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By R. M. G. Machunter

Computer models of metallurgical circuits are extensively utilised in design, and prediction of overall plant separation performance. As always, in the final optimised flowsheet design there is a compromise between the recovery of values and grade of the circuit output. This paper mathematically examines various circuit options for the generation of a heavy mineral concentrate from a generic mineral sands wet concentrator plant. Based on data derived from testwork, any particular mineral sand has a unique release curve describing the separation performance of the heavy mineral suite. Models are constructed that combine particular stage output streams as feed to other stages in various configurations. The ramifications of each option are discussed and reasons why one particular option is selected over the others examined.

Jan 1, 2005

By H. Lyn Bourne

This is the fifth annual tabulation of mineral royalty information. This tabulation, like the previous editions, includes all of the earlier data. There are also 60 new entries. In addition to royalty information, some entries show the cost of purchased reserves. Two articles relative to the evaluation of minerals reserves appeared in December, 1986. R.H. Paschall (In the Aggregate, ME, p 1139) wrote a good article, although it lacked specific royalty data. However, this author takes exception to two ideas in Paschall's column. Paschall contends that the royalty rate does not affect the mineral appraisal. This present author could not disagree more. Paschall also claims the location of an industrial mineral deposit is the only serious or practical caveat in comparing royalties. Comments received here from readers helped make the list of caveats that are cited each year. The readers will decide which factors influence the mineral value for their commodities of interest. Paschall's point about the relationship of the parties who are negotiating a royalty agreement is important and very well taken. The other article, written by William Vernon (Pit and Quarry, December 1986), gives a step-by-step process on how to conduct a mineral appraisal. The article cites three reasons for an appraisal: transfer rights; collateral for a loan; and taxation. A fourth reason is to establish adequate compensation for reserves lost through condemnation. This author has appeared in several court cases where compensation for loss of mineral reserves through condemnation required a mineral appraisal. Royalty rate becomes an important factor in the appraisal process, although the Pit and Quarry article refers to that part of the valuation equation as "profitability." Both articles were well written and are recommended reading for those interested in mineral valuations. The same caveats that have appeared in previous columns relative to factors that influence the value of reserves are restated. The data in the table serve as a general reference and are not intended to present an absolute scale of reserves costs. Several factors influence the value of any mineral commodity. The factors include: • location/transportation - especially significant for most industrial minerals; • market conditions both for the commodity and the company acquiring the right to mine; • quality or grade of the deposit as it reflects the amount of processing/beneficiation necessary to produce the finished product; and • legislative restrictions and conditions for mining a given material at a given site. These and other factors will significantly influence the cost of reserves. The table shows the range of royalties and the variations in the basis of computing the fees. In a few instances, it gives the cost of acquired reserves based on purchase price and quantitative estimates of the material in place. The table lists the commodity in column 1, and the subsequent columns give: • the location by state, US geographic area, or Canadian province ; • the cost - R for royalty and P for purchased cost; • the year in which the lease or purchase agreement occurred; and • comments or footnotes for more information. The cost column requires more explanation. The units may refer to cubic yards (/cu yd), cubic feet (/cu ft), short tons (/T), or acres (/A). 1 cu yd = 0.76 M2 1cuft = 0.28m3 1 st = 0.9072 t 1 acre - 0.4047 ha or m2

Jan 7, 1987

By David D. DiPonio, John DiPonio, Roger Van Omen

The greatest opportunity to attain the most favorable prices for underground construction projects occurs during the competitive bidding period. The Owner, Engineer, and Contractor need to have a thorough understanding of the project’s design concept, how it is going to be built, and what it is going to cost. This paper presents the opinions of three engineering professionals who are well established in their sectors of heavy civil underground public works construction. This paper will outline and summarize the considerations that the three principal contracting parties of Owner, Engineer, and Contractor need to make to obtain good price quotes at the time of bids.

Jan 1, 2005

By understanding the ore breakage characteristics and utilising comminution circuit modelling, the benefits of SAG mill feed size manipulation can be readily assessed. This paper looks at the benefits and problems of total secondary crushing of mill feed and advocates partial secondary crushing as a means of improving grinding efficiency for certain hard ores. Case studies are reviewed for new and existing SAG mill circuits.

Jan 1, 2007

By I. R. Firth, D. L. Taylor

The ability to accurately model the movement of grade boundaries during a typical bench blast would be of significant benefit to mine operators. The movement of grade boundaries during the blasting process may have significant impact on grade recovery if not accounted for. The paper describes the preliminary stages of a research project which combines numerical blast models and conventional mine planning software to allow practical use of the results of a numerical blast simulation. A two dimensional bench blast is simulated using the Universal Distinct Element Code (UDEC). The results of the simulation are used to provide input to the Surpac mine planning package which can be used to generate grade control plans.

Jan 1, 2001

By Syd S. Peng, C. Y. Chen

Introduction The Surface Mining Control and Reclamation Act was passed in 1977 to put an end to the uneven environmental standards and enforcement established by various coal mining states in the US. One year after passage, the Secretary of Interior was directed to promulgate permanent program regulations implementing Title V provisions of the Act. These regulations were to form the basis for state regulatory programs under the Act and to allow states to assume primary regulatory responsibility. To implement the underground mining provision of the Act, the Secretary is directed in section 516(a) of the Act to promulgate rules. But they must take into consideration the differences between surface and underground coal mining. The regulations developed by the Secretary in 1979, with respect to subsidence control, included: • a subsidence control plan as part of the permitting process; • substantive provisions establishing performance standards for subsidence control; and • provisions providing for the prohibition of mining under perennial streams, impoundments larger than 24.6 dam3 (20 acre ft), and certain public facilities and urban areas. Industry challenged the Secretary's regulations implementing the subsidence control provisions. The litigation was consolidated with all other challenges to the permanent program regulations. (Permanent Surface Mining Regulation Litigation, Civil Action No. 79-1144; D. D.C., May 16, 1980.) In two opinions dealing with the multitude of issues raised by both industry and environmental groups, the US District Court of the District of Columbia ruled against industry. It upheld the promulgated subsidence control provisions. The US Court of Appeals stayed the appeals pending the agency's review and revision of its regulations. The regulatory reform was conducted in accordance with the order from the Secretary of Interior and the President in 1981. As a result, the subsidence control regulation was revised, reproposed, and finalized in June 1983 (Federal Register). This regulation, along with others, was again challenged by environmental groups and industry. On Oct. 1, 1984, the court, after finding the Secretary's reading of the Act reasonable, upheld the 1983 subsidence regulation except for one section. On Feb. 21, 1985, in accordance with the court's ruling, the Office of Surface Mining (OSM) suspended a portion of 30 CFR 817.121 (c)(2) concerning the damage correction requirement for structures or facilities and reproposed it on July 8, 1985 (Federal Register). In conjunction with a request for public comments, OSM now also seeks comments on the omission from the 1983 rule of the requirement in the 1979 rule regarding a presubsidence survey and monitoring of damage to structures. Technical issues related to subsidence regulations The subsidence regulations of June 1, 1983, consist mainly of two parts: a subsidence control plan as part of the permitting process; and substantive provisions establishing performance standards. The latter also includes provisions for prohibiting mining under certain areas. Under subsidence control plan section 784.20, minimum require-

Jan 2, 1986

The HB 2500 hydraulic breaker attachment from Atlas Copco Construction Tools features Krupp percussion technology and a service weight of 2,500 kg (5,512 lbs). This makes it ideal for demolition, tunneling, trenching, foundation work, rock mining and secondary breaking. Suitable for carriers in the 29-to 43-t (32-to 47-st) weight class, the HB 2500 accepts maximum hydraulic flow of 219 L/min (58 gpm) at 18 MPa (2,610 psi) of pressure, and delivers an impact rate of up to 550 blows per minute. The HB 2500 features PowerAdapt, a new system that automatically shuts down the breaker in the event that it receives excessive oil pressure from the carrier.

Jan 1, 2006

By James A. Madonna

Several universities and colleges in Australia, Canada and the United States, recognized the need to establish or reorganize geoscience programs at the vocational and two-year technical levels. To accomplish this, a cooperative study with industry in one state or province from each country was initiated. These include Tasmania in Australia, British Columbia in Canada and Alaska in the United States. The objective of this work was to develop a diversified geoscience educational model at the two-year technical, vocational and continuing education level that provides the source from which any nation can draw the proper educational tools to create a new geoscience program, modernize an existing program or strengthen a weak program. More than 100 mineral industry organizations provided information regarding their specific training needs for support personnel. The data gathered have been used to identify training areas for semi-skilled, skilled and technical personnel in geochemical mineral exploration. These included geological, geochemical, environmental and computer skills. With the areas of major educational need defined for geochemical exploration, an educational field skills model was developed for the three levels of training. The international need for a study of this type was recognized as early as 1980, following numerous contacts from domestic and foreign geoscience educators regarding the structure and success of training within the University of Alaska Fairbanks Mining Extension program, located in Fairbanks. Contacts, correspondence and visits included educators from Australia, Brazil, Canada, Mozambique and the conterminous United States. The need was magnified between 1985 and 1989 at the University of Alaska Fairbanks, when an attempt failed at initiating and developing a two-year geoscience vocational-technology program leading to an Associate of Science degree.

Jan 1, 1994

By Michael Karmis

The International Conference on Sustainable Development Indicators in the Mineral Industries(SDIMI 2003) was held May 21-23, 2003 on the island of Milos, Greece. The objective of this international conference was to assist the mining and minerals industries in their global transition to sustainable development. The main theme of the conference was the development, monitoring and assessment of sustainable development criteria for mineral operations. Emerging technologies were also presented that can assist in the development of criteria and performance indicators.

Jan 1, 2003

By F. Djahanguiri

The U.S. Bureau of Mines is conducting research on underground stope leaching of metalliferous ores to minimize the environmental impact of ore processing on the surface and to maximize minable reserves. An aspect of this research is to minimize pregnant solution losses and/or groundwater contamination by leach solutions. Bulkhead design and construction are important issues for solution control and stope leaching economics. A reinforced concrete bulkhead was designed, constructed, and tested for solution control in an underground mine near Idaho Springs, CO. This paper presents the bulkhead design methodology and construction procedure.

Jan 1, 1995

By Timothy D. Arnold

Barrick Gold’s Bullfrog underground mine, near Beatty, NV, was established as an after-thought to the main open pit. As open pit ore dwindles, the underground mining became increasingly more important as an ore source, and to the overall economics of the project. This situation is similar to many other mines. Today, there is a move towards underground mining in the United States. There are inherent problems in starting and operating underground mines in the shadow of any open-pit operation. Several attempts were made at the Bullfrog Mine for productive increases. Some were successful, while others were not.

Jan 1, 1996

By T. Yegulalp, R. Farrauto, A. Belova, M. Castaldi

Calcium oxide, with its ability to capture and store carbon dioxide, is emerging to become an important tool in carbon management. Calcium oxide can be used reversibly to capture carbon dioxide in situ and then be regenerated to produce a pure stream of carbon dioxide ready for sequestration; the Zero Emission Coal Alliance technology utilizes this property of calcium oxide to generate emissions-free electric power from coal. The reaction of calcium oxide with carbon dioxide resulting in thermodynamically stable carbonate is well understood; however, the efficiency of this reaction is limited by the diffusion of gases through bulk calcium oxide and through a layer of carbonate formed on the surface. In addition, the regeneration of calcium oxide is an endothermic reaction and is favored at high temperatures. The purpose of this paper is twofold: first, to address the diffusion limitation, a thin layer of calcium oxide, supported on a high surface area material, is studied thermogravimetrically to demonstrate the physical adsorption of carbon dioxide on calcium oxide. The adsorption is to take place at temperatures lower than required to form calcium carbonate.

Jan 1, 2007

By M. A. D. Azevedo, J. D. Miller, R. Kobler, E. Mutkowska

Up to 3 million tons of automotive shredder residue (ASR) is generated annually in United States. Plastics are the most valuable component of the ASR material making up approximately 25% of the total weight. In order to reuse these plastics, a selective process is required to separate each plastic, with very high purity and favorable economics. One separation technique, which may be able to selectively recover the plastics from ASR, is flotation. The preliminary flotation results have demonstrated that acrylonitrile-butadiene-styrene (ABS) and polycarbonate (PC) can be selectively separated from the model ASR feed, the PC/ABS/PA 66 system, by appropriate control of the surface chemistry.

Jan 1, 2001

By Warder A. Shaver

INTRODUCTION Mine No. 95 of the Consolidation Coal Co., a divi¬sion of Continental Oil Co., is located in Shinnston, WV. Mine No. 95 was opened in January 1967 and has been operating continuously since that time. The mine has 544 hourly and salaried employees. In 1974 this mine was the largest coal producer in West Virginia. Opera¬tions include two longwall sections and eight Marietta Borer continuous miners. Originally there were also two Jeffrey Heliminer continuous miners, but use of these was discontinued in June 1978. Production averages 12 727 t/d (14,000 stpd). Two portals now service the mine with a third portal proposed for 1978. GEOLOGY Mining takes place in the Pittsburgh coal seam which dips 0.04 rad (21h°) northwest. The working seam is steam coal with an average height of 2.2 m (86 in.). Immediately above the seam, the overburden is com¬posed of 6.1 m (20 ft) of shales followed by 6.1 m (20 ft) of weak sandstone and 51.8 (170 ft) of shaley limestone. The shale weathers when exposed to mois¬ture. Surface water has been a problem due to a fire¬clay floor which turns to paste when mixed with water. GENERAL DESCRIPTION The present mining plan for Consol's No. 95 mine has a proposed life of 35 to 40 years. One set of main headings has been driven in the north direction and one set in a western direction. These are the areas in which the majority of the virgin coal lies. The panels are driven off of the mains at 2134 to 3048-m (7,000 to 10,000-ft) intervals. The continuous mining sections are set up 133 x 610 m (600 x 2000 ft). Longwall sections are 259 x 1220 m (850 x 7000 ft). All of these mining sections are retreating uphill. Development is with continuous miners. A Marietta Borer miner is used for the main development work. These miners cut 4.1 m (13 ft 6 in.) wide and 2.1 m (7 ft) high, leaving an arch cut. Along with the arch, a 152.4-mm (6-in.) thick layer of roof coal is left in place to protect the roof from weathering. The combi¬nation of the self-supporting arch and the roof coal al¬lows the entries to be driven 30.5 m (100 ft) without bolting. Full bolting, however, is carried outby the last open crosscut. This bolting is done on 1.5 x 1.5-m (5 x 5-ft) centers using 1.5-m (5-ft) long roof bolts with tandem anchors and 152.4-mm (6-in.) steel washer plates. When needed, crib blocks with a 50.4-mm (2-in.) drill hole are used as supplemental supports. Three-hole steel channel with the ends curved down¬ward to conform to the borer arch may also be used. In the mining sections, a 14 BU loader is used be¬hind the Marietta miner to load two Torkar shuttle cars. These haul to Long-Airdox feeder breakers which load onto a 1.07-m (42-in.) wide conveyor belt. With the exception of the shuttle cars, which use 275-V d-c power, the face equipment operates on 550-V a-c power. The main power line is 7200 V ac. The same loading and hauling equipment is used with the Marietta Borer miner for driving both panel and development entries. In the retreat sections, the same equipment with the exception of the miner is utilized. Because of the advantage of a variable cutting hight, Jeffrey Heliminer ripper-type continuous miners were employed in these retreat sections. For longwall mining the first unit uses an Eickoff EDW 300-L double-drum ranging shearer. Drum di¬ameter is 1.69 m (66 in.) with a drum rotation speed of 45 rpm. The drum is equipped with 48 Carmet 34-L-1 bits. The 300-kW (400-hp) motor operates on 1000 V. Tramming of the shearer head is done through a hydraulically driven 26-mm wide chain. Maximum cutting speed is 7.93 m/min (26 fpm), and maximum tram speed is 12.8 m/min (42 fpm) with an average cutting speed of 3.66 m/min (12 fpm). Web cutting depth is 762 mm (30 in.) and the cutting height is 2.13 m (7 ft). The conveyor width is 686 mm (27 in.). It uses a single chain, 26 mm thick, with flights on 0.91-m (3-ft) spacing. The face conveyor speed is 64.3 m/min (211 fpm) when driven by two 112-kW (150-hp) 1000-V Louis Allis motors through torque convertors. The roof support used is Gullick Dobson chocks [6-legged 77.3 t (85 st) per leg, 463.6 t (510 st) total capacity] with double-acting leg jacks. The hydraulic pressure for the legs is 13.8 MPa (2000 psi), and the hydraulic pressure for rams is 6.8 MPa (1000 psi). The yield valves are set at 57.8 MPa (8500 psi). Each chock weighs 4.09 t (4.5 st). The stage loader is an Eickoff drive, with a 68-kW (90-hp), (00-V Reliance motor. A 26-mm thick single strand chain with flights on 0.6-m (2-ft) centers runs at 75.89 m/min (249 fpm). The stage loader is self-advancing and is connected to the belt tail piece with a special fifth wheel attachment so that the stage loader can advance and guide the tail piece. The hydraulic pressure for the roof support is supplied by two 45-kW (60-hp) Reliance motors op¬erating on 699 V, with 1.26 L/s (20 gpm) constant volume Gullick hydraulic pumps. Each pump services alternate chocks, with the two pressure lines connected at the tailgate by 25.4-mm (1-in.) hydraulic hoses, rated at 27.21 MPa (4000 psi). The 6.8 MPa (1000 psi) pressure rams advance is supplied by a 45-kW (60-hp), 600-V Reliance motor and a 1.26-L/s (20-gpm) con¬stant volume Gullick pump. Return lines are 35.5-mm (1.4-in.) diam hydraulic hoses with rated capacity of 27.21 MPa (4000 psi). The fluid used in the hydraulic system consists of 95% water and 5% Hy-chock fire resistant water soluble oil. Water is stored and mixed in a 6096-L (2000-gal) track-mounted tank and goes to a 655-L (180-gal) reservoir at the pump. The belt conveyor structure is 1.07 m (42 in.) wide with 0.47 rad (27°) idlers, bottom stands located on 3.05-m (10-ft) centers, top stands on 1.5-m (5-ft) centers, and is supported on a 19-mm (3/4-in.) diam

Jan 1, 1982

By Manuel Pérez-Tello, María M. Salazar-Campoy, Óscar Rodríguez-Hoyos

A common problem during the operation of the Kaldo furnace is the formation of accretions and blockage of the injection lance due to splashing. This typically occurs during the processing of anodic slimes for the production of doré. In this work, a 1:7 scale model of the Kaldo furnace currently in operation at Metalúrgica de Cobre, Mexico, was built in transparent acrylic plastic. Water and colored oil were used to represent the doré and slag, respectively. The model was used to determine the maximum penetration distance of the injection lance under a variety of experimental conditions at room temperature. Input variables included two types of oil, rotation speed, inclination angle, and the fraction of the vessel volume occupied by the liquids (φ). A mathematical correlation for the dimensionless maximum penetration distance of the injection lance d ∗ as a function of φ, the Froude number of the liquids in the vessel (Fr), the Reynolds number of the gas jet (Re), and the oil-to-water kinematic viscosity ratio (η)was developed. Reasonable agreement between the calculated and experimental values was obtained. Digital images showed that splashing was caused by the turbulent gas jet impinging upon the free surface of the liquids, whereas the cascade effect was negligible. The results of this work have been useful to improve the operation of an industrial Kaldo reactor.

By Bahram Sameit

Blast engineers have been taking advantage of image-based rock fragmentation analysis methods for the past three decades to estimate rock size distributions. These methods utilize a range of image segmentation techniques, from manual to fully-automated, to delineate the edges of rocks in images. Rock delineation results are then converted to physical rock sizes (and percentage passing curves) by scaling the image pixels to correspond to physical measurements. The size scaling process is critical for acquiring physical measurements, but specifc to the limitations of different applications. This article starts by presenting a shovel bucket-based fragmentation analysis system, where scaling is done using the known width of the shovel bucket within the image. It will then describe methods that use pictures captured by digital cameras with scaling reference objects in the view. Finally, the article presents a method that uses three-dimensional imaging to capture images without the need for scaling reference objects.

By Xiaoting Li, Richard Adasiak, Benjamin Mirabile

Pumpable J-Crib stopping walls and Jennmar high-strength (HS) stopping panel walls are widely used for resisting high-ventilation air pressures and roof support in underground mines due to their installation efficiency, convergence allowance, and better performance than traditional concrete block stopping walls. Based on the different entry dimensions and required maximum ventilation air pressures of up to 36-in. water gauge, 24-, 27-, 30-, and 36-in.-diameter pumpable J-Crib stopping walls, or 18- to 30-in. thick Jennmar HS stopping panel walls can be designed, and their support capacities and convergence allowances are evaluated for better roof control. Pumpable J-Crib stopping walls consist of 0.187-in.-diameter steel wire reinforced, JCrib material-filled bags, and non-wire-reinforced filler bags in between. Jennmar HS stopping panel walls consist of two, 20-gauge (0.04-in.-thick), galvanized, cold-formed steel panels with J-Seal materials pumped in between. Based on the fundamental structural design principles and specifications of ACI 318-19(22), the compression and flexural strength, and shear resistance of the walls can be checked to be compliant with the codes and specifications following the load and resistance factor design (LRFD) method. Specifically, the 0.187- in.-diameter steel wires in the pumpable J-Crib walls and 20-gauge steel panels in the HS stopping walls are integral parts of the walls and are considered to effectively confine the J-Crib or J-Seal materials. These steel elements play a significant role in reinforcing and improving the strength or support capacity of the stopping walls from a structural design perspective. Because of the high ductility of low-density J-Crib/J-Seal materials and steel elements, these walls are good at accommodating or absorbing the roof convergence (≥5% entry height), which cannot be matched by a traditional concrete block stopping wall (Figure 1). Based on the successful design and application of these stopping walls in different mines for improving roof stability and avoiding the air leakage, the authors have formulated the design methodology and verified independently by the finite element analysis (FEA) simulation technique. This design methodology combines the fundamental structural design principles with specific wall dimensions and material properties, facilitating mine engineers and government agencies to understand the in-depth working mechanism of these stopping walls.

Jul 1, 2023

The support of the roof in coal mines remains a critical safety issue with 3 fatalities from roof falls in 2007 and 430 injuries from roof falls in 2006. A snapshot survey by the authors of room design and support details as well as unplanned falls of ground in the industry revealed some useful data concerning support use, seam details, fall dimensions, bolt lengths and spacing and the stand-up time.

Jan 1, 2009

By W. L. Hollinshead, Alwin Poltinger

GENERAL Tunneling by hand digging is as old as the history of humans. Mechanized tunneling did not begin to replace manual methods until around the beginning of the 19th century(Brunel patent, 1818). It brought with it a higher degree of safety for the workers and a much higher rate of boring. But for any kind of tunneling, an accurate and feasible method of survey must be accomplished during mining. That means, that for each tunnel project the most efficient survey method must be devised. The method to be defined depends on the spatial curve, diameter, speed of tunneling method, purpose of tunnel and contractual accuracy. The final accuracy of a finished tunnel is heavily dependent upon three critical elements: 1. Quality of the above ground survey 2. Tunnel traverse and directional transfer from above ground 3. Guidance for the mining job. Combining The Survey And The Guidance System Since the guidance system (GS) is one of the critical factors in the determination of the tunnel final accuracy, it is necessary to analyze how the directional control and tunnel traverse measurements can be effectively integrated into a GS. Any guidance system, whether manual or high technology, must continually detect and compensate for movement of its theodoliteor laser station standpoint. Movement beyond a preset allowable amount must generate a warning or error message to the user. Additionally, optical refraction can interfere with the results. Refraction will have an effect on any kind of optical GS. Therefore, careful selection of target and standpoint locations can minimize, but not completely eliminate, the influence of refraction. Some GS technologies, such as Gyro-based systems, are immune to the effects of refraction, but have other, methodical errors which will be described later in this paper. GUIDANCE Overview Why is a guidance system necessary? A GS, whether manual or automatic, compares the design direction and slope of a tunnel with the current position of the mining

Jan 1, 1997

By L. Lacerda, C. Rose, E. Zellers

Approximately six hundred air sample measurements were taken at several U.S. underground metal/nonmetal mines in 1998 and 1999. This air sampling effort was organized by the Nevada Mining Association to evaluate mining conditions and to evaluate the Mine Safety and Health Administration’s (MSHA) Proposed Rule on Diesel Particulate Matter Exposure of Underground Metal and Non-metal Miners. MSHA’s proposed air sampling method, NIOSH Method 5040, was tested under controlled conditions and under normal mining conditions. Unfortunately, NvMA sampling results have indicated that carbonaceous rock dust, oil mists and even cigarette smoking create severe sampling interference problems for NIOSH 5040.The main conclusion was that NIOSH 5040 did not distinguish between DPM and other airborne carbon particulate. Thus, it is not a reliable method of accurately and unequivocally measuring DPM in underground metal/nonmetal mines.

Jan 1, 2000