Search Documents

By Daniel W. Kappes

This paper examines the issue of heap leach cash flow based on some recent operational examples. After the first six months of operation, single-lift continuous heaps produce into salable bullion only 70% of the total recoverable gold stacked. This compares with "batch" heaps and lab columns which achieve 100% in two to three months. Multiple-lift heaps often experience a second cash flow depression as a result of the holdup of gold in the heap. Heap leach projects seldom fail based Oil technical pe1for111ance, but they often fail in the sense that they gmerate cash flow at a much slower rate than expected.

Jan 1, 2000

By G. F. Willis

Gold mineralization in the Mesquite district is hosted by a series of Jurassic amphibolite grade metamorphic rocks which have been intruded by a Cretaceous leucogranite pegmatite. A series of northwest trending, high angle, normal faults have apparently acted as a structural locus for mineralized epithermal fluids and thus economic concentrations of gold in the Big Chief orebody. Late Tertiary fault reactivation has bounded the fault and fracture controlled mineralization. Popular theories of ore genesis call on, (1) Tertiary detachment faulting, (2) the Cretaceous intrusive, or (3) the opening of a dilational jog between regional dextral strike-slip faults as a source powering the gold bearing hydrothermal fluids.

Jan 1, 1988

By Leo G. Martin, Richard G. Sherman

Recent sewer construction in downtown New Haven, CT required avoiding open trench methods because of near surface utilities and traffic congestion. The .73 km long alignment is in sands with the water table at the tunnel's springline. The designers felt the situation was ideal for an EPB machine; however, the tunneling method was left to the Contractor. The winning bid surprised everyone. The tunnel is being excavated by hand using an open face shield with breasting boards. The review of the bids indicates hand mining was approximately 1/2 the bid price of TBMs. Currently, excavation rates average 14 m/day with no excessive settlements.

Jan 1, 1993

By Tor L. Brekke, Alberto Marulanda P.

The paper describes hazardous water inflows during construction that have taken place over the last three years in tunnels in Colombia. The projects involved are the Chivor II Hydroelectric Power Project (penstock tunnel), and the Chingaza Water Project (two tunnels). In all cases, loss of face control led to significant loss of ground; in one case some 70,000 m3 of slide material was involved, much of which came out through the portal situated 1.6 km from the face. Emphasis is given to: (a) the geologic setting, (b) the remedial measures taken, and (c) the main efforts that can be made to prevent hazardous conditions from developing.

Jan 1, 1981

By Francis B. Gularte, James E. Monsees, John-Paul Whyte, Gary E. Taylor

In a short section of the Los Angeles Metro Rail system approximately 222.5 m (730 ft) of 6.4 m (21 ft) diameter twin-tube tunnels separated by a 1.8 m (6 ft) narrow pillar were designed to pass under a major eight lane freeway and a metropolitan intersection with only 3.7 m (12 ft) to 5.8 m (19 ft) of loose fill and alluvium cover. The design considerations and the chemical grouting program used to stabilize the alluvial materials is discussed. Also included is a discussion of the novel shield used for tunnel excavation and the effects of the subsequent fire.

Jan 1, 1991

By M. L. Leming

Introduction In the decades since silica gel was first identified in material exuding from cracked concrete, a great deal of research has been conducted regarding the chemical reactions between the alkalies found in portland cement and silica found in aggregates. The reaction is complex and one that is not yet completely predictable, especially from the point of view of developing specifications that are appropriate to all situations. This paper is not intended to be a rigorous review of the research findings but is an attempt to provide a simplified review of the mechanisms of the alkali-silica reaction (ASR), so that one can better understand the implications of the specifications, test results and effects on structures. In addition, the contractual relationships between the aggregate supplier and one of their major clients, the concrete supplier, will be examined with regard to the ASR. ASR basics Silica. Silica (silicon oxide) may exist in naturally occurring aggregates in various forms and in combination with a number of other elements. When the silica is completely crystalline, such as in quartz, it is chemically and mechanically stable. Quartz silica is impermeable and reacts only on the surface of the crystal, where the silicon and oxygen bonds are broken. Because the surface area per unit volume of most quartz is low, the reactivity is also low. Completely amorphous (noncrystalline) silica is, on the other hand, more porous and very reactive. The less "crystalline" the silica is in the aggregate, the more reactive. Silica that has melted and cooled quickly without recrystallizing, creating a glassy material (such as in certain volcanic aggregates), has a very low state of crystallization and will be much more reactive in an alkaline solution. Crystalline silica that has been transformed by heat and pressure may have a large quantity of strain energy stored in the crystal lattice. The presence of this higher energy will make the silica more likely to react. The "strained quartz" found in many metamorphic aggregates means that these aggregates are potentially susceptible to deleterious alkali silica reactivity, although the rate of reaction is typically much slower than with aggregates composed of or containing glassy or amorphous silicas. Another problem may exist with aggregates in which the silica is primarily crystalline. In aggregates such as chert, in which the silica exists as very fine crystals (i.e., crypto- or microcrystalline), the very high surface energies between the crystals contribute to alkali sensitivity. Therefore, the potential reactivity of an aggregate is seen to be a function of both the degree of crystallization of the silica in the aggregate and the amount of energy stored in the crystal structure, whether due to a large quantity of microcrystalline silica, a high strain energy stored in the crystals or some combination of these factors. The surface area per unit volume of the reactive silica will also affect the rate of reaction, because a larger surface area of reactive silica will have more opportunity to react. Obviously, the reactivity of the aggregate is also affected by the silica content. However, in this case, the results are not quite so obvious. A discussion of the effect of silica content will be postponed until after a discussion of the contribution of the cement paste. Paste characteristics. Hydrated portland cement is a very alkaline material with a pore solution pH typically in excess of 12. The alkaline environment of moist concrete provides an ideal place for noncrystalline or cryptocrystalline silica to react. However, not all alkalies are equal in their effects. Calcium compounds react with glassy silica to form calcium silicate hydrate, commonly abbreviated C-S-H a poorly crystalline material that can occur in several forms and chemical compositions. C-S-H was at one time called tobermorite gel, because it was chemically similar to the naturally occurring crystalline mineral tobermorite and because it had a gel-like (noncrystalline) structure when viewed under an optical microscope. The formation of C-S-H is the basis for both portland cement hydration and reaction with, for example, fly ash. C-S-H is relatively stable. Although drying will cause some shrinkage and rewetting will cause some expansion, the volume stability of the C-S-H is very good compared to the volume stability of most alkali silica gels. Alkali silica gels with high sodium contents, for example, are nonstable compared to C-S-H.

Jan 1, 1997

By Sharanappa Yalal, Rakesh Kumar Khali

"Himalayan geology poses the most challenging ground conditions almost anywhere in the world, one of the reasons thereof is their evolution as the youngest mountain chains. They are demonstratively rising faster than any mountain elsewhere. The composition is also younger, and consequently less well consolidated than the other older fold belts. This is consistent with the fact that they constitute one of the most active of the plate margin zones rising at a rate that is almost double the Andes, which is in turn three times that of the Alps. Thus, stress conditions (Magnitude and variations) can potentially be more extreme and adverse on a Himalayan tunneling project. Tunnel excavation under the Himalayan Mountains will pose significantly more challenges than an equal length and equal cover drive almost anywhere in the world. These in-depth-tunneling difficulties through high mountainous terrains have posed major challenges for application of traditional drill and blast (D&B) and NATM methods. Handling such adverse geology at any depth is always problematic and generally leads, if not adequately foreseen, to the significant tunneling delays. The author wants to highlight the challenges faced and on the spot solutions adopted while executing most prestigious 273Km long Udhampur –Srinagar- Baramula Rail Link Project (USBRL) in the Himalayas in the state of Jammu & Kashmir, India. The main focus of the paper is on main stretch of USBRL Project i.e. T-80 (India’s 2nd longest railway tunnel & 3rd longest in Asia) where the team of the author has recently completed a 11.2 Km long tunnel and where the train track has gone operational from Qazigund to Banihal. The construction of this stretch posed several challenges akin to highly adverse geological conditions. Executing on -the -spot solutions, the project has been completed successfully. 1.0 INTRODUCTION Construction of Udhampur-Srinagar-Baramulla Broad gauge railway line is one of the Mega projects undertaken by Indian Railways to establish a dependable transportation in the state of Jammu and Kashmir. In engineering terminology, it is a challenging project negotiating breakthroughs against the complex Himalayan geology. The Project which has been operational over last 12 years , stretch by stretch, is in advance stage of construction . The purpose of the prestigious project is to connect the Kashmir valley to the mainland of the state of the country. On the date owing to the hurdles of the tuff terrain, the transport of this part is totally dependent on the surface transport, i.e., through a road which connects NH-1A from Delhi to Srinagar. The altitude of the region varies from 500.00 to 1700.00, ranging through Udhampur to Katra - Reasi –Sangaldan- Banihal-Quazikund-Ananthnag-Srinagar-and Baramulla. The main features of this project are its Pirpanjal Tunnel and T-48 & 49 tunnels which cover the maximum length of tunneling through thick hostile geological conditions, the total length being 138km. Details of Pirpanjal tunnels are shown in Figure-1."

Jan 1, 2016

By R. L. Miller

"Miners worldwide use a tremendous amount of energy in producing finished minerals for the end user market. In accomplishing this goal, they incur energy costs that can range from 20% to 30% of the total cost of production, so there is fertile ground for cost savings. To unlock this huge value in cost reduction, energy analysis is the key and is broken into several segments: 1. Fuel analysis for power generation 2. Energy supply & demand 3. Opportunity cost for lack of power 4. Backup power 5. Power contracts 6. Energy efficiency 7. Renewable energy 8. Continuous energy management FUEL ANALYSIS FOR POWER GENERATION What is the first thing we evaluate when looking at power supply alternatives for the mining industry’s large power requirements? Is it 1) site, 2) engineering, procurement, and construction companies, 3) independent power producers, 4) turbine vendors, 5) grid connections, or 6) fuel supply, logistics, and infrastructure? In evaluations of power supply alternatives for mining companies with special emphasis in developing countries, there is one over-riding conclusion: reliable fuel supply drives the energy generation economic engine. Evaluating several different fuels including liquefied natural gas (LNG), natural gas, diesel, heavy fuel oil, and liquefied petroleum gases (LPG) among others, the best project outcomes are those where fuel supply, pricing, logistics, and infrastructure are given an extensive amount of time during the analysis stage. In an assessment of the twenty-year costs of generating power from multiple fuels, each fuel’s costs were placed into three separate net present value (NPV) buckets of: 1) capital, 2) operating and maintenance, and 3) fuel. Within a range of 65% to 85% for all fuels, fuel cost was the largest component of the NPV of the life cycle power generation costs. So if 65% to 85% of your life cycle power generation cost is one item, shouldn’t miners spend a corresponding amount of time in the analysis stage fully understanding this key determinant for a successful power project? In developing countries where power generation infrastructure may not exist or does not exist at the same level as is found in North America and Europe, fuel analysis becomes even more important. One of the key questions for fuel importation in developing countries is the draft at existing ports to allow larger vessels to dock at the jetty. Can larger ocean-going tankers dock at the port for off-loading, or is the port’s draft too shallow, thus forcing a Single Point Mooring (SPM) system or Conventional Buoy Mooring (CBM) system to be constructed at substantial capital expense? Once fuel can be imported, sufficient onshore storage capacity should exist, or must be built to off-load the entire fuel cargo from the larger ships. Additionally, the throughput capacity of the pipeline from SPM/CBM must allow a rapid unloading of the fuel from ship to shore to eliminate steep demurrage costs for the project."

Jan 1, 2017

By Gail L. Achterman

Most mining companies today perform site assessments and operational or compliance audits regularly. The purpose of these audits is to review company facilities and operations to determine whether they comply with applicable federal, state and local environmental laws and regulations. Why Audit? Companies adopt assessment and audit programs for legal and business reasons. The business reasons include: (1) ensuring efficient and cost-effective compliance through early problem identification and resolution; (2) planning and implementing business transactions such as real estate conveyances, loans, mergers and acquisitions, and waste disposal service contracts; and (3) employee training and relations. Legal reasons for audits include: (1) avoiding enforcement actions and penalties; (2) meeting environmental permit certification requirements; (3) avoiding corporate officer liability; and (4) meeting Security and Exchange Commission disclosure requirements. The Environmental Protection Agency ("EPA") encourages environmental auditing by regulated entities. The EPA Environmental Auditing Policy Statement was published in the Federal Register on July 9, 1986 (51 Fed Reg 25,004). Although the existence of an auditing program does not create a defense to or otherwise limit a company's regulatory responsibilities, EPA considers compliance programs in setting inspection and enforcement priorities. If a company makes an honest effort to identify and correct its own problems, EPA's enforcement approach is influenced, especially if violations are promptly reported. Perhaps more significant to many executives, the U.S. Department of Justice issued guidelines last year on when it will pursue criminal prosecution for environmental violations. See Department of Justice, "Factors and Decisions on Criminal Prosecutions for Environmental Violations in the Context of Significant Voluntary Compliance or Disclosure Efforts by the Violator," (July 1, 1991). It is now commonplace to read about major companies pleading guilty to criminal violations and paying millions of dollars in penalties. For example ALCOA recently agreed to pay criminal fines and civil penalties totaling $7.5 million for unauthorized excavation of PCB-contaminated soil and improper dumping of spent caustic solution. No executive wants to explain to the Board of Directors or shareholders why the company was criminally indicted. The Department of Justice considers factors such as voluntary disclosure, preventive measures and compliance programs, such as a regular compliance auditing program, the effectiveness of the disciplinary system for employees who violate the company's environmental policies, and subsequent compliance efforts by companies to remedy ongoing noncompliance. The Department of Justice looks closely at whether audit recommendations were implemented in a timely way and whether environmental compliance was a standard by which employees were judged. The Lawyer's Role. The role of the lawyer in an audit team deserves special attention. When enforcement action or litigation has begun or is contemplated or a government agency has asked for information, attorneys always should be directly involved in environmental audits. In most situations, however, environmental audits are management tools, and the attorney's role should be to advise the team. The attorney can help make sure the team understands the laws and regulations the facility must

Jan 1, 1992

By A. D. Davis, C. J. Webb, A. Heriba

Introduction The disposal of spent ore from cyanide heap-leach processing facilities is of concern to the mining industry, the regulatory agencies and the general public. The disposal of an additional several hundred million tons of spent ore from heap-leaching operations is planned in the western United States during the next decade, and this amount could increase if gold prices rise. In the past decade, more than fifteen million tons of spent ore have been disposed of in the Black Hills of South Dakota. The criterion for offloading spent ore was set by the South Dakota Department of Environment and Natural Resources at 0.5 ppm weak acid dissociable (WAD) cyanide. South Dakota's nitrate offloading limits vary for each mine, depending on ore type and other factors. Concentrations of nitrate in the leachate from spent ore depositories could be the result of conversion from other nitrogen-containing species. One likely source of nitrate is residual explosives from the blasting of ore. Previous work on the combustion products of nitrogen explosives has shown incomplete oxidation (Johansson and Persson, 1970). Another likely nitrate source is the degradation of cyanide to ammonia, followed by oxidation to nitrite and nitrate. A third possibility may be leaching of background nitrate from soil or rock. Objective The primary objective of this research was to investigate laboratory procedures for predicting nitrate concentrations in effluent after the disposal of spent ore at heap-leach gold processing operations. Laboratory tests were performed to define the concentrations of nitrate and nitrite in leachate as a function of time. Sources and species conversions were examined by investigating the nitrogen contributions from cyanide neutralization and from blasting with ammonium nitrate/fuel-oil (ANFO) explosives. Ore processing at cyanide heap-leach facilities Gold mines in the northern Black Hills of South Dakota normally use the following steps in their heap-leach cyanidation processing: Blasting. Surface gold mines in the Black Hills use ANFO explosives (approximately 95% ammonium nitrate and 5% fuel oil). Field evidence and studies show that explosives often fail to ignite or burn completely in shot holes. One reason is that ammonium nitrate absorbs moisture readily, which reduces blast efficiency. Ferguson and Leask (1988) cited a study conducted at the Fording coal mine in southeastern British Columbia. In that study, the total nitrogen released to surface and ground water was estimated to be about 6% of the slurry explosives. However, Ferguson and Leask (1988), in a later analysis of water quality near coal mines in the Kootenay coal fields, indicated that this is likely an overestimate of the nitrogen release. Mines that used ANFO explosives in dry conditions released 0.2% of the nitrogen from blasting. Cyanidation. At surface gold mines in the northern Black Hills, cyanide species from heap leaching can be classified into free cyanides and simple ions, weak complexes, moderately strong complexes and strong complexes (Scott and Ingles, 1981). Degradation of cyanide and disposal of spent ore. Methods of destroying cyanides include hydrogen peroxide oxidation, natural degradation and evaporation, water leaching and alkaline chlorination. Hydrogen peroxide is used to oxidize cyanides in the northern Black Hills mining area. The hydrogen peroxide oxidation reaction forms cyanate, which can hydrolyze to form ammonium and carbonate ions. After formation of ammonia or ammonium ion, further oxidation to nitrite (NO2-) and nitrate (NO3-) can occur. Spent tailings from gold processing in the Black Hills normally are disposed of in tailings facilities constructed in nearby drainages. Infiltrating rainwater with a typical pH of about 5.6 is an additional component introduced into the spent ore after disposal. Methods Field sample collection. Spent ore samples were collected from conveyor belts (or other convenient points) at heap-leach mines by personnel from the South Dakota Department of Environment and Natural Resources. The samples were placed in black plastic bags that were contained within 5-gal buckets and sealed to prevent exposure to ultraviolet light. The samples had a minimum head space to limit evaporation and contact

Jan 1, 1997

By D. Akers, R. Dospoy

Introduction The geological processes that form coal can also concentrate trace elements in the coal. For example, the average concentration of arsenic in bituminous coal (20 ppm) is ten times the average concentration found in all the other rocks that make up the earth's crust (2 ppm). Similarly, other elements, such as antimony, cadmium, mercury and selenium, are more concentrated in coal than in the earth's crust. When coal is burned, trace elements can be further concentrated. Although no new constraints on trace clement emissions were placed on the power generation industry under the 1990 Clean Air Act Amendments, the act does mandate a three-year study of air toxics. Any new regulations aimed at electric utilities as a result of this Federally-mandated study will almost assuredly be very costly-an estimated $1 billion per year. Many trace elements in coal are associated with mineral matter. For example, arsenic is commonly associated with pyrite, cadmium with sphalerite, chromium with clay minerals, mercury with pyrite and cinnabar, nickel with millerite, pyrite and other sulfides, and selenium with lead selenide, pyrite and other sulfides (Finkelman, 1980). There are also cases in which some of these elements are organically bound. Just as both organic and pyritic sulfur can be found in the same coal, the same trace element may be both organically bound and present as part of a mineral in the sane coal. Physical coal cleaning techniques are effective in removing mineral matter from coal and can potentially remove at least some of the trace elements associated with specific minerals, thereby reducing the release of these elements into the atmosphere. Conventional coal cleaning to remove trace elements As part of a project funded by the Electric Power Research Institute, CQ Inc., a wholly-owned EPRI subsidiary located in western Pennsylvania, has demonstrated that large reductions in the concentration of many trace elements are possible if conventional coal cleaning techniques are properly applied. Four examples are given in Tables 1 to 4. In each example, the results shown were generated by cleaning the coal at CQ Inc.'s commercial-scale cleaning test facility. Cleaning results for Upper Freeport Seam coal from Northern Appalachia are provided in Table 1. Data are presented in the table in two ways: as a weight-based concentration (parts per million) and as a concentration per heat unit (grams per billion Btu). Grams per billion Btu is analogous to pounds per million Btu, but avoids the use of numbers with many decimal places. The heat-based concentration provides a better measure of boiler impacts, because the increased heating value obtained through coal cleaning reduces the number of tons that must be burned to produce a given thermal output. Reducing the quantity of coal burned reduces the quantity of trace elements entering the boiler. This raw coal is relatively high in several trace elements of environmental concern, including arsenic, cadmium and chromium. Cleaning provided large reductions in the quantity of arsenic, barium, cadmium, chromium, fluorine, lead, mercury, nickel, silver and zinc. The results for tests with a Powder River Basin coal, Rosebud/McKay, are presented in Table 2. Large reductions in arsenic, barium, cadmium, fluorine, mercury, nickel, selenium and zinc were observed with cleaning. The concentration of chromium increased with cleaning, while lead concentration increased in one test and decreased in another. Table 3 presents test results for Croweburg Seam coal from Oklahoma. Large reductions in arsenic, barium, cadmium, chromium and zinc were obtained with cleaning. Smaller reductions were obtained with lead and nickel, while chromium, fluorine and mercury increased in at least one of the tests. Table 4 presents cleaning test data for Kentucky No.11 Seam coal. In this case, large reductions were obtained with all elements measured. In general, these data indicate that physical coal cleaning is effective in reducing the concentration of many trace elements, especially if they are present in the coal at relatively high concentrations. The degree of reduction achieved is coal-specific, relating in part to the degree of mineral association of the specific trace element and the degree of liberation of the trace element-bearing mineral. The extent of trace element removal also depends on the method of cleaning the coal. Figure 1 is a washability plot, by size fraction, of arsenic vs. ash content for Upper Freeport

Jan 1, 1994

By J. Guzman

Introduction The Carlin Trend is North America's premier gold producing district. It is located in northeastern Nevada's Elko and Eureka Counties along a northwest trending belt about 65 km (40 miles) long and 8 km (5 miles) wide (Thorstad, 1989; Jones, 1989). This trend is the worldwide reference site for epithermal, sedimentary rock-hosted microscopic gold deposits. At least 19 deposits have been discovered to date, varying in size from 933 kg to 1.08 kt (30,000 to 35 million contained oz) of gold (Fig. 1). Newmont Gold Co. and its parent, Newmont Mining Corp., jointly constitute the largest mineral right holders in the district. They own or control more than 1000 km' (386 sq miles) in and around the Carlin Trend and own all or part of I6out of the 19 mines and prospects identified to date. Since the initiation of Newmont's exploration activities in the Carlin Trend in 1961, 2.24 kt (72 million oz) of cumulative gold resources have been identified. Cumulative production from all mines since the start-up of Newmont's Carlin mine in 1965 to the end of 1989 was about 202 t (6.5 million oz) (Jones, 1989). The incentive of sustained high gold prices and innovation in processing technology resulted in a significant acceleration of gold output over the last few years. Newmont Gold alone produced more than 43.5 t (1.4 million oz) in 1989. That is equal to 22% of the cumulative 1965 to 1988 output, and an almost 200% increase over its 1986 output. The same incentives produced even more spectacular exploration results. In each of the last five years, net additions to reserves and resources outpaced current production by substantial margins. These facts demonstrate the spectacular past prospectiveness of the Carlin Trend and the success of focused, multi-disciplinary exploration methods that made it possible to more than offset the recent accelerated depletion of gold resources. However, is this situation sustainable? How long can the mining companies along the Carlin Trend keep on finding resources faster than they deplete them? These are some of the questions that motivated this study. The authors have not quantified the future potential for gold exploration in the Carlin Trend nor established a deposit discovery path. But strong indications were discovered that the [ ] Carlin Trend remains a relatively immature exploration district and that the potential for significant new discoveries is high. Methodology and data The approach chosen to address the above questions was simple. The authors compiled deposit size data, measured in contained ounces of gold resources, for all known deposits along the Carlin Trend (Table 1). The resource information was obtained by adding cumulative historical production (adjusted for mining losses and metallurgical recovery) to 1989 year-end published resource inventories. In a mature exploration area, where most deposits have been discovered, this distribution would be expected to approximate lognormality and would plot along a straight line on a lognormal probability scale. This result was found in previous work by Allais (1957) and recently confirmed by Cox and Singer (USGS, 1986) in regard to various types of mineral deposits in several regions of the world. It was also found to hold true for oil and gas pool size distributions (Arps and Roberts, 1958; Kaufman, 1962; McCrossan, 1969). [ ] The data used were compiled by Newmont Exploration geologists. The purpose of the study is to make inferences about the underlying geologic processes in the district and the maturity of the exploration effort. Therefore, deposits were not classified according to ownership but according to geologic occurrence as known from current information. Newmont's Post and Barrick's Goldstrike and Betze deposits, for example, are shown as a single occurrence to reflect the actual geologic setting. The cumulative frequency distribution of deposit sizes was plotted on lognormal probability paper (Fig. 2). The abscissa shows the cumulative fraction of deposits at or below a certain deposit size and the ordinate shows the deposit size in thousands of ounces of contained gold resources.

Jan 1, 1992

By Jozef S. Stachulak

Inco Limited operates 11 underground mines in the Sudbury District. The mines are located on the rim of the Sudbury Basin, an oval with the axis in the range of 27 and 60 km. The ore dips to at least 3000 m below surface. The ores are mined primarily for nickel and copper. Total ore production from underground is in excess of 55,000 tons per day. Over 40 surface fans have been installed since the late 1960's. All of the fans are adjustable pitch, axial flow units. A major factor influencing ventilation design in the last 30 years has been the introduction of diesel equipment underground. Volumes per fan have ranged from 60 to 330 (cubic metres per second), with motors from 100 to 2500 hp. Fans of the axial flow type have been in common use for main fan installations at Canadian mines for many years. The standard arrangement has been to mount these fans horizontally, i.e. with the fan shaft and the long axis of the housing horizontal. This is a natural arrangement for an underground fan, but for a surface installation, a vertically mounted fan has definite advantages. The surface area taken up by a typical vertical fan installation is generally about one quarter of that with a horizontal fan of the same capacity. (1) This is not a problem with isolated fans and flat surface outcrop sites, but where the installation is to be near existing buildings, or where there are poor surface soil conditions, space and cost considerations greatly favour vertical fans. MAIN FANS INSTALLED There are 43 main surface fans in operation at 11 mines. Twenty-seven of these fans are supply units, and 75% of them are vertical installation. The remaining 16 units are main surface exhaust fans, with predominantly horizontal installation. Within the last five years, some 20 main booster fans have been installed underground at several mines. Axial flow fans, with adjustable pitch blades, are used for both surface and underground installations. Exhaust fans are equipped with stainless steel or cast aluminum blades. Main underground fans are arranged horizontally and the majority of them have a floating shaft between the fan shaft and the motor. (2) The size of the fans in service varies from 1.8 m to 5 m in diameter, with the majority ranging from 1.8 to 2.5 m. The pressure produced by these fans varies from 0.25 kPa to 2.0 kPa. At Inco Limited, two main fans in parallel are preferred, rather than a single fan, so that if one fan fails, the remaining fan can still supply up to 70% of normal air quantities, while the damaged fan is repaired. This requires closure doors on each of the fans so the fan can be isolated in case of failure. It is more expensive than a single fan, but results in less production interruptions. The fan installations are well away from sharp inlet and outlet bends. FAN DESIGN INTEGRITY Both the mine operator and the fan manufacturer must understand that the main fan is critical to the mine operation, and that everything technically possible must be done in design and manufacture to ensure the highest degree of reliability. Some of the design parameters and criteria, based on Inco experience, are outlined and discussed below. RESONANT FREQUENCIES AND HOUSING MODEL ANALYSIS Any fan assembly will have many different resonant frequencies. It is a challenge to the designer to arrive at a design in which forcing frequencies do not coincide with any of these resonant frequencies to produce unacceptable vibration levels in operation. Finite element analysis is a useful tool that can be used to identify the most critical of these, so that housing and blade stiffness can be adjusted to change any resonances that might be close to forcing frequencies. Shaft critical speeds should be at least 25% above the fan operating speed, and there should be sufficient separation between other resonant and forcing frequencies to avoid excitation that might result in high vibration levels. QUALITY ASSURANCE Radiographic Blade Examination Mine fan blades are normally cast in small lots. To ensure that the castings are sound, a full radiographic examination is recommended in the highly stressed lower 1/3 of the blade.

Jan 1, 1995

By K. Mulukhov

Introduction The depth of open-pit mines continues to increase. The typical designed depth for large quarries is now in the 300 to 600 m (1,000 to 2,000 ft) range, with some quarries even reaching depths of 700 to 800 m (2,300 to 2,600 ft). The cost of handling material from such deep mines can be 60% to 70% of the total cost. The capacity of using a cyclic means of transport, such as trucks or skip hoists, continues to drop as the depth of the mine increases. In addition to the problem of maintaining relatively low slope angles for haul roads, increases in environment pollution and increased fuel consumption make the use of trucks ineffective below 100 to 150 m (300 to 500 ft). In handling large quantities of bulky materials, i.e., more than 15 to 20 Mt/a (16 to 22 million stpy), under fixed terminal points, belt conveyors are preferable in almost in all cases. An exception is when the lump size of the load is too high for the conveyor to handle. Belt conveyors in mines can handle loads having a maximum lump size of no more than 250 to 300 mm (10 to 12 in.). The use of a movable crusher prior to the belt conveyor in an open-pit mine would make the process very expensive. Such an operation could not compete with truck haulage on a cost basis. Belt-car conveyor The main factor that restricts the lump size for conventional belt conveyors is collision impacts between the load-carrying belt and idlers. These impacts are completely eliminated in the belt-car conveyor, which was previously described by Mulukhov (1977). The unique feature of the conveyor is that the load-carrying belt is supported by moving cars. Each car consists of a troughed cross strap that supports the belt with the load and two rollers, on which the cars move along upper and lower rails. Cars are connected by two endless chains revolving over sprockets. The system uses conventional drive pulleys, tail pulleys and return idlers. There are no fasteners between the carrying belt and the car cross straps. The conveyor belt simply drives the cars along by the force of friction between the belt and the straps. A belt-car conveyor was first installed in 1970 at the Karatau mining operation in Kazakhstan. The system was an elevating conveyor with an inclination of 20° and was located on the final part of the truck haul road. The conveyor transferred blasting phosphate ore from the quarry to the crushing plant. A bunker was mounted between the trucks and the conveyor. The maximum lump size was typically 1,200 mm (48 in.). However, there were some lumps as large as 1,500mm (60 in.), which was greater than even the belt width (Fig. 1). Because the prototype belt-car conveyor was intended only for industrial trails, it had a relatively short length of about 50 m (160 ft). Nevertheless, the conveyor yielded a substantial profit over the cost of truck haulage. Naturally, the first question one might ask is why such a system is not now widely used. The reason was the further development of a huge transporting complex for the whole quarry at the base of the belt-car conveyor. After successful industrial trials from 1971 through 1973, which demonstrated the systems engineering feasibility, the State Planning Committee of the USSR included the development of the belt-car conveyor in its plan for the development of the national economy. However, mistakes were made. First, a single organization was not in charge of the development. Second, a plant that was not specialized in belt conveyors was put in charge of the design and manufacturing of the conveyor system. In addition, the crusher was installed before the stacker, which negated the major advantages of the belt-car conveyor. For loading the conveyor, a special blade feeder with pendulum suspension was developed, but it was found unreliable. The main conclusion was that, in-

Jan 1, 2003

By Dennis S. Kostick

Introduction Soda ash is known chemically as sodium carbonate, an important inorganic chemical. It has been produced for several centuries by processing certain vegetation and minerals. The US soda ash industry has evolved from several small sodium carbonate mining operations in the West. Now, a nucleus of six companies produce about one-fourth of the world's annual soda ash output US producers currently dominate the world market. But certain international events are occurring that will reshape the domestic soda ash industry in the next decade. Historical perspective Soda ash is used mainly in the manufacture of glass, soap, dyes and pigments, textiles, and other chemical preparations. All of these are the first basic consumer products produced by developing societies. About 3500 BC, the Egyptians became the first society to use crude soda ash. The soda ash was used to make glass containers. It was most likely obtained from dried mineral incrustations around alkaline lakes. Soda deposits were virtually nonexistent in western Europe. So people resorted to burning seaweed to obtain the ashes. The ashes were then leached with hot water and the solute was recovered after evaporating the solution to dryness. The solute, a crude "soda ash" was impure. But, it could be used to make glass and soap. These two products and industries were important to the population and economic growth of the region. About 11.5 t (13 st) of seaweed ash was required to produce about 0.9 t (1 st) of soda ash. Along the coasts of England, France, and Spain, seaweeds with varying alkali contents became important items of commerce and sources of soda ash before the 18th century. The LeBlanc process used salt, sulfuric acid, coal, and limestone. It became the major method of production from about 1823 to 1885. In the early 1860s, Ernest and Alfred Solvay, two Belgian brothers, successfully commercialized an ammonia-soda process to synthesize soda ash. It used salt, coke, limestone, and ammonia. The Solvay process produced a better quality product than the LeBlanc method. In 1879, Oswald J. Heinrich presented to the Baltimore meeting of AIME, a paper entitled "The manufacture of soda by the ammonia process." The paper compared the two processes and foretold the demise of the LeBlanc technique. World production of soda ash in 1880 was 680 kt (750,000 st). Of that, 544 kt (600,000 st) was produced by the LeBlanc process. Of the 2.8 Mt (3.1 million st) of soda ash produced worldwide in 1913, only about 50 kt (55,000 st) was by the LeBlanc method. The LeBlanc process was never used successfully in the US, except for a brief period from July 1884 to January 1885 in Laramie, WY. Previously, soda ash had been produced by burning certain plants, as exemplified by the early Jamestown colonists, or by recovering small quantities of natural sodium carbonate found in alkaline lakes, such as those found near Fallon, NV, and Independence Rock, WY. Before the 1884 startup of the first synthetic soda ash plant in the US at Syracuse, NY, most of the domestic soda ash demand in the East was met by imports, primarily from England. Large-scale commercial production of natural soda ash began in California in 1887 from surface crystalline material at Owens Lake. Production from sodium carbonate-bearing brines at Searles Lake began in 1927 (Fig. 1). In 1938, during exploration for oil and gas in southwestern Wyoming, a massive buried trona deposit, presumably the world's largest, was accidentally discovered. Recent mineral resource evaluation by the US Geological Survey and the US Bureau of Mines indicates that the Wyoming trona deposit contains 86 Gt (93 billion st) of identified trona resource in beds 1.2 m (4 ft) thick or greater. Additionally, there is about 61 Gt (67 billion st) of reserve base trona. Of this 36 Gt (40 billion st) is in halite-free trona beds and 24 Gt (27 billion st) is in mixed trona and halite beds. In 1953, the Food Machinery and Chemical Corp. (later shortened to FMC Corp.) became the first company to mine trona in Wyoming. Soda ash demand increased.

Jan 10, 1985

By Archie M. Richardson, Carl E. Brechtel, Tom R. Kelly, Frank Feero

Recent work to upgrade the ventilation system for the el Mochito lead/zinc mine in central Honduras is discussed. Network modeling and underground measurements were used to evaluate cost-effective alternatives for achieving satisfactory ventilation in a complex and expanding underground operation. Both interim and long-term solutions were implemented to make mining possible under difficult conditions. INTRODUCTION Mining operations at the el Mochito Mine, located in the central highlands of Honduras, Central America, have been virtually continuous since the opening of the mine in 1948. Initially a high-grade silver mine, the mine has expanded along a westward trend of pipe-like orebodies to a distance of roughly one mile from two centrally located shafts. Production of the relatively large San Juan orebody at the western most extension of the mine (Paddock, 1981) led to the introduction of diesel-powered equipment; however, the ventilation infrastructure was insufficient to meet the needs of mechanized mining. A succession of owners/operators in the 1980s allowed the existing ventilation infrastructure to decay to the point that ventilation in the production areas of the mine was very poor. Environmental conditions in some working areas were not conducive to efficient ore extraction because of high dry bulb temperature, high humidity, and diesel emissions. Upon acquiring the mine in 1990, Breakwater Resources of Tucson, Arizona began an aggressive program to refurbish the mine infrastructure to complete extraction of the San Juan orebody and to allow the extension of the mine another 2500 ft to the west for extraction of the Nacional orebody. The program included increasing the capacity of the main ventilation system. This article presents a case history of the process of upgrading the ventilation system in a mine where extensive old workings cause large air leakage. This process has been one of selecting solutions to difficult technical problems that are compatible with the existing mine infrastructure and economic constraints. The initial ventilation system is described in the background section, along with ventilation projections for the mine expansion. Field characterization of the ventilation system for design verification and fan specification is then discussed. The paper describes a series of interim changes to the system to improve ventilation pending completion of new ventilation boreholes. In addition, the temperature/ heat problems in the mine are described. BACKGROUND Initial Condition of Ventilation System The ventilation system is illustrated in Figure 1, which shows the extent of mining with the main ventilation paths superimposed. Early mining around the two shafts opened up vertical connections (stopes and raises) over the entire 2420 ft (737.6 m) of vertical extent, and mining progressed to the west primarily using compressed air and electric-powered equipment. Since ventilation was not a complex problem in the original mining system, the stopes and interconnecting raises were not sealed. The San Juan orebody was much larger than the silver ore zones mined previously, being primarily zinc and other base metals. Its geometry, size, and grade allowed the use of vertical crater retreat (VCR) stoping with diesel mucking and haulage equipment. Its depth, along with the existence of warm groundwater, resulted in a mine climate problem on the lower levels. To establish a complete ventilation circuit, two vertical boreholes (Bonanza Nos. 1 and 2) had been drilled by previous owners from the surface in the vicinity of the San Juan orebody. The system design called for air to be drawn down the intake shafts, across the lower mine levels to the San Juan workings, then up through the San Juan ramps and ore passes to these two exhaust boreholes (see Figure 1). In practice, however, only the Bonanza No. 1 borehole was drawing air through the desired path. Leakage across the old upper levels from the intake shafts, the Caliche tunnel, and from intervening abandoned stopes and raises supplied most of the air flowing to the base of the Bonanza No. 2 borehole. In effect, there were two ventilation circuits in semi-parallel through the mine, of which only one was delivering appreciable air to the San Juan workings.

Jan 1, 1993

By Jeffrey B. Smith

Introduction The gas shortage is going to be with us for some time to come. If we can set aside political and industry rhetoric (along with subjective personal opinions), we still are confronted by two serious "facts of life": (1) for almost a decade the U.S. has been consuming natural gas at a greater rate than we have been finding new reserves; and (2) there is a finite amount of natural gas present within the earth&apos;s crust. Much of the known and easily exploitable sources of gas (the so-called "conventional" sources, such as the high permeability sand reservoirs of the Tertiary sequence along the Gulf Coast) already have been developed; their production is declining rapidly. The total producible reserves from conventional gas reservoirs amount to only 216 Tcf, less than an 11-year supply. However, several large potential resources of natural gas remain to be developed. These "unconventional sources" have low permeability and/or peculiar producing characteristics. The DOE program for development of these unconventional sources of gas is called the enhanced gas recovery (EGR) program. The primary goal of this program is to provide a data base of resource characterization and production technology that will lead to commercial development. DOE will encourage and support industry participation in developing and demonstrating technologies needed to reach this goal. Unconventional Resources Four major unconventional resources of gas have a high potential for commercial development. There are other unconventional sources (such as gas hydrates) that are too poorly defined to warrant a major development thrust at this time. The four unconventional sources of gas currently included in the EGR program are: 1. The carbonaceous shales of Devonian age in the Appalachian, Illinois, and Michigan sedimentary basins are the targets of the Eastern Gas Shales Project (EGSP). 2. The low permeability, low porosity so-called "tight" gas sandstones of the Upper Cretaceous/Lower Tertiary in the Rocky Mountain areas constitute the resource target for the Western Gas Sands Project (WGSP). 3. The free methane trapped in coal beds of both the eastern and western U. S. constitute the Methane from Coal Beds Project (MCBP). 4. The abnormally high pressured, high-temperature saltwater aquifers of the Texas-Louisiana gulf coast are targets of the Geopressured Aquifer Project (GPAP). Basic implementation strategy for these EGR projects involve (1) assessing and characterizing the resource potential of the resource; (2) conducting cost-shared field testing with industry to improve, develop, and demonstrate various stimulation and production technologies; (3) coordinating EGR activities within DOE and with other federal agencies (such as the Bureau of Mines) to minimize duplication; and (4) aiming all projects toward commercial development of the gas resources. EGSP What type of "geological animal" is the EGSP dealing with? While gas undeniably is related to the occurrence of natural fracture systems within the shale, the overall producing mechanism and precise location of fractured, gas-bearing locales within each basin is still poorly understood. By developing reliable resource characterization techniques and applying effective stimulation technologies we intend to elevate the Devonian shale from the status of a potential gas resource to that of a proven gas reserve. Once we have done this the private sector can take over the large-scale commercial development of the Devonian shale gas resource. WGSP The second largest project (both in terms of complexity and level of funding) is the WGSP. The primary targets for this project are the low permeability (< 1 md) gas sandstones of the Piceance, Uintah, and Greater Green River basins and the Northern Great Plains Province. Project success in these four primary geologic locales will permit investigating additional low permeability sandstones in 16 other sedimentary basins. It appears that the only practical means of increasing permeability and resultant flow rates from these sandstones lies in the use of massive hydraulic fracturing techniques. Unfortunately, it is still too early to design such jobs with predictable results. MCBP The MCBP is to be involved in producing and utilizing methane derived from coal beds. The coal, like portions of the Devonian shale, is impermeable, highly fractured (termed "cleat" by

Jan 9, 1980

By Jeffrey B. Smith

Introduction The gas shortage is going to be with us for some time to come. If we can set aside political and industry rhetoric (along with subjective personal opinions), we still are confronted by two serious "facts of life": (1) for almost a decade the U.S. has been con¬suming natural gas at a greater rate than we have been finding new reserves; and (2) there is a finite amount of natural gas present within the earth&apos;s crust. Much of the known and easily exploitable sources of gas (the so-called "conventional" sources, such as the high permeability sand reservoirs of the Tertiary sequence along the Gulf Coast) already have been developed; their production is declining rapidly. The total producible reserves from con¬ventional gas reservoirs amount to only 216 Tcf, less than an 11-year supply. However, several large potential resources of natural gas remain to be developed. These "uncon¬ventional sources" have low permeability and/or peculiar producing characteristics. The DOE program for development of these unconventional sources of gas is called the enhanced gas recovery (EGR) program. The primary goal of this program is to provide a data base of resource characterization and production technology that will lead to commercial development. DOE will encourage and support in¬dustry participation in developing and demonstrating technologies needed to reach this goal. Unconventional Resources Four major unconventional resources of gas have a high potential for commercial development. There are other unconventional sources (such as gas hydrates) that are too poorly defined to warrant a major development thrust at this time. The four unconventional sources of gas currently included in the EGR program are: 1. The carbonaceous shales of Devonian age in the Appalachian, Illinois, and Michigan sedimentary basins are the targets of the Eastern Gas Shales Project (EGSP). 2. The low permeability, low porosity so-called "tight" gas sandstones of the Upper Cretaceous/Lower Tertiary in the Rocky Mountain areas constitute the resource target for the Western Gas Sands Project (WGSP). 3. The free methane trapped in coal beds of both the eastern and western U. S. constitute the Methane from Coal Beds Project (MCBP). 4. The abnormally high pressured, high-temperature saltwater aquifers of the Texas¬Louisiana gulf coast are targets of the Geopressured Aquifer Project (GPAP). Basic implementation strategy for these EGR projects involve (1) assessing and characterizing the resource potential of the resource; (2) conducting cost-shared field testing with industry to improve, develop, and demonstrate various stimulation and production technologies; (3) coordinating EGR activities within DOE and with other federal agencies (such as the Bureau of Mines) to minimize duplication; and (4) aiming all projects toward commercial development of the gas resources. EGSP What type of "geological animal" is the EGSP dealing with? While gas undeniably is related to the occurrence of natural fracture systems within the shale, the overall producing mechanism and precise location of fractured, gas¬bearing locales within each basin is still poorly understood. By developing reliable resource characterization techniques and applying effective stimulation technologies we intend to elevate the Devonian shale from the status of a potential gas resource to that of a proven gas reserve. Once we have done this the private sector can take over the large-scale commercial development of the Devonian shale gas resource. WGSP The second largest project (both in terms of complexity and level of funding) is the WGSP. The primary targets for this project are the low permeability (< 1 md) gas sandstones of the Piceance, Uintah, and Greater Green River basins and the Northern Great Plains Province. Project success in these four primary geologic locales will permit investigating additional low permeability sandstones in 16 other sedimentary basins. It ap¬pears that the only practical means of increasing permeability and resultant flow rates from these sandstones lies in the use of massive hydraulic fracturing techniques. Unfortunately, it is still too early to design such jobs with predictable results. MCBP The MCBP is to be involved in producing and utilizing methane derived from coal beds. The coal, like portions of the Devonian shale, is impermeable, highly fractured (termed "cleat" by

Jan 1, 1980

By Frank J. Foster

This Paper will deal with Bucketwheel Suction Dredging and its particular relationship to mining applications. There will be comparisons with other methods of suction dredging and descriptions of two recent bucketwheel installations in the mining area. The Directory of World Dredging Fleets published in the `larch 1984 edition of World Dredging magazine shows that there are about 1.000 conventional Cutterhead Dredges listed throughout the world with 289 in the United States. There are 242 Bucketline Dredges listed throughout the world of which only 6 are in the United States. These figures are not completely accurate because several units are known to the author that are not on the list and, of course, there may be others.

Jan 1, 1984

By Richard J. Fantel

The U.S. Bureau of Mines performed engineering and economic analyses on 130 domestic phosphate mines and deposits, all containing demonstrated resources, of which 73 are in the Southeastern states of Florida and North Carolina, to determine the domestic availability of phosphate rock. Of the total domestic resources studied (6.4 billion tons of recoverable rock) over 60 percent is found in these two Southeastern states. At an estimated cost of $30 per ton, in January 1981 dollars, approximately 1 billion tons of rock is potentially recoverable from Southeastern deposits, increasing to 3 billion tons at $45. In addition, there is nearly 6 billion tons of recoverable rock at the inferred level (some presently unmineable) in Florida and Georgia.

Jan 1, 1983