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By James D. Hervol, Josepph W. III Leonard, Barbara J. Arnold

INTRODUCTION Dry concentration or pneumatic cleaning uses air currents as the separating medium. Pneumatic cleaning in the United States today is limited to three or four operations. In fact, the quantity of bituminous coal cleaned pneumatically has declined from its peak in 1965 of about 25.4 million tons (23.1 Mt) to its current level (1988) of less than 4 million tons (3.64 Mt) annually, as shown in Fig. 7-9 1. In 1940, pneumatic methods accounted for 15% of the bituminous coal and lignite cleaned in the United States as given in Table 7-32. Currently, pneumatic methods account for less than 2%. The largest air plant was built in Pennsylvania about 1968 with 14 units processing 1400 tph (1280 t/h) of 3/, in. X 0 (19 mm X 0) raw feed and two additional units retreating middlings. This same plant is one of only three or four plants still operating air-cleaning devices in the US. A wet plant was added in 1985 to increase the efficiency of cleaning the fines. The air plant is still used as a scalping operation for the coarse material. Five additional operations are either no longer in production or have discontinued the use of air tables within the last year. Wet cleaning methods are now being employed at the facilities that have discontinued air table operation. he advantage of dry cleaning techniques is the low moisture content of the clean coal product. However, to decrease respirable dust, mining operations must use water sprays on modem mining equipment. This water adds to the moisture content of the run-of-mine coal and decreases the effectiveness of dry cleaning processes that rely on low surface moistures in the feed to allow for good separations. Wet cleaning methods are preferred, especially since they also allow for much sharper separations. The advantage of dry cleaning techniques is the low moisture content of the clean coal product. However, to decrease respirable dust, mining operations must use water sprays on modem mining equipment. This water adds to the moisture content of the run-of-mine coal and decreases the effectiveness of dry cleaning processes that rely on low surface moistures in the feed to allow for good separations. Wet cleaning methods are preferred, especially since they also allow for much sharper separations. Despite dry concentration's drawbacks, research has continued into dry concentrating methods, including both electrostatic and magnetic separation techniques. The real application for future dry cleaning processes is probably at the power plant between the pulverizer and the boiler feed where particles are already dry and ground to fine sizes (- 200 mesh [ -0.075 mm]), increasing the effectiveness of the separation and also, possibly, increasing the liberation of clean coal from finely disseminated impurities. From a utility's viewpoint, pneumatic cleaning offers the best delivered Btu cost, based on the premise that a percent of moisture is just as detrimental as a percent of ash. Additionally, several other advantages are inherent in dry concentration techniques. The main advantage is that dry concentration re-

Jan 1, 1991

By J. Burrows

The Chamber of Mines of South Africa is an association of major mining corporations, gold mines, coal mines, diamond mines and other metalliferous mines. It looks after those interests of its members which can best be handled on a co-operative basis. The Chamber's activities include the engagement of labour; the negotiations of all labour matters, including minimum wage rates, with recognised trade unions and associations; the carrying out of advanced technological research; representing the industry in dealings with the government and the administration of miners' training colleges and rescue training stations. Subsidiary companies of the Chamber handle matters such as the refining of gold and other metals; the processing and marketing of uranium and the marketing of Krugerrands. The thirty-two gold mines members of the Chamber of Mines of South Africa produce approximately 694 tons of gold per annum. To achieve this they employ 430000 persons and break about 8400 kilotons of rock per month. The working revenue is approximately 4 225 million US dollars per annum and the annual working profit is approximately 2 050 million US dollars. The electric power consumed by the gold mining industry is 1650 MW, about 20% of the total power produced in South Africa, and the annual costs of this power is approximately 200 million US dollars per annum. It is estimated that about 25% of this power is used for ventilation and refrigeration appliances. Thus the annual power costs for ventilation and refrigeration is approximately 50 million US dollars. The gold bearing reef is a narrow tabular deposit situated at varying depths throughout the gold fields. The weighted mean rock breaking depths and, virgin rock temperatures range from 1 700 - 9 000 ft (520 - 2 740 m) and 68 - 113°F (20 - 45°C) respectively. The maximum virgin rock temperature at which mining currently takes place is 130°F (54.5 C). Obviously at such depths and temperatures and with such a large labour force the provision of an acceptable environment is critical and to achieve this 59 000 000 cfm (28 000 m3/s) of air is downcasted during the main shift and 122500 tons (430 MW) of refrigeration are provided. Legislation in South Africa requires that on metalliferous mines when more than 1000 persons are employed underground during any one shift the manager shall appoint one or more ventilation officers. The legal responsibilities of the ventilation officers are concerned mainly with the measurement and control of dust. In practice, this is not the main part of their work and ventilation personnel perform many other duties such as measuring and controlling the distribution of air, checking the performance of fans and cooling plants, planning air and refrigeration requirements and, more recently, measuring and controlling noise and illumination levels. Ventilation personnel are specialists who provide management with advice on all matters relating to the mine environment. At the present time there are 506 persons engaged as ventilation personnel in gold mines and a further 1744 persons are employed in gold mine ventilation departments in a non-technical capacity, for example as construction workers, filter cleaners, fire patrollers, etc.

Jan 1, 1980

By M. Murillo, J. Botín

"Mining has always played a key role in the improvement of the quality of life, but also, it has been a source of environmental and social conflict. Despite the great progress in social responsibility and sustainability experienced in recent decades, the public image of mining remains low and mining projects and operations must face growing difficulties and demands, which results in increasing capital and operating costs. In practice, the objective of a mining company is to create value for its shareholders and therefore, the best approach to Sustainability would be to foster those drivers aligning sustainability and shareholders value, this is, the “Business Case for Sustainability” (BCS). Only under this assumption, convergence of corporate strategy and sustainability may be assured. This paper discusses partial results of a Research Project being carried out at the Catholic University of Chile, aiming to develop a social and economic model for the effective integration of the values of sustainable development in the strategy of mining companies and their stakeholders. INTRODUCTION Mining is fundamental for the wellbeing of our everyday life. But as important as these benefits are for humanity, the mining Industry has had negative environmental and social impacts. For these reasons, a closer relationship between “Sustainable Development and Mining” is more necessary than ever. The Brundtland definition of Sustainable development does not have a clear relationship with mining, but it provides the context. Under a strict interpretation of the Bruntlandt definition, stating that “Sustainable development is development that meets the needs of the present without compromising the ability of future generations to meet their own needs” [10], mining would not be sustainable because the minerals are non-renewable resources. However, this research focuses on how mining can contribute to the sustainable development in the territories where the projects take place. In other words, this research follows the publication titled “Berlin II Guidelines for Mining and Sustainable Development” [9], which stated that “A mining project that is developed, operated and closed in an environmentally and socially acceptable manner could contribute to sustainable development. Critical to this goal is ensuring that benefits of the project are employed to develop the region in a way that will survive long after the mine is closed”. The potential conflicts between mining stakeholders arises from the interpretation of the concept “developing, operating and closing a project in an environmentally and socially acceptable manner”, and the impact on the economic objectives of mining companies. Nevertheless, the ability of mining companies to control new demands and expectations of stakeholders has threatened the implementation and continuity of their projects and operations. To deal with these issues this research analyses the “Business Case for Sustainability” (BCS). It refers to the process leading to a business and social-economic environment whereby shareholders value may only be optimum under a business strategy which fully integrates the values of sustainability. In the model proposed, the Business Case will start when the benefits are higher than the costs. Benefits not only associated with efficient use of resources, also benefits associated with avoiding difficulties to continue with the mining projects such as strikes, heavy penalties, loss of investment, and so on."

Jan 1, 2016

By Henry T. Miller

INTRODUCTION The basic purpose of this paper is to suggest procedures and practices to insure that no employee working in a uranium milling operation receives exposure to radioactive, toxic, or other materials or agents that might produce a permanent, deleterious effect upon his physical health and well-being. This program is also designed to insure that each employee can carry out his assigned duties without risk to his health or to that of his fellow workers. The assumptions underlying this program are by necessity, conservative, with emphasis being placed upon prevention rather than correction or reaction to an undesirable situation. The limits specified-either as Threshold Limit Values (TLVs), Maximum Permissible Concentrations (MPCs), or Radiation Protection Guides (RPGs)--represent both standards of best practice and those of law and regulation. The total program is envisioned as a balanced combination of occupational hygiene and radiation monitoring. This includes surveying, air sampling, personnel dosimetry, bioassay, medical surveillance, epidemiology, and training--all backed by a thoroughly tested and evaluated set of emergency procedures. The program, as presented, is keyed to the results of monitoring, surveying, air sampling, medical surveillance, and epidemiology--it being obvious that no problem can result when no hazard can be identified. The lack of an identified hazard, however, does not negate the need for such a program and prudence suggests the need for routine surveillance programs in the best-designed facility. [The Model Mill] The model mill used in this paper is that presented by the U. S. Environmental Protection Agency in its study of the uranium fuel cycle1. The model mill processes 600,000 metric tons of ore per year into 1,140 metric tons of U308 as yellowcake per year using an acid leach process. Other characteristics of the mill such as location, tailings dam design, number of light water reactors supported, and geographical location are not pertinent to this paper. The usual remote western location of these facilities produce some difficulties in delivering the required occupational medical services and is not within the scope of this paper. The operating lifetime of most mills is commonly estimated as from 12 to 15 years in actual practice; however, the U. S. Environmental Protection Agency estimated a 20-year lifetime for its model. The principal steps in the acid leach process include :2 (a) Ore is blended and crushed to pass through a 2.5 cm (one inch) screen. The crushed ore is then wet ground in a rod or ball mill and is transferred as a slurry to leaching tanks. (b) The ore is contacted with sulfuric acid solution and an oxidizing reagent to leach uranium from the ore. The product liquor is pumped to the solvent-extraction circuit while the washed residues (tailings) are sent to the tailings pond or pile. (c) Solvent extraction is used to purify and concentrate the uranium. (d) The uranium is precipitated with ammonia and transferred as a slurry. (e) Thickening and centrifuging are used to separate the uranium concentrate from residual liquids. (f) The concentrate is calcined and pulverized. (g) The concentrate or yellowcake is packaged in 208 liter drums for shipment. This process is illustrated in Figure 1.3

Jan 1, 1981

By George H. K. Schenck

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

Jan 1, 1982

By Leonid Pevzner, Valerie Dmitriev, Yuri Babitchev

Performance of belt conveyors, service life of belts and mechanical elements can be radically improved if the automatic adjustment of the belt speed movement is made in such a way, that the unit load over the conveyor length be constant and close to the rated one under random input freight traffic. Such an adjustment is possible when the system of a controllable thyristor electric drive on the basis of induction squirrel cage motor is used. Computer-aided adjustment of the belt conveyor movement is effected via bicontour digital control of the main drive in order to redistribute loads when the loading of conveyor and contour of the control of tension drive are changed. The system envisages control and condition monitoring of the conveyor electromechanical system, as well as metering of electrical power and the amount of conveyor freight. The results of the simulation of speed control of the conveyor belt are given in the report. Mathematical model of the conveyor electromechanical system describes the movement of the belt with cargo and the tension drive. The input cargo traffic is assumed as stationary ergidic random process with the assigned characteristics. The solution of the problem of the synthesis of a closed-loop control system, the structure and parameters of adjustment, the conditions of asymptotic stability of the control back-up system and the condition of minimization of mean square error of shadowing are also given in the report. Microprocessor controller is used to realize the function of the adjustment unit. Under this control system standard mode of conveyor operation, different modes of start-up, operating mode with decreased speed when transporting people, service and emergency braking of a conveyor and some other modes are realized. The digital model of a conveyor as an instrument for decision making in case of emergency is used in those modes. MASTER TEXT Long-standing practice of belt conveyor operation in mines of Russia has proved their inefficient use with 40-70% capacity and 20-25% availability factor. It is associated with unreasonable power intensity and reduction of service life of mechanical components, and as a result, with the growth of total transportation costs. Low efficiency of conveyor application can be explained by the extremely irregular distribution of conveyor freight traffic. Usually, the width of a conveyor belt is identified in accordance with the possible maximum traffic value, engine power capacity for long-term operation under rated loads. However, actually a conveyor often runs without any load and its average load in most cases is lower than the rated one. The problem can be eliminated by way of the adjustment of the operational mode of a conveyor according to an actual input freight traffic via belt speed control providing for the conveyor regular linear load which is nearly equivalent to the rated one. Experimental research into freight traffic under working conditions of the operating underground and open-cast mines proves the fact that the actual freight traffic can be adequately presented through the model of stationary random function of time, for ex- ample, as a J (t) continuous random Gaussian process that is modulated with a G(t) elementary pulse flow:

Jan 1, 1996

By W. Jacobi, K. Eisfeld

The primary objective of dosimetric modelling for the inhalation of radon daughters is the evaluation of the dose distribution and of the mean dose to sensitive cells or target tissues in the respiratory tract which are assumed to be responsible for the initiation of lung cancer. The results of such dosimetric models are influencing the practice of radiation protection in a twofold manner: Firstly, they enable us to identify appropriate operational quantities for air monitoring and for the surveillance of individual exposures to radon daughters. Secondly, they can be used to derive secondary and operational limits for these radionuclides from the basic dose limits. The dose to target tissues in the lung from inhaled radon daughters depends on various physical and biological parameters: The unattached fraction of daughter atoms, the AMAD of the carrier aerosol for attached atoms, the geometry of the respiratory airways, the breathing conditions, the parameters for the retention and translocation of daughter atoms in the different lung regions, and the location of the target cells or target tissues, respectively. Therefore, dosimetric modelling for radon daughters has to include a sensitivity analysis showing the influence of these parameters. To meet these objectives we have recently described in a comprehensive report an improved dosimetric model for the inhalation of 222Rn, 22ORn and their short lived daughters (Jacobi and Eisfeld, 1980). In the following this model is shortly described and the main results are summarized. DOSIMETRIC MODEL [Model Structure]. For the geometry of the bronchial airways the model A of Weibel (1963) was used. Each bronchial generation i (i =0.1 .... 16) was considered as a separate compartment i, for which the following transfer pathways for radon daughter atoms were taken into account (see fig. 1): (1) the uptake by deposition (deposition probability wi); (2) the translocation by ciliary clearance (rate constant [A]i); (3) the transfer to blood by diffusion through the bronchial epithelium (rate constant [A]b). To distinguish between attached (Index a) and free (index f) daughter atoms each bronchial generation was divided in two subcompartments (i.a and if) with a transfer rate constant [a]s (ia- if) characterizing the desorption or dissolution rate of attached atoms in the mucus fluid. Thus for daughter atoms primarily deposited in attached form the transfer to blood is a twostep process with the rate constants [a]s and [A]b which were assumed to be independent of the generation number i. [Deposition]. The deposition probabilities wi in each bronchial generation were calculated on the basis of the laminar diffusion model of Gormley et al. (1949); however the deposition in the upper airways was corrected for turbulent air flow, taking into account the results of our deposition experiments (Martin et al., 1972). Figure 2 shows the resulting deposition probabilities per unit surface area of the bronchial generations for nose breathing at a rate of 20 1/min. The values for attached daughter atoms refer to an AMAD of 0.2 - 0.3 µm ([O]g =3) for the carrier aerosol; the deposition values for free atoms were derived with a diffusion coefficient of 0.05 cm2/s, assuming a prefiltration of 50% in the naseopharyngeal region.

Jan 1, 1981

By John M. Ackerman

INTRODUCTION In the spring of 1988 the Joint Venture of Freeport-McMoran Resource Partners, Limited Partnership, IMC Fertilizer Inc, and Homestake Mining Co. was awarded the sulphur lease for Main Pass Block 299, lease number OCS-G 9372. This was one of eleven leases awarded to the Joint Venturers. Following the initial discovery wells in December 1988, a program of delineation drilling confirmed proven sulphur reserves of 67 million long tons. Proven, recoverable oil and gas reserves are 39.4 million barrels and 7.2 billion cubic feet, respectively. Development of the sulphur mine was given approval as a fast-track project by the Joint Venturers in 1989. The oil and gas reserves were acquired from the original hydrocarbon lessee in 1990 and given fast- track status immediately. Fast-track for these projects meant design and engineering would not be complete before construction began JOINT VENTURE The Main Pass Joint Venture is one of thirteen joint ventures formed between Freeport and three other companies - IMC Fertilizer, Inc., Felmont Oil Corporation, a wholly owned subsidiary of Homestake Mining Co., and Texasgulf Minerals and Metals Co. Felmont has subsequently changed its name to Homestake Sulphur Co. and Texasgulf assigned their interests in the three leases they participated in to their parent company. Texasgulf did not participate in the Main Pass 299 joint venture. Joint venture negotiations, anticipating a lease sale in mid-1988, began in December of 1987. This Joint Venture was the first of its kind; created to acquire offshore leases, explore, develop and mine sulphur. The Minerals Management Service announced a lease sale date of February 24, 1988 on the first business day of 1988. This greatly accelerated the joint venture negotiations. Of the thirteen leases the Joint Ventures bid on, eleven were awarded. Freeport holds its interest in all the leases through its ownership interest in Freeport-McMoRan Resource Partners, Limited Partnership, and Resource Partners is the operator on all of the leases. The Main Pass joint venture percentage ownership is: 58.33 Freeport-McMoRan Resource Partners 25.00 IMC Fertilizer, Inc. 16.67 Homestake Sulphur Co. 100.00 The Main Pass Joint Venture paid $3,401,757 for the Main Pass lease and will pay a royalty of 12.5 percent of production value to the MMS . LOCATION The Main Pass mine complex is located in the Gulf of Mexico, in U. S. federal waters, about 27 kilometers (17 miles) east of Venice, Louisiana or 140 kilometers south east of New Orleans, Louisiana. It is situated in about 63 meters (210 feet) of water. HISTORY The Minerals Management Service (MMS) of the United States Department of the Interior administers the mineral properties of the United States located in federal waters. Periodically the MMS holds lease sales of blocks under its jurisdiction. Prior to the offshore discoveries of sulphur associated with salt domes leased for hydrocarbons, the MMS leased blocks with all minerals including hydrocarbons awarded to the successful bidder for the block. In 1963 MMS held the first lease sale of blocks specifically for sulphur. This sale was followed by a similar sale in

Jan 1, 1992

By G. Le Bel

Discussion by G. Le Bel The paper proposes three methods to determine optimum lifetime of a mining project 1. The conclusion is that the three methods arrive at the same optimum production rate (by coincidence). It is also stated that: "The stochastic and option pricing models, in addition to providing a more realistic optimum mine life than the traditional NPV (net present value) calculation, showed that the example project has the potential for substantially higher return. By explicitly considering a project's potential risks, the stochastic or option pricing techniques can more accurately describe its probable investment performance." This discussion is intended to demonstrate that: •It is no coincidence that the three methods arrive at the same optimum production rate. • The stochastic and option pricing models do not provide a more realistic optimum mine life than the traditional NPV calculation. •The potential for a substantially higher return is misrepresented. First, a review of investment decision theory in mining is needed. Net present value is the preferred performance measure because of the flexibility offered when comparing various alternatives. However, it presents the following disadvantages: • The discount rate selection is not straightforward. • The NPV obtained represents a deterministic characterization of the variables included in the cash flow calculations, i.e. a single-point estimate. Discount rate selection The opportunity cost of capital represents the rate of return that would be obtained if the capital was instead invested in other opportunities considered to exist. Such opportunities must be bypassed if the available capital is invested in the project studied 2. If these other opportunities carry risk similar to the project under analysis (for example, if they are other mining projects), there is no need to add incremental [ ] risk factors. On the other hand, if these opportunities are represented by the acquisition of government bonds yielding 7%, a risk premium factor must be included. The types of risks associated with a mining venture are shown in Table 1. Together they represent the "business risk" of being a mining company. The individual project risk is important to investors if it differs substantially from the business risk of the company and that of the other investments in mining ventures that we must pass up to invest in this particular project. A risk premium or discount would then be required. This can be illustrated by the following example. Consider that a company has two investment alternatives: one is the Cavender hypothetical project located in Nevada, and the other is the same project located in a civil war zone of central Europe. The second project requires a risk premium factor on the discount rate. Conversely, a project showing highly uncertain returns that are negatively correlated with the returns of the firm's other assets may decrease the company's corporate risk. Deterministic character of NPV calculation The cash flows used in NPV calculations are obtained from a set of variables. These variables are estimated using one parameter obtained from the engineering studies. Each parameter is thus a single value that is often termed a "realistic" or "best estimate" as opposed to "optimistic" or "pessimistic." However, considering that each variable can assume a range of values between the

Jan 1, 1994

By Patrick J. Maessen, Bart H. M. Hendrix

"In the Netherlands, the Dutch Tunnel Standard has been developed for the execution of road tunnel projects. The Dutch Tunnel Standard is based on the Dutch Tunnel Law (WARVW) which sets the minimum safety standard for road tunnels in the Netherlands. Uniformity of tunnels and the uniform perception of a tunnel by its users leads to a higher safety level. Therefore, several items in the Dutch Tunnel Standard are specifically developed to ensure safety in emergency situations through the plane of human factors. Taking human factors into account is important in regular traffic situations, but in an emergency situation, this can be vital. In this paper, the development and contents of the Dutch Tunnel Standard are explained in the first part and in the second part, the systems related to human factors are described. DUTCH TUNNEL LAW AND THE NEED FOR THE DUTCH TUNNEL STANDARD In the 90’s of last century and the early 2000’s, several serious accidents occurred in tunnels in Europe. These accidents involved fire and led to many casualties. After these accidents, the European Union decided to change the regulations for tunnel safety, leading to the EU directive 2004/54/EG. The primary goal of this directive is, that casualties are minimized in case of a calamity in a tunnel (e.g. fire). In the Netherlands, the EU directive led to the development of the tunnel law and a tunnel regulation (in 2006) to prescribe the safety features required for a safe tunnel. The new Dutch tunnel law – cited in the Building Act – was meant to make tunnel design easier – but resulted at first in the opposite effect. Tunnel design in the Netherlands and design standards of Rijkswaterstaat Rijkswaterstaat – the executing organisation of the Dutch Ministry of Infrastructure and the Environment – is the owner of the tunnels in the Dutch national motorway network, thus owner of the majority of the road tunnels in the Netherlands. In the early years of tunnel building in the Dutch national highway network (approximately 1950 – 1970), Rijkswaterstaat was not only owner, but also designer of its own assets. Most Dutch tunnels are submersed tunnels, because these type of tunnels provide a cost-effective way to make river crossings. Another type of tunnel that is used, is a TBM tunnel. Because of the geological conditions in the Netherlands (merely clay/peat and sand), there are no NATM or drill-and-blast tunnels in the Netherlands. Since the last decades, land tunnels (covered roads) are becoming more common to provide for multiple use of surface space and to meet more modern environmental regulations as traffic increases near densely populated areas. Tunnel lengths in the Netherlands vary from 500 m to 1.5 km for river crossings up to 6 km for the Western Scheldt (bored) tunnel. The lengths of land tunnels vary from 500 m up to 3 km. With these increasing lengths, some safety topics and design topics arise, topics that were non-issues in the past, when tunnels were only used as river crossings."

Jan 1, 2016

By Purnima A. Shivdasan

The fluorspar deposits at Okorusu, Namibia are closely related to an alkaline igneous ring complex of late Cretaceous age (125±7 Ma). The fluorite ores consist of relatively fine-grained purple replacement fluorite and subsequent coarser grained purple and green vug-filling fluorite crystals. Homogenization temperature measurements for 38 primary fluid inclusions shows that main stage purple to green fluorite crystals were deposited over a temperature range from 166°C to 144°C, and minor late yellow fluorite was deposited at lower temperatures from 132°C to 128'C. Homogenization and freezing temperatures measured for 18 primary fluid inclusions shows that the fluorite-depositing fluids were less saline than MVT ore fluids and overlap the lower temperature range of fluids that deposited epithermal ore deposits. Mining in two open pits (A and B) at Okorusu has gradually provided exposures which reveal that the fluorite orebodies have been emplaced primarily by the preferential replacement of carbonatites. Evidence for the replacement of carbonatite includes: 1) replacement remnants of carbonatite within the fluorite ores, 2) goethite pseudomorphs after the carbonatite minerals, pyroxene, pyrrhotite, and magnetite that are disseminated in the fluorite ores, 3) partial replacement by fluorite of a titaniferous magnetite rimmed carbonatite dike in the A pit, and 4) partial replacement by fluorite of a large sill-like carbonatite body in the B pit. Electron microprobe analyses of minerals in unreplaced portions of the carbonatites show that their mineralogy consists of titaniferous magnetite, diopside pyroxene, apatite, biotite, and calcite. Magnetite contains 6.97% to 10.46% Ti02, averages 8.44% Ti02, and has an average V203 content of 0.35%. At high magnifications under the SEM, the magnetite shows abundant fine ulvöspinel exsolution lamellae arranged in a cloth texture. Apatite contains moderately high fluorine contents that range 3.21 to 5.0% and average 4.2%, and apatite crystals range in total rare earth oxides from 1.7% to 2.61%. Pyroxenes in the carbonatites are iron-rich diopside, and they contrast to pyroxenes in the fenites that are aegirine-augite. All of the microprobe analyses for mica crystals in the carbonatites give biotite compositions. The fluorite ores also have replaced marble locally in the A pit, especially along the A band in the B pit, and dominantly in the orebody proposed for the new C pit. Carbonatite and marble were preferentially replaced by fluorite because the calcite was readily soluble in the presence of the fluorite-depositing fluids and the dissolution of calcite provided calcium for the formation of fluorite.

Jan 1, 2005

By E. W. Cokayne

GENERAL DESCRIPTION OF THE METHOD The original application of sublevel caving was in ground so weak that it would collapse even in small headings when the support was removed. Heavily tim¬bered drifts were driven across the ore body; the timber was removed at the end of the drift, and the ore caved and slushed out. When dilution became excessive the next set was removed and so on. This method gave high dilution and poor recovery and was slow, but it was the only way to mine this type of ore body at that time. In more recent times the method has been adapted to stronger ground which needs to be drilled and blasted, so it is not a caving method as far as the ore is con¬cerned. However, the method does rely on the walls to cave and so the name sublevel caving has been retained. The method as presently generally applied is shown in Figs. 1, 2, and 3. Sublevels are established at approximately 7.6 to 12.2-m (25 to 40-ft) vertical intervals. They are usually accessed from a ramp system. A haulage drift is driven down the strike of the ore body, but in the waste. Production drifts are turned off the haulage drift and driven across the ore body on about 10.7-m (35-ft) horizontal centers. This divides the ore body into a geometrical pattern. Slot raises are driven at the end of the drifts near the ore-waste contact and the slot is expanded to the shape of the fan drilling. Up holes are drilled to break towards the slot at burdens of 1.2 to 1.8 m (4 to 6 ft). They are drilled past the next sublevel to the production drift im¬mediately above. When a sufficient number of fans has been drilled, production can start by blasting one or two fans at a time towards the slot. Load-haul-dump (LHD) units are used to muck the ore, which rills out into the drift. At first, when only a small area is mined, the walls may not cave, and the operation will be similar to open stoping. Later, sloughing from the walls will fill any voids created, and the ore after blasting will be con¬tained by the caved waste and gravitate downwards as mucking proceeds. This process is repeated in each of the production drifts as the face is retreated towards the haulage drift. When the top level has been retreated back a safe dis

Jan 1, 1982

By Mritunjoy Sengupta

The Green River formations in the Western U.S.A. contains vast amounts of petroleum reserves in oil shale deposits. Exploitation of this petroleum reserve calls for large scale mining of the oil shale. Applicable mining methods include conventional underground mining using roan and pillar, sub-level caving and block caving; underground insitu retorting of oil shale and open pit mining. Roam and pillar mining offers one of the most economic methods of underground mining of oil shale. A 2.4 x l09 litre per year (50,000 barrels per day) of crude oil production would require a daily oil shale production of 100,000 tomes per day from underground operations. For this large production, a highly mechanized room and pillar mining muld utilize large capacity diesel equipment such as 85 tonne trucks and 8m front end loaders. The total air requirements of an oil shale mining operations can. be as high as 28443/s. Ventilation planning for such a large scale underground mining operation presents a formidable task. The major objectives of providing ventilation air to underground workings are to dilute airborne .dust and diesel fumes to harmless levels, and to insure that the mine environment is compatible with the health and safety of mine personnel. Regulation of air velocities and air temperatures is a major factor in mine environmental control. Colorado State law requires that a minimum of 0.047m3/s of fresh air be supplied for each kW of diesel power operating underground. Additionally, 0.047m3/s of fresh air must be supplied for each underground employee, and a minimum air velocity of 0.15m/s must be maintained in all working places. It is desirable to exceed this minimum velocity in order to prevent stratification of ventilation air. Moreover, Colorado law also stipulates the maximum allowable concentration of noxious gases mine air, below: ? C02 - 0.5% by volume, 5,000 ppm. ? CO - 0.005% by volume, 50 ppm. ? Oxides of Nitrogen - 0.0025% by volume, 25ppm. ? Nitrogen Dioxide 0.005% by volume, 5 ppm. This paper examines ventilation requirements for the underground mine, with respect to air quality requirements and distribution of air in mine workings in light of the above regulations. Analyses were made for sane typical mining plans to determine airflow resistance or pressure drops and fan horsepower requirements. In addition, the possible occurrence of high concentrations of toxic gases, was examined. Expected temperature levels in mine workings were approximated.

Jan 1, 1980

By Dale R. Ralston, Roy E. Williams, Gerry V. Winter, George L. Bloomsburg

GENERAL STATEMENT Data frequently can be collected by several dif¬ferent methods. The method required for the nec¬essary precision for inflow prediction to mines should be selected and other methods avoided. Pre¬cision monitoring or measuring techniques fre¬quently require more time, expertise, and expense in their implementation that may not be justifiable. The temporal distribution of the data needs also may dictate the method used for data collection. Continuous recording devices may be required for time sensitive data whereas point measurements of other parameters may be adequate. Data needs will vary as the hydrologic investigation proceeds. In¬itial data needs may require only a qualitative col¬lection of data such as the areal distribution of precipitation. Subsequent data needs may require the measurement of ground water levels to the near¬est hundredth of a foot. Data requirements can be difficult to fulfill for ground water systems. This is particularly true for anisotropic, heterogeneous hydrostratigraphic units since the hydrogeologic parameters vary both ver¬tically and areally. Techniques exist for determining the storage characteristics of both unconfined and confined hydrostratigraphic units but these pro¬cedures require the installation of multiple wells in the same unit which may be impractical. Specific tests are discussed subsequently that permit the determination of vertical hydraulic conductivity. Data collected from field procedures or tests are generally preferred to laboratory derived data. This is especially true for hydrogeologic parameters such as hydraulic conductivity. SURFACE WATER SYSTEMS Precipitation Precipitation data can be collected from sources that maintain a historical record. The US Depart¬ment of Commerce, National Oceanic and Atmos¬pheric Administration (NOAA), maintains weather observation stations across the country. Such a sta¬tion frequently can be found near a proposed mine site but the type of weather data collected varies from station to station. Fortunately, precipitation data are collected ubiquitously. The data collected by NOAA can be purchased for the area of interest in the form of monthly and annual tabulations. Private weather stations constitute a second source of precipitation data. Existing data usually are analyzed by one of three methods. The station average method merely sums the depth of precipitation at each station and divides the sum by the number of stations as fol¬lows: [ ] where Pavg is the average precipitation (L), P, is the depth of precipitation at station i (L), and N is the total number of stations. This method is applicable where there are a relatively large number of stations uniformly distributed across the area (Hjelnfelt and Cassidy, 1975, pp. 71-72). The second method, called the Thiessen poly¬gon method, weights the precipitation at each sta¬tion in direct proportion to the area it represents without consideration of topography. The stations are plotted on a map and perpendicular bisectors are drawn on the lines connecting the stations. The polygons formed by these bisectors constitute the area that the precipitation for that station repre¬sents. The overall precipitation can be calculated with the following equation:

Jan 1, 1986

By C. P. Newman, T. Gray, G. Beale, T. Cluff

Predictive geochemical and limnologic modeling of pit lakes is an important aspect of modern permitting. One of the key assumptions of many predictive pit-lake models is that highwalls and pit backfill become unreactive once submerged by the filling pit lake. Existing pit lakes provide useful data to test this assumption. The Sleeper pit lake (northwestern Nevada) is near hydrologic equilibrium and contains good-quality water that generally meets regulatory requirements for pit lakes. Despite the overall stable geochemical composition, seasonal trends in the hypolimnion indicate the generation of solutes and fluctuations in pH. This work applies mass balance, analysis of subaqueous pyrite oxidation, and trends in solute concentrations to evaluate the potential causes of long-term solute generation in the hypolimnion of the Sleeper pit lake. Results will aid in refining the effect that inundation of highwalls and waste rock has on ultimate pit lake chemistry and long-term closure planning. In the case of the Sleeper pit lake, ongoing solute loading does not substantially change the closure plan at this time as influent groundwater contains abundant alkalinity. INTRODUCTION As mine pit lakes become more widespread, regulatory agencies are increasingly requiring advanced predictive modeling of pit-lake geochemistry during permitting. The end goal of all those involved in these permitting actions (i.e., the mining company, the regulatory agencies, and the public) is to have pit lakes become geochemically and hydrologically stable for the long term, which may then allow mining permits to be relinquished. The modeling endeavors therefore are important to inform long-term closure of pit lakes. The conceptual models and numerical methods used to predict pit-lake water quality are complex and require substantial amounts of data and a variety of assumptions (Castendyk et al., 2015). One of the common modeling assumptions is that solute generation via acid mine drainage is eliminated (or at least decreased) by the filling of the pit lake and inundation of highwalls and any in-pit waste rock. However, few studies have reported on observations that support or refute this assumption. The water-quality record of the Sleeper pit lake is an ideal dataset with which to test these assumptions. The former Sleeper mine is located in northwestern Nevada (Fig. 1), and has been inactive (in terms of ore extraction) since 1996. The mine extracted Au and Ag from the famous high-grade Sleeper vein and associated widespread stockwork mineralization (Nash et al., 1995). Water quality in the pit lake has been monitored since the cessation of dewatering in 1996. The current pit lake actually occupies three separate sub-pits, the Sleeper, Wood, and Office pits (Fig. 1). The Sleeper and Wood pits undergo regular water-column monitoring using a conductivity-temperature-depth (CTD) probe, and sampling for dissolved inorganic constituents at several depths (~10, ~130, and ~330 ft. below lake surface). Groundwater quality is also monitored using a series of wells (Fig. 1).

Jan 1, 2019

By Ruifeng Li, Yanghui Ren, Hua Gao

"Since entering the new century, China’s coal industry has witnessed the rapid development, and provided a reliable energy supply for the national economic and social development. But after the “golden decade” of rapid development, the coal industry now is facing overcapacity, low industrial concentration and production efficiency, high pressure of production safety and environmental protection and other prominent issues. Based on defining the connotation of highly-efficient recovery of coal resources, this paper established the relevant 8 indicators to reflect the highly-efficient recovery level of coal resources, summarized the present situation of China's coal resources recovery and the main problems, analyzed the development environment of China's coal industry under the new situation, and finally put forward the strategic measures to promote highly-efficient recovery of coal resources in China.1. Background Coal is the main energy in China. Since entering the new century, China's coal industry has made significant progress, and the level of production safety, resource recovery rate and mining mechanization increased significantly, which provided a reliable energy supply for the rapid development of the national economy. However, due to the changes of China's economic environment, slow growth of energy demand and other factors, some prominent contradictions in the coal industry have significantly restricted the healthy development of the industry. The first problem is serious overcapacity. In the past few years, 16 industries including coal, electric power, chemical and others have invested in coal mines. It is reported (Han 2016) that only during the nine years from 2005 to 2014, the fixed assets of coal industry increased totally more than 3.3 trillion RMB, which brought a large number of production capacity. Affected by the slowdown of economic growth and energy structure adjustment, Chinese coal demand has reached the peak platform and caused the overcapacity of coal supply, so shutting down the backward capacity is the most important task in the future. The second problem is the low efficiency of coal production. There are about 4.3 million coal mine workers in China, and the annual coal production per capita is less than 900 tons, while it was reported (EIA 2016) that the annual production per capita in the United States is more than 10 thousand tons. The third problem is that the green coal resources are not rich. Among the 1.53 trillion tons proven reserves, only about one third are green coal resources in accordance with the principle of ""safety, technology, environment and economy"". The coal resources with high quality and easily mining conditions are not abundant."

Jan 1, 2017

By Keith Kern, Joseph P. Matoney, Kenneth E. Harrison

INTRODUCTION "When there is an economic need for cleaning fine coal to low ash, the technical difficulties will be solved. The utilization of fines below 28 mesh will determine the degree to which beneficiation will be carried. Whenever these "fines" will be processed into smokeless fuels or used as new materials for the chemical industry or for hydrogenation, a higher degree of cleaning will be economically justified”.(l) This quote from the 1943 edition of Coal Preparation correctly anticipates the use of fines as "smokeless" fuel or as a future component of boiler fuel and a1 so as a raw material for the chemical industry or possible new uses such as for internal combustion engines. Fines utilization depends upon an economic need and possibly neglects to say an environmental need or a government funded development program as the prime mover for development. In any event, this conference is faced with the "need" to beneficiate fine coal and undoubtedly among the processes that will be discussed are the unit processes of future coal preparation facilities. HISTORY OF DEVELOPMENT It is interesting to look at where we have been, where we are now and where we possibly will be in the future. In order to put some limitations on this presentation, we will for the most part confine our remarks to this century and particularly developments since World War II. We also will limit our remarks to the preparation of bituminous coals in the United States, and for the most part, to wet processes. In the 1943 edition of Coal Preparation, clues to the processes being used to beneficiate coal are contained in the Chapter Titles. An extensive treatment of "Hand Picking" was contained in Chapter 8. Chapter 7 contained a description of methods for "Preparation at the Face". Chapter 18 dealt with "Slurries, Sludges, Slimes and Water Treatment" and defined this size of coal as being the size at which "Jigs and Chance Cones are usually dispensed with and other methods such as the hydrator, tables, and froth flotation are employed. Generally the limit is 1.7mm to 3.4mm. In the days before World War I1 the then modern bituminous coal preparation plant contained jigs or Chance Cones for the coarse material +6.35mm, concentrating tables, launders or some classifier type of coal washer for the fines (6.35 x 0.15mm), and froth flotation as the only method of cleaning minus 0.15mm coal which was used very sparingly.

Jan 1, 1988

By Wenbing Guo, Gaobo Zhao, Erhu Bai, Brijes Mishra

"Violent damage to overlying strata can occur above the large mined-out space left by high-intensity longwall mining. For the safe mining under the aquifer and surface water body, it is imperative to investigate the criteria of overburden failure degree due to high-intensity mining. In this paper we present the characteristics of high-intensity longwall mining panels and overburden critical failure. Based on the geological and mining conditions of a high-intensity mining panel, 3D numerical models that consider the maximum horizontal principal stress and minimum horizontal stress and their orientations with the panel advance direction were established by using the 3DEC software. Sizes of the panel reaching the critical failure of the overburden under different mining thicknesses (3.5 m, 6.5 m, 9.5 m, 12.5 m, and 15.5 m) were analyzed. Further, criteria of overburden critical failure were proposed for the high-intensity mining. The coefficient of overburden critical failure (Cf) is the ratio of the strike length (Ls) to the mining depth (H). Through curve fitting analysis, a fitting equation between Cf and mining thickness was obtained. The results showed that overburden failure criterion with high-intensity mining panel is 0.38 = Cf = 0.97 with different thicknesses. Field measurements validated the results from the numerical models.INTRODUCTIONHigh-intensity longwall mining refers to a coal mining method with thick coal seam, large panel width, fast face advance rate, and severe damage to the overburden and surface in China (Guo et al., 2018). Violent movement of the roof rock and severe damage to overlying rock strata occur in the large mined-out space left by a high-intensity longwall mining panel, which is a 3D (three-dimensional) structure layout with highly non-uniform stress distribution. Recognition of such fact a will facilitate this overlying rock strata failure problem investigation (Peng et al., 2019).The studies of overlying strata failure due to longwall coal mining have been reported by some scholars with different methods. Kang et al. (2018) studied the sudden massive roof collapse during longwall coal retreat mining by a large-scale two-dimensional physical model and a UDEC numerical model. Ghabraie et al. (2015) researched the mining-related strata movement by applying the three-dimensional laser scanner, optical transducer, and digital image processing techniques to two-dimensional physical models. Wang et al. (2015) analyzed the structural characteristics and fracture failure process of overlying hard and thick main roof by a 3DEC numerical model; this study also investigated the geostress reduction and permeability enhancement redistribution of the adjacent coal measure strata by FLAC3D. Guo et al. (2012) performed a three-dimensional finite element analysis and shared some insights into the dynamic interaction between mining induced strata stress changes, failures, and gas flow patterns. Meng, Shi, and Li (2016) proposed a 3D coupling model between stress and permeability, investigating the dynamics of a stress, permeability coefficient using FLAC3D. Khanal, Guo, and Adhikary (2019) used the 3D numerical modelling approach (COSFOW) to investigate the strata deformation and the change in permeability in the overburden strata due to the underground extraction. However, few studies considered the influence of the horizontal stress on overburden failure in a high-intensity longwall mining panel, and the critical failure criteria of overlying rock strata is unclear."

Jan 1, 2019

Drainage and erosion control often become problems at mine reclamation projects. An example of this was an abandoned surface uranium mine on the south Texas plains. Engineers with the Texas Railroad Commission solved these and other problems with a three-dimensional geomatrix erosion-control matting at the 69-hm2 (170-acre) site, about 70 km (45 miles) southeast of San Antonio. Enkamat, manufactured by Akzo Nobel Geosynthetics, was installed to reinforce the grass waterways that handle runoff from the reclaimed mine. Reclamation was completed in the spring of 1992. Bill Walker is regional manager for Erosion Control Systems. He worked with the contractor to identify an erosion control product that complied with the project's design and installation parameters. "Use of erosion control matting enabled us to reduce erosion control costs," said Mark Rhodes, with the Texas Railroad Commission. "Also, the erosion control material was more aesthetically compatible with the reclaimed landscape." The commission's design called for a terrace ditch to be constructed midway up the reclaimed spoil slope to catch and collect water. Commission engineers selected spoil slopes, ranging from 5% to 8%, to build reinforced grassed waterways to drain the collected water from the terrace ditches. This was done instead of dropping the water at steeper slope location in rip rap- or concrete-lined channels. Enkamat's three-dimensional geomatrix of heavy nylon monofilaments is 95% open space that aids root growth. Its root-reinforcement system resists the hydraulic lift-and-shear forces created by high-volume discharges of water. In addition, fine eddy turbulences are prevented from isolating and uprooting individual roots and clumps of vegetation. Enkamat is flexible and lightweight. This makes it easy to install. And, because grass grows into it over time, it "disappears" after installation. Between 12,540 and 14,200 m2 (15,000 and 17,000 sq yds) were installed in the waterway using 150-mm (6-in.) staples to fasten the matting, said contractor Dale Blackstock. "It was easy to use; no problem at all to put down." Blackstock selected Enkamat from a list of eight different channel lining materials described as acceptable in the commission's design specifications. The design specifications were based on the results of channel-lining performance tests conducted by the Texas Department of Transportation. Rhodes said that the commission's Abandoned Mine Land Reclamation program had installed Enkamat on other abandoned uranium mines and on an abandoned surface lignite mine near Austin. "Pretty much all of our installations of Enkamat have worked out well," he said. "There have been no failures. The reinforced grass waterways have performed as designed."

Jan 1, 1998

By Steve Kral

"Gold rush" is usually associated the western United States - California, Nevada, and Colorado. The southeast section of the nation is known more for its rice, textiles, tobacco, indigo, and cotton. This area, though, is where it all began with gold in this country. The United States' first gold rush, in fact, is said to have started with the discovery of a 7.7-kg (17-1b) nugget in the Piedmont section of the Carolinas in 1799, some 50 years before the rush to California. The Carolina gold rush continued intermittently through 1942 when the Federal Government shut down all gold operations. Strong gold prices throughout the early part of this decade and new extractive technologies spurred a small resurgence of gold operations in South Carolina. Mind you, this is not Nevada with its flood of geologists and mining companies. Nor are there any 7.7 kg (17 lb) nuggets being discovered. But Ridgeway Mining Co., a joint venture between Kennecott Corp. and Galactic Resources, has set about mining 51 Mt (56 million st) of low-grade ore containing an estimated 43.5 t (1.4 million oz) of recoverable gold. The open pit, carbon-in-leach operation is located near the town of Ridgeway in South Carolina, about 32 km (20 miles) north of Columbia. Carolina gold North and east of Ridgeway, Piedmont Mining Co. operates the historic Haile gold mine near Kershaw, SC. And Brewer Mining Co. operates the nearby Brewer deposit. Between 1829 and 1942, South Carolina gold mines produced 10 t (318,800 oz) of gold. It all began in 1799 when Conrad Reed, son of a German immigrant farmer, found the 7.7 kg (17 lb) nugget in Little Meadow Creek, NC. His father thought the rock was worthless and used it as a doorstop, so the story goes. It took nearly three years before Conrad's father decided his doorstop was indeed gold. Gold was later discovered in the Greenville district of South Carolina. And, in 1828, Benjamin Haile discovered gold in a stream that ran through his property in Lancaster County, SC. Haile's mine became South Carolina's most famous and most productive gold mine, producing more than 7.8 t (250,000 oz). At first, miners worked the stream beds of Haile's property. Later, lode deposits were discovered on slopes alongside the creeks. They would lease 4.6 ml (50 sq ft) lots and mine as deep as 7.6 m (25 ft). During the Civil War, the Confeder-

Jan 1, 1989