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By James S. Goff, Stephen J. Navin, Wesley W. Moulton

The White River Shale Oil Corporation (WRSOC) is developing a prototype oil shale facility on Federal Lease Tracts Ua and Ub located in northeastern Utah. This paper covers the successful driving of 1,525 m (5,000 ft) of dogleg downgrade decline with a 70-ton Paurat Roadheader mining machine and a conveyor muck-handling system. Ground support is with rock bolt and shotcrete; welded wire mesh is optional. The Upper Decline of 3.89 m by 8.69 m (12.75 ft by 28.5 ft) was mined through the Uinta Formation sandstones and shales in 210 days. The sandstones have a compressive strength in the 48 MPa (7,000 psi) range with a few bands of shale up to 137.8 MPa (20,000 psi). Green River Formation dolomitic marlstone with some sandstone lenses was mined in the Lower Decline 3.89 m by 5.64 m (12.75 ft by 18.5 ft) in 175 days. The Green River marlstones are tough massive formations ranging up to 206.8 MPa (30,000 psi). Roadheaders for use on steep slopes must be modified to accommodate the particular problems associated with mining downdip. The paper concentrates on the application of Roadheaders to steep slopes and the problems of mining tough materials such as marlstone. The paper also shows the advantage of using the method of excavation on long drives.

Jan 1, 1985

Several years ago, North American mining companies began to realize they needed to expand their spheres to remain viable. This strategy has been successful, according to Robert W. Schafer. Today, companies from other continents have also recognized the need for an international view for long-term growth and viability, he said. Schafer is vice president for exploration with Kinross Gold. He was the program chairman and moderator of the keynote session at the 1997 SME annual meeting, held in Denver, CO, Feb. 23-27. The theme of the meeting was "Minerals - Serving the Global Community." A record attendance of more than 6,000 people participated in the meeting, which featured 415 exhibitors occupying 582 booths. The title of the keynote session was "Corporate Perspectives on a Global Minerals Industry." Each of the three keynote speakers discussed a different perspective of a global minerals industry.

Jan 1, 1997

By Owen E. Owens

INTRODUCTION This paper deals with the main components in exploring for and developing a gold ore deposit. It addresses geology, exploration drilling, ore reserves, and evaluation criteria in an integrated manner to present an overall picture. Gold forms such a small component of a gold ore body, that particular care should be taken in dealing with the data from a developing deposit. The paper lists the main gold ore deposit types and their character. Their differences, as well as differences to other types of ore deposits, are important to understanding exploration approaches, and problems of ore reserve definition. Gold exploration has enjoyed a great boom in recent years. Many new ore deposits have been developed but experience shows that definition of gold is not a simple straight forward process and this outline attempts to point out some of the important aspects of ore reserve definition including some pitfalls. GOLD DEPOSIT TYPES The Witwatersrand deposits in South Africa produce about 50% of the world's gold. The gold occurs as fine grains in thin beds within gently dipping quartz pebble fan conglomerates usually with associated pyrite grains. It is characterized by rather uniform distribution of gold along thin layers within the conglomerate.

Jan 1, 1988

By G. E. McCormack

The Development of Water Gel Explosives Water gels are the most recent development in the evolution of commercial explosives. They are replacing dynamite, having proven themselves to be excellent alternatives to this standard explosive. Water gels have steadily displaced dynamite over the past two decades. In 1959, they accounted for only 1 percent of the market they share with dynamite. By 1965, the share taken by water gels had grown to 21 percent. In 1970, it was more than 40 percent, and today it is nearly 70 percent. Black powder, the first widely used explosive, revolutionized mining and quarrying in its day. However, when nitroglycerin dynamite became available about a century ago, it rapidly displaced black powder because it provided greater energy for a given weight, much higher velocity for better rock shattering performance, and water resistance for easy use in wet conditions. During the past 50 years, ammonium nitrate has played a dramatically increasing role in explosives. First, it was used as an ingredient in dynamite, and then as the main component in nitrocarbonitrate blasting agents. About a quarter of a century ago, it began to find use in a simple and inexpensive mixture with fuel oil that has revolutionized the explosives industry and today fills about 80 percent of the requirements for explosives in the U.S., and certainly the bulk of blasting requirements worldwide. Water gels and dynamites share the remaining 20 percent. Water gel explosives based on ammonium nitrate have also been developed during the past quarter century. These explosives contain

Jan 1, 1979

By Keith G. Wallace, Gary K. Rogers

As a supplement to an analysis involving methodologies for the estimation of airflow resistance in shafts used for both ventilation and hoisting, a detailed shaft characterization survey was distributed worldwide. The survey provided detailed information which enabled verification of recommended velocities and established limitations on fill factors and relative velocities. This paper evaluates the information obtained for 37 shafts and presents a statistical evaluation of the results. Apart from airflow rates, conveyance rates, and physical characteristics, qualitative information regarding means of conveyance stability and velocity control techniques at unloading and loading points are also presented. Calculations of fill coefficient, free area velocity, and most importantly, relative velocity between the airflow and hoist, are conducted. These results are presented by frequency diagrams which illustrate the trends, as well as provide limitations for these parameters. The data illustrates that, commonly, the fill coefficient for two or more conveyances, does not exceed 30 %. A limit of 50% is suggested for a single conveyance shaft. A recommended free area velocity of 10 m/s (1,970 ft/min) for combined ventilation and hoisting shafts, is sup- ported by these results. Relative velocities of 30 m/s (5,905 ft/min) are most frequent and 50 m/s (9,840 ft/min) being a practical maximum.

Jan 1, 1987

By John E. Frost

Duval's (now Battle Mountain Gold Co.) copper and gold deposits in the Battle Mountain District in Nevada are in two subdistricts - Copper Basin and Copper Canyon - about 18 km (11miles) apart. Location of the two subdistricts is shown in Fig. 1. The two centers of mineralization have distinctive geologic and ore characteristics that required different innovative approaches to exploration and development. Nearly all of the copper from Copper Basin has been recovered by heap leaching. Because of relatively high gold and silver values in the Copper Canyon ores, a significant part of the sulfide copper ores has been recovered by a conventional flotation mill at Copper Canyon. Recently, this mill has been converted to the recovery of gold that is mined from zones peripheral to the copper deposits. Prior to Duval's entry, small amounts of copper, gold, silver, lead, zinc, and antimony had been mined in the district over a nearly 100 year period, beginning in 1866. The bulk of the $13 million total production to 1962 was derived from copper and gold lode ores and gold placers. Duval's interest in Battle Mountain-arose directly from a submittal by Asarco, which controlled key claims in both the Copper Basin and Copper Canyon subdistricts. Asarco acquired its holdings in the district in 1959 and undertook limited exploration until mid-1961. At that time, the company decided to dispose of its interests. Duval was invited to evaluate Asarco's properties and accepted. On Dec. 6 and 7, 1961, J.E. Frost, R.C. James, R.W. Sayers, and H.A. Schmitt, representing Duval, and K. Richard and E.W. New- man, representing Asarco, visited Copper Basin and Copper Canyon. Frost concluded that the properties were of considerable interest and prepared, by Jan. 12, 1962, a voluminous report recommending that Duval acquire and explore Asarco's properties. Furthermore, he recommended a large initial exploration program that included not only geologic mapping, geochemistry, and geophysics, but also the extensive drilling that was to involve, as the first phase of exploration, 49 drill holes in Copper Basin and 21 drill holes in Copper Canyon. Following property agreements between Duval and Asarco, and Duval and several other claim owners, geologic mapping, geochemical surveys, and drilling began in August 1962. Exploration drilling was completed in July 1964 and engineering studies in April 1965. Development of the Battle Mountain properties was authorized in July 1965, and production commenced in the first quarter of 1967. Because of the unusually high oxide copper content of the ore bodies, particularly in Copper Basin, a much larger than typical proportion of the copper in the deposits has been recovered by heap leaching. The generalized geology and ore zones of the Copper Basin are as shown in Fig. 2 and of Copper Canyon in Fig. 3.

Jan 1, 1991

By Donald E. Hulse

Selection of a block size for a computerized 3 dimensional ore body model is an exercise in compromises. Larger blocks decrease the use of computer resources and decrease the variance of the block estimates. Smaller blocks increase the resolution of the model and the selectivity which can be used in the planning, but increase the variance of estimates in the individual blocks. A common rule is that blocks can be as small as one fourth the average sample spacing, however this ignores the interrelationship between the block size, the sample spacing, and the variogram model. This paper will explore this relationship Kriging estimation variance will be used as a comparative measure of estimate quality. Graphical techniques will be used to show that there is a point where smaller block sizes cease to contribute to the resolution of the model. A method of evaluating the consequences of block size decisions will be examined.

Jan 1, 1992

By D. Vavadakis

Abandoned bauxite surface mines in the Parnassos area may be rehabilitated using bauxite residues as filler. The objective of this study was to estimate the percentage of precipitation that may infiltrate through the bauxite residues to the lower geological formations when they are disposed off in mine open pits. Two scenarios were investigated simulating: a) available field tests, and b) a demonstration test that was planned to be performed in an abandoned open pit close to the Aluminium of Greece plant. It was indicated that the annual infiltration from bauxite residues disposed off in the demonstration test for the most probable scenario is very low, equal to 13.17mm corresponding to 3.1% of the annual precipitation. Furthermore, it was estimated that the annual percolation from the field tests is approximately 4.4%, very close to the infiltration values measured in the field tests for the same period.

Jan 1, 2006

By M. Misra, J. Nanor

The removal and stability of arsenic over a wide pH range (3.0-10.0) from synthetic solutions and industrial waste effluents (including mining waste tailing slurry) can be accomplished with addition of impure lanthanum chloride alone or in conjunction with ferric sulfate. The long term stability of these correspond­ing precipitated compounds is critical from the disposal stand­point. In this paper test results of this new process for arsenic removal and stabilization from the above mentioned systems are presented. Also, precipitated arsenic compounds (lanthanum arsenate, lanthanum-ferric arsenate, ferric-lanthanum arsenate and ferric arsenate) from synthetic solutions have been charac­terized. The characterization technique used was XRD. XRD patterns which shows that the lanthanum arsenate and lan­thanum-ferric arsenate formed in acidic and alkaline pH at room temperature are crystalline. On the contrary, ferric arsenate and ferric-lanthanum arsenate compounds formed under the same conditions as lanthanum arsenate were amorphous.

Jan 1, 1998

By E. Simoncini, S. Ulgiati, M. Raugei

A thorough Life Cycle Assessment of ilmenite extraction and processing to yield synthetic rutile (ERMS/EARS process) was performed according to ISO norms 14040 and updates. The LCA includes the parallel application of selected environmental impact assessment methods (as illustrated in Ulgiati et al. [1]), the results of which are compared and discussed. A MI value of 16 kg of abiotic matter per kg of output rutile is calculated, which is largely determined by the amount of coal that is employed in the reduction and drying steps. The Gross Energy Requirement is 3.33 * 107 MJ/kg, a value which we found to be comparable with that previously calculated by us for alumina. Finally, the increase factor of the Emergy Intensity from the raw mineral to the marketed rutile is 3.19, which points to a comparatively medium-low requirement for indirect environmental support. GWP and AP were calculated to be respectively 1.89 kg(CO2eq)/kg and 12.7 g(SO2eq)/kg.

Jan 1, 2005

By Steven G. Vick

Not too long ago a mill superintendent, when asked where his tailings were being disposed, could simply shrug his shoulders and point to the open end of the thickener discharge pipe. This “out of sight, out of mind” approach to tailings disposal is a practice of the past. Today there are very few planned hardrock or energy-related mining facilities where tailings disposal has not been identified as the major area of impact and controversy. In fact, development of some ore bodies has been stymied altogether by lack of suitable sites or methods for tailings disposal. The bottom- Iine result is that tailings disposal often controls the permitting lead time and has a significant influence on construction cost (up to 2070 in some cases) of new minemill complexes. The following article, the first in a series of three, explores technical background and historical perspectives on siting and design of tailings impoundment facilities.

Jan 6, 1981

By B. Das, B. K. Mohapatra, K. S. Narasimhan, S. K. Bhaumik

The iron ore slimes (d80 ~ 50 µm) generated at Noamundi iron ore washing plant (TISCO), assaying 57% Fe, 4% Si02 and 8.3% A1203, have been subjected to hydrocyclone studies. The effects of different design and operating parameters have been studied to get a product suitable for subsequent use in the sinter mix. The studies indicate that it is possible to obtain a product containing 64% Fe, 1.4% Si02 and 3.5% A1203 with a recovery of 49% iron. The process involves classification followed by separation in a hydrocyclone. By this method, the overall yield of iron ore production in the plant is expected to increase by 8%. Microscopic examinations and X-ray diffraction studies were conducted to elucidate the mineralogy and liberation characteristics of Noamundi iron ore slimes. This paper deals with the findings of the investigations.

Jan 1, 1993

By A. Pavlides

The purpose of this study is to compare estimation methods for lignite reserves at the Amyndeo mine and to provide a map of the estimated thickness of the coal seams. The reserves are first estimated with the method of Voronoi diagrams, using a group of mirror points to close the polygons around the boundaries of the mine. The second estimation method is based on Ordinary Kriging using the Spartan variogram model. Finally, the reserves are estimated by the mean lignite thickness value of the drill-hole data Two data sets are investigated. The first one corresponds to an exhausted part of the mine. The derived reserves estimates are compared with the actual data obtained by the Public Power Corporation (PPC) to validate the analysis. The second data set extends over the entire mine, and it is used to derive an approximate estimate of the total lignite reserves. A geostatistical analysis of the total layer thickness is performed. An interpolated kriging map of the total thickness variation is produced accompanied by a map of the kriging uncertainty.

Jan 1, 2008

By Ren Bryce

The ninth annual Mineral & Metallurgical Processing Division (MPD) student poster contest was held in February at the SME Annual Conference & Expo in Minneapolis, MN. A total of 33 students presented their work at the contest. The competition began in January when the students submitted a one-page abstract for judging. In February, they prepared and presented a poster that explained their research. A panel of judges went to each student and listened as the students explained their research in five minutes. The students worked hard at their research and then set aside time to complete all the necessary steps to enter the contest. The MPD recognizes the effort required to enter and, as a result, provides monetary rewards to all participants. As in past years, there are separate levels of competition for undergraduates and graduates, and the prizes are designated accordingly. At the graduate level, first place receives $1,000, second place $500, third through seventh $200 each, and all others $125 each. At the undergraduate level, first place receives $1,000, second place $500, third place $250, and all others $125 each.

By J. Partyka

Respirable dust produced during blasting operations in underground mines exhibits both diffusion and convection properties of the airflow. However, the application of the mathematical model presently used to calculate the change in time in the concentration of blasting pollutants is mostly limited to a single drift and cannot be easily utilized for blasts taking place in stopes. This paper discusses an analytical solution based on considerations that enable us to reliably identify the initial condition. Such an approach may be used for any blasting area regardless of its position, its size, or the ventilation system in use during blasting. In addition, a test procedure using field data is established to determined those parameters not measured directly. As well, equations are derived to find the diffusion coefficient, and the position and size of a dust source from the field data. Mathematical considerations The convection-diffusion problem, for a one-dimensional case, is described by the following partial differential equation (PDE): ct. = Dcxx - Vcx + source - sink(1) where C = the concentration of a pollutant T = time V = the mean air velocity in x-direction D = the diffusion coefficient X = distance away from the pollutant source source = any additional pollutant source specified by an initial value problem sink = decay of a pollutant inside the region specified by the boundary conditions. Respirable dust produced during blasting is best considered as an initial value problem (IVP). To derive a mathematical model, the following assumptions are made: •At time zero (t = 0), the dust source is at point x = 0 of an infinite drift. •The problem is considered a one-dimensional case. •The mean velocity of air and the diffusion coefficient are constant. •There is no source or sink of dust other than specified at point x = 0 and time t = 0. We have to solve the following problem [PDE ct=Dcxx-Vcx-m<x...m;0<t<m(2)] ICc(x,0)=f(x)t=0, the initial condition Where c = dust concentration in mg/m3 V= the average mean velocity in a drift in m/s t= time in s D= the diffusion coefficient in m2 /S x = distance coordinate in m. We observe that in problem (2), we have moved the boundaries to[ -m and +m] in order to have an initial-value problem. We introduce new coordinates, s = t and z =x - Vt. To rewrite the PDE in terms of (z, s), we use the chain rule ct = czzt + csst = -Vcz + cs cx = czzx = cz cxx = (cz)x = czz. We now substitute our computed ct, cx. and cxx into the PDE to get new IVP in terms of z and s :

Jan 1, 1991

By C. A. Moseley, M. B. Kahle

The Gulf Coast Region has been the center of extensive lignite exploration and development activity in the past 10 years. Currently, 27.2 Mt (30 million st) of lignite are mined annually, the majority of which is uncovered by large walking draglines. Projections indicate that production could reach 125 Mt (138 million st) by 1990. As mining moves to deeper depths, removal of overburden becomes the decisive factor in mine planning and cost. The geology of a region plays a key role in the selection of an overburden removal system. While alternate stripping methods must be analyzed to determine the most economical and operationally sound method, so far bucketwheel excavator systems seem to be the best method for removing large volumes of soft, unconsolidated material that is found at projected mining depths in the Gulf Coast region. Geology of Gulf Coast Lignites Extent Distribution of the principal tertiary lignite-bearing deposits in the Gulf Coast Lignite province extend from the Texas-Mexico border to the Alabama-Georgia state line (Fig. 1). The lignite-bearing strata includes portions of eight states: Texas, Louisiana, Arkansas, Missouri, Kentucky, Tennessee, Mississippi, and Alabama. Lignites occur primarily in the Wilcox, Claiborne, and Jackson groups in the Gulf Coast. The uppermost strata of the Midway group in Mississippi and Alabama also contain significant lignite resources. Reserves Commercial lignite seams in the Gulf Coast Lignite province range from 0.6-3 m (2-10 ft) in thickness. The estimated lignite resources throughout the Gulf Coast in seams greater than 0.9 m (3 ft) in thickness and up to 61 m (200 ft) in depth currently are estimated at 32 Gt (35 billion st). Table 1 shows these resources by state. Quality Lignite quality can vary within individual mining areas and seams. However, trends in quality occur because of the depositional environment, age, and west-to-east geographical location. Better quality lignites occur in the Wilcox group in Texas and Louisiana. Poorer quality lignites occur in the northern regions of the Wilcox group and those of the geologically younger Claiborne and Jackson groups. Environmental Conditions Gulf Coast province environmental conditions are ideal for successful mine reclamation. The lignite occurs in flat to moderately rolling terrain, a topography easily reproduced from the spoil dumps or spoil rows developed during the mining operation. High to moderate rainfall facilitates rapid revegetation. Evidence from current mine reclamation programs and ongoing test plot studies indicate that mined lands can be reclaimed to reach premining productivities. Sometimes productivities are even greater than that of premining conditions. Existing Mining Methods Present Mines & Production There are seven active mines in the Gulf Coast province, all located in Texas (Fig. 2). Lignite

Jan 8, 1983

By Ignacio Salazar

Somebody sure gave me the wrong title. I can talk much more freely about the impact of the anti-nuclear movement on the DRILLING industry in New Mexico. In a word, it is being "clobbered." It doesn&apos;t seem like very long ago that my brothers and I were busy trying to find enough new equipment (and trying to talk bankers into lending us enough money) to keep up with drilling work we could get. Seemed like no matter how hard we worked, and how much money we borrowed, we were always one step behind where we wanted to be. We had so many holes to drill that all our crews were working three shifts a day, and over several states. The nation needed energy, and we were ready to help supply it. Exploration and development drilling jobs were all going full-blast. We had trouble keeping up with them. With luck, we figured we&apos;d be able to get all our equipment out of hock before all the necessary uranium mines were developed.

Jan 1, 1979

By William Tattersall

On behalf of my agency, the Mine Safety and Health Administration, I would like to add my appreciation that all of you could be here today. Previous dust symposia over the years have proved the value of close cooperation among universities, government agencies, and professional societies in advancing knowledge of dust research and development in the interest of miners and the industry. I&apos;ll just comment briefly on MSHA&apos;s part of the overall effort by all parties to control miners&apos; exposure to respirable coal dust and other airborne dusts. As you know, our agency&apos;s congressional mandate is to do what is necessary to assure that miners will have safe and healthful workplaces. In the area of coal mine health, this has meant joining with the mining community in a continuous campaign to keep respirable dust and quartz levels below applicable standards to prevent or reduce disabling coal workers&apos; pneumoconiosis and silicosis arising from excessive exposure to respirable quartz. Currently, our agency is processing about 120,000 coal dust samples taken by MSHA and company people during the year and about 8,000 samples taken from metal and nonmetal mining operations. Our dust program, like other MSHA programs, has benefited from increased computerization of records. The attack on CWP has had the highest priority of government and industry. The goal has been to sharply reduce social costs of suffering victims and their families as well as the associated future payments under the black lung benefits program. During the 1989 fiscal year, about 104,000 former miners received about one and a half billion dollars in benefits. Since the program began, nearly 25 billion dollars have been paid out. Much progress has been made in lowering dust levels in coal mines since the standard of two milligrams per meter of air was established in 1972. Before that time, there was a 4-to-10-percent probability of a miner developing coal workers&apos; pneumoconiosis at typical pre-1972 exposure levels. However, under the maximum respirable dust levels allowed today, this probability has been reduced to about 1 percent. Since average dust levels are lower today, we can expect fewer miners to develop CWP in the future. Out goal continues to be elimination of both coal workers&apos; pneumoconiosis and silicosis in both coal and nonmetal mining. In 1985, MSHA revised its coal mine quartz policy to allow mine operators to submit dust samples used in setting the quartz standard. Previously, only single MSHA samples were used. Even though less than half of the mine operators have taken advantage of this change, it has provided a larger population of samples for analysis then before. Also the policy allows operations that are on a reduced quartz standard to have the standard reviewed every six months. In underground coal mining, the job of roof bolter typically has higher than normal exposures to quartz. About 45 percent of respirable dust samples from roof bolting operations contain more than 5 percent quartz and therefore are put on a reduced quartz standard. One objective of the policy revision was to improve miners&apos; health by follow-up monitoring, and this is being done with roof bolting jobs. MSHA studies of surface coal mining operations have provided data showing that bulldozer operators and highwall rotary drill operators often have high exposure to quartz dust, and our agency is directing enforcement efforts to reducing this hazard. One of the most important responsibilities of MSHA&apos;s metal and nonmetal mine health division is the respirable dust sampling program. In the last few years, nearly 3 out of 10 samples have

Jan 1, 1991

By Jon J. Kellar

The adsorption of oleate at the surface of a fluorite single-crystal internal reflection element has been investigated using Fourier transform infra-red/internal reflection spectroscopy (FTIR/IRS). The application of this innovative experimental technique has added a new dimension to the investigation and understanding of collector adsorption phenomena. It is now possible, for the first time, to measure in-situ and in real time the adsorption density and bonding nature of collector species adsorbed at mineral surfaces. In this regard, oleate adsorption densities were calculated from spectral data, and the equilibrium adsorption isotherms were found to agree quite well with other adsorption isotherms reported in the literature. At low equilibrium oleate concentrations(<10-5 M), monolayer-type cover-age was observed in the chemisorbed region. It was found that in this region the adsorption density increases with an increase in temperature, which sup-ports experimental results from previous ex-situ FTIR studies and thermochemical studies in which it was shown that the chemisorption reaction was endothermic. At higher equilibrium oleate concentrations (>10-5 M), the adsorption density increases dramatically at the point of surface precipitation. The oleate adsorbed in the chemisorbed region has a distinctly different spectrum than the oleate adsorbed in the surface precipitation region of the adsorption isotherm. In the chemisorbed region, the oleate appears to be mono-coordinated at the fluorite surface.

Jan 1, 1989

By Ted D. Johnson

INTRODUCTION The tailing dam is part of the SOHIO L-Bar Uranium Mine and Mill located about 70.8 kilometers (44 miles) west of Albuquerque in northwestern New Mexico. The dam location is shown on Figure 1. The tailing dam is one of the last upstream method dams permitted in the industry. Design and construction aspects of the project, as well as operation and maintenance considerations, are reviewed herein. Since initial construction, SOHIO Western Mining Company and Woodward-Clyde Consultants have developed several modifications and improvements in the dam&apos;s operating and monitoring procedures. This work has been done in coordination with the New Mexico State Engineer&apos;s Office and in accordance with recommendations in a report done for the Corps of Engineer&apos;s Phase I Dam Inspection Report. Modifications and improvements described should be of particular interest to owners and operators of other tailing dams utilizing the upstream method of deposition. The tailing facility was designed for a 10-year life to provide an ultimate storage capacity of 6 1/2 billion kilograms (7 1/2 million short tons) in an area of about 180 acres. A plan view of the tailing dam is shown on Figure 2. The tailing consists of approximately 75% sand size material and 25% fines (material passing the No. 200 sieve). The tailing dam includes a small earth starter dam with raises of the dam accomplished using tailing deposited by upstream method techniques. The starter dam is a homogeneous embankment about 610 meters (2000 feet) long, a maximum height of 12 meters (40 feet) and 2 1/2:1 (horizontal to vertical) upstream and downstream slopes. The starter dam contains an internal drainage system which includes a near vertical chimney drain connected to horizontal finger drains. A typical section through the maximum height of the starter dam is shown on Figure 3. A small confining dam, approximately 1.6 meters (5 feet) high, is constructed beyond the limits

Jan 1, 1980