical characteristics of iron ore fines being mined from these different deposits vary considerably. Overall, ores now have highly complex mineralogy, are becoming more porous and friable, and contain increasing amounts of gangue minerals, which will have important implications for the sintering process and sinter quality.Blending is one of the most effective options being explored by iron ore producers to deal with multiple small complex deposits. This allows the iron ore producers to reduce the number of products and take full advantage of the unique characteristics of each individual deposit to achieve long-term product consistency while increasing production. Blending is also a common practice in sinter plants as it is almost impossible to produce from one single ore a high-performance sinter mixture with the required processing characteristics. A typical sinter mixture used in Japanese steel mills (Nakano and Okazaki, 2007; Matsumura et al., 2013) often contains five or more ores to achieve balanced sinter quality and productivity with relatively low cost of raw materials."

Jan 1, 2015

By V. B. Achanti, A. W. Khair

Introduction Laboratory studies were carried out at West Virginia University to investigate the rock-fragmentation mechanism of continuous miners using an automated rotary cutting simulator. The primary factors influencing the fragmentation process were found to be bit spacing, bit geometry, depth of cut and cutting-drum rotational speed. This paper presents a discussion of the effects of these parameters in achieving optimum energy consumption and minimizing dust generation during rock fragmentation. The removal of rock ridges/walls between adjacent grooves is analyzed with three hits mounted simultaneously on the cutting head, while the bit tip angle was varied from 600 to 900. Bit spacing was varied from 25.4 to 50.8 mm (1 to 2 in.) while the cutting process was assessed for varying cutting depths. Respirable dusts generated during the course of the experiments were analyzed utilizing cascade impactors. Assessment of these parameters has led to a better understanding of the cutting mechanism of continuous miners in terms of energy consumption and dust generation. A review of the literature revealed that a considerable amount of research has been carried out on rockcutting processes. Many authors agree that the mechanical cutting efficiencies of mining machines (e.g., continuous miners, shearers and road headers) are affected by a host of parameters. Some of these parameters are machine controlled, some are operator controlled, while others are uncontrollable. Efforts were focused on understanding the influence of parameters such as bit spacing, cutting depth, attack angle, bit type, drum speed, bit geometry (i.e., tip size, shape and tip angle) and rock type on the cutting process efficiency in terms of specific energy consumption and respirable dust generation (Strehig?? et al., 1975, Hanson et al., 1979, Khair et al., 1989). Roepke et al. (1976) in an attempt to study the dust and energy generated during coal cutting using point attack bits found that the dust and the specific energy consumed both decrease as the depth of cut increases. The four fundamental stages of dust generation luring rock fragmentation are identified by Zipf and Bieniawski (1989). Coal breakage by various types of wedges was assessed by Evans and Pomeroy (1966) in an extensive experimental study on British coals. Yet the industry today requires further attention and guidance to optimize the energy consumption and dust generation during the rockbreakage process. This paper attempts to give a better understanding of the influence of some of the parameters listed above and focuses on further improvement in the rock-cutting process. The specific energy consumed for different types of bits used and the respirable dust generated are analyzed in the context of the variation of a few other parameters. Laboratory investigation The experiments were performed in the Rock Mechanics Laboratories located at West Virginia University. A rock-cutting simulator designed and fabricated by Khair (1984) was utilized for this purpose. The details of this machine are available in the literature (see Khair 1984). For this study, a series of preliminary experiments was carried out to determine the optimum cutting frame advance speed. This was intended to facilitate a maximum cut depth of 31.75 mm (1.25 in.) at an advance rate of 0.525 mm/s (0.0207 ips) for all types of bits being used and various bit spacings being considered. To look into the cutting-process efficiency of a continuous miner in the laboratory, several parameters of influence are being considered. Besides the bit-geometry parameters, machine- and operator-controlled parameters, such as spacing of bits on the cutting head, the cutting head rotational speed and the total cutting depth during an experiment, are varied. At the time this paper was written, only part of the completed experiments were ana¬lyzed, and a number of experiments were still being carried out following an orthogonal fractional factorial experimental plan to assess the effect of all of the above¬mentioned parameters on the cutting efficiency in terms of energy consumption and dust generation. Three different types of tip angles, namely, 60°, 75° and 90°, and two different tip sizes, namely, diameters of 7.94 and 24.61 mm (0.313 and 0.969 in.), were used. At

Jan 1, 1999