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By R Lloyd

Coal Services (Health and Safety Trust) funded ARRB Transport Research Ltd to investigate fatigue and performance of truck drivers over consecutive shifts. The main questions were to assess: What is the most important contributor to acute fatigue in mining? Is it length of shift (eg eight hours versus 12 hours) or is it time of day (eg circadian effects)? What is the limit of successive day or night shifts before chronic fatigue affects operator performance in open cut mines? The project methodology included utilising ARRB Fatigue Monitoring Devices (forced-choice stimulus reaction) in eight haul trucks, testing 24 subjects over several weeks. Each subject filled in a Fatigue Risk Questionnaire on lifestyle and health. Operators worked a 14-night, one day off, 13-day roster of 12-hour shifts in a fly-in/fly-out operation with excellent opportunities for restorative sleep between shifts. A total of 3500 hours of real-time data was collected, representing the first set of objective, real-time driver performance data for Australian mining. The data showed that for a single night of driving haul trucks, there appears to be no effect on operator fatigue from the amount of time on task (for the 12 hours). In fact performance was worse through the first eight hours than the last four. The circadian influence had a stronger affect than working more than eight hours. Of the 14 nights worked at the site, nights 13, 14, seven and two respectively, were the worst. This presents a conundrum for policy makers who may like to place caps on working hours or rosters in the hope that it will alleviate the fatigue issue. It will not. The major finding within this study is the combined influence of circadian rhythms (human body clock) and individual variability on performance. The lifestyle habits and the health of individuals have a profound effect on operator performance as shown by the strong positive correlation between high fatigue risk scores from the Fatigue Risk Questionnaire and poor performance. An individualistic approach would appear to have the best chance of reducing high potential incidents due to fatigue through successive nightshifts.

Jan 1, 2005

By A. J. A. Janse

The Argyle diamond mine is located in the Kimberley region in the northern part of Western Australia at approximately 16°43'S, 128°23'E in subtropical bush/shrub savannah. The 46 ha (113 acres) pipe is situated in the Halls Creek mobile zone, which was cratonized about 1800 Ma ago and adjoins the Kimberley Block, a plateau upheld by Early to Middle Proterozoic rocks underlain by presumably Archean basement. The crater and upper pipe zones of the diatreme have been preserved within tilted and faulted strata of Middle to Late Proterozoic age. Its age of intrusion is about 1178 Ma. The diatreme rock is not kimberlite but olivine lamproite which was not recognized as a primary diamond host rock until 1978. The Kalumburu Joint Venture, a syndicate of local and overseas mining companies, started in 1972 to prospect the Kimberley Block with a modest annual prospecting budget of $125,000. Annual expenditure doubled each year and reached $1 million in 1975 after which CRA took over management of the newly-formed Ashton Joint Venture with an annual expenditure of $2 million (all figures are in current US dollars). The Argyle diatreme was found near the end of the eighth field season and after a total expenditure of $12 million. Annual production is about 34 million carats which places Australia as the world's number one diamond-producing country, although, because of the low value per carat, it is number four in terms of total annual value. The global cost required for a major diamond mine is between $300 and $400 million, with a discovery every 10 to 12 years. Because Canada contains large areas of prospective platform rocks underlain by Archean basement it is likely that a major diamond deposit does exist there. It is probable that it will be found before the end of this decade and exploration expenditure may be in the order of $100 million.

Jan 1, 1992

By Allan J. Clark, W. J. Sharwood

When the cyanide process came into general use, late in the nineteenth century, chlorination was quickly supplanted, but amalgamation yielded place more slowly, being still the major process at many plants, where it is followed by cyanidation of the tailings. The research work of the Denny brothers in South Africa' was probably the most important contribution of its time to the metallurgy of gold. By grading analyses of mill products, they pointed to the necessity of fine grinding to accomplish maximum extraction by cyanide. The work here described is merely an extension of the Dennys' method. Developing the fine-grinding principle, the steps of "crushing in solution" and of "all-sliming" were logical sequels, though both had been attempted early in the history of the cyanide process. Crushing in solution introduced certain complications in amalgamation; when "all-sliming " was adopted it was observed that the omission of amalgamation did not materially affect the total recovery. Thus the field of the older process has become much restricted, being at present confined to ores which, yielding freely to amalgamation, either are too low in grade to show an economic advantage in "all-sliming," or present chemical or other difficulties making the use of cyanide solution unadvisable during the crushing and preparation of the ore. South African ores are typical of the former class; Homestake ores are of both classes, but even in these instances the amalgamation work is being curtailed and greater burdens thrown on the cyanide plants. It is obvious that the time required for the dissolution of gold depends on the size of its particles. In an effort to determine the critical size, the following laboratory method was evolved. A standard amalgamation test2 had already been established, which involves the agitation of about 100 gm. of ore with 20 gm. of pure mercury, the latter then being separated and its gold contents determined. This method was applied directly to certain mill products with interesting results, a few of which may be noted here. The column "Free Gold" gives the value per ton thus amalgamated, as against the " Total Gold " determined by fire assay.

Jan 1, 1923

By Abdurrahman Tosun

It is very important for after blasting processes that muck pile fragmentation is predicted and controlled during blasting. Therefore, a significant research effort has been devoted to the development of mathematical equations that predict the value of average muck pile fragmentation. The specific charge parameter used in all of these equations is determined as the ratio of the total mass of explosives used to break a volume of rock. The length of blast surface and the row number of the blasthole are ignored in the calculation of the specific charge. However, in a blasting operation, blasthole geometry may result in additional material being impacted by a blasthole. This is due to effect shares of in both end sides of blastholes and the back region of blasting surface. In this study, a total of 14 blast tests, 8 at the first quarry and 6 at the second quarry of the Bati Anadolu (Western Anatolia) Cement Factory, were performed. In each study area blast tests were designed in the same direction continuously to provide stable structural properties of the rock. Therefore muck pile fragmentation depended on only specific charge values for each blast tests. Specific charge values were calculated according to both an average blasthole and also the total volume of material occurred as a result of blasting for each blast tests. During loading until the end of blasted material the front, the back, the arm off and the bucket off pump hydraulic pressure values occurred in hydraulic pistons of the loader were determined with a lot of data received for each blast tests. After that the relationships between specific charges values obtained according to two different methods and hydraulic pressure values of the loader were established with two correlations. The relationships between specific charges values obtained according to the total volume of material occurred as a result of blasting and hydraulic pressure values of the loader occurred as the best results especially the blasting operations that have the total length of blasting surfaces are short.

Aug 1, 2013

By J. R. McWilliams

THE task of recovering most of the contents of an oil and gas reservoir economically presents many and varied problems. In order to attempt intelligently a solution to these problems, an understanding of fundamental physical laws is essential. As the pertinent facts needed in applying these laws to recovery problems are not readily obtainable, observation must be made and the data tabulated. Here, good judgment must be exercised, for at best the information is meager. The engineer is well fitted to undertake the task of solving this problem. In my opinion, the decision on the part of some production executives to allow the engineer to cope with the oil recovery situation and other problems, is an outstanding example of improvement in production methods. The executive head must make this decision or the qualified engineer will be working under a severe handicap with the present system of supervising the production of oil and gas. A qualified engineer should have not only the required technical training but also enough field experience to enable him to understand the difficulties encountered in producing oil and gas. Such a man tactfully cooperating with the practical oil man can greatly improve present methods of oil recovery. The realization that nature's forces have not been efficiently controlled in past recovery of petroleum deposits has recently brought distinct improvement in production methods. Papers are to be presented at this meeting discussing in detail improved methods of conserving gas in both flowing and pumping wells, and also methods of lifting oil after Nature's supply of gas has proved inadequate. The application of back pressure to producing oil wells has not, to my knowledge, been extensively applied in Oklahoma or Kansas up to the present time. Most of the flowing wells are allowed to flow through the casing until the rock pressure has been dissipated. However, some flowing wells have been tubed before the well stopped flowing through the casing. Recording pressure-gages can be installed to study well performance.

Jan 11, 1926

At first glance the staid Joralemon background does not seem to equate with the life of adventure that irresistibly called me for over 70 years. But my ancestors had their adventures too. My father's family were Dutch Reformed Protestants who had energy or daring enough to come to America to escape the Catholic persecution that followed the Massacre of St. Bartholemew. The first Joralemon came from Friesland to New Amsterdam m 1636, just after Peter Stuyvesant. My father's mother's family, the Berdans, were French Huguenots who came to northern New Jersey only a little later to escape the Catholic domination in France. Some say that the first Berdan to come to America was given all the land between the Passaic and the Hacken-sack Rivers by King James. Whatever, the Berdan family settled there and was able to make a scanty living by farming and selling their produce in Manhattan. The spirit of adventure in my mother's family was just as tenuous. Gamaliel Beaman was pastor of a shipload of Puritans who came to Massachusetts in 1636. Gamaliel's descendents were peaceful farmers and some ministers in New England and later New York State. They had got all of the adventure out of their systems in the long and dangerous voyage across the Atlantic to escape religious persecution. The Joralemons and the Berdans married into families that had been in America almost as long as they had. On my father's side the Berdans, Doremus, DeGreys, Vanderbeeks, and Varicks showed by their names their Dutch or French Huguenot stock. The only one who achieved any fame was Peter Varick, who was George Washington's military secretary during the American Revolution. After this he was nine times mayor of New York and was one of the founders of Kings College, later Columbia University. He was responsible for one of my few ancestral claims to distinction. When the Marquis de Lafayette made a fare-

Jan 1, 1976

By Gregory Witte

This paper will explore the technology behind and applications of smart motor control centers (MCCs) being used today in mineral processing facilities. The two technologies to be discussed are solid-state motor controllers that can be installed in MCCs and the complementary digital communication networks that can transmit operational parameters of the solid-state devices to the supervisory system. The toughness and reliability of solid-state technologies have led to a dramatic increase in their usage throughout the mining industry. Products such as solid-state motor protectors, controllers, and variable frequency drives have demonstrated greater flexibility and performance in a variety of applications than their electromechanical and mechanical counterparts. The three key trends apparent in the solid-state devices are: 1. A shift to smaller, smarter, more integrated devices, which allows a single solid-state device to serve the purpose of multiple electromechanical devices but in a much smaller form factor. 2. Distribution of control logic in motor control devices rather than conveyed through a hard wired connection to a central automation system. 3. Embedded network communication that allows supervisory systems to have an expanded view into critical loads and sensors. In applications, the use of smart actuators and sensors with integral local intelligence and communication allow a supervisory controller to send high level commands and let the smart device coordinate its own work The ability of smart field devices to return noncontrolled, nonpoled information opens a whole array of new capabilities for preventive maintenance and diagnostics and forms the basis for an effective motor management program. The smart motor control devices this paper will explore are the smart motor protectors (SMP), smart motor controllers (SMC), and variable frequency drives (VFD).

Jan 1, 2000

By L. B. Anderson, G. A. Carmody

We wish to comment on the paper by G.W. Heunisch (MINING ENGINEERING, January 1987), which raised several important points regarding treatment of acidic waste waters. This paper confirms, in part, tests carried out by Allyn Corp. Allyn is the manufacturer of the proprietary product Alka-Ment, one of the materials that Heunisch tested. The first point is the extreme sensitivity of the sedimentation rate on the nature of the treatment material used. This apparently results from differences in nucleation of the metal hydroxides, and is consistently higher for Alka-Ment. This decreased sludge volume is also observed when Alka-Ment is combined with flocculants, such as organic polymers (see Table 1). Our second point is to emphasize that Heunisch's work shows how dependent the sludge volume is on pH. It is common practice to overtreat AMD by raising pH to 9 or 10. This practice, in addition to being wasteful of material, actually decreases the density of the resulting sludge, probably by creating charged particles that do not readily coalesce. When treating with Alka-Ment, we recommend that the treatment be monitored by testing both pH and dissolved iron. For both tests, it is important to allow adequate time to elapse between treatment and measurement, so that dissolution of alkali, aerobic oxidation of iron, and precipitation of the metal hydroxides are nearly complete. Our tests have indicated that from 30 minutes to one hour is commonly required. Only in this way can a minimum pH be attained, with adequate treatment, denser sludge, and lower treatment rates achieved. Thirdly, Heunisch correctly recognizes that ASTM titration of Alka-Ment does not reflect the true treatment value available in a unit of Alka-Ment. Our work at Allyn indicates this is a most important consideration when evaluating various materials for the treatment of acidic waste waters. While we were unaware of Mr. Heunisch's work, we welcome comparisons between Alka-Ment and any competitive products readily available for field use. Allyn has hundreds of hours of laboratory and field data we are ready to share with interested parties treating acidic waste water.

Jan 1, 1988

R. M. Raymond, New York, N. Y.—The filling and drawing down of the overhead material was done at considerable depth, which is not the usual method in which it starts at the surface.. R. J. Mechin.— It is a little more than 1000 ft. at El Bordo. The vein above had been caved practically to the surface and some of the ore had been excavated, using underhand and overhand square-set methods. This was far enough away from the shaft to prevent any bad effects. When the sill floor was taken out, to insure the old cave material following down well, we put chutes in the square sets for several weeks until we were sure that the material from above was moving down; on each level above, several crosscuts were driven into the old fill by which we could see the material coming down and thus be sure it was following the first. We had to dynamite it very little because it was dry and dusty. The only trouble was that timber starting from the old square sets would get across the chute and hang up. R.. M. Raymond.—HOW many floors did you operate at once? R.. J. Mechin.—We would operate one floor through the whole length and prepare for the second. F. W. SperR, Houghton, Mich.—Could you have worked it in steps? R.. J. Mechin.—Yes; but working it all on one floor was much better for ventilation. The passage between steps would tend to become heavy and blocked and, as the passages from the mining floor down are usually filled with ore, there would be no way for the air to pass up. By means of doors and fans, air was driven to the extreme end of the orebody through the stope and up the other end; but it could have been worked in steps. F. W. SpeRR.—D the following-up material ever become so heavy that you could see weight on the caps before you had to take out the posts?

Jan 1, 1927

RAYMOND M. Howe (author's reply to discussion*).-Dr. Unger states that magnesite will hardly supersede dolomite in fettling after a heat; that it is hard to believe a slag composed of oxide of iron containing about 6 per cent. of phosphorus is the active corroding, agent in a boil in a furnace; and that boiling cannot be said to be the result of the chemical action of a high-phosphorus slag on a basic bottom. These are well-established facts that no one would venture to question. The authors were misinterpreted in the statement that "in the summary, it is claimed that magnesite low in lime, or of greatest purity, ranks first in resistance to slaking," etc. On the contrary, the position has been taken throughout that while a low lime content is advantageous, the presence of several per cent. of iron oxide (an impurity), is necessary in order to confer the proper bonding and sintering properties in dead-burned grain magnesite to be used for furnace hearths. We cannot subscribe to the view that pure magnesia is no better fitted for a refractory than pure lime,' nor that the pure natural varieties of magnesite; such as Grecian or Washington, must necessarily be burned with silica and oxide of iron to produce a satisfactory material. On a small scale, pure magnesia has been satisfactorily used as a refractory in the form of crucibles. In the production of dead-burned grain magnesite from the Grecian or Washington mineral, iron oxide is added but the addition of silica is neither customary nor desirable. It is possible to manufacture brick of good quality from both the Grecian and the Washington material without adding either silica or iron oxide, providing they are burned at a sufficiently high temperature. This is regularly done on a commercial scale both in America and abroad when brick of. exceptionally high refractoriness or special purity are desired. The following analysis is of an English brick of Grecian magnesite which is extremely low in all impurities.1

Jan 6, 1919

By K. -J. Shou

A special three-dimensional boundary element method has been developed for the simulation of the development of fracture zone in rock mass. This approach comprises a boundary element method and a three-dimensional tessellation mesh generator. By using the tessellation generator, a set of pre-defined potential fractures is defined, and the fractures activate when the failure criterion is exceeded. This approach is justified by the reproduction of features observed in laboratory simulations of fracturing in blocks of rock subjected to compression. An attempt is then made to explain some of the attributes of fracture zone surrounding underground openings. Eine spezielle drei-Dimesionen Grenze Element Methode ist entwickelt zu simulieren die Entwicklung der Fraktur Zone in der Steinmasse. Diese Untersuchung benutzt eine Grenze Element Methode und ein drei-Dimensionen tessellation Masche Generator. Durch das Benutzen des Tessellation Generator, ein Satz der vor-bestimmt potentielle Fraktur ist festgelegt. Die Fraktur wird betaetigen, wenn das gescheiterte Kriterium ist uebersteigt und sind in dem Verscheiben-Abbrechen Element eingeschlossen. Diese Untersuchung ist von der Reproduktion der Fraktur justiert, die ins Labor beobachtet wird, um die Fraktur in der von der Kompression unterworfenen Steine zu simulieren. Ein Versuch ist gemacht, um einige Eigenschaften der Fraktur Zone um die untertageoeffnung zu erklaeren. Une méthode tridimensionelle spéciale de élément frontière a été dévélopée pour la simulation du dévélopment d’une zone fracturée dans la massive roche. Cette approche comprend une méthode de élément frontière et un générateur de maillage de la mosaïque tridimensionelle. En utilisant le générateur de mosaïque, une série des fractures potentielles predéfinies est définie avec d’une procédure de la triangulation de Delaunay. Les fractures activent lorsque le critère de rupture est dépassé et sont inclues dans les éléments du déplacement discontinuité. Cette approche est justifiée par la reproduction des caractéristiques observées dans le laboratoire, les simulations des blocs de roche fracturé et sujetés en compression. La tentative est alors fait pour expliquer quelques arttributs d’une zone fracturée alentours les ouvrages soustérrains.

Jan 1, 2003

By David Segura

THE Huitzuco mercury deposits have been described by C. W. Vaupell,1 therefore the general information such as location, history, production, etc. will not be repeated here. The present owners acquired the property when the deep levels of the Cruz mine were flooded, and it was believed that a relatively large tonnage of low-grade ore and fills would be encountered in the flooded section, because of the early mining methods used in the Cruz mine. The mine was partly unwatered and about 187,000 tons of fills was pit-sampled in the San Cayetano stope. These fills average 2.4 kg. Hg per ton. During the period of maximum production at Huitzuco, the ores were sorted in the mine and cobbed into 3-in. pieces, which were treated in Scott furnaces. Large pieces of raw ore are found in the sinter, and it is indicated that both recovery and costs were not very satisfactory. Recent practice has been to treat the ores in closed retorts made of 10-in. pipes. Because of the presence of combined antimony in the mineral livingstonite (HgS.2Sb2S3) and considerable native sulphur, little or no metallic mercury is recovered in the condensers and a second retorting of the condenser muds is necessary. This makes the treatment of the ores expensive and tends to low recoveries of mercury. The principal costs of retorting are, in their relative order: (1) fuel, (2) low recovery of mercury, (3) replacement of retort pipes, (4) labor. METHODS TESTED Three series of experiments have been made, with the object of reducing the costs of recovering mercury and antimony from the ores, by: (1) wet metallurgical methods, (2) bulk roasting in large mechanical furnaces, (3) concentration to reduce the bulk to be followed by roasting in some mechanical furnace. The wet metallurgical methods did not appear feasible because of the high cost of the alkaline solvents used, and investigations of these were not carried beyond the preliminary stage.

Jan 1, 1938

By H. F. JR. Miller

(Butte Meeting, August, 1913.) THE major cause of the deterioration of the open-hearth furnace as its length of service increases, is the melting down, or rather the slagging, of the checker-brick, together with the deposition of dirt in the form of iron oxide, etc., which helps to close the openings, as well as to slag the brick by chemical action. Any construction that will do away with the necessity of having the waste gases come down through the checkers will eliminate the main reason for their deterioration. There are some other well-defined reasons which are subordinate to the main one given above. Some of the defects of the present type of regenerator chambers are as follows: 1. The checkers cannot be cleaned while the furnace is in operation, nor can they be cleaned at the end of the run without taking out all of the checker-brick. This causes high costs for labor and brick. The labor cost is high because all checker-brick have to be handled twice. The brick cost is high because many of the bricks are spalled in the double handling, rendering them useless because a spalled brick will not stand up. 2. The top courses of brick become coated with iron oxide, etc., melt clown, and large amounts of dirt are deposited throughout the checkers, especially at the bottom. The draft of the furnace is decreased by the closing of the top and bottom openings, especially the top ones, and the entire working of the furnace is affected. If an effort is made to clean the checkers the dirt and brick falling down into the holes still further diminish the draft. 3. The distribution of the air or waste gases through the checkers is in many cases uneven, resulting in a lower temperature of pre-heated air and therefore a slower working furnace. Also, the passage of air and waste gases is so crooked that sometimes the chimney is not powerful enough to overcome the friction.

Jan 6, 1913

Zircon is extensively used as moulding sand in the production of steel and high-duty iron castings. To achieve a satisfactory grade of zircon concentrate, the zircon particles may need ,to have surface coatings removed, as such coatings can hinder electrostatic and magnetic separation of zircon from associated minerals. Xn addition the presence of surface coatings on the zircon. particles can cause sinterins and fusion of the zircon to the east metal resulting in "burnt-on" sand or fritting of the casting. "Burnon" necessitates cleaning of the casting by grinding or chipping, which is expensive and results in loss of definition on the casting face.Hudson (1966) has pointed out ,that some coatings on mineral sands are lightly-adhering clays which can be removed quite readily by agitating the sands in water for a few minutes-thus their removal constitutesno problem. However, the heavy gingery coloured coating which adheres strongly to the surface of the particles in some deposits, and which has generally been considered to be a form of iron oxide (Baker, 1962; Hudson, 1966), cannot be removed effectively by attrition cleaning in water (Hudson, 1966). Such coatings limit the usefulness of these concentrates in the foundry industry.An investigation of strongly-adhering coatings on "ginger zircon" concentrates from the Yoganup deposit of Westralian Sands Ltd has shown that the stain consists of a cemented mixtvre of hydrated iron and aluminium oxides and kaolin minerals, and that it can be removed by a relatively low cost treatment.IDENTIFICATION OF THE STAIN BY XRDX-ray powder diffraction (XRD) patterns were obtained from small samples of stain scraped, individually, from the surface of hand-picked zircon and ilmenite crystals. These scrapings were mounted on quartz fibres and exposed to CoKa radiation in a 57 mm Debye camera. To avoid heavy fogging of the x-ray films, a substantial amount of the iron oxides was preferentially leached from the scrapings (before mounting) with an ammonium oxalate-oxalic acid solution (Rooksby, 1961). The residual material gave a fairly well resolved powder pattern (Table 1). A...

Jan 1, 1975

By Sidney L. Hansen

This report was written because of concerns of the possible error involved when stopping leakage is incorrectly assumed to be turbulent. It is common practice to use turbulent flow to characterize each airway branch for computer models of mine ventilation circuits. This report addresses computer methods to accurately model air-leakage paths. A graphical method is described for deriving the equations of leakage through stoppings. The significance of modeling leakage according to the proper flow regime (turbulent or laminar) is demonstrated by the following method. A hypothetical mine was developed to represent present-day conditions. The stoppings of the mine were chosen to leak according to laminar flow properties. The mine was represented by constructing two base models. One base model used turbulent leakage equations for each stopping, while allowing the corresponding resistances to vary. The other base model used the correct laminar leakage equations and held corresponding resistances constant. The mine circuit was then expanded to demonstrate future development. The expanded portion was designed with stoppings that were identical to those in the base mine. The mine in the future state was represented by two projection models. Each projection model was an extended version of its associated base model; one used turbulent leakage equations and one used laminar leakage equations for stoppings. The laminar projection was theoretically correct and is used as a basis for comparison of differences in airflow quantities. Rather than modeling every leakage path for stoppings, a more practical approach is also discussed. A good level of accuracy is maintained by combining leakage paths if the proper flow equations are used. This method was used on the hypothetical mine projection and the flows were very similar to the theoretical values.

Jan 1, 1989

By William L. Barnes, Maurice Leiser

INTRODUCTION The literature on project management considerations and techniques for large rail rapid transit projects in mid-1979 was fairly large and getting larger. UMTA was establishing special organizational elements to deal with the subject as it related to those properties requesting capital assistance, and comprehensive project management plans were being requested from those properties as an early step to build confidence in UMTA that the particular project could be planned, designed, and executed efficiently and effectively, thereby using critically short capital assistance funds to best advantage. In the late 60's and early 70's, the choices of models for heavy rail construction would have been quite limited.., and perhaps much simpler than was the case in 1979. One of the writers was with an agency in 1979 in the process of finalizing its organization for project implementation after several years of study and had the opportunity to review the experiences of others at length, presumably to build on the success and to try to avoid the pitfalls and failures. The properties that were surveyed all reported more or less success stories. One could only conclude that there are many ways to organize for a major transit (construction) project and that there is a great variety of viable management structures, but it seemed that chances for success were best where the executing property had prepared a comprehensive project management plan, then resolved to follow it. Perhaps more important, a commitment or prior approval had been sought from UMTA as to the validity of the project management plan, and the property was allowed to proceed without detailed prior UMTA review or approval of all specific intermediate steps. Most important, however, an early overall UMTA funding commitment had been obtained that permitted some reasonable expectation of the rate of funding, perhaps by commitment

Jan 1, 1981

By M. A. Mallinger

A series of laboratory tests was conducted to determine the limits of practicability in regulating the weights of various cement slurries. It was found that slurry weights of 12 to 19 lb per gallon could be obtained by adding bentonite to decrease the weight and barytes ore to increase the weight. It was also found that barytes ore was not usable in all cements. LIGHTWEIGHT CEMENT SLURRIES INTRODUCTION During the drilling and completion of oil wells the magnitude of the hydraulic pressures exerted on certain formations is critical. In many cases considerable care must be exercised to prevent drilling mud weights from becoming too high, and in those wells where loss of circulation is a problem, the use of normal weight cement slurries will generally result in lost returns during casing cementing operations. While mud weights can easily be controlled, it is necessary to exercise care in the selection of weight reducing additives for cement slurries in order that slurry weights may be reduced sufficiently to prevent the loss of a large portion of the cement to the thief formations without seriously affecting the setting properties of the cement. EXPERIMENTAL RESULTS An extensive search revealed that a large number of materials .that might have been otherwise practicable as a weight reducing additive were not acceptable for use with oil we cements because the imposition of pressure on the slurries caused the additive being tested to be compressed or to absorb water from the slurries, resulting in negligible decrease in the slurry weight. This search for lightweight additives disclosed that bentonite, a material very commonly used for reducing the weight of drilling muds and cements, produced the best results without undue harmful effects on the properties of oil well cements. It was found that there was a limit to the amount of benton-

Jan 1, 1950

By Alfred C. Maevis

On October 27, 1972, ground was broken for the first section of the Second Avenue Subway, a two track underground rail system that was to extend from the south end of Manhattan at Broad Street, north to the Harlem River at 126th Street, a distance of eight miles. This ground breaking was the culmination of 67 years of planning for an east side subway in Manhattan. Following the successful opening of New York City's first subway on October 1904 there was a mass planning for additional subways throughout the City. In 1905 approval was requested of the Board of Estimate for a First Avenue route and Third Avenue route. The Third Avenue route was approved but when action was not taken to start construction, the approval in accordance with the conditions under which it was given was withdrawn. Further reports and plans for an east side subway, usually to be constructed on Third Avenue, were issued in 1920, in 1922, in 1924, etc. Those of us who had worked on the Sixth and Eighth Avenue subways in the thirites and early forties left to join the armed services of World War II in the expectation that we would return and start work on designing and constructing that east side subway. The reason that much of the thrust in those years was to construct on Third Avenue was that there was a Third Avenue "EL" and business interests of the City were in favor of its demolition. This demolition was accomplished in 1955 so one of the incentives to construct on Third Avenue was removed; but an east side subway was always part of the master plan of the City.

Jan 1, 1976

By T Martin, M Martin

Mining of sulfide-based ore deposits typically creates a potential for release sulfate and acid to groundwater, and many operating mines are thus predicted to require long-term post-closure treatment of groundwater. Toward the goal of selecting an optimal treatment, this paper compares costs of four demonstrated technologies for long-term treatment of sulfate-contaminated groundwater: in situ biological reduction (ISR), in situ permeable reactive barrier (PRB) composed of compost and zero-valent iron (ZVI), and two ex situ æpump-and-treatÆ technologies ù ex situ biological reduction, and ex situ nanofiltration with lime precipitation. Costs for each technology are estimated for treatment of a model aquifer system containing sulfate-contaminated groundwater, with the goal of treating 1 m3/min of groundwater from 1500 mg/L sulfate to a target level of 500 mg/L. (Metal removal is typically achieved by these technologies, and thus excluded from this analysis.) Costs are estimated for each technology as a function of treatment duration (ten to 60 years) and, for PRB and ISR, required replacement frequency (ten to 60 years). Estimated costs include capital and operation and maintenance, using a seven per cent net discount rate. Results indicate that ex situ technologies (biological reduction and nanofiltration with lime precipitation) are the most expensive. In situ technologies may be much less expensive, with PRBs having by far the lowest potential cost, followed by ISR as the next least expensive. However, actual costs for the in situ technologies are very sensitive to poorly constrained parameters, particularly the ZVI content of the PRB, and the effective life of PRB and ISR systems. This large cost uncertainty supports delaying remedial actions where viable, and suggests that feasibility analyses are generally warranted for specific applications to focus on these sources of uncertainty.

Jan 1, 2003

By Charles E. Weaver

Discussion of the paper of CHARLES E. WEAVER, presented at the San Francisco meeting, September, 1915, and printed in Bulletin No. 103, July, 1915, pp. 1419 to 1427. MILNOR ROBERT, Seattle, Wash.-A year ago at this time, 12 wells were being drilled for oil in western Washington. Some excitement attended the operations; many thousand acres of land were leased to oil companies and stock was widely advertised. The location of the wells may be described by referring to the conclusions in Dr. Weaver's paper, in which the State is divided into four areas. In the area described first, which includes the Cascade Mountains and eastern Washington, where conditions are decidedly against the existence of commercial oil fields, no new drilling was reported. In the second area, the Puget Sound Basin, where conditions are unfavorable, but to a less degree, five wells were being drilled. The third region, covering certain parts of southwestern Washington, contained four drilling wells and several others projected or beginning operations, in spite of the absence of petroleum seepages. None of the wells mentioned is fully active today. On the western side of the Olympic Peninsula, along the Pacific Ocean, are three wells which have been the scene of operations for the past two to three years. The region they are prospecting contains seepages of light oils and is the fourth area described in the paper. The difficulties of transporting both men and supplies to the well sites have delayed operations and brought the costs up to high figures. The drilling of these three wells is interesting on account of the presence of oil in nearby seepages and its occasional appearances in the material penetrated. A. C. McLAUGHLIN, San Francisco, Cal.-The most important part of Washington is on the Olympic Peninsula. The geological conditions on the Peninsula are not similar to the conditions in California, and it would not prove to be a continuous belt of oil-bearing territory from there to here if oil should be found at either end.

Jan 12, 1915