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By E. D. Gardner

"The Bureau of Mines has studied blasting practices in metal mines of the Southwest. The investigation of blasting drift rounds showed among other results, that cut-holes in ground difficult to break could be chambered to advantage under some conditions.In open--cut mines, chambering nearly all holes before blasting is the common practice. Two or more free faces are generally maintained in this type of work, and the amount of explosive necessary to break any bench or bank can usually be estimated from the yardage to be moved. However, in blasting drift or tunnel rounds only one free face is available before the cut-holes are shot, and the quantity of explosive required is generally ascertained best by trial.In addition to being the most difficult to break, the cut-holes are the most important of a drift round. When the cut-holes ""break bottom"" less work is required of the other holes, and a full break of the round is generally obtained. In ground difficult to blast, extra care and expense are justified to insure that the cut-holes ""break bottom.""At many mines the work is so organized that all blasting is done at the end of the shift, and a decided advantage would have to be gained to make chambering of the cut-holes economical. Also, in many mines blasting during the shift would not be practical. One or two holes could probably be drilled during the time necessary to do the chambering. However, under certain special conditions, such as in small operations or those where ""leasers"" or contractors are working, drill holes could be chambered should there be an advantage in doing so. Moreover, in some mines it might be practical to chamber the cut-holes at the lunch hour. In ordinary rounds the cut-holes are generally drilled by noon. It would not be necessary to tear down the drill or remove tools from the face to do this work of chambering."

Feb 1, 1928

By J. -B. C. Varnier, J. A. Cremeens, D. D. Overton

"Slope failures that occurred in 2008 and 2009 at the Pikeview Quarry in El Paso County in Colorado in the United States were investigated using three-dimensional discrete element modeling methods. Large-scale block sliding failures occurred along weak bedding planes that dip steeply into an excavation located at the toe of the slope. Several mining sequences at the toe of the slope were simulated to evaluate the effects of various degrees of extraction and backfill on slope stability. Slope stability back-analysis was conducted using prefailure and post-failure conditions to determine the joint shear strength properties. Laboratory testing was also conducted to determine intact rock properties. Three-dimensional discrete element analysis was uniquely applicable to this evaluation because it allowed representation of shear stress of the basal sliding surface and afforded evaluation of several scenarios of extraction and backfill at the toe of the slope. IntroductionThe Pikeview Quarry is an aggregate quarry located approximately 8 km (5 miles) northwest of Colorado Springs in El Paso County, CO, USA. The quarry has been operated by several owners since the early 1900s. According to the U.S. Mine Safety and Health Administration (MSHA, 1986), the quarry has been a major source of construction aggregate for the Colorado Springs area for many years. In the 1950s, the site was mined by Peter Kiewit and Sons to provide limestone for the construction of the nearby Air Force Academy.A large-scale slope failure occurred on Dec. 2, 2008, which stopped mining activity at the Pikeview Quarry. A second, related slope failure occurred on Sept. 13, 2009. The second slope failure was located immediately adjacent to the initial December 2008 failure.The objective of this paper is to discuss the three-dimensional slope stability evaluation. The causal relationships of slope instability are not within the scope of this publication. However, omitting the precursors to slope failure would limit understanding of the nature of the slope stability evaluation. The regional and local geology and the history of slope instability at the Pikeview Quarry aid in understanding these precursors. Two main failure mechanisms were hypothesized during the failure investigation to explain the triggering of the 2008 and 2009 large slope failures at the Pikeview Quarry, and both of these mechanisms were given equal credence in the present study. Quarry excavation and backfilling was a key factor in the slope instability for both failure mechanisms. The excavation was backfilled in stages in an effort to provide slope support. The objective of the three-dimensional model was to evaluate the effects of extraction and backfill at the toe of the slope with respect to slope stability."

Jan 3, 2018

By Y Cui, G Cao, H Zong, J Zheng, J Wang, S Yue, J Zhao, B Li

The iron (Fe) concentration in nickel slag is high, making its recovery and utilisation a hot and challenging topic. In the nickel smelting process, using CaO instead of SiO2 to adjust slag’s composition, and then directly reducing Fe from the output of molten slag is a more feasible method in practice for Fe recovery in nickel slag. The adjustment of slag’s composition in the nickel smelting process will have an impact on the distribution of impurity elements such as Pb, Zn, and As. In this paper, the slag adjustment of nickel smelting process was carried out with the oxygen-enriched topblowing system. The distribution of impurity elements in the smelting products with each composition slag was conducted by using thermodynamic software. Experimental verification was performed with rich oxygen top-blowing, and the distribution characteristics of impurity elements in the smelting products were characterised to explore the differences in mineral phases of impurity elements with each smelting conditions, elucidating the distribution patterns of Pb, Zn, and As in each phase. The results indicate that after slag adjustment, impurity elements are more easily to volatilisation and removal into the gas, reducing their total concentration in slag and matte. In the slag, the existence form of Pb, Zn is changed from silicate to ferrate, and As is changed from arsenate to low-valent arsenic oxide. It is easier to decompose the impurity-containing mineral phase and remove the impurity elements in the subsequent iron extraction process.

Jun 19, 2024

By Leonard Kellogg

Introduction The Mine Manager or Superintendent is a man faced with a difficult problem. He must direct the activities of his mine in such a manner that, over a period of time, the greatest possible profit will be realized. In this respect, a mine is no different from any other business organized for profit. What makes the mine manager's task so difficult then? The profit of a mine is determined by -among other things -the grade of ore, the market price of the metal or metals produced, and the cost to operate the mine. The mine manager cannot do much about the first two of these, but he has considerable control over the operating cost. This cost will depend in large measure on the nature of the programme on which he plans to operate his mine. This operating plan may be broad or specific, on a long-range basis or on a month to month basis. The development of the plan, however, is made very difficult by the existence of so many factors which must be considered. These may be (1) factors requiring decisions based on the best judgment and interpretation of engineering and geological information available, or (2) items which must be considered but which are largely non-controllable. To be more specific, examples of such items would be: (1).-(a) The amount and type of development work to be done in relation to amount of ore to be mined. (b) Mining methods to be followed and type of equipment to be used. © Amount of exploration diamond drilling to be done if the orebody is not clearly defined. (2).-(a) Rock type and degree of hardness. (b) Width, depth, and pitch of veins. © Amount of work requiring timbering. (d) Faults encountered. (e) Water and gas conditions met with.

Jan 1, 1952

By G. J. S. Govett

World mineral production is dominated by six countries - the U.S., the U.S.S.R , Canada, South Africa, Australia and China - which also hold most of the reserves of the world?s minerals. For the developed and centrally planned countries, there is a strong positive correlation between the mineral production of a country and its land area: there is no correlation between mineral production and land area for the less developed countries. These facts imply that, after taking due account of geological controlling factors, a country may expect to produce minerals in approximate proportion to its land are a provided that political, economic, and historical factors are conducive to mineral exploration and mining. Contrary to popular political belief in many parts of the world, the developed countries are the major source of mineral imports for the industrialized countries. Moreover, exports of ores and minerals account for a relatively small par t of world trade, and are an important factor in the export picture of only a few of the less-developed countries. The looming resource confrontation between the developed and the less-developed countries is based largely on the false premise that the industrialized countries of the world are minerals-poor and owe their wealth to cheap minerals from the less-developed countries. The data presented refute this. Geological controls are relatively minor in the present distribution of world mineral production; political and economic factors are dominant. It is argued that an expansion of world trade in minerals - based on increased exploration and new forms of finance for mining ventures - rather than cartels, price-fixing and trade restrictions is by far the preferable solution to the problem of redressing the legitimate grievances of the poor countries of the world.

Jan 1, 1977

By H. C. Chang, N. J. Grant

Creep of very coarse grained high purity aluminum was studied at 400°, 700°, and 1100°F at initial stresses of 50 to 1200 psi. Local strain measurements were made across the grain boundaries and in different regions of various grains during creep, thereby permitting the drawing of component creep curves. The relative amounts of elongation as measured across a grain boundary and in the grains adjacent to the boundary depend on the temperature of testing. The significance of creep curve equations and activation energies is discussed. THIS paper presents some of the results of a research program devoted to the study of the mechanism of creep in high purity aluminum. Previous publications have dealt with the cyclic processes of alternating grain boundary sliding and migration, and the deformation in the grains caused by these processes;' the direction and the driving force of grain boundary sliding and migration;' the deformation in the grains and the restricting effects of the grain boundaries on this deformation." These previous papers were based mainly on either microscopic observations or X-ray data, or both. Sometimes creep curves across two closely spaced reference marks (about 0.6 to 0.7 mm apart) were also presented with a view to correlating the elongation values with the optical observations in general, and with certain deformation processes in particular. These curves shall henceforth be referred to as component creep curves. No attempts were made, however, to compare either the component creep curves which were obtained in different regions of a specimen, or those obtained from different specimens subjected to different creep test conditions. It is the purpose of this paper to present these component creep curves. Among the factors influencing the behaviors of these curves, those studied were stress, temperature, and grain arrangements in the specimens. It is proposed to show how greatly the deformation can vary from grain to grain and within an individual grain in a coarse grained specimen. The experimental technique and procedure have been presented.' It remains necessary to emphasize that the reference marks were produced by pressing a thin sewing needle into the surface of the specimen before the final annealing at 900°F for 24 hr and at 1150°F for 12 hr. The cold work around the needle indentations was therefore removed. The depth of these holes after electropolishing the annealed specimens was about 0.01 mm. This depth and area were negligible in comparison with the thickness of the specimens, 2.3 mm, and hence would not affect the creep behavior. Results Fig. 1 shows a development diagram of specimen P-8. The dots on the surface locate the needle point reference marks used to obtain the component creep curves. A curve obtained between two dots across a grain boundary measures not only the boundary deformation but also the deformation associated with boundary sliding occurring in the grains in the region between these indents. The original distance between two reference marks was about 0.6 to 0.7 mm. The maximum error of measurement was 0.01 mm, or in terms of percentage elongation, of the order of & 1.5 pct. Whereas the component creep curves shown in the figures to follow were drawn as a gross average of the points, in view of the fineness of some of the steps of grain boundary sliding and migration, as shown in the previous publications,¹,2 it appears logical to expect that some of the finer cycles of movement have not been delineated for the strain sensitivity of the measuring system. It seems justifiable to say that the curves are not as smooth as shown. 700°F Tests: Fig. 2 shows component creep curves across four of the grain boundaries in specimen P-8, tested at 700°F and 85 psi. These curves clearly show alternate cycles of increasing and decreasing

Jan 1, 1954

By T. L. Myron

In recent years increasing attention has been given to improving both the chemical and physical properties of sinter. The first step toward this goal was the production of a self-fluxing-sinter. More recently, blast-furnace operators have found that significant improvements in blast-furnace performance can be achieved by improving and controlling the physical size of the sinter. The results of full-scale blast-furnace tests have shown that blast-furnace production can be improved by using a closely sized agglomerate. The improvements obtained by the use of pellets, a closely sized agglomerate, are well known. In contrast, blast furnaces charged with sinter having a wide size spectrum have generally not yielded results as satisfactory as those from furnaces 1) * charged with pellets. Beginning in 1958, Appleby-Frodingham charged plus 3/8-inch sinter to their blast furnaces. Further tests by other companies in Canada, 2) the United States, 3) and Australia showed that blast-furnace production could be improved by about one percent when the plus 3/8-inch material in the burden was increased by one percent. Because of the promising results obtained with screened sinter, an extensive test program was initiated to determine an effective means of producing sized sinter in U. S. Steel's sinter plants. This paper describes the work done by the Applied Research Laboratory in this program.

Jan 1, 1965

By Seedsman R

The fundamentals concepts of roof support are similar between coal mining, metal mining and tunnelling and yet there are different design approaches, different hardware, and different levels of site investigation. This should not be the case and this paper discusses the key geotechnical issues that should underlie all roof support design - the need to reinforce defects. A design method to build a beam in laminated rock under high horizontal stresses is presented. The method allows for changing rock strengths, defect properties and varying horizontal stresses. At one level the design can be implemented with design charts and tables. Lessons learnt during the implementation of the designs at 5 different mine sites are Discussed.

Jan 1, 1998

By Frederick B. Kuhnow

Harmonic energetic delamination a building is being considered for demolition at a formerly known atomic power lab. The building is within a containment structure under a negative pressure, so that no radioactive material is leaked out to the environment. A new approach was developed to make this work possible.

Feb 1, 2020

By J. Fletcher Smoak

Consumption of abrasive materials in the United States in 1981 was approximately $340 million, of which 37 % was industrial diamond (natural and synthetic), 39% manufactured abrasives, and 24% natural abrasives. Production and shipments of natural abrasives, excluding emery and industrial diamond, decreased in quantity 9% and 7%, respectively, when compared with that of 1980. Emery showed the largest change in output, decreasing 30% in quantity and 20% in value. Production of nonmetallic .manufactured abrasives material plus shipments of metallic abrasives material decreased 5% in quantity but increased 4% in value. Nonmetallic manufactured abrasives consisted of aluminum oxide (fused) and crude silicon carbide produced in the United States and Canada and accounted for 63% of the value of all manufactured abrasives. Metallic abrasives shipments included chilled and annealed iron shot and grit, steel shot and grit, plus cut wire, aluminum, and stainless steel shot and equaled 37 % of the value of all manufactured abrasives.

Jan 1, 1982

By M. E. Lombardi

IN comparison with the mining industry the petroleum industry is new and inexperienced, and until now it might have been called the fortunate industry. Its great good fortune consisted in two things; a bountiful supply hidden by nature in the rocks of the earth, and a steady and large increase, during the past twenty-five years, in the demand for petroleum products. It is a remarkable fact that consumption of gasoline in the United States has shown an increase every year since automobiles first became important, and even in 1930 there was a four per cent increase, in spite of the universal depression; the estimated increase if the times had been normal, was ten per cent. This industry has, therefore, not passed

Jan 1, 1931

By Jim Galvin

4 and 6 point timber chock constructions are used extensively on both a systematic and spot basis by Australian longwall operators to support tailgate roadways. In 1996, the School of Mining Engineering at UNSW published design and performance criteria for timber chocks constructed from Australian hardwoods. The research was funded by the Australian Coal Association Research Program (ACARP). In 1997, ACARP awarded funding for a stage 2 testing program. This testing focussed on examining the effect of using thinner timber elements in chock constructions to improve OH&S and comparing the full-scale performance of traditional 4 point chocks to Link-n-Lock constructions. It was established that chocks constructed from thinner elements have failure loads much less than thicker elements. However up to the failure loads of the thinner elements there was no significant difference in the response to convergence of chocks constructed from 25, 50 or 100mm thick elements. Chocks constructed from 150mm thick elements have a similar ultimate strength to the 100mm thick elements. However, they have a significantly lower resistance to load, thereby permitting more convergence to occur. There is more potential for chocks comprised of thin elements to be constructed 'out of plumb' and so suffer eccentric loading. The full scale testing program established that at any given level of convergence up to 10% strain, the timber in a Link-n-Lock chock generates twice the support resistance of the same quantity of the same timber in an equivalent size 4 pointer chock. On the basis of support resistance per cubic metre of timber, there is only a marginal difference in the performance of 1.0m Link-n-Lock chocks and 1.2m Link-n-Lock chocks constructed from select Blue Gum. The cost per tonne of support provided by a Link-n-Lock chock is effectively half of that provided by a 4 pointer chock of equivalent dimension and timber. Up to a strain level of 4%. Link-n-Lock chocks constructed from 1.2m long elements are slightly more cost effective than the same chocks constructed from 1.0m long elements. Thereafter, there is no significant difference in cost per tonne of support resistance. Beyond a strain level of about 4%, Link-n¬Lock chocks constructed from landscape grade Blue Gum are just as cost effective as Link-n-Lock chocks constructed from structural grade 4 Blue Gum.

Jan 1, 2000

By Jing Zhong, Yugang Li, Chen Yan, Shaoda Zhang, Qian Wang

Online surface quenching technology has been developed for the hot charging process to prevent the surface cracks in high strength low-alloy steel slabs. In this paper, a two-dimensional heat transfer model of surface quenching process was presented. This finite element model includes nonlinear thermodynamic properties, by which the slab temperature distributions were computed. The model predicted temperatures show reasonable agreement with the measurements. The effects of the water flow rate and slab movement velocity on temperature variation during the quenching and subsequent tempering process were investigated. The result shows that the temperature drop increases but the tempering temperature changes slightly with increasing water flow rate and decreasing slab velocity. Keeping the slab movement velocity at 1.2-2.1m/min and the water flow rate at 55-70m3/h, the slab surface experiences a temperature drop of 400-600 firstly, then recovers above 650, the quenching and energy-saving effect are remarkable.

Jan 1, 2015

It is well known that the crushing and grinding steps in the mineral processing of ores represent a substantial proportion of total energy, cost and carbon emission in mining. For instance a recent study calculated that 36% of the energy consumed by gold and copper producing mines in Australia was attributable to comminution (Ballantyne et al, 2012). Understanding this component of a business is essential in a world seeking to lower production costs, improve earnings and reduce its carbon footprint. The issue commands the attention and support of industry leaders in the mining sector, government, equipment manufacture, engineering design and research organizations. Furthermore, several researchers have often stated that globally, comminution processes consume between 3-4% of the world?s electrical energy production (Daniel and Lewis-Gray, 2011). In 2009-10, Australia?s Federal Government Energy Efficiency Opportunity (EEO) participants in the mining sector used 336.5 PJ, representing approximately 6 percent of Australia's total energy use. It is evident that the current practices of breaking rocks in tumbling mills are relatively inefficient (Fuerstenau and Abouzeid, 2002). Research and improved engineering design has established that a range of improved blasting, crushing and grinding techniques and flow sheets may lower project costs and carbon footprint. These include relatively simple process flow sheet strategies such as gangue rejection or pre-concentration prior to size reduction, a better combination of grinding technologies (both current and new systems), and selection of coarser grind sizes where ore and waste mineralogy allow. Yet, despite this systematic and detailed research, the industry remains somewhat reluctant to implement these strategies. CEEC (Coalition for Eco-Efficient Comminution) was established in 2011 to support knowledge sharing and change in an area of high-energy consumption for the global mineral industry. CEEC is a not-for-profit industry company, funded by grants from world wide industry companies, whose mission is to accelerate knowledge and technology transfer in the field of energy-efficient comminution. The key benefit to the mining sector arising from CEEC?s activities will be an increase in the rate at which more energy-efficient comminution data sets, processes, and technologies are developed, demonstrated, shared and applied. This will contribute directly to the industry?s target of reducing operating costs and carbon emissions. CEEC International Ltd and JKTech Pty Ltd teamed together to facilitate the inaugural CEEC Workshop, in June 2012. The purpose of the workshop was to produce a roadmap to provide guidelines for how (1) operations can achieve best EEC practice in the short to medium term, in collaboration with vendors, engineering companies and other service providers, and (2) what R&D is needed to achieve paradigm change in EEC in the medium to long term, e.g. 75% reduction in comminution energy in 10 years.

Sep 1, 2012

By D Osborne

Current mineral extraction and valorisation activities still have a strong focus on the ‘delivered product(s)’ which might be a bulk commodity such as metallurgical coal, hematite, bauxite, etc., and mineral concentrates such as metal sulphides, oxides, and carbonates; precious metals, such as gold, silver and PGMs, REEs or battery metals; halogen salt forming elements; or industrial minerals, which are non-metals including crushed rock, sand, and gravel etc. With the current focus on tailings disposal the pushback has been to seek alternatives to the currently ‘conventional’ disposal method via increasingly higher risk tailings settling facilities (TSFs) with a changing focus towards ‘dry disposal’. Whichever option is selected, the alarming trend is the increasing cost and ESG impacts. An emerging approach is to pose a focusing question ‘Is there anything else passing into the tailings stream and/or disposable waste streams which could be recovered and pass into the product revenue stream? All of the above ‘primary’ process streams offer opportunities for increasing recovery and offsetting some of the disposal costs. Some secondary opportunities are already being pursued, such as REEs and battery metals from base metals deposits, but other opportunities are still being overlooked, such as building materials and soil conditioning pellets. These potential ‘new’ product streams can also utilise other mining and domestic wastes such as smelting slags, glass waste, plastic waste etc. This paper will address these opportunities and provide examples including process approach, economics and some of the current barriers that exist, which are holding back progress on such initiatives.

Sep 1, 2024

By A. B. Hooker, L. C. IlsLey

In lare-scal e blasting, it is customary to fire me.l'\Y shot.!:l simu1. taneously. The electric !atonators ara usually conneced in seies, ard are fired either from a power circuit or :from a blasting ma.chine with large capacity, operated by hand. "!hen :'::ri111; with a blasting ca.chine, trouble with missed shots has been freqi.,.ently e>:pert0nced.. Unit1:i:' thes ,:::rcu:ns tances, the shots that fail to fire are often a gr"up in the t.:E"nter r,f t>e series, Various explaz,.a.tions hs.ve been given as to the cause of the failurs. Tis paper covers a few tests made cy the electrical section of tl:.e Dti..'t'cau. cf Mines, wh;.r.;h may be helpful in understanding the -probable cuse of tl1e misfiring of c.:anter shots, Wh3tever th9 source ot the firir. current generator, magneto, or batter, the CUl"rent delivered th:·01.1.gh a shot-f:?·ing circu.it is determined by t.ae total resistance of ,;he clrcili +.; c:1nd all cure:it leaving th9 p,si tive terminal must return to tre negative te1-minal. If a. metallic circuit is well i1,.sul.:tted, praeti­ccl. l;1 a.ll of the t.'Urrent keepo within the circu.it; but in a. shot-!11·.irig ci.rc1li t, Nhere the ·.vires f:ll·e often 1n c.c,i.cact with the eru:"!ih as i11 wet holer., end whero the ins".lla.tion of the leg wires :i. inadecpa.te to prc·,ent lea.l.tage, the current ma.v tu.e v a.r ious pa the: thro-.. .. gh the nr r. earth.

Aug 1, 1922

By E. N. Harshman

Uranium mining and milling are rather new industries in Wyoming and South Dakota, but, in the past 15 years, ore valued at more than $400,000,000 has been mined and milled in the region. The major uranium deposits are in structural and/ or erosional basins or on the flanks of uplifts. The principal host rocks are medium- to coarse-grained sandstones that were deposited in continental or brackish-water environments and were derived in large part from the granitic cores of the uplifts and the sedimentary rocks that flank the cores. The ages of the host rocks range from Cretaceous to Miocene, but more than 95 per cent of the known reserves in the region are in rocks of early Eocene age. The bulk of the ore in the region is unoxidized and lies below the water table; some oxidized ore has been mined in the Powder River Basin and Black Hills districts. The unoxidized ore comprises pyrite, uraninite and/ or coffinite, marcasite, hematite, jordisite(?), and selenium. Minerals in the oxidized ore include hydrous uranium carbonates, phosphates, and sulfates and hydrous uranium and/ or vanadium silicates. The deposits are in the form of ore rolls that contain from a few hundred to a few hundred thousand tons of ore, ranging in grade from 0.1 per cent to 1.0 per cent U3O8. The rolls are at the margins of large tabular bodies of sandstone that has been altered by the orebearing solutions. The deposits are believed to have resulted from the interaction of an alkaline; oxidizing, ore-bearing ground water and a reducing environment. Exploration guides include the spatial relations of ore to altered sandstone, favorable lithology in the host rock, the content of certain elements and/ or compounds in ground water, and the relation of deposits to major paleodrainages.

Jan 1, 1968

By Brochot. Stephane, Manuel Gonzalez Fernandez

"HPGR (High-Pressure Grinding Rolls) is widely used today for grinding of various kinds of ores such as iron ore or base and precious metal ores. It can be used as first stage of grinding replacing tertiary crusher and SAG mill, but also as regrinding stage to enhance mineral liberation. In addition to better energy efficiency, the advantage of HPGR is to produce much more internal particle fractures allowing faster leaching as it has been demonstrated for some gold ores. This grinding mode of compression and fracture generation also promotes mineral liberation without overgrinding. Mineral processing modelling and simulation associated to detailed ore characterization constitute a powerful tool for plant design and optimization. The mathematical models of the units of operations (grinding and separation) have to reproduce the mineral deportment for which the mineral liberation plays a central role. The present paper is focused on the HPGR modelling with the objective of plant optimization or advanced design from pilot plant tests. The traditional approach of the population balance linked to the energy dissipation and the residence time distribution is generalized in association with a mineral liberation model. The effect of the specific HPGR breakage matrix and grinding function on the mineral liberation rate is discussed and compared to the tumbling mill. The implementation of this mathematical model into a plant simulation tool is illustrated by the example of the optimization of an iron ore concentrator."

Jan 1, 2014

By Ivan F. Avery

OUTPUT of carbon black in 1959 increased in all producing States except New Mexico. The national increase was 20 percent. Domestic sales and exports increased 22 percent and 16 percent, respectively, resulting in a 21 percent gain in sales. Sales to the rubber industry, which purchased 95 percent of the domestic output of carbon black, increased 23 percent. Sales to all other major consumers also increased. Stocks decreased 82 million pounds during the year. [ ]

Jan 1, 1960

By R. V. Higgins

The, large volumes of low A. P. I. gravity oil that will remain in many shallow fields of California when such oils no longer can be recovered profitably by pumping are important potential sources of oil when ways and means are found to remove them from the reservoir sands. Estimated ultimate oil recoveries in many of the "low-gravity fields" of the San Joaquin Valley in California indicate that less than 25 percent (on the average) of the original oil will be recovered. These oils have viscosities of 1,000 or more centipoises at reservoir temperatures and, because of their high viscosity and low intensity of the reservoir pressure, the quantities that will be retained in the sands ultimately are so large and their recovery is such a challenge to petroleum engineers that the Bureau of Mines has made the study of their possible recovery one of its major projects. The fact that a slight increase of temperature materially reduces the viscosity of low A. P. I. gravity oils has 'prompted the use of various means for increasing the temperature of the oil in the reservoir sands and hence the rate of flow and Completeness of recovery of the oil. Hot oil, hot water, heated air, and natural gas, steam, super-heated steam, and the products of combustion have been circulated in wells to raise the temperature of the oil it the sands; electric heaters have been installed in wells, and butane and air have been intermittently exploded and the flames propagated through the tubing to the "pay" zones. None of the methods, however, have been entirely satisfactory for the recovery of large quantities of oil.

Jan 1, 1949