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By R. J. C. MacDonald

"The first recorded discovery of placer gold in Canada was on the Chaudiere River in Quebec's eastern townships in 1832 and the first major production came from placer deposits in British Columbia in 1858 (Verity 1974). Placer deposits were the major source of gold until the late 1890's, although lode deposits were found in Nova Scotia in 1862. In 1896 the Klondike Gold Rush commenced with the discovery of gold by Henderson and Carmack in the Yukon. The peak was reached in 1900 when over one million ounces of gold valued at approximately $22.3 million were recovered. With the exhaustion of the easily accessible deposits, more capital intensive operations were necessary such as those initiated in 1906 by the Yukon Gold Corporation. This giant dredging operation continued to produce until 1966. Today only a few small operations remain.The development of Canadian lode gold mining began in earnest around 1900 when the railroads were opening up the interior and western part of the country. At this time many discoveries occurred, particularly where mineral outcroppings were near surface. Tremendous impetus for exploration and the reopening of former producers followed the elimination of the gold standard by Great Britain around 1931, and more importantly the United States in 1933, with subsequent stabilization of the price of gold at $35 per ounce in 1934. Prior to this the price of gold was basically unchanged since 1717, approximately, $20 per ounce (Verity, 1974; Green, 1973).The depression of the 1930s adversely affected most industries, however for the gold producers it was a very profitable time. Prices of materials and equipment declined and labour was cheap and plentiful. Interest in gold mining was further stimulated in 1936 when a three year tax exemption was declared for all new metal mines. In 1938 a decline in the value of the Canadian dollar relative to the U.S. dollar benefitted the industry. The second world war provided yet again further incentive for gold production and in 1941 Canadian gold production reached a high of approximately 5.4 million ounces. This figure was not surpassed until 1991 when approximately 5.6 million ounces were produced (Natural Resources Canada, 1996 )."

Jan 1, 1998

This paper reviews the current techniques employed in the prediction of subsidence and how they have been applied to partial extraction systems. The method using influence functions to describe the shape of the profile is investigated in detail and has been adopted to model the subsidence over workings in Huntly. It is compared with the profile methods developed in the United Kingdom and found to predict the subsidence profile more accurately than these methods.

Jan 1, 1991

Silver is one of the three major metals produced from the Broken Hill mines. The accurate determination of the silver in these ores, therefore, is of economic importance.With due appreciation of the divergence of opinion of many who are interested in the subject of the determination of silver, the writer has been prompted to present the results of some recent experiments on silver losses during cupellation and conclusions from these experiments in the present paper. The losses of silver in the fire assay for this metal are due to silver retained in the slag during the fusion or scorification of the ore, and to silver volatilized from, and absorbed by the cupel during the process of cupellation.The literature describing the results of experimental work on silver losses due to the above causes is extensive, and the general conclusion of these experiments is that the silver lost in the slag is small, but that lost by cupellation is larger and more variable.The term cupellation loss in this paper refers to the total cupellation loss, no distinction being made between the absorption and volatilization losses.Volatilization loss is shown by King to be 0.5% of the total cupellation loss and, as it is generally accepted among assayers as being only a small fraction of the cupellation loss (3) (5), it will not receive further consideration in this paper.The factors causing the loss of silver by cupellation and the precautions to be obs~rved to reduce the loss in this operation are described in (1), (2), (3), (4) and (5). The...

Jan 1, 1940

ignitions caused by cutting tools have occurred at a rate of some 14 per year in UK coal mines since 1965, mostly on the coalface. The rate has held steady despite reducing output and number of collieries. Investigations were carried out over this period into factors affecting both the concentration of firedamp in the cutting zone and the igniting source. A study into detecting and suppressing ignitions on shearers has been undertaken and it is feasible to prevent certain ignitions from spreading outside the cutting zone. Practical measures have been adopted by British Coal to identify frictional ignition sites and to lay down working procedures to minimise risk.

Jan 1, 1988

By M. Melashvili

"The SART Process was introduced to the gold mining industry in the late 1990s and several commercial plants were built over the following 10 to 15 years. Most of these plants have operated well and enjoyed plant performance in line with expectations based on the underlying chemistry of the process. However, two of the plants have consistently suffered from high reagent consumption and poor product quality, which has motivated a re-examination of the process fundamentals in the hope of understanding the reasons for poor performance when it occurs, and developing mitigating strategies.This testwork along with a review of the relevant literature has demonstrated that the amount of sulphide added to the primary reactor, relative to the copper concentration, is the most critical parameter dictating plant performance, with pH and residence time important secondary factors. The primary product in the SART Process is synthetic chalcocite, Cu2S, and the stoichiometric requirement for sulphide addition is therefore 0.5 moles per mole of copper in solution. Testwork shows that when sulphide is added in excess of this amount (generally with the express purpose of lowering residual copper in solution) it not only fails in its intended purpose, but can have a negative impact on the settling and filtration characteristics of the copper sulphide product, as well as the final product quality. Moreover, these problems are exacerbated the higher the selected operating pH in the pH 3 to 5 range, and the longer the residence time of slurry in this pH range.Freshly precipitated copper sulphide compounds are known to be very fine and slimy, with poor settling and filtration characteristics. It is therefore important to engineer the SART Process in such a way as to encourage growth in the average particle size of the precipitate crystals. This was the rationale for recycling a portion of the thickener underflow to the primary reactor when the SART Process was developed. The theory of crystallization predicts that compounds of extremely low solubility, such as chalcocite (Cu2S: KSP = 10-48), are likely to experience growth in crystal particle size through aggregation rather than heterogeneous nucleation. On this basis, the SART practice of recycling a portion of thickener underflow to the primary reactor in the commercial plants is unlikely to lead to crystal growth via nucleation, and could in fact promote breakage of the aggregates that formed in the thickener. Recycling thickener underflow could therefore have the opposite of the desired outcome. The result of testwork that is reported here lends credence to this theory and questions the wisdom of recycling in the SART process.There is evidence to suggest that operating the copper sulphide thickener in the SART Process without flocculants may be beneficial. Thickener underflow densities could improve from the current 10 - 15% range to 40 - 50%. This will not only reduce the flowrate of slurry to the filter, but the filtration properties of the slurry (filtration rate, filter cake moisture content) will also improve. This is one of the most critical engineering and design aspects of the SART Process, and the achievement of high solids density in the thickener underflow will not only improve the health and safety aspects of the operation, but also has the potential to lower capital costs and improve the quality of the final product."

Jan 1, 2016

By D. E. Pollock, Andrew B. Cecala

It has long been known that ventilation is a cost-effective method to lower respirable dust concentrations in all types of mining applications. The National Institute for Occupational Safety and Health has been working on a number of different research techniques that use ventilation as a critical component to lower respirable dust levels at surface and metal/nonmetal operations in the United States. This article presents five such research efforts. The first research area discusses how improving the ventilation flow patterns in an iron ore mill facility lowered respirable dust concentrations by 31 pct throughout the primary grinding area. The second research area examines how respirable dust levels were lowered at a dimensional stone shop by using air cleaning units suspended from the ceiling. A third area discusses how improvements in ventilation at underground limestone mines lowered respirable dust concentrations. The fourth area examines how improvements in dust filtration and pressurization systems significantly impacted the air quality in enclosed cabs of surface mining equipment. The last area presents a newly developed clothes cleaning technique. This technique uses an air spray manifold to blow dust from a worker’s clothing in an enclosed booth, which confines the dust for capture and removal by a baghouse dust collector. These research areas represent an array of different control technologies to lower respirable dust concentrations. Ventilation is an integral part of these control technologies.

Validation of the JKMRC Rotary Breakage Tester (JKRBT) Ore Breakage Characterisation Device

Sep 13, 2010

By R. S. Dean

In addition to the normal work and that pertaining to strategic minerals, the Metallurgical Division has been made responsible for carrying on major investigations in the development of processes' for the treatment of low-grade domestic manganese ores that heretofore have failed to yield products suitable for the preparation of ferromanganese, an alloy essential to every ton of steel produced. These investigations are particularly important at this time because of their part in tucking this country self-sufficient with regard to manganese when the stock-piled ore becomes e exhausted and as foreign shipments of ferro-grade ore are cut off. The results indicate that there is a good probability of accomplishing this objective, while at the same time conservation may be promoted through the use of normally useless material from the large, low-grade ore deposits.

Jan 1, 1941

By Nilce Alves Santos, Vânia Lúcia de Lima Andrade, Keila Lane Carvalho Gonçalves, Harold Wyslouzi1

"When confronted with the difficult decisions on advancing an exploration project to the feasibility and construction stages, it is important to know how various regions of the ore body will react in the milling and concentration processes. While bench scale test work is common and inexpensive, the results and information provided are somewhat limited due to the difficulty in accurately predicting the effects of recirculating middling streams or gradual changes in solution chemistry.Conventional Pilot Plants, processing 100 to 1,000 kg/hour of sample ore, provide the detailed engineering data required to develop a final process flowsheet and to size the equipment. They are, however, expensive to set-up and operate, and are really only viable for major projects that are well advancedIn the latter part of 1999, the Mineral Development Center at CVRD (MDC) purchased a new piece of mineral processing equipment which can run continuous flotation tests using drill core samples. To evaluate its performance, and to compare it with conventional testing methods, various types of flotation tests were carried out using two different types of copper ore. The limitations and advantages of open circuit and locked cycle bench tests, as well as the Mini Pilot Plant and the Conventional Pilot Plant were evaluated.Open circuit tests carried out at the bench scale only provide basic information. More complete and trustworthy information is obtained, in terms of the recirculation of flows and reaching a permanent and stable flotation regime, as the dynamics and complexity of the tests increases. The data obtained showed that the MPP’s metallurgical performance was equivalent to that of a conventional pilot plant."

Jan 1, 2002

By Damon P. Vandermaat

Coupon testing allows for the rapid understanding of a material?s likely behavior in service conditions, allowing engineers and operators to make informed decisions about the likely service-life of parts when subjected to field conditions. Bend testing has been identified as the best approach for exploring rock bolt susceptibility to stress corrosion cracking (SCC) in underground mining. The bend test specimens used were 2?3 mm thick slivers of steel, cut from the outer edge of a rock bolt. For the specimens, 1350 grade steel was chosen due to its wide use in the industry. The jigs themselves are small assemblies made of stainless steel that are capable of imparting a bending load on the specimens. The jigs were installed at a coal mine located in the Western Coals West of Sydney, Australia, in such a way as to expose the test specimen to water dripping from the mine roof. It is expected that this approach will yield results pointing to the likely SCC mechanisms affecting the service life of rock bolts, as well as demonstrate the jigs? usefulness as an early warning system for SCC. Furthermore, the coupon tests will be correlated against the results of large-scale laboratory tests.

Jan 1, 2012

By Michael Arnold, Bob Faulhaber, Joe Nichols

"Center Hill Dam is a critical part of the infrastructure controlling floods at the Caney Fork River and reducing flood levels at the Cumberland, lower Ohio and Mississippi Rivers. The Center Hill Dam Foundation Remediation Project was necessary to address Karst features in the foundation rocks and observed signs of distress of the dam. From 2011 to 2014 BAUER Foundation Corp. (BFC) performed deep foundation works for this project. The executed features of work include the installation of two walls. The first wall extends throughout the clay embankment into the underlying rock to provide encasement during the construction of the second wall. Its width of 7.4 feet makes it the widest diaphragm wall ever installed. The second wall is the actual seepage barrier and provides a continuous 2- feet wide wall. It extends through the first wall into the underlying rock and is up to 300 feet deep. The paper introduces the project, the requirements specified by United States Army Corps of Engineers (USACE) for encasement and barrier wall and the approaches taken by BFC. It discusses the ground conditions and describes the excavation equipment as well as the excavation methods used. A special focus is put on the steering capabilities of the hydrocutter equipment and the verticality performance.PROJECTBeginning in 2011, BFC was tasked by the USACE with the installation of a seepage barrier/cutoff wall at the Center Hill Dam Foundation Remediation project under a partially performance based specification. The remediation was necessary due to karst features in the foundation rocks and observed signs of distress, which led USACE to rate the dam with the most critical dam safety action classification (DSAC I). The executed features of work include the installation of two walls. The first wall, an encasement wall, extends vertically through the clay embankment and is embedded minimally into the underlying bedrock to provide embankment stability during the construction of the second wall, the barrier wall. The encasement wall is constructed of overlapping panels 7.4 feet (2.25 m) wide, which is the widest diaphragm wall ever installed, and up to 210 feet (64 m) deep. The second wall is the seepage barrier and provides a continuous 2 feet (0.6 m) wide wall consisting of overlapping panels. It extends through the encasement wall into the underlying bedrock and is up to 307 feet (93 m) deep. BAUER BC40 and BC50 hydrocutters mounted on MC96 and MC128 foundation cranes were used to perform the majority of the works (Figure 1). Excavations were performed under bentonite slurry (encasement wall) and water (barrier wall) and concrete was poured with the tremie method. Adding to the challenges, the project had to be executed in an environmentally sensitive area. Based on the high risk, a special focus was put on monitoring, quality control and geometric accuracy. The WallTracker information management system was used to share all relevant data between BFC and USACE (Roff et al. 2014)."

Jan 1, 2017

By P. Mark. Hawley

Development of Barrick Gold?s Pierina Gold Mine in north-central Peru requires construction of a waste dump for disposal of non-ore grade materials stripped from the open pit to expose the ore. The waste dump site is located near the open pit in a U-shaped valley that is underlain by thin colluvial and morainal soils with locally thick peats and glaciolacustrine deposits. Bedrock underlying the site consists of variably weathered volcanics. The site is subject to moderately high seasonal rainfall and significant seismicity. The geotechnical quality of the waste rock is highly variable, ranging from low quality, degradable, soil-like materials containing appreciable quantities of silt and clay, to high quality, hard, durable, blocky waste rock. A hybrid dump design concept was developed that requires in-pit segregation of high and low quality waste, and selective placement of materials in the dump to maximize stability and enhance internal and subsurface drainage. The design includes a rock drain beneath the dump along the axis of the valley and a series of connected finger drains. Extensive peripheral diversions and stormwater control measures to inhibit infiltration of precipitation are also included. This paper discusses the unique challenges presented by the site and waste materials, the design methodology and stability analyses, and planned monitoring systems.

May 1, 2002

By R. M. Waterstrat

X-ray diffraction and metallographic examination of binary Mn-rich alloys with Ti revealed the presence of intermediate phases in this system. A binary R phase has been identified and also a phase having an a-Mn type structure. The X-ray pattern of the phase TiMn3 is presented and a tentative phase diagram for the Mn-rich portion of this binary system has been constndcted. MANGANESE-rich alloys with titanium have been subjected to thermal analysis at temperatures down to 1100°C by Hellawell and Hume-Rothery.1 Their results indicate the existence of an intermediate phase TiMn3 in this binary system. This phase appears to coexist with the Laves phase TiMn2, and with terminal solid solutions based on the allotropes of manganese, at least down to 1100°C. The above investigators did not attempt to determine the crystal structure of this phase, however. Since, by analogy to the binary systems Mn-V and Mn-Cr3 one might expect to find a phase in this region of the diagram, it seemed desirable to obtain X-ray diffraction patterns of alloys in the concentration range in question, in order to ascertain whether the sigma phase or related structures occur. EXPERIMENTAL PROCEDURES Alloys were prepared by arc-melting electrolytic manganese and iodide titanium in a water-cooled copper crucible under a helium atmosphere. Excess manganese was added to each melt in order to allow for losses incurred during melting. The alloys were then wrapped in molybdenum foil and sealed in evacuated quartz tubes for annealing at various temperatures, as shown in Table I, followed by quenching in cold water. Although the inside of the tubes showed evidence for some loss of manganese from the specimen during annealing, there was no evidence of any reaction between the specimen and either the molybdenum foil or the quartz tube at these temperatures. It was later found that the manganese loss was considerably reduced by annealing these alloys in sealed quartz tubes containing one atmosphere of purified argon. As an added precaution the surface of each specimen was ground off before crushing the alloys to -270 mesh powder, using a hardened steel mortar and pestle. X-ray powder patterns were obtained using either FeKa or CrKa radiation in an asymmetrical focusing camera, which gave excellent dispersion of the closely spaced lines. Pictures were also obtained using a Debye camera of 5-cm radius which provided data in the angular region not covered by the focusing camera. The alloy specimens were polished and examined metallographically, using an etchant consisting of 60 pct glycerine, 20 pct nitric acid, and 20 pct hydrofluoric acid. The 10-g buttons were in each case broken in half and found to be well-melted and free of any gross segregation. One half of certain alloy buttons was submitted to chemical analysis in the "as-cast" condition. No appreciable change in composition seemed to occur during annealing . RESULTS The X-ray pattern of alloy Ti1.17Mn0.83 was found to agree well with that of the ternary R phase which was first found in the Mo-Cr-Co system,' and re-

Jan 1, 1962

By Innes J. A

The structural variations occurring in pellets of finely divided mixtures of iron ore and coke breeze, metallized by firing at elevated temperatures, have been examined microscopically. For any given pellet composition, time and temperature are the two most important factors in determining the development of the bonding processes.INTRODUCTIONThe success of sintering and pelletizing processes in the agglomeration of fine iron ores for blast furnace use has brought about rapid changes in thinking relative to methods of preparing furnace burdens. Increases in productivity with the use of sinter emphasize the advantages of a stable, uniformly sized furnace feed. Current practice is to produce a self-fluxing sinter containing most of the lime requirements of the furnace incorporated as a stable medium (Figgis and Charker, 1962). The elimination of fine material from the burden allows greater stack permeability and smoother operation.Sintering and pelletizing of iron ores with or without the incorporation of fluxing compounds has conditioned operators to the extent that it is now considered possible that all components of the burden may eventually be beneficiated in one form or another. Investigations are in hand to improve the quality and sizing of coke, which occupies the major part of the burden volume (Callcott, Innes and Brown, 1961; Innes, Gregory and Worner, 1963). Interest is also being revived in methods of pre-reducing iron ores prior to charging to the furnace (Hansen, Melcher and Fine, 1961). Recent experiments in a small experimental blast furnace (Melcher, 1963) show that charging of semi-metallized pellets can give rise to a production increase of 55 per cent...

Jan 1, 1963

By S. Bang

In recent years, suction piles and suction anchors have been used as underwater foundations for various offshore structures. They have been proven very versatile, effective, and environmentally friendly. This paper describes the load resisting mechanism, installation, and design of suction piles and suction anchors. Also included are brief descriptions of three offshore structures that have recently been constructed using suction piles and suction anchors. They include: a new breakwater system utilizing gravity-type concrete caisson superstructure on top of concrete suction piles; a floating breakwater system utilizing embedded suction anchors; and a foundation system utilizing suction piles for temporary mooring of underwater tunnel elements.

Jan 1, 2009

The Bega granite at Whipstick, N.S.W., consists of a granite metasomatically altered to leucogranites and a spessartite-granite. The ore occurs in pipes controlled by jointing and shows evidence of post-depositional tectonic forces.The principal ore minerals are molybdenite and bismuthinite, with minor amounts of uraninite, pyrite, gold, bismuth, joseite A, joseite B, tetradymite, chalcopyrite, arsenopyrite, and skutterudite. The gangue minerals are microcline, albite, quartz, muscovite, and spessartite, with fluorite, chlorite, magnetite, and biotite as accessory minerals.The intrusion of granite was followed initially by ascending metasomatic fluids which penetrated along joints to be trapped in the domed roof of the intrusion to form the leucocratic granite. Iron-manganese-rich fluids metasomatically altered the granite to a spessartitegranite which is associated with the later deposited ore.Quartz veins and dykes of varying composition post date the ore deposition.

Jan 1, 1968

By Manqiu Xu, Songtao Yang, Zongfu Dai, Robert Pelton, Carla Abarca

"Fine mineral particles are difficult to process by flotation because of their low collision efficiencies with air bubbles. Decreasing air bubble size and increasing mineral particle size through flocculation are the most common approaches to improve flotation efficiencies. Herein we demonstrate that nanoparticle flotation collectors can be engineered to selectively induce fine mineral particle aggregation and promote flotation.Flotation tests of clean fine pentlandite (Pn) and Clarabelle Mill (CBM) rock tails and pyrrhotite (Po) tails that contained significant amount of fine pentlandite have confirmed that the nanoparticles can serve as both collectors and selective flocculants for fine pentlandite. In the case of clean fine Pn, the nanoparticle collectors virtually can float almost all Pn fines. Scoping tests on CBM tails have shown significant improvement in nickel recovery. Addition of nanoparticles as a flotation aid can recover 20% of the nickel contained in the rock tails and 40% of the Pn contained in the Po tails compared to current mill operations giving 0% nickel recovery from rock tails and no Pn selectivity from Po tails. Ongoing work includes attempts to identify optimal strategies for employing nanoparticles in fines flotation.IntroductionFine particle flotation is an industry-wide challenge in mineral processing. Fine mineral particles are difficult to process by flotation because fine particles have low mass and thus low momentum, which results in their low collision efficiency with air bubbles (Dai, Fornasiero, & Ralston, 1999). In addition, fine particles have not only high specific surface area but also more active surface area, which causes high reagent consumption. For base metal sulphide minerals, particle size between 10 µm and 70 µm is a typical range where high flotation recoveries can be achieved with conventional flotation chemistry and acceptable cell residence times. For pentlandite, the flotation recoveries decrease sharply when the particle size is below 10 µm and hydraulic entrainment is the dominant mechanism, giving recoveries proportional to water recovery and no Pn enrichment. (Jameson, Nguyen, & Ata, 2007)"

Jan 1, 2015

By Michael E. Berndt

Asbestos first became an issue to Minnesota?s iron industry when it was revealed that mineral fibers similar to those in Reserve Mining?s tailings were being found in drinking water for several communities that used Lake Superior as their primary water source. This discovery turned what had largely been an environmental court battle into a case concerning public health. The courts listened to much conflicting and uncertain scientific testimony on the size and distribution of the mineral fibers and on the potential health effects imposed by them. In April 1974, the plant was ordered to shut down by a federal judge but the company quickly appealed the decision. The appeals court granted a stay and ultimately ruled that the plant?s closure could not be justified based on the unknown health effects of the mineral fibers since the consequences of such an action would have immediate and severe social and economic impacts. The plant was allowed to continue operation, but ordered to abate emissions to air around the plant and to switch to a land-based tailings disposal system. Much of the scientific uncertainty and public concern over mineral fibers in Minnesota?s taconite industry remain today. ©2007 Elsevier Inc. All rights reserved. Keywords: Minnesota; Iron ore; Taconite; Asbestos; Amphibole; Reserve Mining; Tailings; Mesabi Iron Range

Jan 1, 2007

By Paul Billingsley

THE purpose of this paper is a comparison, based on genesis, of the ore deposits associated with the igneous rocks of the central Montana Rockies. Considered separately, without attention to their origin, the orebodies appear to be distributed haphazard, with no fundamental laws controlling their position and character. It is believed, however, that if all the factors be weighed, this apparent confusion will to a large degree disappear. The, first essential to such a comparison is detailed knowledge of the individual mines, as to their geographic position (see Fig. 1), their mineralogical character, and in particular the derivation of their ores.1 The authors of this paper have pursued such an investigation for the past 6 years, and have accumulated a mass of evidence that enables a preliminary correlation to be attempted. It appears that the majority of ore deposits are genetically related to one or another of the igneous

Jan 4, 1917

By A. S. Odeh

During the last few years, several authors1-6 have advanced mathematical solutions, both exact and approximate, to the pressure behavior of a well producing from a region bounded by a circular discontinuity. The most recent solutions were reported by Carter, and by Bixel and van Poollen. Varter solved the appropriate equations analytically. Bixel and van Poollen reported a numerical solution and provided correlation charts for the interpretation of flow tests; their method uses an overlay technique and is valid for wide ranges of mobility and storage capacity ratios. The mobility ratio is defined as the ratio between k/p of the two zones, and the capacity ratio is defined as the ratio between $c of the two zones. We have developed independently an analytic solution, programmed for the CDC 1604, that agreed with Carter's.'1 The same differential equation, assumptions, boundary, and initial conditions were used. We ran 30 drawdowns and 30 buildups of a unit capacity ratio and of varied mobility ratios. We also varied the radial distances to the discontinuity, and the radii of drainage. We examined the results and obtained correlations that were simple to apply and useful in flow test interpretation. The results and the correlations reported here in graphical form allow the determination of the mobility of the two zones and the radial distance to the discontinuity. They eliminate the need to fit field data to a curve as is required in the overlay technique.6 Thus, the advantages of the present correlations are that they require less effort and they provide a unique answer. They also provide guidelines for selecting the parts of the flow test plot to be used in the interpretations. Moreover, the results obtained from the programmed analytic solution define the time intervals where the approximate solution reported by Hurstl and discussed by Larkin2 are applicable. Normal buildup and drawdown plots resulted in an early straight line, a long transition zone and a late (second) straight line. However, the second straight line will not occur if the drainage radius is not at least 10 times the radial distance to the discontinuity. The mobility of the imer and outer zones can be calculated from the slopes of the first and second straight lines, respectively, by the normal methods of the flow test analysis. We found that the radial distance to the discontinuity may be determined from the abscissa value for the point of intersection of the extensions of the two straight lines. Results similar to some of these were reported by Adams et aL7 The approximate solution compared favorably with the exact one during the first and second straight-line periods. However, during the transition zone, deviation occurred between the two solutions. Discussion of Figures Figs. 1 through 7 summarize the results of this investigation. On Figs. 3 through 7 there appears the equation of each respective plot. In this paper the mobility ratio M is defined as klp2/kzpl, where subscripts 1 and 2 refer respectively to the inner and outer zones, and dimensionless time = t, = 6.32 X In the investigation, M ranged from 0.25 to 50;

Jan 1, 1970