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By W. B. McCarter

Broadly speaking, the year 1934, one of the most active in the history of the Coast, has been both disappointing and revealing. Of the 18 new oil-producing areas, only two have shown the qualities of a first class oil field. Many of the new areas are discouraging for the horizons exploited, because of close association of gas and salt water, and the absence of any appreciable oil zone. In spite of amply fulfilling the expectations of drilling and geophysical development, the year has fallen far short of the results that normally should be expected from such a campaign. In 1934, the Cockfield zone has been disappointing because its exploration down the dip from the Conroe and Tom Ball fields revealed only such distillate fields as Splendora, Eureka, and the great gas-sand thickness at Katy. The zone proved good at Raccoon Bend. Newer Oligocene and Frio fields in Texas, with the exception of Hastings, were also disappointing, as Van Vleck arid Old Ocean attest. The discovery at Dickinson is not startling. The development of Hastings, Tom Ball, Dickinson and the deep zone at Raccoon Bend will be slow. A new strip of prospective territory of minor importance lying inland from the Conroe trend was opened by the Kittrell discovery. During the year 206 unproductive wildcats were drilled on the Texas Coast, compared with 207 during 1933 and 141 during 1932. During 1934, there were 835 oil and gas wells and 229 dry holes drilled in proved fields, in contrast with 1100 in 1933 and 512 in 1932. The past year's production was actually 60,868,260 bbl., or 2,020,764 bbl. less than in 1933. The present annual rate of Texas Gulf Coast production is about 61 million barrels per year. The daily Texas Coast production decreased 5537 bbl., or 3.2 per cent, during the year. The total daily production at the end of the year was 173,598 bbl. Eighteen wildcats opened up new oil and gas pools during 1934, so that there are now 78 active pools and 3753 producing oil wells. The production of the average well is 45 bbl. per day. The price of crude oil was almost uniform in 1934,

Jan 1, 1935

By J. S. Hardison

In the fall of 1912, the appearance of water in the oil of the Nevada Petroleum Co., Coalinga, Cal., made necessary the installation of a dehydrating plant to reduce the water below the 3 per cent. limit prescribed by the agency. Unlike the mining industry, technical literature of the oil industry is limited and extremely unsatisfactory. Until the recent efforts of the Petroleum and Gas Committee of the American Institute of Mining Engineers, no concerted movement has been instituted to secure publication of papers dealing with the problems of the oil business, and because of the additional fact that dehydrating of oil has not been practiced in California to any extent up to comparatively recent time, it was found necessary to experiment in order to determine the most satisfactory plan for this purpose. It is easy to write regarding successful enterprises, but, while not SO pleasant, it is equally desirable to write of failures so that others may be saved the loss incident to such investments. The water in this oil occurs both free and as an emulsion. Free water easily settles out, but the emulsion requires treatment. So far as can be determined, the emulsion consists of globules of water surrounded and enveloped by a film of oil. Starting with this hypothesis the theory has been evolved that in order to break up this emulsion the water must be heated at least to boiling point, when an explosion takes place destroying the globule. Unfortunately, this condition was not recognized in the installation of the first plant, which was planned on the following lines: A heater was arranged so that all the oil from the wells could be heated before reaching the shipping tanks, these being already fitted with coils for steam heating. Tanks were filled with about 4 ft. of water, live steam was turned into the coils and the water brought nearly to the boiling point, when the heated oil was run in. This for a time was partly successful, but the length of time necessary to apply this heat was too great for practical operating.

Jan 1, 1915

By Sidney L. Wise, Walter Strache

While the installation of mine bulkheads to retain water under high pressure is by no means a rarity, the following points which arose in the designing and placing of two of these bulkheads may be of interest: The Hibernia magnetite mine, located about 40 miles west of New York, in the State of New Jersey, is partly fillcd with water. This mine is located on an ore lens, the outcrop of which is over a mile long. The ore has been mined from this lens from the surface to a depth of more than 1,500 ft. It was held desirable to separate the old workings from the new, and to allow the former to fill to the 850-ft. level. Two weak places existed between the new and old workings below this Ievel: namely, a temporary bulkhead on the 10th level and a rock bulkhead of indeterminate thickness on the 16th level. When the matter of allowing the old workings to fill was investigated, the fact was developed that if these old workings were filled, the bulkhead on the 10th level would be subjected to a water pressure of about 50 lb. per square inch, and that 200 lb. per square inch would act on the 16th-level bulkhead. As the above-mentioned barriers were not deemed of sufficient strength to permit of these pressures, it became necessary to design and install new bulkheads. Design As the 16th-level bulkhead was required to withstand a pressure of 200 lb. per square inch, it presented some difficulties. A careful consideration of the various types of mine dams now in use led to the adoption of a design of the form of a truncated wedge. In this, the pressure side of the dam is of greater area than the hack, so that the resultant action is similar to driving the wedge. By cutting generous skewbacks in the walls, roof, and floor, this type in reality becomes an invisible arch. The wedge feature tends to compress the materials in the bulkhead, thereby adding to its imperviousness.

Jan 1, 1915

By Pol Duwez, Howard E. Martens

INTRODUCTION THE high temperatures encountered in the operation of jet engines have imposed most drastic requirements upon the materials used in their construction. There are two different approaches to the materials problem connected with the design of parts in contact with high temperature gases. The first approach is to use a material which will perform satisfactorily at a temperature comparable to that of the gas. The second approach is to use a cooling system so that the material is maintained at a temperature appreciably lower than that of the gas. Besides the conventional method .of cooling by means of a liquid in contact with the material exposed to a high rate of heat transfer, a less orthodox method consists of making the part to be cooled of a porous material, so that the cooling fluid can be forced through the pores. In this scheme, the temperature of the coolant moving in a direction opposite to that of the heat flow increases gradually while passing through the porous material, and the coolant forms a protective layer on the surface exposed to heat. This method, referred to as "sweat cooling," was proposed at the jet Propulsion Laboratory in September 1944, and the experimental work on the preparation of porous metals was started early in 1945. The sweat cooling method opens new lines of research in powder metallurgy by extending the field of application of porous metals. The porous metals available at the present time are used mostly as bearings or filters, and a rather limited number of alloys are made for these applications. In addition, the fabrication of bearings or filters does not require a very dose control of the permeability, while in the case of sweat cooling such a control is of primary importance. In designing a sweat cooled part it is imperative to assure a given rate of flow of coolant under certain conditions of pressure drop. Generally, the proper pressure drop and the rate of flow of coolant are deduced from heat transfer and design considerations, and consequently the permeability of the porous metal to be used is also determined. It is therefore necessary for the designer to have available a continuous series of porous metals covering a range of permeability as wide as possible. Contrary to the requirements established in the fabrication of filters, the actual pore size of a porous metal intended for sweat cooling is of little consequence, except for its effect on permeability. The tensile strength is likely to be of great importance for some applications. Since the tensile strength of a porous alloy decreases with increasing porosity, an important problem is to develop porous alloys having the maximum permeability with minimum porosity.

Jan 1, 1948

By M Frenzel

At the last MassMin Conference in Sudbury, Caterpillar’s vision of a continuous mining concept for block caving operations was presented on the background of global raw material demand and its continuous increase. By analysing today’s structure of mining processes in block caving, the paper (Krings, Steinberg and Frenzel, 2012) highlighted the main bottlenecks in the process chain and identified conceptual solutions by help of a case study. The main focus was the paradigm shift from discontinuous batch type operation to a continuous mining process of the blocks.In 2016, the strategic vision has taken major strides forwards and remote continuous operation has become a reality. Caterpillar’s continuous operation system – the Rock Flow system – was successfully installed in Codelco’s Andina mine and start of operation is scheduled in the first half of 2016.Based on continuous product developments, this paper sets the focus on two main aspects of continuous mining. First, field experiences of the Rock Flow system will be shared, highlighting the Andina project course from system installation, to commissioning and successful system start. The second aspect concentrates on the system and mine layout as discussed in the case study scenario in MassMin 2012 (Krings, Steinberg and Frenzel, 2012).Based on Caterpillar`s continuous product development, the Rock Flow system set-up and panel design were modified for a general roadway design and drawpoint layout. The new design allows an optimised use in block preparation phase (spreading) and full production phase (reaping), operating a fleet of load-haul-dumps (LHDs) and Rock Flow systems. The combined use of the two technologies optimises the mining process in every stage of block life and significantly increases the performance of the continuous block caving operation in the production panel to pave the road for the next generation of super caves.CITATION:Frenzel, M, 2016. New continuous production system in block cave mining – the rock flow system, in Proceedings Seventh International Conference and Exhibition on Mass Mining (MassMin 2016), pp 543–548 (The Australasian Institute of Mining and Metallurgy: Melbourne).

May 9, 2016

By Theodore H. Eyde

The porphyry copper province of the southwestern United States and northern Mexico, is the largest copper producing region in the world. Most of the deposits occur in Arizona which now produces over 54% of the newly mined copper in the United" States. As the many recent discoveries are brought into production, Arizona's share of the nation's copper production will increase. The term "porphyry copper" as used in this paper includes all deposits of generally low grade disseminated copper mineralization amenable to bulk mining techniques. Therefore, copper deposits such as Pima-Mission, Twin Buttes, and Vekol where the copper mineralization occurs principally in the intruded rocks will be considered as "porphyry copper" deposits. Most of the deposits have been discovered since World War II, consequently, the southwest, and particularly southern Arizona is perhaps the most intensively explored area for copper deposits on earth. Most of the major natural resource companies in the United States and Canada maintain exploration offices in southern Arizona. To illustrate the success of exploration efforts, the location of porphyry copper deposits either known or producing prior to World War II are shown on figure 1 and the deposits discovered since World War II which are either in production or in the exploration and development stage are shown on figure 2.

Jan 1, 1973

By E. C. Skinner

Cardox, which consists essentially of a steel tube containing carbon dioxide compressed to the liquid state, is a trade name designating a device used principally in coal mines to break down coal. The use of Cardox to break down coal was developed in the Middle Western coal fields around 1927 or 1928. At first, a large, heavy tube was used, requiring many parts. Through experimentation and technologic improvement, the weight of the tube has been brought down from 88 lb. to 20 lb. and the new, light tube makes use of larger volumes of CO2 at lower pressures than formerly was thought possible. This blasting method is extensively used today in coal mines and probably one tenth of the bituminous coal produced in the United States is prepared with Cardox. Construction and Application As used in the coal mine, Cardox looks like a piece of steel pipe 2 to 3 in. in diameter and 3 to 5 ft. long. When it is necessary to fire a shot, the tube is wired with two leg wires that are long enough so that when the tube is pushed the desired distance into the hole the wires will protrude 5 to 6 in. It has been found that No. 18 insulated copper annunciator wire is the best material for leg wires. A two-conductor rubber-covered shot-firing cable of sufficient length is then attached to the leg wires. A 2 to 6-volt current is applied, either by means of shot-firing batteries or the standard type of shot-firing generator. Upon the application of the current, the charge in the cartridge is set off. Because the breaking action upon the coal is supplied by the relatively slow physical expansion of the carbon dioxide gas rather than by the rapid chemical formation of gases as in more familiar explosives, the result is a slow heaving action, somewhat similar to that produced by black powder, rather than a shap detonating effect. Permissible powder does its work at roughly ten times the pressures in use with Cardox. The escaped gas from a Cardox shot apparently works along the planes of weakness in the coal face, such as the coal cleats, the edges of the cut, and along the top of the seam, breaking off the coal toward the free faces and heaving the mass of coal outward. Cardox tubes are manufactured in various sizes to take care of variable conditions of seam size, undercut depth, cleatage, coal hardness, and so forth. The tubes range in diameter from 2 to 3 in. (Table I) and the normal CO2 charge is from 24 to 96 oz. For any one tube size, several sizes of shear disks are available, thus excellent control of the type of shooting desired is possible. In general, local experience and experimentation dictate the size of tube chosen for any particular job. For high coal, and deeper or wider cuts, a larger size is chosen to start with, and the opposite is true for low coal, softer coal, or small cuts. After the tube size has been experimentally selected, tests are run in the mine to determine how many shot holes are needed and where the holes should be placed. The size of the disk,

Jan 1, 1944

By E. C. Skinner

Cardox, which consists essentially of a steel tube containing carbon dioxide compressed to the liquid state, is a trade name designating a device used principally in coal mines to break down coal. The use of Cardox to break down coal was developed in the Middle Western coal fields around 1927 or 1928. At first, a large, heavy tube was used, requiring many parts. Through experimentation and technologic improvement, the weight of the tube has been brought down from 88 lb. to 20 lb. and the new, light tube makes use of larger volumes of CO2 at lower pressures than formerly was thought possible. This blasting method is extensively used today in coal mines and probably one tenth of the bituminous coal produced in the United States is prepared with Cardox. Construction and Application As used in the coal mine, Cardox looks like a piece of steel pipe 2 to 3 in. in diameter and 3 to 5 ft. long. When it is necessary to fire a shot, the tube is wired with two leg wires that are long enough so that when the tube is pushed the desired distance into the hole the wires will protrude 5 to 6 in. It has been found that No. 18 insulated copper annunciator wire is the best material for leg wires. A two-conductor rubber-covered shot-firing cable of sufficient length is then attached to the leg wires. A 2 to 6-volt current is applied, either by means of shot-firing batteries or the standard type of shot-firing generator. Upon the application of the current, the charge in the cartridge is set off. Because the breaking action upon the coal is supplied by the relatively slow physical expansion of the carbon dioxide gas rather than by the rapid chemical formation of gases as in more familiar explosives, the result is a slow heaving action, somewhat similar to that produced by black powder, rather than a shap detonating effect. Permissible powder does its work at roughly ten times the pressures in use with Cardox. The escaped gas from a Cardox shot apparently works along the planes of weakness in the coal face, such as the coal cleats, the edges of the cut, and along the top of the seam, breaking off the coal toward the free faces and heaving the mass of coal outward. Cardox tubes are manufactured in various sizes to take care of variable conditions of seam size, undercut depth, cleatage, coal hardness, and so forth. The tubes range in diameter from 2 to 3 in. (Table I) and the normal CO2 charge is from 24 to 96 oz. For any one tube size, several sizes of shear disks are available, thus excellent control of the type of shooting desired is possible. In general, local experience and experimentation dictate the size of tube chosen for any particular job. For high coal, and deeper or wider cuts, a larger size is chosen to start with, and the opposite is true for low coal, softer coal, or small cuts. After the tube size has been experimentally selected, tests are run in the mine to determine how many shot holes are needed and where the holes should be placed. The size of the disk,

Jan 1, 1944

By John Greenawalt

IN view of the increasing importance of sintering in the beneficiation of iron ores preparatory to their reduction in the blast furnace, the writer believes the time is opportune for an up-to-date, though brief, discussion of the general principles of sintering, as well as a discussion of some of the recent improvements and mechanical developments that have been achieved in this industry, its present trend and future possibilities. Sintering may be defined as the agglomeration of fine mineral particles into a porous mass by incipient fusion caused by heat produced by com-bustion within the mass itself. Its first successful application, in connec-tion with the treatment of lead sulphide ores, was known as blast roasting. In this process the lead ore was given a preliminary roasting in a reverber-atory furnace, then the incandescent mass was placed in a pot with a grate and an updraft was applied. It was soon discovered that much better results were obtained by cooling the ore, moistening it, and placing it upon an incandescent layer of fuel previously placed upon a grate, and then applying the updraft blast of air until the charge was converted into a more or less agglomerated mass. The fuel was sulphur and the process one of desulphurization, which was the principal result desired. The application of downdraft to blast roasting brought about tremen-dous results; the principal one of which was that an air blast of high pressure could be passed through a layer of ore deposited on a grate, or on a grate covered by a layer of coarse porous material, without disturbing the particles of which the layer is composed. The air current is compelled to filter through the pores of the charge, whereas if the same pressure applied to the top surface of the charge with downdraft were applied to the bottom of the charge with updraft, the charge would be completely blown off the grate. Another result of reversing the draft was that a flame could readily be applied to the top surface of the charge for ignit-ing purposes.

Jan 1, 1938

By Brian J. Gorval, Robert L. Kemeny

INTRODUCTION Financing, or providing adequate capital at low cost for developing and bringing a natural resource property into production, is a fundamental requirement for profitable operation. Under the present economic environment the cost of financing any project is gaining more importance than ever. The fundamental keys for success in mining are "the Four M's": Mineral, Market, Management and Money. In recent years, with capital and operating costs skyrocketing, increased emphasis on obtaining adequate and low cost capital funds for bringing in a prospect is essential. For a small mining company there are limited sources for raising funds. In most cases these companies have neither a substantial asset base, nor any cash flow, and they are deemed as risky investments. The issue of shares of capital stock of such a company is probably the most common method of raising funds via the public purse. Alternatively, debt could be used and a source of such capital might be private investors or financial institutions. The company which intends to raise funds from the investing public by way of the sale of equity, subsequently lists those shares for trading on a stock exchange. Going and staying public commits a company to a multitude of complex laws, rules and reporting requirements. The governing bodies are the watch dogs for the public. The largest of such an organization of this kind in the world is the Securities and Exchange Commission (SEC) in the United States. In Canada, the Securities Laws come from within provincial jurisdiction. Most provinces have their own Securities Act, and Securities Commission. It is the Commission which administers the Act, and hence controls the distribution and trading of securities within the province. Extra provincial registration is required for trading and distribution in more than one province. In British Columbia, prior to a public offering, the approval of the Superintendent of Brokers (SOB) and of the VSE must be obtained. THE VANCOUVER AND TORONTO STOCK EXCHANGES There are five stock exchanges in Canada which supply a trading forum in which mining companies may operate. The stock exchanges are located in Montreal, Toronto, Winnipeg, Calgary, and Vancouver. Of the five stock exchanges, the Vancouver Stock Exchange (VSE) and the Toronto Stock Exchange (TSE) are the most heavily used for resource-based issues and are the focus of this paper. The objective of this section is to provide a brief background on the VSE and TSE. The Vancouver Stock Exchange The VSE is Canada's predominant stock exchange for financing resource-based, junior issues. Since 1907, the VSE has acted as a source of funds for mining, exploration and energy companies throughout Canada and the world. Currently, there are over 1400 companies listed on the VSE. During 1983, over 3.1 billion shares traded at a value exceeding CDN$3.96 million. The VSE has developed a reputation as one of the few exchanges where a speculative resource venture may obtain a listing and subsequent financing in a relatively short time, generally five to ten months. The SOB, the Investment Dealer's Association of Canada (IDA), and the VSE all regulate trading at the Exchange. Any irregularities in a company's stock dealings are investigated and trading may be halted by the VSE until a satisfactory explanation has been received from the suspended company. Companies can be suspended from trading or delisted for various reasons. One reason is misrepresentation of facts in documentation submitted to the SOB, VSE, or to the public, through news releases.

Jan 1, 1985

By K. W. Stöckelhuber, A. Wenger, H. J. Schulze

"An important elementary step of the flotation process is the forming, drainage and rupture of a thin intervening liquid film between bubble and mineral particle.After the encounter of the bubble with the solid surface a wetting film is formed. The initial thinning of this film is driven by the capillary pressure inside the bubble. When the film has reached a certain thickness rupture can take place.There are two possible mechanisms describing this process:-, Growing fluctuation waves (spinodal dewetting) on fluid interfaces under the influence of any kind of attractive forces (electrostatic, van der Waals). -, Nucleation inside the film.Metastable wetting films on negatively charged, hydrophobic glass surfaces (gaseous phase methylated), or on hydrophilic, positively charged glass surfaces (with Al3+ ions), have been analysed by means of the kinetics of film thinning according to the Reynolds law. These experiments demonstrate that both mechanisms can be responsible for the rupture of a thin wetting film. In the case of hydrophobic surfaces the nucleation mechanism is preferred; in the case of opposite charged silica surface the capillary waves mechanism due to an attractive electrostatic double layer force between silica and air bubble takes place.Hereby the existence of a long range hydrophobic force. on a hydrophobic surface can be excluded. The apparent interaction can be explained by the presence of gas nuclei formed on heterogeneous surface sites. The results provide deeper insight into the mechanisms of wetting film stability, adhesion process in flotation and droplet coalescence."

Jan 1, 2003

By Cayo Soto

Mining activity is important for the development of humanity. This development must be understood from the Academy where it allows showing alternatives for the industry performance. Likewise, these mining companies need to improve their results. Historically, a few years ago the application of the ESG (Environmental, Social and Governance) criterion was verified, which is positioned as a factor/regulator in the design of the performance of industries in general. ESG as an integral part of the concept of sustainability serves as a tool to conceive mining without conflicts and with a reputation based on trust in the interrelation of the university, industry and society. Therefore, the academy must strive for new knowledge in a way that allows it to continue in the virtuous circle of constantly feeding it. In this context, the best mining practices are implemented in constant questioning of their activities. The scientific essence of the industry supported by the academy developed in an operational dynamic is addressed. Finally, the objective of this research work is to interrelate 3 variables: ESG, Academia, and the Mining Industry, where the transition of knowledge and the role of each variable in relation to each one can be understood. This allows for an additional understanding of mining in the future with new mental attention.

Feb 1, 2025

By Edmund Cardoza

In the spring of 1993, construction began on an ambitious building project in Boston, Massachusetts. The Shriners organization committed to offer state-of-the-art adolescent burn treatment and rehabilitation in its Boston facility by expanding in a most unusual way. In short, the concept included installation of caisson foundations and perimeter slurry wall panels to enable construction of a completely independent structure around and above the existing, continually functioning, hospital. Two distinct phases of work were detailed in the Contract Documents. At the completion of the first phase, patients will be moved into the new building, above, and demolition of the old structure, beneath, will be performed. In Phase I, sections of the slurry wall and caissons, located outside the footprint of the existing building, were planned. The next phase of the work includes the installation of the remaining caissons and slurry wall. Once excavation is completed, the balance of the new hospital structure will be constructed from the bottom upward, connecting to the under side of the new structure.

Jan 1, 1994

By T. J. Ballard

The Tallapoosa Copper mine in Haralson County, Ga., has been operated, or under development at various times since before the Civil War. The geology of the deposit is similar to that of many of the pyrite deposits occurring in the belt of ancient crystalline and metamorphic rocks that extends diagonally northeast-southwest across the State. The Bureau of Mines, in its program of investigation of, strategic mineral deposits, diamond-drilled four holes totaling 1,570.3 feet at the deposit during the period June to September 1947 to verify the old drill records and the continuation of the deposit at depth.

Jan 1, 1948

By J. E. Danoczi

Shore Gold Inc. (Shore) is presently designing a diamond processing plant for the Star and the Orion South kimberlite deposits in the Fort à la Corne (FalC) area in central Saskatchewan. The first step in the plant is diamond liberation and occurs in the comminution section. Diamond liberation is the process of breaking up the kimberlite rock, as gently as possible so as not to damage the diamonds. Autogenous grinding mills (AG mills) use a combination of impact crushing, abrasion and attrition to break up the kimberlite. The forces that occur during the actions of abrasion and attrition cause the kimberlites to break up preferentially along lines of weaknesses which offers a ?diamond friendly? method of comminution. Hard competent kimberlite is commonly found in the diatreme zone as well as the root zone of the volcanic pipe and is the source of kimberlite for most of the Southern African diamond mines. Softer kimberlite is typically found in and around the crater area of a volcanic pipe and it is this kimberlite that makes up the majority of kimberlite for both the Star and Orion South deposits. These softer kimberlites can be readily broken up using abrasion and attrition imparted by AG mills. In order for Shore to benefit from the softer kimberlite found in the Star - Orion South diamond project, two large (10.97 m diameter x 7.10 m long) AG mills were designed for the comminution section based on extensive test work and simulations by Metso Minerals (Metso).

Jan 1, 2012

By Francis R. Conley, John A. Sievert, John N. Dew

A brief review is presented of the past performance of a number of large, thin, highly permeable reservoirs with low dips in the Bolivar Coastal fields of Venezuela. The performance of these reservoirs indicates that the fluids are segregated vertically within the sand section by gravity. With this assumption, equations are developed which describe the performance under pressure maintenance operations. Methods of solving these equations and results of simple example calculations are presented. Example calculations indicate that under pressure maintenance conditions injected gas tends to {tow preferentially along the top of the sands and that encroaching water has a tendency to flow preferentially along the bottom. The expected performance of segregated fluids is discussed and compared with that of fluids which are uniformly distributed in any sand section. INTRODUCTION The field performance of a number of reservoirs described herein indicates that the fluids are segregated vertically within the sand section by the force of gravity. Thus, it was felt that any method used to predict the future performance of these reservoirs should consider the effects of segregation in the sand sections. A study of the available information on some of the reservoirs suggests that increased recoveries may be expected if the reservoir pressure is maintained by either crestal gas injection or flank water injection. To predict future performance of reservoirs under pressure maintenance operations, a method of analysis was needed which would account for segregation of fluids in the sand sections. Several methods of analysis have been developed to take into account the segregation of fluids in the reservoir as a whole1,3 To our knowledge no method considers segregation of the fluids within sand sections in the manner indicated by the past performance of several reservoirs in the Mara-caibo Basin. This paper outlines part of the work done in studying some of these reservoirs and contains a description of their performance characteristics. The analysis presented is restricted to pressure maintenance conditions, since space limitations prevent a full discus- sion of the development of the mathematical relations and the various methods of solving them. Only a sketch of the development of the mathematical relations is given. The various methods of solving these relations are pointed out, but the actual determination of solutions to various problems are omitted except for one example. It is felt that these results may aid in the study of reservoirs outside the Maracaibo Basin. Some concepts on which the present analysis is based are outlined by D. N. Dietz. However, the analysis presented herein includes several factors not considered by Dietz: (1) variations in permeability and in the cross section, (2) various shapes of the cross section, and (3) the production of fluids. Also, the mathematical development presented by Dietz differs considerably from our corresponding analysis. RESERVOIR CHARACTERISTICS The reservoirs under consideration, large and thin, with low dip and high permeability, are large to the extent that they contain from 0.5 to over four billion bbl of oil initially in place. The section thicknesses vary from 100 to 400 ft and the lengths from 10,000 to 40,000 ft. Thus, a length-wise cross section of the producing formation appears to be long and thin. The dip angles vary from 0 to 6 degrees and the average permeabilities vary from 0.5 to over three darcies. The reservoirs contain from 20 to 50 per cent shale inter-laminated with the productive sands. Most shale breaks within the major sands do not correlate from one well to the next, and correlation is often difficult, even between wells drilled from the same location. However, it is often possible to correlate the shale breaks between major sands for some distance. Herein we are concerned with the performance of reservoirs containing oil of gravities over 20" API. Past reservoir performance indicates good pressure communication and, in cases where pressure sinks have developed, large amounts of fluid migration have occurred. Few of the reservoirs have had initial gas caps, but also, few have been found to be highly under-saturated. Due to gravity segregation, secondary gas caps usually form before the pressure has been reduced more than 25 per cent of its initial value. The effect of gravity has not only been apparent on a reservoir-wide basis, but has also caused segregation of the reservoir fluids in productive sections. Proof of this is found from the results of selective well tests, workovers, and electric and various other types of logs. Evidence that

By T. C. -M. Cheng

Iron removal and disposal in zinc plants has imposed a major environmental challenge to the industry. Among the existing options of precipitating iron as jarosite, goethite, and hematite, the Hematite Process seems to be the best candidate for the future development of a waste-free technology, because of its potential of producing saleable iron oxide residues for feed in iron-steel making industry. Unfortunately, under the current practice of the Hematite Process, it is not possible to produce iron residues clean enough to meet the specifications of the said industry. In this paper, the practice of industrial hematite precipitation and the relevant technical literature are reviewed with the objective to identify the critical processing parameters which are responsible for the contamination of hematite, particularly with sulfur and zinc. The analysis is conducted from a crystallization theory standpoint with emphasis given to the role of supersaturation with respect to the cleanliness of hematite precipitates.

Jan 1, 1997

By A. Redfern, B. G. D. Smart

The ultimate objective of the development work presented in the paper is to devise an appropriate method for relating powered support specification to geological conditions and mining practice, thus enabling the longwall method of coalmining to be successfully implemented throughout the varied coalfields of the world. The paper begins with a review of various methods previously used to determine support rating, each method being analysed with regard to its source, the logic and theory which constitute the method, and subsequent recommendations. An alternative approach is then proposed, the Powered supports being considered to be one of three foundations, sharing load with coal ahead of the faceline which acts as another foundation, and the waste behind the supports which acts as the remaining foundation. The load/deflection characteristics of the coal and waste, which in turn are related to geological conditions and mining practice, are shown to affect the load transfer on to and hence the specification of the powered support. In particular the characteristic of the waste foundation is shown to be affected by the caving mechanism, an alternative form to that mechanism accepted for British conditions being proposed and designated "parting-plane-controlled" caving. A finite element solution for a beam on elastic foundations was used to translate the above concepts into numerical results, these results being compared with data from Indian and Chinese operating conditions. The Indian case study examines the effect of using a conventional chock type of support in a one-web back mode when mining towards a roof break generated ahead of the face, while the Chinese case study compares measured subsidence over a longwall with "blocky" roof conditions with the model results. The comparisons are encouraging, suggesting that the approach used is worthy of further development.

Jan 1, 1986

By J. E. Dutrizac

The dissolution of low-iron sphalerite, (Zn,Fe)S containing 0.08% Fe, in ferric sulphate-sulphuric acid media was investigated using closely sized fractions of crushed sphalerite crystals and sintered disks made from the crushed crystals. Linear kinetics are observed, and the predominant reaction products are ZnSO4, FeSO4 and elemental sulphur. The leaching rate increases systematically with increasing temperature, and the calculated apparent activation energy is 56 kJ/mol for the sized sphalerite and 54 kJ/mol for the sintered disks. The relatively high apparent activation energy suggests that the rate is chemically controlled, a conclusion supported by the insensitivity of the rate to the disk rotation speed. The rate increases as the 0.22 power of the Fe(SO4)1.5 concentration, and as the 0.22 power of the H2SO4 concentration. However, in the absence of ferric ions, the rate increases more significantly with increasing H2SO4 concentrations, and relatively rapid rates are observed in solutions containing >0.5 M H2SO4. In 0.1 M Fe(SO4)1.5-0.3 M H2SO4 media, the measured rate decreases with increasing sulphate concentrations, added as Li2SO4. The rate of dissolution of low-iron sphalerite is promoted by dissolved copper concentrations as high as 800 mg/L. In contrast, dissolved silver, even in the 0-2 mg/L concentration range, significantly retards the dissolution of the low-iron sphalerite.

Jan 1, 2008

By M. T. Abasov

The work suggests an approach to the problem of optimally planning the drilling and development of gas fields. The approach is based on systems analysis and allows a joint optimal choice of the total number of wells in the field, their spacing in the reservoir, the diameter of production tubing, drilling dynamics, the number and spacing of drilling platforms and well pattern schemes for each plat- form as well. The optimal solution satisfies the set of technical- economical requirements and provides a minimum value of discounted total expenditures (capital costs plus operating costs). A mathematical model of the problem is presented. This work represents an extension of the authors&apos; previous investigations that formulated and solved problems of optimal drilling of oil and gas deposits by directional wells [1] and problems of optimal control of the number of active wells in multi-reservoir oil and gas fields [2], [3]. For construction of the mathematical model we choose a range (Nmin, Nmax) of possible values of the total number of wells, N, which should be drilled in the deposit. Here Nmin, Nmax are minimum and maximum numbers of wells correspondingly, that should be drilled into the reservoir. N[€](Nmin = N1, ... N[o], ..., Nm = Nmax). For simplicity a one-reservoir gas field is considered. It is assumed that: (a) wells are spaced in the reservoir uniformly, (b) there is a certain algorithm providing a uniform spacing of the given number N[o] ([0]=l, . . . , m) of wells in the reservoir, (c) the diameters of the production tubing for all the wells are the same. The relation Qz (x), i.e. the dependence of the cumulative output of gas from the cumulative "well-years" of operations is constructed for each diameter dZ from the given range of diameters d1< ...< dr.

Jan 1, 1977