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By J B Tucker, A T. McGrath

Situated in north-west Queensland, Australia, the Ernest Henry Mine (EHM) has been producing copper and gold since 1996. Owned and operated by Glencore Xstrata, ore is currently sourced from a newly developed underground sublevel cave (SLC) mine below the old EHM open pit (the pit). Ore feed to the concentrator is supplemented by satellite open pits comprising the Mount Margaret operation.The transition from open pit to a producing underground mine involved a complex extraction sequence for the first two production sublevels on RL1650 and RL1625. Underground mining during cave initiation was further complicated when south wall instability prevented completion of the final open pit shell. The transition mining phase included mass firing of an inclined crown pillar between the initial SLC void and the pit followed by a combination of transverse and longitudinal SLC mining.The inclined crown pillar mass firing was the first significant milestone in the extraction of the transition ore block. Comprising approximately 890 000 t of in situ material; 70 000 m of drilling was charged with 351 t of emulsion and 6026 electronic detonators to initiate the single mass firing. The firing contained five direction changes over 19 seconds making it one of the largest and most complex underground blasts ever conducted in Australia.Risk management was integrated into every facet of the technical and operational planning and execution of the transition mining phase, whilst maintaining daily mine construction and production commitments. This paper provides a detailed overview of the transition objectives and constraints, mining method selection, post mass blast outcomes and the journey ahead.CITATION:McGrath, A T, Campbell, A D and Tucker, J B, 2014. Transition of the Ernest Henry Mine from open pit to an underground sublevel cave, in Proceedings 12th AusIMM Underground Operators’ Conference 2014 , pp 297–308 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Mar 24, 2014

Equilibrium calculational software programs can be effectively utilized to illustrate the importance of activity or concentration, solubility, temperature, solution pH, solution potential, and ligand complexation to students studying hydrometallurgy. The calculational program STABCAL will be demonstrated. illustrative examples for the creation of diagrams such as EH-pH, activity or concentration-pH and activity or concentration-EH will be presented to demonstrate the fundamentals of real hydrometallurgical systems. Speciation, titration and mixing will be used to simulate or predict operational processes. Applications will include treatment of Acid Mine Drainage (Berkeley Pitlake water and sediments), mass balance potential/pH diagrams, leaching and selective separation of metals from electroplating sludge, metal adsorption at femhydrite surfaces, and flowsheet development for treating mercury sulfide K106 sludge.

Jan 1, 2002

By J Player, R de Vries, E Villaescusa, R Hassell

Concerns regarding the uncertainty of stope dilution at MMG Dugald River Underground Mine led to a trial stoping program to acquire full-scale comprehensive geotechnical information and to test the validity of the geotechnical and mining parameters. Given that stope production drilling and blasting is a key parameter that influences dilution a detailed stope blast vibration project was implemented. The stoping trial consisted of 19 stopes, a combination of both sublevel open stoping and longitudinal and transverse stopes, of which five were monitored for blast vibrations.The blast vibration project utilised 20 triaxial geophones and five triaxial accelerometers installed in the hanging wall of five monitored stopes. For each stope fired, full blast waveforms were recorded along with detailed information of the blast design. Analysis of the blast vibrations used the Holmberg-Persson approach, with potential damage determined by the relative (Holmberg-Persson) K values.The project identified the benefit of parallel, breakthrough slot holes in reducing blast vibration damage. The results show that radial rings toeing or stabbing into the stope wall lead to higher vibrational energy yield which contributed to increased overbreak. Hanging wall overbreak, by comparison, with parallel, breakthrough blastholes with 1 m standoff, combined with lower density explosive charging (0.8 EP) and increased delays limited blast vibration damage. A number of ring burdens from 1.4 m to 2.5 m were monitored with no difference in terms of energy yield being identified. Changes to the stope drill and blast design are suggested to better control the large blast vibrations and recommendations from this project have been implemented into future mine design and planning.CITATION:Hassell, R, Villaescusa, E, de Vries, R and Player, J, 2015. Stope blast vibration analysis at Dugald River Underground Mine, in Proceedings 11th International Symposium on Rock Fragmentation by Blasting, pp 221–232 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 24, 2015

By L. C. IlsLey

Cooperation between the United States Bureau of Mines and the Safety in Mines Research Board of Great Britain, continuous since 1924, has made possible this and other papers on safety subjects. Grateful acknowledgment is made to representatives of the Board for their assistance in arranging visits to several mine safety stations, and to F. R. Wynne, Deputy Chief Inspector of Mines, for arranging visits to mines in Great Britain and other countries in Europe. During the summer of 1927 the writer had the privilege of visiting the mine safety testing stations in great Britain, Belgium, Germany, and France, in the order named. All of these countries have large coal mines, many of which are rated as gassy. Therefore, when the installation of electrical equipment is contemplated, each of these countries is confronted with the same safety problem as is the United States - the development of electrical equipment that will not ignite gassy atmospheres which through neglect or accident might surround the equipment. The means employed by these countries in safeguarding gassy mines should therefore be of general interest to safety engineers in American coal mines, and it is proposed to give a brief survey for each of the four countries mentioned. The third of these surveys covers conditions in Germany.

Jan 1, 1929

By A. A. Cochran

Bureau of MinesThe Bureau of Mines studied the development of practical methods for recovering manganese from the extensive domestic resources of low-grade ores and open hearth slags. A one-step operation was developed on a laboratory scale in which manganese and iron compounds are treated with chlorine, volatilized, and separated by selective condensation. Manganese recoveries of 90 to nearly 100 percent were obtained from 9 of the most important domestic ores. Analyses of the manganese chloride product indicated that it was suit-able for aqueous electrolysis with little additional purification. A complete, integrated process that provides for the recovery and recycling of chlorine is proposed. This process appears to have distinct advantages over previously proposed manganese recovery processes, and, in sharp contrast to leaching procedures, it is applicable to practically any manganese-bearing material.

Jan 1, 1967

By P. L. Sibrell

The U.S. Bureau of Mines has tested cyanide leaching for recovery of palladium (Pd) from spent petroleum processing catalysts. Three different catalyst samples were supplied by a spent-catalyst processor. These catalysts consisted of a zeolite base and contained 0.4 to 0.7 pct Pd. During alkaline cyanide leaching, the catalysts exhibited ion-exchange properties due to their zeolite matrices. Hydrogen ions were released from the zeolite in exchange for sodium ions in solution, resulting in a significant decrease in solution pH. This could present a safety hazard because of the potential for release of toxic hydrogen cyanide gas. A pretreatment step where the catalysts were contacted with a 1.0M sodium hydroxide solution was found to prevent the pH shift from occurring. Following the sodium hydroxide pretreatment, two stages of leaching at 160 °C with solution containing 1 pct sodium cyanide and 0.1[M] sodium hydroxide gave at least 75 and up to 95 pct Pd recovery. The Pd was quantitatively recovered from the leach solution by thermal decomposition in an autoclave at 250°C for 1 h. The Pd content of the precipitate was over 50 pct. Thermal decomposition also decreased the total cyanide content of the barren solution to less than 0.2 mg/L. The catalyst leach residues passed the Federal Toxicity Characteristic Leaching Procedure and the California Waste Extraction Test, indicating that landfill disposal of the leach residues would be acceptable.

Jan 1, 2010

By E. J. Eisenach, W. E. Jones

SAFETY and industrial hygiene have always been recognized as highly important in company policy, and the co-operative support of the company officials and entire plant personnel has contributed largely to the present success of the Company's program in this field. Growth has necessarily been slow, and dependent, at times, on trial and error but the program in its present form combines the regulations, working methods, and equipment that have proved their value under actual operating conditions. Every effort is made to keep the program on a practical basis. Continuous experimental work is carried on in conjunction with the health and safety program, with resultant periodic changes for improvement. As a proof of accomplishment, in 1945 the frequency and severity rates for accidents were the lowest in the history of the operation. A full-time safety engineer was employed in 1933 and an assistant safety engineer in 1936. During the following year a stenographer was employed in the office and a safety inspector was added for the mine. During the gradual expansion of operations that was climaxed by

Jan 1, 1946

By Gavin Beer

The Chelopech Copper/Gold Mine, located in west-central Bulgaria, has been in operation since 1953 producing a low-grade copper/gold concentrate containing up to six percent arsenic by weight. The penalties and restrictions imposed on smelting this concentrate, have seen the past operators of the mine strive to find a balance between achieving acceptable metal returns from the smelters and installing novel or unproven technologies in the face of changing political and economic climates. This paper documents the colourful history of the concentrate processing alternatives that have been investigated and implemented over the 45 year life of the Chelopech Mine, culminating in the high pressure oxidation option that is now the focus of a Definitive Feasibility Study (DFS) presently being undertaken by GRD Minproc Limited. The fundamental chemistry of the preceding alternatives is outlined followed by a discussion as to why each was ultimately not employed. A detailed description of the technology proposed in the DFS is given including: pressure oxidation, arsenic fixation, SX/EW copper recovery, CIL gold recovery and AVR/SART cyanide/copper recovery processes. Keywords: Arsenic, Scorodite, HPOX, Chelopech, Dundee

Jan 1, 2005

By J S. J van Deventer

"Conventional wisdom is that gold not assayable is not recoverable or does not exist. There has been no formal study on the possibility that gold assays using advanced spectroscopy such as mass spectrometry and neutron activation could grossly underestimate the gold recoverable by proprietary extraction techniques. Such gold is not considered here to be æinvisibleÆ to detection but rather æabnormalÆ or ædistortedÆ at atomic level. Large-scale extraction of gold from æabnormalÆ ores has not been demonstrated, hence this emerging field still lacks credibility and acceptance. For fear of harming their reputation, few academic researchers or companies are expected to venture into this field until it is well proven. This paper aims to create awareness amongst metallurgists of the potential of æabnormalÆ gold ores by reviewing unconventional observations, the pseudoscience mainly available in non-peer reviewed web sites, and hypotheses for the recovery of ænon-assayableÆ gold, including: the occurrence of high levels of nano-sized gold in clays detectable by high-precision analytical instruments, which represents æinvisibleÆ gold accounts of ambient transmutation of elements, mainly using thermal methods orbitally rearranged monoatomic elements (ORMEs), which are virtually undetectable by conventional means, and their conversion to normal metals the possibility of a æhigh spinÆ state of transition metals the formation of microclusters altering the chemical behaviour of gold and the possibility of clustering with other elements. It is suggested here that gold in ores displays a range of clustering, from ænormalÆ gold amenable to conventional extraction to the ORME state. The concepts of æintermolecular waterÆ in forming complex salts locking up precious metals, electro-dissolution in strong acids, and electrowinning at both the anode and cathode are discussed as a means of treating æabnormalÆ gold ores. It is also possible that the gas phase extraction of gold chlorides under controlled temperature and pressure conditions, as well as transferred-arc plasma fuming of gold through carbothermic reduction of sulfides, could be used as a basis for extraction of gold from æabnormalÆ ores.CITATION:van Deventer, J S J, 2013. Is there gold that we do not recognise?, in Proceedings World Gold 2013, pp 27-34 (The Australasian Institute of Mining and Metallurgy: Melbourne)."

Sep 26, 2013

By S. Fazluddin, K. C. Sole, M. R. Serwale, T. Coetsee

The CSIR-Ti process employs lithiothermic reduction of titanium tetrachloride feedstock to produce titanium sponge. The product is therefore contaminated by a range of lithium and chloride species. In this study we examine the effects of particle size, temperature, and HCl concentration as input leaching variables on the removal of chlorides from the crude titanium sponge. A review of the aqueous chloride chemistry of Li and Ti provided initial conditions for leaching of impurity species from the sponge. Experimental results confirm that the effectiveness of leaching and removal of dissolved impurities from the sponge are dependent on leaching kinetics, which are influenced by temperature, particle size, and morphology. Of the variables tested, reaction temperature had the strongest influence on the oxygen content of the leached product. The HCl lixiviant concentration had a negligible effect under the conditions tested. Leaching of crude titanium sponge (−10 mm size fraction after crushing) at 14°C in either 1 M or 0.032 M HCl yielded a titanium sponge product that met the ASTM standard specification for commercially pure Grade 1 titanium, i.e., oxygen content < 0.18 mass% and chloride content < 0.15 mass%.

Dec 13, 2022

By C. W. Davis

THE name fuller&apos;s earth, which was derived from its early use in "fulling" or removing grease from woolen goods, is a term that is generally considered to designate mineral matter, containing hydrous aluminum silicates of the clay group, which has a high capacity for removing colors from oils, whether of animal, vegetable, or mineral origin. Fuller&apos;s earths are usually dried and may be "tempered" by heating or may be treated with water or reagents to improve their efficiency. For our purpose the term "acid-treated earths" will include those products from the partial decomposition of mineral water by acid which are equal to or more efficient than an untreated fuller&apos;s earth from Quincy, Fla., when used in bleaching mineral oil by the "contact" process. In general, the contact process for treating oils consists of heating and agitating a mixture of oil and finely divided adsorbent, the latter being subsequently removed by filtration. The term "adsorbent" is used because, although there has been a great deal of controversy over the mechanism of removal of impurities from oils by mineral matter, there is no doubt that adsorption, which involves the surface of the solid, plays a major part.1 The properties that make fuller&apos;s earth useful in the treatment of petroleum products by "percolation" and in the purification of edible oils by the contact process have long been studied, but the comparatively recent adaptation of the contact process to the bleaching of petroleum products has made desirable the further investigation of fuller&apos;s earth and other adsorbent earths.

Jan 1, 1929

By Stanley Rajnak, John E. Dorn

The saddle-point activation energy for the nu-cleation of a pair of kinks is estimated as a function of the applied stress, the lattice constants, and the height and shape of the Peierls&apos; hill by employing the line-energy model of a dislocation. The kinetics of disloration motion by the Peierls&apos; mechanism are discussed and the microscopic dislocation velocities and macroscopic strain rates are formulated by the Arrkenius type of approach in terms of the activation energy for the formation of pairs of PEIERLS was the first to illustrate that a straight-dislocation line has its lowest energy when it lies in a potential valley parallel to lines of closest packing kinks. The log of the dislocation velocity is found to be approximately proportional to the log of the local applied stress. It is also shown that the low-temperature mechmical properties of several bcc metals might be explained in terms of the Peierls mechanism, good agreement being obtained between theory and experiment for the flow stress us ternperature, activation energy vs stress, and activation volume us stress curves for these metals. of atoms on the slip plane. When such a straight dislocation is moved rigidly from one valley toward the next, the atoms in the vicinity of the core of the dislocation change their positions and bond angles, causing the energy of the dislocation to increase. Midway between two adjacent valleys, the disloca-tion energy reaches a maximum value and any additional small displacement will cause the dislocation to fall down the hill into the next valley. The maximum shear stress necessary to promote such

Jan 1, 1964

By S. C. Lindsay, D. J. I. Evans

With each year that passes hydrometallurgical processes are being more widely used to recover base metals from ores and concentrates. Generally these processes involve liquid-solid separation of metal-bearing liquors from barren residues. This may be done by countercurrent decantation, thickening and filtration, or filtration or centrifuging only. Choice of method is governed by chemical and physical characteristics of the pulps; cost of labor, power, materials, and waste disposal; and even by the amount of capital available when building plans are under way. Since all these controlling factors change with time and location and with technological advances, the separation method must be re-evaluated when new plants or expansion programs are contemplated. At the Sherritt Gordon refinery in Fort Saskatchewan, Alberta, an ammonia pressure leaching process recovers nickel, copper, and cobalt from the nickel concentrate produced at Lynn Lake, Manitoba.&apos; Pregnant liquors are separated from the barren leach residues in a thickener-filter liquid-solids circuit. The decision to use this type of separation was based on a summation of economic factors during the pilot plant and design stages. Although countercurrent decantation was not incorporated in the refinery flowsheet, the problems solved in thickening the leach residues and the experience gained in operating the commercial plant are pertinent to countercurrent decantation, particularly where it is applied to slow-settling leach slurries or slurries having a relatively high vapor pressure of a valuable, obnoxious, or poisonous vapor. The nickel sulfide concentrate is leached under pressure with air and ammonia to dissolve the nickel, copper, and cobalt and most of the sulfur, leaving a residue of iron and siliceous material to be discarded.2,3 During a subsequent distillation stage to recirculate part of the ammonia used for leaching, copper is recovered from the solution as a high-grade sulfide. After further purification to remove traces of copper and undesirable sulfur compounds, the nickel is precipitated by hydrogen under pressure at an elevated temperature. The cobalt, together with unreduced nickel, is precipitated as a mixed metal sulfide by hydrogen sulfide for eventual cobalt recovery. The end solution is then evaporated to produce a crystalline ammonium sulfate byproduct. As shown in Fig. 1, the nickel concentrate is leached in a continuous, two-stage, countercurrent circuit. In the first stage, or adjustment leach, the fresh concentrate is partially leached (60 pct) in the liquor produced in the second stage, or final leach. This system is necessary to produce a solution suitable for treatment in the copper sulfide precipitation step. Both stages are carried out at 170" to 180°F under a total pressure, with air, of 100 to 120 psig in large, horizontal, water-cooled autoclaves. The autoclaves are divided by overflow weirs into four compartments to give a certain amount of staging during leaching. In each compartment a relatively slow-speed (100 to 150 rpm) agitator maintains the nickel concentrate in suspension. Adjustment leaching, carried out in two autoclaves in parallel, is followed by a thickener-filter liquid-solids separation. Final leaching is done in two parallel trains of three autoclaves using aqueous ammonia to extract metal values from the adjustment leach residues. This too is followed by a thickener-filter circuit. Here the residues are repulped once with aqueous ammonia and twice with process water and refiltered after each repulping stage, to insure a maximum recovery of soluble metal, before being discarded to the tailings pond. The adjustment leach is followed by a seven-disk Dorr-Oliver American filter, with an agitator in the filter boot to keep the residues in suspension. Agitators have also been installed beneath the four ten-disk filters following the final leach thickener. Slurry Characteristics: A typical analysis of the Lynn Lake nickel concentrate, produced by differential flotation,&apos; is given in Table I. The concentrate comprises 35 to 49 pct pentlandite, (Fe, Ni)S; 3 to 6 pct chalcopyrite, CuFeS,; 1 to 4 pct pyrite, FeS2; and 24 to 42 pct pyrrhotite, Fe1-2S (where x varies between 0 and 0.2) together with various siliceous minerals such as actinolite, hornblende, and talc. During leaching, the sulfur in the pentlandite, chalcopyrite, and part of the pyrrhotite is successively oxidized through the thiosulfate and poly-thionate radicals to sulfate and sulfamate." Simultaneously the nickel, copper, and cobalt are converted to soluble ammine complexes and the iron in the pentlandite, chalcopyrite, and part of the pyrrhotite is oxidized to a spongy hydrated iron oxide. The hydrated iron oxide is readily broken down by the agitation needed to sustain the concentrate in suspension and to maintain a satisfactory mass transfer of the oxygen from the air to the slurry. Table I shows typical chemical, screen, and

Jan 1, 1961

By J. A. Ehler

Speed and torque control of standard A.C. induction motors can be accomplished using a variety of devices ranging from mechanical, hydraulic, and electrical. Also, the motor can be specified as an A.C. wound rotor or D.C. motor using the associated controls. Normally, any one of these drive types can be adequately used to provide the torque control necessary to accelerate a conveyor belt system. The selection will depend on the economics, experience, and preference of the designer or end user. These drive types will be presented from a functional and performance viewpoint to assist in the understanding of the available technology.

Jan 1, 1998

By D. E. Emerson, C. G. Kirkland, E. M. Frost

An apparatus has been developed for determining the helium content of gases containing 40 percent or more helium . It utilizes activated coconut charcoal at liquid nitrogen temperatures to adsorb constituents other than helium in the mixture . Series of analyses have been run where the mean deviation was less than 0.05 percent helium . Twenty or more analyses can be made in an 8- hour day .

Jan 1, 1964

By Junnan Chen, Hongwei Zhang, Wei Wang, Feng Xie

Thiosulfate is a green gold leaching agent that shows great potential to replace the highly toxic cyanide. Ammonia is traditionally used as the copper ligand in thiosulfate solutions to reduce the reactivity of copper ions towards thiosulfate, but it is a toxic reagent. As a result, efforts to replace ammonia with another complexing agent have received increasing attention. In this study, a novel alternative to ammonia is proposed to explore the potential use of the copper-malic acid-thiosulfate system. Experimental parameters include temperature, initial pH, copper, malic acid, and thiosulfate concentration. The feasibility of the copper-malic acid-thiosulfate system was discussed using Eh–pH and species distribution diagrams. Themodynamics analysis indicates that the mechanism of gold leaching in copper-malic acid-thiosulfate solution is mainly catalyzed by the Cu2( Mal)2H-2 2−/Cu(S2O3)3 5− redox couple. The capability of gold leaching in this system for processing the refractory gold concentrate is comparable to that of cyanidation and the traditional copper-ammonia-thiosulfate leaching processes. Thiosulfate consumption is greatly reduced in the presence of malate due to its strong chelating effect on the copper ions. Moreover, the copper-malic acid-thiosulfate system demonstrates a significant enhancement in leaching kinetics, making it a viable option for processing refractory gold concentrates.

Aug 1, 2023

By J. U. MacEwan

Introduction The equilibrium between copper, zinc, and chlorine may be considered as the simultaneous equilibrium of the two separate chemical reactions of formation of the stable chlorides of copper and zinc at the temperature of the reaction i.e., l440°K. (1,167°C., 2,133°?F.)§ [Zn(g) +Ciz(g) ZnCl2z(g). ... (1) ] [2Cu(l)+Cbi&apos;g) Cu2Cl2(l 1 ....... (21 ] Subtracting, [Zn(g) +Cu2(Clz(l i ZnClz(g) +2Cu!l) ... . (3) ] If this reaction is carried out with a copper-zinc alloy at some constant temperature above the melting point of brass, and at a total pressure of one atmosphere, the equilibrium concentrations obtained will depend on the free energy decrease .of the reaction as written in (3) and on the activities of the various components in. the reaction mixture. If the equilibrium is examined from the point of view of the phase rule, it is a three-component system, and if it exists in two phases (liquid and gas) it will have three degrees of freedom. If the restrictions of constant temperature and pressure are imposed, the degrees of freedom are reduced to one. This means that, in a two-phase system under these conditions, each different alloy phase will be associated with a gas phase of constant composition.

Jan 1, 1951

By Earl C. Redman

The Juneau Gold Belt, centered on Juneau and approximately 120 miles long by 10 miles wide, stretches from Berners Bay on the north to Windham Bay on the south (fig. D-l). Access within the Juneau Gold Belt is provided by the Glacier, Thane, and North Douglas Highways. The Glacier Highway extends northwest from Juneau for approximately 40 miles to Echo Cove, and the other two roads end 5.5 miles southeast and 12.6 miles west from Juneau, respectively. None of these roads connect with other highway systems and access to other areas is restricted to airplanes, the State ferry system, and commercial barge lines. BUREAU ACTIVITIES The majority of lode gold produced in Alaska came from the Juneau Gold Belt portion of the Juneau Mining District. To evaluate the economic potential of the Gold Belt, the Bureau located 173 mines and prospects within the area. A total of 263 ad its and shafts were found and 56,650 feet of underground workings were mapped. The Bureau collected and analyzed 1,557 samples during the study. In addition, considerable detailed geologic map-ping was done in the vicinity of the mines and prospects from which ore control models were developed.

Jan 1, 2012

By Harry R. Nicholls

A simple economical method has been developed for determining the in situ longitudinal and shear velocities in rock from which its elastic constants can be calculated. By detonating small charges of high explosives in shallow drill holes near a free surface, both longitudinal and shear waves were generated simultaneously. Accelerometers and velocity gages mounted to the rock surface were used to record the arrival time for both waves over travel path lengths from 50 to 450 feet, Measurements were made in two rock types. From the arrival time measurements, the longitudinal and shear wave velocities were calculated. From these velocities and the density of the rock, the elastic constants were calculated. The dynamic elastic constants of drill core samples of the two rock types were determined in the laboratory by the resonant-frequency method, The lab- oratory and in situ elastic constants are compared and significant differences noted. In general, there is less spread in the in situ data than in the laboratory data, and Poisson&apos;s ratio does not fall outside the theoretical limits for the in situ data as is often the case for laboratory-determined elastic constants.

Jan 1, 1961

By Kendall E. Born

Production of crude oil in Tennessee during 1941 was slightly more than 15,000 bbl., about the same as in the preceding year. Two thirds of the total was from the Mississippi limestone pools in Scott and Morgan Counties. About a dozen small scattered wells in Clay, Pickett, and Fentress Counties, pumped irregularly, produced approximately 5000 bbl. Natural gas was marketed from wells in Morgan and Fentress Counties for consumption in the Sunbright and Jamestown areas, respectively. The production by counties is shown in Table I. Drilling activities in the state fell to the lowest stage since 1933. Twenty-seven wells were spudded in during 1941, of which six were drilling or only temporarily suspended on Dec. 37, 1941. Of the 22 completions, four were gas wells of undetermined quantity, but are classed as commercial, and 17 were dry or near-dry holes. One well encountered oil, but it has not been tested. The total footage drilled during the year was 29,916 ft.; the average depth of wells was 1359 ft. The more important wildcats are listed in Table 2; the distribution of tests according to physiographic divisions is given in Table 3. East Tennessee.—One well was drilled to 3480 ft. in northwestern Anderson County, in the highly disturbed Appalachian area, a region long considered the most unfavorable in the state for the accumulation of oil and gas. Another test is now drilling below 2000 ft. in this general area. Cumberland Plateau.—There were two completions in the Cumberland Plateau area. One test, drilled in northwestern Scott County in an area of former Mississippi limestone production,&apos; encountered a considerable amount of free oil at 850 ft., about the same horizon as in the old abandoned wells. No pumping tests have been made to date. A well at Oneida, Scott County, found gas at 1579 ft. in a sandy limestone of lower Mississippian age. The volume was reported doubled by a shot, although no open-flow measurement is available. Eastern Highland Rim.—Six completions on the eastern Highland Rim resulted in two small gas wells in Fentress County, which will be used to augment the supply for Jamestown. The producing horizons are shaly limestones of Ordovician (Trenton) age. Two wells in this area tested the upper part of the Knox dolomite group, one of which encountered some free oil in a sandy dolomite at 1670 to 1681 ft. A test high on a well-defined structure in Clay County found a considerable amount

Jan 1, 1942