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By J. R. Allan

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

Jan 1, 1996

By Luther J. Russell

Mineral development in South Dakota is aggressively encouraged by both state statute and county ordinances. Yet, an increasingly conservative philosophy is taking hold, resulting in expanded regulatory controls and concomitant increases in direct and indirect costs to permit and operate a mining operation. This paper examines the hidden costs associated with protracted permitting timeframes, land buy-outs, attorney's fees, etc., associated with the increasing emphasis for stiff controls on mineral development in South Dakota.

Jan 1, 1998

By M. Liu

In-situ soil washing using surfactants shows great potential to remove hydrophobic organics from subsurface environment. Lab scale soil washing experiment were performed using Sodium Dodecylsulfate (SDS), an anionic surfactant, and anthracene, a highly hydrophobic compound. The results of this research show that solubility of anthracene increases linearly with an increase in SDS concentration above the critical micelle concentration. Furthermore, SDS can be effective in removing anthracene from contaminated sand.

Jan 1, 1992

By David Brown, Adrian St. John, John Comins, Mike King

"INTRODUCTION The Crossrail project provides an East-West link across London enabling an additional 200 million journeys each year linking up major business centres and existing transportation interchanges, while reducing congestion at street level and within the current underground system. The project has built major structures in Sprayed Concrete Linings (SCL), including 5 new stations, 6 crossover caverns and numerous passages, adits and ventilation tunnels. Construction has taken approximately 4 years, and has involved the spraying of over 200,000m3 of concrete. At the start of the project there was a lack of available records from previous projects concerning concrete falls, which hindered risk assessments and the ability to make meaningful comparisons around spraying or material performance. This paper redresses this lack of data by providing information on falls experienced on this project and makes recommendations to assist future projects. Construction Methods Table 1 below summarises the main contracts with elements of permanent SCL. The listed primary lining lengths include pilot tunnels, side drift tunnels and enlargements as separate tunnels.A range of construction sequences were used in order to deal with the design requirements, geometry, and the ground and ground water conditions. The following briefly summarises the build up by layers, with typical design dimensions provided for a platform sized tunnel (approximately 9m internal diameter), excluding tolerance allowances. Some linings were sprayed in more than one layer to allow for inclusion of reinforcement layers or to avoid a too aggressive build-up of thickness: •,Initial shotcrete, or sealing, layer (ISL) of 75mm followed by a Primary Lining (PL) of 300mm. •,Regulating layer of 40mm followed by a Waterproof membrane (Sprayed membrane, 3-6mm). •,Secondary lining (SL) of 250mm with a separate sprayed fire protection layer (FPL) of 50mm."

Jan 1, 2016

By Behzad Khorshidi

"The ultimate objectives of the post-tensioned strands are to connect the precast segments into a ring and provide pre-stress in the tunnel circumference. The method involves inserting the strand through the preliminary embedded duct in the precast concrete segments and then tensioning the strand by jack and fastening it. The post-tensioned strand eliminates the necessity of bolt connections between the segments and may reduce the labor cost and increase speed of lining construction. Pre-stressing of the segmental lining reduces the steel reinforcement or fiber volume in precast segments and/or thickness of the precast segments which would result in reducing the manufacturing cost of them. Pre-stressing would reduce cracking and increase water tightness, and would provide smooth intrados due to eliminated bolt joints and pockets in the tunnel lining. Pre-stressing would increase resistance to internal pressures as well. Further, this system could work very well structurally in tunnels with squeezing rock or tunnels with high seismic loadings. INTRODUCTION OF THE METHODPost-tensioning is commonly used in the construction of bridges, large buildings, nuclear power stations, and water tanks but its utilization in tunnels is still in its infancy.The precast pre-stressed concrete segmental lining can be used for shield tunnels. It provides a reinforced lining ring by the following construction stages and as shown in Figure 1:1) Segments to be assembled by TBM to complete a full ring2) Steel strand to be inserted through embedded duct (sheath) within the segments3) The strand to be tensioned by Jack after anchorage of the other end4) The strand to be fastened after reaching required tension load5) The horizontal strands (if necessary) to be installed, tensioned and fastened6) The recesses and gap between strand and duct to be filled by groutingFurther, Figure 2 shows an unbonded post-tensioning strand model; Figure 3 shows an anchor sample (Xanchor); Figure 4 shows a duct (sheath) and anchor within reinforcement; Figure 5 depicts the posttensioning of a strand by jack; and Figure 6 is an image of the Midorigaoka Trunk line tunnel in Sagamihara, Japan which was constructed using this system."

Jan 1, 2016

By S. K. Sharma

The Bureau of Mines' rapid method of dewatering fine-particulate slurries uses extremely high molecular weight polymers. These polymers form large, strong flocs which can be dewatered by sieving or sedimentation. However, sometimes these flocs deteriorate rapidly if water is not removed immediately. During field tests, clay slurries have been encountered that could have been dewatered economically if it were not for the rapid floc deterioration. A kinetics study was conducted to determine the rate of kaolin-PED floc degradation in which polymer addition, to bring the flocs back to their original condition, was used to determine the extent of floc decay. Results showed that the amount of polymer added, to replace polymer lost by spreading, was less for the lower PEO concentrations. For this clay-polymer system, floc breakage as judged by polymer additions followed first order kinetics. It was also found that the addition of fluoride ion stabilized the kaolin flocs.

Jan 1, 1992

By W. E. Crandall

The Star-Morning mine is a. 1000-ton-per-day lead-zinc producer located in the easterly part of the Coeur d'Alene Mining District of North Idaho. The property includes both the Star and Morning mines which are located on the westerly and easterly portions of the same complex of veins. The mine is 70% owned by the Bunker Hill Company (Gulf Resources, Inc.) of Kellogg, Idaho, 30% owned by Hecla Mining Company, and operated by Hecla Mining Company with head-quarters in Wallace, Idaho. One of the characteristics of mines in the Coeur d'Alene District is that many of the ore bodies extend to very great depths. The Star is no exception, but rather a leading example. The complex of near vertical veins and ore shoots extend some 7000 feet below the surface at present, with a maximum strike length of about 4000 feet. Ore has been mined on 20 separate veins and ore structures. More than 50 miles of headings have been driven and nearly 23 million tons of ore mined since the Morning mine started in.189S. The problems of mining at such depths are formidable temperature on the bottom level is 1100 F and high humidity problem. Pressure from the sheer weight of rock is as high foot. This paper will deal with providing a new access to the deeper levels of the mine and the installation of a hoist and related items to service these levels. By the time development had reached the 7100 level (Figure 1), it became apparent that a different 'hoisting system was going to be necessary if we were going to continue to mine at a profit. The main shaft extended to the 4000 level at which point a ?:3100"'foot winze shaft provided access to the 7100 level. The two hoists were limited to handling 3-ton payloads at 1200 to 1500 feet per minute. Some ores are still produced from above the 4000 level but as these ore reserves become depleted the hoisting of the desired tonnage (1000+/day-) from the lower levels of the mine would exceed the capacity of the existing hoisting system. A decision was made to sink a new vertical shaft at a point well in the south wall of the ore bodies where strong rocks and a minimum of faulting was expected and to purchase and install a new hoist capable of hoisting over 7000 feet. This new' shaft, designated as the No. 4 Shaft, was planned to reach 2000 feet below the 7100 level which was the deepest level at that time. The hoist would be in-stalled on the ,main 2000 Adit level which is used as an access and haulage level to the concentrator and surface plant in Burke.

Jan 1, 1971

By Sheldon P. Wimpfen

SHELDON P. WIMPFEN The measure of the success of any mining operation is the size of the difference between sales realization and cost of product, including operating, sales, administra¬tive, capital and others. Evaluation of a mining project necessitates projecting costs and sales realization as accurately as possible. Possible variations in specific cost elements and sales prices must be considered. To accomplish this, project studies should include giving effect to possible fluctuations to determine the break¬even points with varying costs and prices. Development of cost data is never an easy task. Mining is characterized by a great variety of conditions which result in still more variables. Achieving realistic figures as of any specific time involves consideration of changes, which may be frequent, in labor-contract terms and equipment used, variations in working condi¬tions and mining practices, and changes in mineralization and metallurgy. The purchasing power of the dollar, which has been declining persistently, also affects the situation at any given time. The only thing that can be said of costs, with certainty, is that they will change. With this in mind, the "guideline" data in this section are offered for the use of the mining engineer. It was provided for the most part by representative companies, whose contributions are gratefully acknowledged. In reviewing the tabu¬lations and other material it is of course necessary to remember that the data presented are valid only for the time periods specified, and will change with changing economic conditions in any other. 31.1-CAPITAL COST OF PLANTS Plant capital costs are set forth in Tables 31-6 and 31-7 by year of construction. To estimate capital cost of plants built in future years, various cost bases and indexes are utilized. Design and Construction Manpower: Monthly Salary Ranges, January, 1970 Project manager $1,400-82,600 Project engineer1,000- 1,500 Project design engineer1,050- 1,800 Senior engineer1,000- 1,700 Senior design engineer1,100- 1,700 Senior engineering designer1,000- 1,300 Layout draftsman750- 1,000 Detail draftsman650- 1,000 Material engineer900- 1,400 Construction-project manager1,500- 2,500 Project engineer1,500- 2,500 Construction superintendent 1,500- 2,500 A rule of thumb in design work is one man-month per drawing. The number will vary for each specific plant. For a recently constructed 30,000-tpd (tons per day) copper concentrator, 1,800 drawings were made. Design should provide for the addition of milling units when these may become justified in relation to investment markets and life of the deposit.

Jan 1, 1973

By ERIK CHARRIER, ADITYA JUGANDA, Jürgen Brune, Gregory Bogin, Patrick Maier

Confined explosions are a major risk in underground coal mine operations. Depending on confinement, ignition location and obstacles in the flame path, methane explosions can transition to detonations. This research focuses on explosion experiments in a reactor with diameter of 71 cm and length of 30.5 m, along with developing the corresponding computational fluid dynamics (CFD) model that captures the observed flame behavior. The results show that the 2D and 3D CFD models are able predict the experimental flame front propagation velocity, but the 3D model shows more explicit conformity than the 2D model.

May 1, 2024

By M. Misra

The interaction of xanthate with SO2 in the flotation of sphalerite and chalcopy.ite is discussed. 'The SO depression of sphaledte is due to the decomposition reaction of xanthate by SO2 The decomposition reaction is dependent on the cole ratios of xanthate/SO2, oxygen concentration and pH. Xanthate adsorbs at a faster rate by chalcopyrite than sphalerite. The faster rate of xanthate adsorption by chalcopyrite and the specific action of S02 on the surface of chalcopy.ite facilitate the selective separation of chalcopyrite from sphaledte.

Jan 1, 1989

By W. L. Hollinshead

For more effective and cost-saving production in tunnels, the machinery used on construction sites in the future will be equipped with guidance systems combined with working-procedure management systems. This paper describes sensors for these systems and vari¬ous system configurations. A project is described that uses a computerized guidance system on an underground mining machine for guidance and profile control.

Jan 1, 2000

By Michel Grenier, Mahendra Gangal

Two methods used to measure the exposure of mine workers to diesel exhaust particulate were compared. One method was developed by the U.S. Bureau of Mines and the University of Minnesota's Particle Technology Laboratory. The method which was originally developed for use in coal mines is a size selective sampling technique. The other method uses an “ashing" approach to assess the exposure to diesel particulate. This second method is based on the principles of sampling according to the ACGIH respirable dust sampling criteria and the use of a silver membrane filter as collection substrate. Because a silver membrane filter is used, both respirable combustible dust (diesel particulate) and respirable silica can be measured from one sample. The comparison study took place in a Canadian lead-zinc mine during tests which were primarily designed to evaluate the gaseous and particulate emissions characteristics of new diesel engine technology. The study has shown that under the test conditions encountered, some differences were observed between the results as obtained by the two methods. These differences seem to depend on the characteristics of the respirable dust found in the work- place, i.e., size and composition. The impact, on sampling and analysis of diesel particulate, of the proposed guideline (ACGIH) for a TLV® of 0.15 mq/m3 is discussed briefly.

Jan 1, 1997

By J. Sidhom, W. Fourie

"As the mining industry continues to face difficult challenges the need to persistently focus on productivity and utilization of labor and equipment costs will intensify. With these challenges in mind, Joy Global is helping to bring mining performance and analytics to the next level. By evaluating underground haulage, a significant component of the Room and Pillar mining value stream, variables have been identified and sensitivities have been performed to show opportunities to balance and streamline throughput. As a result of applying operational excellence tools to the mining environment, Joy Global is seeking to optimize mining performances by limiting waste, simplifying the process, increasing productivity and machine utilization, as well as removing workers from harm’s way. These days it’s not enough to work harder; mines have to work smarter. Improving efficiency, reducing costs, increasing safety: these are no longer ‘nice to haves’, they are ‘must haves’. Starting with the best products and the right people is a first step, but then what? How do mines take their operations from just OK to being a top performer? By using operational excellence principles aimed at eliminating waste, simplifying processes, automating where possible and removing people from harm’s way. The fundamentals of operational excellence (OpEx) are based upon the DMAIC process, which stands for define, measure, analyze, improve and control. DMAIC is an improvement cycle that helps with, for example, haulage optimization by: defining the issues and problems in the process as well as the areas of improvement; measuring and collecting mining operational data by time study and equipment monitoring system; analyzing through sensitivities and production modelling to determine root causes; then improving the performance and controlling the operation. Joy Global uses Joy OpEx or ‘lean’ principles inside its organization to drive efficiency, but it also helps its customers employ them to achieve new levels of productivity and safety. These same principles can be applied in most situations to streamline processes, improve operations and deliver value. As applied to mining, OpEx is seen as moving the needle on value creation by first identifying the value stream, then analyzing the areas that are not optimized. Once defined, these areas are ‘designed out’ of operational processes to create an automated operational flow. At a US coal mine, Joy Global was brought in to apply Joy OpEx principles to optimize mining and haulage to improve productivity and reduce costs. This multi-phase process started by defining the existing systems challenges to start finding areas for improvement."

Jan 1, 2017

By M. Bleeck

Calcium carbonate (CaCO3) is one of the most ubiquitous and versatile minerals found in the earth's crust. Its availability, attractive physical properties and relatively low processing cost make CaCO3 the most widely used filler material today. It is mined in three different forms - chalk, limestone and marble. Each physical form of CaCO3 has different qualities due to differences in postdepositional geology. But the chemical composition remains the same, with CaCO3 an inert component of the finished product. In the past, the paper industry largely left CaCO3 by the wayside, as it cannot withstand the acid-based papermaking process. But conversion to an alkaline system by many US mills changed this picture. Carbonate suppliers have put time and effort into research and development, demolishing barriers and creating new possibilities for what is a simple, natural product. By controlling particle size, size distribution and particle charge, the industry uses ground calcium carbonate (GCC) as a performance enhancer and as an extender for more expensive ingredients. It is estimated that the United States uses 3.6 to 4.1 Mt/a (4 to 4.5 million stpy) of CaCO3. Consolidations and mergers are taking place in the industry. Of the 12 major GCC producers in operation nine years ago, seven are left. Mineral Technology (US) is the dominant precipitated calcium carbonate (PCC) producer with more than 50 satellite plants worldwide. Other producers include Georgia Marble (French); Franklin Industries (US); OMYA (Swiss); J.M. Huber (US); ECC (English); and Filler Products (US). Global Stone PenRoc (Canada) is the only newcomer. In addition to this group, there are three small producers left in North America, each with a capacity of less than 100 kt/ a (.110,000 stpy). The trade organization operating as the Pulverized Limestone Division of the National Stone Association renamed itself the Pulverized Mineral Division, to increase its membership pool. The US paper industry is a predominant GCC con¬sumer, using approximately 800 kt/a (882,000 stpy) at an approximate cost of $130/t ($1.43/st). European paper mills pioneered alkaline papermaking. In the early 1960s, they began using GCC as filler and soon thereafter added GCC to their coating formulations. A decade later, the North American paper industry followed suit. The conversion from acid to alkaline paper production benefits the economic and performance aspects of the industry. Less pulp is needed, paper machine maintenance and effluent treatment costs are reduced, and sheet strength, opacity and brightness are increased. Perhaps most important to the reader, the sheet is desensitized to ultraviolet light, extending the paper's archival ability. CaCO3 can provide the papermaker with additional control of his paper. For example, PCC has long supplied the tobacco industry with a means to slow down the burning rate of cigarettes. Due to enhanced performance with regard to bulk and opacity, filler PCC use has risen to 1,500 kt/a (1,650 stpy) in the United States, at an approximate price of $130/t ($143/st). The majority is produced onsite at the paper mill, using "satellite plants." This concept reduces freight cost because only quicklime (CaO) is shipped to the mill, not CaCO3 slurry. The future of CaCO3 is encouraging. The amount of natural ground CaCO3 used is expected to double by the year 2005 to approximately 8 Mt (8.8 million st) worldwide. Acid papermaking practices will feel an increasing pressure to convert to an alkaline process as larger volumes of GCC containing paper enter the recycling market. CaCO3 reserves are plentiful. They will supply the ever growing demand for increasingly sophisticated paper. The plastics industry is supplied with almost 900 kt/ a (990,000 stpy) GCC at an annual growth rate of 4% to 5%. The price of a functional, inorganic filler, surface modified for the plastics industry, has an average selling price of $220/t ($243/st). GCC represents the most common filler, creating a product with higher gloss, better dielectric properties, impact resistance, weatherability and shrinkage control. CaCO3-filled plastics surround us - auto hubcaps and dashboards, shower enclosures, floor tiles, wire coatings, microwave dishes and Tupperware. The caulking and sealant industry is an enormous GCC user, with annual consumption requiring 1.13 Mt (1.2 million st) at about $44/t ($48.50/ st). Caulking and sealant may be highly filled with GCC yet undergo no adverse flow effects, with a narrow particle-size-distribution filler decreasing the binder demand. The CaCO3 industry, as well as the carpet industry, are more or less tied to the growth rates of the construction industry. It is estimated that the carpet industry uses some 680 kt/a (750,000 stpy) of GCC at about $25/t ($27.50/st). The paint industry uses approximately 300 kt/a (331,000 stpy) at a 1 % to 1.5% annual growth rate. Here, too, GCC is the dominant filler. It is used to enhance flow characteristics and color uniformity. It also extends costly titanium dioxide, creates sheen and controls roughness, hardness and tack. More CaCO3 should be used in the future as industry shifts from solvent-free or water-based formulations that can accommodate higher GCC volumes. CaCO3 is not imported or exported in any great quantity. Most areas have reserves of their own and the selling price is relatively low. Even so, quality varies from deposit to deposit. As our needs become morespecific, it remains a challenge to provide varying industries with a fitting product.

Jan 1, 1998

By Kerry Spickelmier

A traveler driving across desolate central Nevada to the Round Mountain property can easily identify with turn- of-the-century prospectors seeking a fortune, or at least an existence, from the high desert. But now a shadow passing overhead is as likely a sophisticated fighter aircraft from Stillwater Naval Air Station as a circling bird, and the famous trails west show only as historical markers along the highway. Even the approximately 400- km (250-mile) trip southeast from Reno or northwest from Las Vegas takes hours now, not days. And the camps of Round Mountain’s early prospectors have been replaced by one of the world’s largest gold heap-leaching operations. While Round Mountain’s uniqueness begins with its scale, it doesn’t end there. Consider the mine’s low cutoff grade of 0.2 g/t (0.006 oz/st) gold, made possible by a new dedicated (not reusable) leach pad. Also, the operation’s degree of automation means that ore dropped in the south crusher system is not manually handled until it is stripped from the pads. There is the bonus of a geothermal field that is tapped to heat the leach solutions in the winter. And, of course, there was that “unexpected” hitch expected in all mining operations. For Round Mountain, it meant buying a gravity plant to process the gold nuggets discovered last summer in a fault gouge.

Jan 1, 1993

By L. J. O’Connor

Mining activities in the small historic mining camp of La Colorada in the state of Sonora, Mexico can be traced back to the mid eighteenth century when gold was discovered and exploited by Jesuit missionaries. Following one hundred and twenty six years of activity, the mines were shut down due to the Mexican revolution. Very little activity took place until 1989 when a company, Minera Las Cuevas, spent US $350,000.00 in mapping and sampling and diamond and reverse circulation drilling over a period of three years. During 1991 HRC Development Corporation entered into a joint venture agreement with Minerales de Sotula, S.A. de C.V. that allowed an earn in of 70% interest. Exploraciones Eldorado, S.A. de C.V. and it's parent company, Eldorado Corporation Ltd. were subsequently incorporated by HRC to hold the La Colorada project. Property evaluation ensued and a thirty thousand tonne Run of Mine test heap was leached from August of 1992 until May of 1993. Full scale commercial production commenced in late November, 1993 and the first gold/silver doré bar was poured early in January of 1994. Pre-production costs were a total of approximately four million dollars U.S. During the fall of 1994 a buyout of the venture partners' interest gave Eldorado one hundred percent ownership of the property. The first two years of mining have utilized a local contract mining fleet and have placed nearly two million tonnes of ore on a run of mine heap. Low cost ripping and dozing mining methods have been employed in conjunction with drip irrigation leaching. An initial pilot plant carbon adsorption circuit was supplemented with a Merrill Crowe recovery circuit in June of 1994 and was completely replaced with the construction of a second Merrill Crowe circuit in July of 1995. Approximately fifty thousand troy ounces of gold and over one hundred thirty thousand ounces of silver will have been poured by year end 1995. Ongoing, on site exploration continues to increase mineable reserves in the known deposits. A feasibility study is ongoing concurrent with the operations and will provide the basis for future elevated production levels. Exploraciones Eldorado continues to explore outward from the current operations and several anomalous outlying zones may play part to the long term economic growth of the company in Mexico.

Jan 1, 1996

By Burton E. Gilpin, Dr. George W. Annadale

A two level risk analysis procedure which is specially designed for use in river environments, is presented. The first level risk assessment (Level I) is often adequate for purposes of most projects, and entails hazard assessment, exposure identification, consequence estimation, and risk characterization. Two unique features of the method are the Composite Hazard Rating system, which is used in hazard assessment, and a formalized decision model for risk characterization. The risk characterization procedure provides information on the acceptability of risks, can be used to distinguish between the relative severity of risks, and to evaluate whether more advanced risk assessment would be required.

Jan 1, 1992

By Brett P. Spigarelli

To meet the growing need for CO2 capture and storage technology, Michigan Technological University is researching CO2 capture and storage using alkali solutions1. Without implementing emission controls, world energy emissions will rise from 29.7 billion t in 2007 to 42.4 billion t in 2035. The first objective of the project was to determine the optimal amount of alkali in solution for CO2 absorption. This was accomplished using theoretical equilibrium calculations and experimental absorption data for alkali solutions of 1, 2, 3, 4 and 5 weight %. As the % alkali in solution increased, so, too, did the ionic strength of the solution. Theoretical calculations took into account the effect increasing ionic strength will have on the solubility of gaseous CO2. The calculations assumed negligible formation of H2CO3, constant temperature and constant pressure. In the experimental setup, a 16% CO2 /balance air gas mixture was bubbled through an Erlenmeyer flask filled with an alkali solution. The gas exiting the flask was fed to a gas analyzer to record the % CO2 in the exhaust gas stream. The % CO2 absorbed was taken as the % CO2 in the exhaust gas minus the % CO2 in the feed. Theoretical calculations found that as the % alkali in solution increases, the solubility of CO2 decreases. This was verified by experimental data, which found that a 2% alkali solution, by weight, was the ideal solution for CO2 absorption. Current research is focused on using the 2% alkali solution for permanent storage of CO2 via mineral carbonation to produce a value-added product.

May 1, 2012

By M. L. Smith

SAG mill throughput is often a bottleneck in increasing metal productivity. Improved size distribution control in the mill feed can be used to increase mill production. This can be accomplished by controlling blast fragmentation as a function of energy input and rock characteristics. Such an approach requires the integration of rock fragmentation and comminution modeling. The optimal mill feed size distribution is determined by adapting a population balance comminution model to a measure of rock hardness or fragility which is also relevant to the rock fragmentation that is achieved by blasting. The relation between the size distribution and fragility of the feed and of the mill's product and the critical size in a SAG mill is examined.

Jan 1, 1993

By Henry O. Ash

On June 16, 1982, Rio Blanco submitted a request to the US Minerals Management Service for a suspension of operations under its federal oil shale lease. The company also wanted a five-year delay in the requirement for further minimum royalty payments and a five-year extension of the primary lease terms of 20 years. The request was granted in late July in the midst of yet another slump in oil shale activity. Rio Blanco holds one of four Federal oil shale leases that were granted under the Interior Department's Prototype Oil Shale Leasing Program in 1974. That year the DOI offered for lease six federal oil shale tracts, two each in Colorado, Utah, and Wyoming. The leases in Colorado and Utah were granted for substantial bo¬nus bids as shown in Table 1, but the Wyoming tracts had no takers. In the eight years hence, no oil shale has been produced commercially. Indeed, the prospects for successful oil shale development have fluctuated widely, continuing a 60-year- old cyclical pattern in the fledgling industry. The prototype program was developed and initiated while this infant industry struggled to begin anew, without the benefit of any federal oil shale leasing regulations, or any history of federal oil shale leasing at all. Permanent regulations were not required because the effort was limited and designed to provide the foundation for any future leasing program. Section 21 of the Mineral Leasing Act of 1920, a Nov. 30,1973 Federal Register notice, and the leases themselves constitute the legal and regulatory framework for leasing and developing prototype oil shale tracts. The prototype leases contain some unusual provisions including installment payments of the bonus bids, credit of early development costs against payments due the government, adjustment of royalty rates according to regional crude oil price fluctuations, and unusually extensive and detailed environmental stipulations. Several of the provisions were intended to provide incentives for early development and financial penalties for failure to develop. The requirement for minimum royalty payments in lieu of actual production royalties beginning in the sixth year of the lease falls in the latter category. After the sixth year, the required minimum payment is increased east year through the 15th year by an increment equal to that due in the sixth year. The minimum royalty then remains the same through the final five years of the primary lease term. Minimum royalty is calculates on a different production level for each prototype lease, based on the quantity of 125 L/t (30 gal/st) shale estimated to be recoverable from. each tract by room-and-pillar mining. That mining system was selected for reference as the only proven extraction method for oil shale at that time (early 1970s) The minimum royalty (production) rate was calculated for each tract by direct comparison of its estimated recoverable 125-L (30-gal) resource to a hypothetical tract containing 1.9 Gt (2.1 billion st) of recoverable 125-L (30-gal) shale. For the hypothetical lease tract, the minimum production for the sixth year was calculated to be 3.2 kt/d (3,500 stpd), increasing by a like amount each lease year through the 15th to a daily production rate of 31.7 kt (35,000 st). Under this formula, the hypothetical lease tract would yield $16 million in minimum royalty payments over the 20-year primary lease term if there were no production, no development credits taken, and no adjustment of the royalty rate due to fluctuation of regional crude oil prices. The production rate of 31.7 kt/d (35,000 stpd) was a somewhat arbitrary figure that was considered, in the early 1970s, to represent a minimum size for a commercial oil shale operation. A production rate of 7.9 dam3/d (50,000 bbls/day) soon supplanted that smaller mine production figure as a rule of thumb for the minimum size for a commercial facility. The proportional relation-

Jan 11, 1982