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By J. R. Woolsey, B. G. Noakes

A prototype remote placer drill (RPD) has been designed and successfully tested in the offshore gold district of Nome, AK. The RPD system is unique in its design and in the cooperative support that was provided during its development by industry, academia and government. The RPD was conceived as a low-cost, rapid drill sampling system for the reconnaissance of marine precious metal placer deposits. The drill functions as a counter- f lush system. It incorporates a dual wall casing and a rotating percussion bit driven by way of a drill rod through the central eductor casing by a heavy duty pneumatic rock drill. The cost of drill sampling in the exploration of marine precious metal placers has traditionally been high due to the technical difficulties imposed by the environment and the remoteness of the sampling target. The objective was to reduce capital and operating costs of a drilling program through the development of a portable system incorporating drill compoments having a high power-weight ratio that could be deployed from work vessels of opportunity. A principal design requirement was to operate the system remotely from the sea floor with only flexible connections to the surface. This includes cable for deployment and retrieval and hoses for air supply and sample recovery, thus making the drill system semi-independent of the vessel's sea-induced motion. The prototype RPD underwent full operational test off Nome, AK in late summer 1988. It was deployed from the landing door of a modified LCT vessel, and rigged for a single pass of 5 m (16.5 ft). More than 50 holes were drilled in a wide variety of bottom materials typical of the off shore Nome submerged beach placers. The prototype RPD proved to be an overall success, recovering representative samples throughout the section in all but the heavy boulder ground. Compared with a conventional drill system on board that was assigned to a production drilling system, the RPD was capable of reliably sampling moderate ground less than 6 m (20 ft) in thickness, using a smaller crew and operating in higher sea conditions.

Jan 1, 1991

By R. T. Hukki

This paper presents a preliminary analysis of the fundamental relationship between the net energy used and the respective product size throughout the entire range of sizes covered by crushing and grinding, and an attempt to find a sensible correlation between the existing theories. Walker, Lewis, McAdams, and Gilliland' have given the following differential equation of a general form for comminution: dE = -C§ [1] where E is the net energy required per unit weight in a certain process of comminution; x is the factor indicating the fineness of the product; n is the exponent indicating the order of the process, and C is a constant related with the material, units chosen, etc. H exponent n in the above equation is replaced by numerical figures 2, 1, and 1 1/2, the integrated form of the general equation leads to the well known fundamental theories represented by the law of von Rittinger,' law of Kick, and the third theory of comminution by Bond, respectively. The total net energies (E) from infinite feed size to a product of size x are as follows: The net energy required in a certain process of comminution is proportional to the new surface developed according to the law of von Rittinger, to the weight or size of the bodies treated according to the law of Kick, and to the length of the new cracks formed which initiate breakage according to the theory and explanation by Bond. On a logarithmic paper, where particle size is presented on the abscissa and energy consumption on the ordinate (see Fig. 21, all three relationships are represented by straight lines. The slope m of the line according to the law of von Rittinger is equal to -1.0; of Kick, 0; and of Bond, -0.5. Experimental evidence in favor of the law of Rittinger has been presented, e.g., by Gross and zimmerley5 on quartz crushed in a drop weight crusher and evaluated for surface by the method of dissolution, by Dean on magnetite crushed by a similar method and evaluated for surface by the determination of coersive force, by piret7 and co-workers on a group of minerals crushed again by a similar method as well as by compression and evaluated for surface by permeability and gas adsorption methods, and by schellinger8 on a group of minerals ground in a calorimetric ball mill and evaluated for surface by gas adsorption. Experimental evidence in favor of the law of Kick seems to be scant in the field of comminution. On the other hand, in the field of mechanical engineering Kick's law seems to be of fundamental nature in processes such as cutting, pressing, shaping, and rolling of metallic substances. Experimental evidence in favor of the third theory has been provided by Bond." To a large extent, data are based on the vast amount of grindability tests performed in the laboratories of Allis-Chalmers Manufacturing Co. In addition to the devoted proponents of one or the other of the basic theories listed above, certain investigators have indicated that one of the theories might be applied for a certain range of sizes, while another theory might be used for other sizes. In a discussion of a paper by Bond," Dobie" presented a statement at the International Mineral Dressing Congress in London (1952) indicating that 1) for large particles, the law of Kick was approximately correct; 2) for finer particles, von Rittinger's suggestion was nearer to the truth; and 3) Bond's new

Jan 1, 1961

By Fábio de Oliveira Braga

Blindagens cerâmicas são frequentemente utilizadas para proteção contra projéteis de alta energia, como os de calibre 7,62 mm e 5,56 mm. Recentemente, foi demonstrado que a modificação da geometria da face de impacto, de plana para convexa, possibilita uma melhor distribuição das tensões geradas pela interação do projétil com o alvo. Este efeito provocaria um aumento na absorção de energia do projétil, proporcionando uma maior segurança ao usuário. No presente trabalho caracterizou-se uma blindagem com a face cerâmica convexa, por meio da metodologia da norma NIJ-0101.06 (2008), ou seja, pelo ensaio balístico com medição do trauma em um material de referência que simula a consistência do corpo humano. Este ensaio visa a aplicação do material em coletes balísticos. Foi observada uma melhora significativa no desempenho balístico após a modificação da geometria.

Oct 30, 2017

"The Australasian Joint Ore Reserves Committee (JORC), which is sponsored by the Australian mining industry and its professional organisations, developed the JORC Code to provide minimum standards for public reporting. JORC was first published in 1989 and later incorporated into the Listing Rules of the Australian and New Zealand Stock Exchanges making compliance mandatory for listing public companies in Australia and New Zealand."

Jan 1, 2003

By K. E. Soroka

More than 500 fires are now burning in abandoned coal waste banks and coal deposits in the United States. Once established, such fires can burn for decades, and extinguishing them by conventional methods such as surface sealing to exclude air, excavation to remove fuel, or flushing to cool the fire zone is usually difficult and always expensive. Burnout Control, a technique developed by the Bureau of Mines for the control of abandoned coal fires, involves the accelerated combustion of coal in place with total management of the heat and fumes produced. A burning waste bank or mine is placed under negative pressure relative to the atmosphere, and heat and combustion products are drawn from the combustion zone through an exhaust ventilation system. Heat produced appears as sensible heat in the exhaust, at temperatures as high as 1,0000 C (1,832° F), and could be recovered for the production of steam, hot water, process heat, or electricity. The Bureau's first held field demonstration of Burnout Control was at Calamity Hollow in Allegheny County, PA (near Pittsburgh). Calamity Hollow was the site of an underground mine in the 1900's and was surface mined in the 1940's. In the winter of 1961-62, a fire of undetermined origin was discovered in the exposed coal. In 1963, the Bureau constructed a trench barrier around the fire and a surface seal over the affected area. The fire was isolated but not completely extinguished. In 1979, when the Bureau began work on the Calamity Hollow Mine Fire Project to demonstrate controlled burnout, the fire was still smoldering on the hot side of the trench barrier. The project, which was begun in December 1979 and ended in July 1982, consisted of the design, construction, operation, and subsequent dismantling of a Burnout Control ventilation system. This report is the fifth in a five-part series that describes the Calamity Hollow Mine Fire Project. The first report, part 1, describes the design and construction of the field installation. Part 2 will present the results of a continuous 4-month burnout operation. (Because part 2 involves the analysis of a substantial body of data, it will not be published until after publication of parts 1, 3, 4, and 5.) Part 3 describes the instrumentation used to control and monitor the progress of the burnout operation. Parts 4 and 5 deal with the closeout phase of the field demonstration. Part 4 describes the procedure used to quench the fire, and this part (part 5) describes the final excavation and backfilling of the heated zones. The reports in this series document the Calamity Hollow controlled burnout demonstration which showed that (1) controlled in situ combustion is a feasible method for controlling underground fires in abandoned mines, (2) the resultant thermal exhaust output is sufficient for energy utilization, and (3) water injection-fume exhaustion is a potentially effective method for cooling large underground fire zones. Further investigations of both Burnout Control and the water-injection-fume exhaustion quenching procedure are planned.

Jan 1, 1986

By André. C. Louw

Diamond production from the reworked beach and river channel gravels along the west coasts of South Africa and Namibia is well known throughout the world, not only for the exceptional quality of the diamonds, but also for the sheer scope of the operations. To date an estimated 140 million carats have been produced from the resource worth some US$40 billion in today's money terms. It is therefore not surprising that it didn't take long before the attention turned towards the possible extension of this valuable diamond resource on similar submerged beaches in the adjacent territorial waters of South Africa and Namibia. This has proved to be a very successful move as the marine production in Namibia today constitutes 48% of that country's total diamond production. Although small scale recovery of diamonds from the sea commenced in the 1950's, the first large scale production came from the operations of the Texan entrepreneur, Sam Collins, in the 1960's. The De Beers Group, which took over Sam Collin's operations in 1969 had all the concession areas, as well as the financial and human resources to tackle the new mining frontier in the correct way by years of data gathering (ie diamond reserves and seafloor terrain studies) before commencing with large scale mining operations in the 1980's. Those who are familiar with the beach mining operations along the South African and Namibian west coasts will have appreciation for the daunting task of surgically removing the payable portion of diamond deposits from the non-payable material on the one hand, and the complex uneven fractured and gullied bed-rock floor of the deposits on the other hand. Not to mention the additional complexities of large boulders and hard calcretised diamond gravel layers (conglomerate) which are sometimes associated with the deposits. It was a natural development that the sea diamond mining operations were approached with the known information of the land operations, attempting to duplicate land mining operations in the sea, but taking cognisance of the variations of the terrain in the different areas. For instance, we encounter the following variations, and combinations, off the Namibian Coast. ? Close inshore - holocene sand cover often masks large areas of potentially diamondiferous gravels. ? In midwater areas (40m - 100m)

Jan 1, 1998

By A N. Danoucaras, S E. Goater, D Hearne, A M. Bansuan

It is globally recognised that existing and emerging mining activities should be conducted in a æsustainableÆ manner for the entire life of mine. Many companies annually publish corporate sustainability reports to demonstrate their actions in accordance with international expectations. This concept of sustainable mining practices also applies to mine water management, with influences often extending beyond site operations, affecting surrounding catchments and community users. This paper explores differences in conceptual and operational terms of æsustainableÆ mine water management within the Philippines. Outcomes of a review of corporate sustainability reports, governance documentation, research programs, and stakeholder knowledge acquired through action-research workshops are presented. The paper aims to inform and support future capacity building exercises delivered by the International Mining for Development Centre (IM4DC), with outcomes tailored to government officials and policy makers of developing countries. The project is funded by IM4DC, which is working with AusAID to identify needs and develop capacity with regard to mining for development; specifically, issues pertaining to governance and regulation, social and environmental sustainability, and operational effectiveness in developing countries.CITATION:Bansuan, A M, Goater, S E, Danoucaras, A N and Hearne, D, 2012. Integrated sustainable water management in mining for development in the Philippines, in Proceedings International Mine Management 2012, pp 205-226 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 20, 2012

By M. Dotto, T. G. Joseph, M. Curley

"Equipment selection is a key decision in mine planning. Equipment size affects decisions from pit size to the total operation cost. Particularly when replacing haulage fleets for long-life surface mining operations, mining operations focus solely on haulage productivity. Road width and specifically ramp width are significantly affected by an increase in hauler size in terms of overall strip ratio through required highwall set-backs or sterilizing of lower ore access due to required overlying ramp fill structures. In addition, in light of ever increasing regulatory and reporting emissions requirements, bigger equipment may not translate into lower emissions per tonne moved. This study analyzes hauler-scale impacts on aspects not currently incorporated into conventional mine planning, including expansion of roads to accommodate larger equipment, road layer thickness variation depending on hauler size, and fuel consumption and emissions. Results indicate that such relationships have a large impact on cost. Therefore, an expanded consideration of the capacities of increasing equipment size in mine planning is highly recommended. ¦ KEYWORDS Equipment selection, Grade resistance (GR), Haul road, Haul truck, Mine planning, Payload, Rolling resistance (RR), Total resistance (TR) RÉSUMÉ La sélection du matériel est une décision importante dans l’aménagement des mines. La taille de l’équipement a une incidence sur la prise de décisions, depuis la taille de la fosse jusqu’aux coûts totaux des activités. L’exploitation minière se concentre uniquement sur la productivité des moyens de transport, particulièrement lorsqu’il s’agit de remplacer les parcs de camions de transport dans les exploitations à ciel ouvert à longue durée de vie. La largeur des routes, et spécifiquement la largeur des descenderies sont considérablement affectées par une augmentation de la taille des camions de transport en termes du coefficient global de recouvrement qui dépend de la taille requise des parois (si les descenderies sont agrandies, les parois devront être creusées de haut en bas afin d’agrandir la taille du fond de la fosse). Il faut aussi pouvoir procéder à l’excavation des gradins les plus profonds de la mine et ainsi préserver l’accès à cette zone, ce qui pourrait devenir difficile si la quantité de stériles utilisés pour la structure des descenderies augmente et s’ils occupent une trop grande partie du fond de la fosse. En outre, au vu des exigences de plus en plus strictes en matière de réglementation et de déclaration des émissions, un équipement plus volumineux ne sera pas nécessairement synonyme d’une réduction des émissions par tonne déplacée. Cette étude analyse l’incidence de la taille des camions de transport sur des aspects actuellement non intégrés dans l’aménagement classique des mines, notamment la création de routes permettant le passage d’un équipement plus volumineux, la variation de l’épaisseur de la couche d’asphalte des routes en fonction de la taille des camions de transport, ainsi que la consommation de carburant et les émissions. Les résultats indiquent que ces relations ont des répercussions importantes sur les coûts. Ainsi, on recommande d’accorder une attention accrue aux capacités d’augmentation de la taille de l’équipement dans l’aménagement des mines. ¦ MOTS CLÉS aménagement des mines, camion de transport, charge utile, résistance au roulement, résistance due aux rampes, résistance totale, route de transport, sélection du materielINTRODUCTION Over the past five decades, the mining industry has progressively attempted to increase production rates through the use of larger haulage equipment. Each new generation of mine is designed to operate with the most recent and often bigger equipment iteration, while older mines are re-evaluated to keep pace with upgraded mixed fleets. Whereas production gains are often realized with larger equipment, there are additional aspects to consider when evaluating the feasibility of mo"

Jan 1, 2019

By T Purvis, Z Liu

Ernest Henry mine (EHM) is an underground sublevel caving mine located in north-west Queensland with a production rate of 6.4 Mt/a. The mine has sublevels spaced with a 25 m vertical interval. The production process on each level starts with the drilling and firing of a blind uphole rise to start the slot extraction. These rises are fired as a single shot and routinely need to achieve a design height of 22 m to 30 m depending on the local mine geometry. If the rise fails to perform as designed then the production phase of the level is delayed with significant rework required to remedy the situation.This paper describes the process that the EHM team has followed to improve the first time performance of each rise. The improvement process commenced with a review of results from previous rises and resulted in the identification of a number of improvement opportunities. A revised holistic approach was established for the process, which included involvement from operators and engineers. Through continued measurement and review of rise performance, inputs to the process were improved to a point where reliable results can now be expected. The successful performance of subsequent rises has been verified through a comprehensive quality assurance quality control (QA/QC) program. Before quality control processes were implemented, the success rate of blind uphole rises was only 11 per cent (one out of nine rises) at EHM. In comparison, the success rates have been vastly improved to 25 per cent (one out of four) and 75 per cent (three out of four) respectively following two stages of improvements. This paper concludes with the ‘current’ series of best practices used at EHM and recommendations.CITATION:Purvis, T and Liu, Z, 2017. Improving the performance of blind uphole rises at Ernest Henry Mine – a case study, in Proceedings 13th AusIMM Underground Operators’ Conference 2017, pp 335–340 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Oct 16, 2017

United Engineering Trustees, Inc. W. D. B. MOTTER, JR., '40 A. L. QUENEAU, '41 ALBERT ROBERTS, '39 The Engineering Foundation GEORGE D. BARRON, '40 F. F. COLCORD. '42 A. L. QUENEAU, '41

Jan 1, 1939

By Darina Manova

Plasma assisted surface modification methods have been used to nitride stainless steels in order to enhance hardness and therefore the lifetime of industrial components. Nevertheless, the corrosion resistance is not always preserved since it is very sensitive to the process parameters, either in implantation or in thermochemical diffusion techniques. AISI 316L stainless steel was plasma nitrided using three different techniques: Low Energy Ion Implantation (LEII), Plasma Immersion Implantation (PI3) and conventional DC pulsed plasma nitriding (DCPN). For each method several parameters were tested and some of them were selected to achieve the so called S phase of about 3 µm thickness and no nitrides precipitation. The nitrided layer was characterized with OM and SEM. Hardness was assessed on the surface with Vickers indenter and 25 g load. XRD was performed to characterize the S phase and to detect other phases if present and electrochemical corrosion tests were carried out in 3.5% NaCl solution. Although the S phase thickness in the three groups of samples was comparable, hardness was not and neither was the corrosion resistance. The PI3 samples were the hardest with the best performance in corrosion experiments, resulting in a higher breakdown potential and lower currents.

Aug 8, 2018

By J. K. A. McLEOD

Address by the President of the Institute (Mr J K A McLeod) in presenting the Institute Medal to Howard Knox Warner, C.B.E. Address by Howard Knox Warner after accepting the award.

Jan 1, 1978

By Richard J. Mainiero, Marcia L. Harris, James H. Rowland

In the past several years, there have been a number of blasting-related carbon monoxide migration incidents. In each case, there have been some common factors that appear to be related to carbon monoxide migration. Geology, heavy confinement of the blast, close proximity to the affected structure, and open pathways into the affected structure are just a few. Prince William County, VA is an area undergoing a boom in suburban development. In the year 2001, more than 4,050 new homes were constructed within the county.1 In the new housing development located in Bristow,VA, trench blasting was utilized to excavate proposed sewage and utility lines. On November 8, 2002, a trench blast was conducted for the installation of utility lines. Later that evening a carbon monoxide detector alarmed in a nearby house. The gas company and fire department sampled for carbon monoxide and measured 100 ppm and more than 200 ppm on the first floor and in the basement respectively. Subsequent efforts to prevent migration of CO into the nearby homes met with mixed results. The only reliable technique for protecting neighbors from CO appears to be the placement of CO monitors in the homes. The events are presented and discussed.

By Doerner H. A.

From thermodynamic calculations Maier predicted that zinc oxide and methane would react at specified temperatures substantially according to equation (1). (1.) ZnO + CH4 = Zn (gas) + CO + 2H2. Other reactions that may take place are (2.) CH4 = C + 2H2. (3.) ZnO + CO = Zn (gas) + CO2. (4.) ZnO + H2 = Zn (gas) + H2O. Maier predicted that 870°C would be the critical temperature above which reaction (1) would take precedence over reaction (2) and that reaction (1), at equilibrium, would be substantially complete at or above 927°C. Theoretical considerations indicate that reaction (1) is reversible: this is supported, but not proved, by experimental data. The other reactions are known to be reversible. At low temperatures reaction (2) requires a catalyst and it is slow below 1000°C. Reactions (3) and (4) are rapid, but the equilibrium constants are such that only small concentrations of CO2 or H2O can be produced below 1000°C. Maier calculated that CO2 should not exceed 0.5 per cent at 870°C., and the calculation for water vapor gives a still smaller order of magnitude.

Jan 1, 1931

By C. H. Attwood

Introduction In presenting this paper on the water-power resources of the Province of Manitoba, it was considered that a brief resume of the power resources of Canada, of the power already developed, and its utilization in the industries 9f the country, would be of particular interest in showing the rapid progress that has already been made. Amongst Canada's national assets, water power occupies a foremost place because of the facilities it provides for the development of her other great natural resources. The rapid development of the mining, pulp and paper, and general manufacturing industries, and the concomitant growth in urbanization, with its numerous demands for domestic, commercial, and municipal service power, have combined to stimulate hydraulic development to a point surpassing even the most sanguine predictions of a few years ago. With the bringing into operation during 1928 of hydraulic turbines aggregating over 550,300 h.p., and of a further 378,400 h.p. in 1929, Canada's total hydraulic development at the end of 1929 amounted to over 5,700,000 h.p., and in this year, 1930, the active construction in hydro-electric plants throughout the country totals in excess of 1,600,000 h .p.

Jan 1, 1930

United Engineering Trustees, Inc. W. D. B. MOTTER, JR., '40 A. L. QUENEAU, '38 ALBERT ROBERTS, '39 The Engineering Foundation GEORGE D. BARRON, '40 F. F. COLCORD, '38 A. L. QUENEAU, '38

Jan 1, 1937

By P. D. Rao, D. E. Walsh, W. W. McClintock

A column flotation system was designed to conduct release analyses on a very finely disseminated, stratiform copper ore from the Denali Copper Prospect in Alaska. Chalcopyrite is finely disseminated in a carbonate matrix, and pyrite is disseminated in both the chalcopyrite and the gangue. Grinding to -20 µm is required for adequate liberation. A 25-mm x 2.4-m glass flotation column was designed. The ore and its flotation tailings were repeatedly floated to obtain a rougher concentrate. The rougher concentrate was then repeatedly floated to produce release fractions. These fractions were analyzed for copper, iron, calcium and magnesium, and release curves were plotted. The procedure was replicated at each of two grinds, i.e., 99% and 88% passing 37 µm (400 mesh). The procedure should find application for flotation characterization and process selection. Bond grindability testing was also conducted.

Jan 1, 1996

By P Ormsby, T Hewitt

Ore in pillars constitutes a significant proportion of the ore reserve at PasmincoÆs Broken Hill Mine. Many of these pillars are narrow (5 -6 m) and high (16 û 35 m) vertical blocks. High-grade ore has been left between primary horizontal cut and fill stopes. These stopes were mined from the late-1960s until the early-1980s. Fill material is a variety of hydraulically stowed sand, sand-cement, and remnant timber support. Many pillar cut-throughs were made between the old stopes. In order to gain access for the current pillar recovery operations, it has been necessary to develop pillar drill drives through, and adjacent to, the hydraulic filled stopes. Whilst long hole recovery of vertical pillars is not new, some notable successes using innovative ideas have been achieved during 1997. These innovations include:The use of steel mesh reinforcing in shotcrete when mining through, or adjacent to, hydraulic fill. Construction of steel sets to support vent rises and haulage ways encountered in the cut-throughs. The use of ground movement monitoring equipment. Survey lasers with chain droppers and convergence monitors were installed in the 18 Pillar to detect any movement of the pillar block and decrease risk to personnel. The development of a three-dimensional computer model of primary stoping, using Vulcan software, to assist in the design of pillars. Maximisation of recovery using different drilling and firing practices.

Jan 1, 1998

By Fazel G. Jahromi, Ahmad Ghahremaninezhad

Enargite (Cu3AsS4) is a major complex copper mineral and often is a carrier of invisible gold particles. Pyrometallurgical treatment of enargite is challenging due to the arsenic release to the environment. Hydrometallurgical treatment, on the other hand, appears to be a promising alternative for enargite processing. In an atmospheric enargite leaching process both the copper and arsenic from the mineral can be leached into a sulphuric acid solution and the arsenic can be simultaneously precipitated as ferric arsenate. The unlocked gold in the residue then can be recovered in a subsequent cyanidation circuit. The downside of the atmospheric enargite leaching in sulphate solutions is the slow leaching kinetics. In this study the effective parameters of the atmospheric enargite leaching process in ferric chloride media were investigated. All the experiments were carried out at temperatures below 100 °C. The process parameters that were investigated in this study included ferric, chloride, sulphate and cuprous concentrations, temperature, mixing rate and leaching slurry pulp density. It was found that the copper recovery rate increases with the addition of cupric chloride and also with raising the temperature. The highest leaching kinetics and the highest overall copper recoveries were achieved when cupric was initially added to the leach solution and at 90 ºC.

Jan 1, 2015

By J. R. Ruark

Because of the importance of oil-shale reserves the Bureau or Mines resumed research investigations in this field beginning in 1944. The Bureau early decided to build a Large-scale demonstration plant to accomplish the dual purpose of obtaining design data and to produce crude shale oil for refining studies. Two 40-ton-Capacity batch N-T-U (Nevada-Texas-Utah) retorts were completed for this purpose in 1947. After a preliminary operation, the plant was used to investigate process variables of air rate, recycle gas rate, and shale particle size and crude. Correlations obtained from these variables are presented in this report. Researchers concluded that recycle gas was not necessary for efficient operation of the process. Short comings of the batch method of processing the oil shale are pointed out in the report. Several periods of prolonged production operations were carried out. During these periods, the retorts were operated to produce as much crude oil per day as possible. The lifetime operation of the N-T-U retorts was divides into three periods, namely, preliminary operation, process variable study, and production operation. During the preliminary operation, 187 runs were made with a total of 7,596 tons of raw shale used to produce 3,688 barrels of crude shale oil. During the process variable study, 179 runs were made with a consumption of 7,494 tons of raw shale and a production of 4,218 barrels of crude shale oil. The production operation consisted of 554 runs in which 22,457 tons of raw shale was used to produce 12,359 barrels of crude shale ail. The overall total for the lifetime operation of the retorts consisted of 920 runs with a total consumption of 37,547 tons of raw shale and a production if 20,265 barrels of crude shale oil.

Jan 1, 1956