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By P. A. K. Wilkinson

This paper is concerned with the development of an in situ recovery of boric acid from a deep-seated colemanite deposit in California. The deposit contains 147 million tons of borate ore containing 6.4% boron oxide (B203). It is the largest undeveloped colemanite deposit in the United States and because of its unique occurrence and grade, in situ extraction appears to be the most economic viable process. The paper covers the geological, mineralogical, hydrological, well-field development, and extraction considerations leading to the initial pilot plant and subsequent demonstration plant phases of the project. Market potential aspects and preliminary economics of the project are also covered.

Jan 1, 1987

By Leland E. Paulson

The Grand Forks Energy Research Center and Commonwealth Edison of Chicago jointly conducted tests in which 400 tons each of subbituminous coal and lignite were dried in a commercial scale dryer, oil sprayed and cooled, then shipped in open-top rail cars from Pekin, IL to Grand Forks, ND and stockpiled. Cars containing raw coal and dried coal that had not been oil sprayed were also transported for comparative purposes. The tests were conducted in August and November of 1974. The subbituminous coal was dried from 26 to 16 pct moisture and the heating value was upgraded from 8,420 to 9,650 Btu/lb. The lignite was dried from 39 to 22 pct moisture and its heating value was increased from 6,420 to 8,300 Btu/lb. Before loading subbituminous coal was cooled to 115° F and sprayed with oil at a rate of from 2 to 6 gal/ton. Similarly, lignite was cooled to 85° F and oil sprayed at rates of from 1 to 2 gal/ton. Both shipments arrived in Grand Forks with a minimum of difficulty. The subbituminous coal was subjected to 2 inches of rain during transit with no change in average moisture content. The cars containing dried and oil treated coal had less material lost to wind than either cars containing raw or untreated dried coal. With both dried lignite and subbitnminous shipments, a moderate increase in temperature was measured during transit. With the subbituminous shipments, ignition occurred near poorly fitted bottom dump doors, but was limited to a very small area. The dried lignite was shipped when ambient temperatures were below freezing, and frozen coal was noted around the edge of the car. At Grand Forks, the coal was stockpiled using proven techniques and has remained stable to this date.

Jan 1, 1975

By R. Al-Hent, M. Aissa, J. Asfahani

"This paper focuses on determining the favorable geological and hydrogeological conditions for forming uranium occurrences in the Froqlos-Jabal Abou Rabah region, Palmyrides, Syria. Eighty rock samples from this area were analyzed by spectrometric gamma-ray technique to determine the radioactive concentrations of eU and eTh, and of %K. Uranium concentrations in shallow and deep groundwaters were also determined. High uranium concentrations were registered owing to the presence of phosphatic outcrops in the study region. The uranium migration trends and its remobilization were analyzed through the analysis of the behavior of eU, eTh, and their ratio. A positive relationship between eU and eU/eTh, and a negative relationship between eTh and eU/eTh in geological formations of different ages supports the secondary remobilization of uranium. The role of groundwater movement in transportation and redeposition of uranium mineralization is discussed: deep waters are less oxidizing than near-surface waters, and do not appear to have significantly remobilized uranium. This indicates that the source of secondary uranium mineralization is surficial or near surficial, and is mainly related to the leaching of outcropping phosphatic layers.Integration of the results with established radioactive and geological sections reveals four radioactive anomalous zones: north Froqlos, Kherbet Hannora, northwestern flanks of Jabal Abou Rabah (Kherbet Al-Hajar), and the northwestern limit of the Ghuntor depression. The high uranium concentration in north Froqlos is because of its original presence in the phosphate beds and extensive fracturing, which allows groundwaters to penetrate and remobilize the uranium. High uranium concentrations in the other three locations are caused by the presence of evaporites and capillary action, which draws solutions upwards and causes redeposition as surface crusts (gypcrete)."

Jan 1, 2009

"The QEMSCAN system is the third generation of automated mineral analysis systems that began with the QEM*SEM at CSIRO 30 years ago (Butcher et al. 2000). It is now a successful commercial instrument with 39 instruments in operation around the world. Back Scattered Electrons (BSE) and Energy Dispersive (EDS) X-Ray spectra are used to create digital mineral images with mineral identification occurring online. Low-count EDS spectra are used preferentially to BSE brightness, allowing minerals to be accurately identified by chemical composition. Individual minerals or groups of similar compositions are identified by comparison to a comprehensive mineral database incorporated into the QEMSCAN software. Optical mineralogy and Scanning Electron Microscopy is a metallurgists’ tool for determining the ore process mineralogy and delivering a better flowsheet. These techniques have limitations, provide limited information and sometimes fail to provide representative information. In contrast, QEMSCAN provides information on the nature and occurrence of all mineral species present plus the liberation characteristics, looking at thousands of particles in the sample presented. Some of the data presented includes:• Particle and grain sizes• Dominant mineral species• Particle shape• Calculated metal content• Density• Deportment of elements of interest between the mineral phases• Quantitative mineralogy• Mineral associations, locking and liberation• Textural relationships (particle images)It can also be used on concentrates, polished sections or drill core. This is a significant improvement in ore characterisation providing lithology, mineralogy, liberation, ore grade and porosity from a sample. Theoretical grade-recovery curves can be created to be mineralogically limited from QEMSCAN analysis which, when compared with metallurgical results, are valuable in evaluating whether liberation or chemical selectivity is driving flotation performance. A number of case study examples are given where the QEMSCAN had a significant bearing on future project development at an early stage of exploration and mine development."

Jan 1, 2012

By Reimer TH

The model described in this paper has been developed for and in close collaboration with the Bong Mining Company Ltd, Liberia/West Africa. When working an irregular deposit a de- tailed short term production schedule becomes necessary in order to ensure an equal and homo- genous concentrator feed. If the irregularity of the ore body exceeds a certain degree, blen- ding facilities have to be provided between the mine and the concentrator, but even then sche- duling of the mine production is inevitable to meet the given requirements of the blending bed and to keep these conditions steady for at least some of the successive periods. To assist the planning department in set- ting up optimal production schedules a computer model has been developed using a linear pro- gramming approach. The deposit is divided into equally shaped blocks containing defined quali- ties and quantities of ore and/or waste; from which one has to select those quantities of ore which will meet the requirements of the blen- ding bed. The objective function of the model

Jan 1, 1977

By J. R. Sturgul

Thompson and Adler (1988) present a new simulator for use in designing conveyor belt systems for underground mines. Their model uses a large Fortran program to do a Monte Carlo simulation. Their approach is similar to that done by other researchers, such as Stanford (1965) and Newhart (1977). All Fortran based simulation models use an imaginary clock that is advanced a time increment. At each advance of time increment the program has to determine what is to take place in the system. Since many things can take place at different places in the system, this approach requires that a large amount of logic is needed at each step of the program. This, then, necessitates a great amount of jumping between different portions of the program. In fact, the Fortran program, CMBCS, takes up to 15 to 20 minutes to run on the mainframe and several hours for the PC. Such execution times are typical of Fortran based simulation studies. There have been other examples of Fortran based simulation studies in mining. Cross and Williamson (1969) presented a model of a working open-pit copper mine. Fosinato (1982) presented a model of the bucketwheel excavators for the large open-pit mines in Southern Victoria in Australia. These excavators discharge into a complex conveyor system, which can be thought of as being similar, but more complicated than the one modeled by CMBCS. All previous simulation studies using Fortran as the source code have several things in common: they all are very lengthy, often being of the order of 10,000 lines of code and take a great deal of CPU time to execute. The program of Fosinato takes more than 4.5 hours to run on a large mainframe. Also not mentioned by Thompson and Adler is the actual time it took to write the Fortran code. This, too, can be considerable, as Fosinato's program took several man years to write and de-bug. In order to change it, as the parameters change or as the mine increases, further large amount of time is needed. An alternate to such Fortran based programs is provided by using a special computer language known as General Purpose Simulation System (GPSS). This language is designed to solve a wide variety of discrete event simulation problems and appears to be particularly suited for mining situations as shown by Sturgul and Yi (1 987). Although GPSS was introduced as far back as 1961, it has not been used in mining to any great extent. However, the original form was quite crude and inefficient compared to modem versions. Perhaps with the increasing popularity and availability of these recent versions, this will change. Two such versions are GPSS/H for the mainframe and GPSS/PC for the personal computer. GPSS/H allows for interactive usage, such as allowing input data to be entered via a menu. GPSS/PC has excellent graphical capabilities and one can actually view the simulation in progress. In addition, one can halt the simulation and change values of various parameters. If a particular portion of the simulation is in progress via a concept known as "microwindows." These new versions have greatly increased the potential for using the language for mining situations. Mining applications of GPSS have been given by Sturgul and Yi (1987), Sturgul and Harrison (1987), Sturgul (1987), and Sturgul and Singhal (1988). Schriber (1987) presents a summary of the recent advances in GPSS.

Jan 1, 1987

By Robert K. Towner, Francis S. Kendorski

Shotcrete drift linings at the Climax Mine of the Climax Molybdenum Company have undergone four years of extensive evaluations as a replacement for thick poured-in-place concrete linings. Any drift lining at Climax must be able to withstand the rigors of undercut and secondary blasting, weight problems induced by panel caving, and the abrasion of slusher mining. Examined here are the conclusions reached so far on shotcrete drift linings which have not yet completed the approximately six-year cycle from development to ore exhaustion. Properly applied shotcrete is an adequate drift lining for all phases of the Climax panel caving system of mining. Undercut blasting will damage shotcrete unless blasts are correctly timed. Weight problems will destroy shotcrete drift linings as efficiently as it will destroy concrete or steel support systems. Abrasion and secondary blasting results are inconclusive as the shotcrete lined slusher drifts have not entered the ore production stage of the mining cycle. In addition to drift linings, Climax has found shotcrete to be invaluable for lining large underground shops, for temporary support systems, for permanent ventilation bulkheads and for the repair of abraded or weight damage concrete lined slusher and haulage drifts.

Jan 1, 1974

As part of the Phase I study activity, a questionnaire was sent to a representative sampling of mining companies to obtain information on the uses, use conditions, personnel policies and other user concerns associated with front-end loaders. On-site visits to several of the mining companies were also parried out and the questionnaire was used during these visits as well. Five major areas were dealt with in the user questionnaire. These included: 1) Front-End Loader Critique - An open-ended form was employed to permit the mining companies to describe the particular things they liked and disliked about front-end loaders, in general, and/ or specific models in particular. 2) Usage - A form indicating the functions performed by front-end loaders and the numbers, makes and models of front-end loaders being used by mining companies. 3) Environment - A form where the user indicated the various types of conditions encountered in their use of front-end loaders. 4) Personnel - A form for describing factors pertaining to the qualifications of front-end loader operators. 5) Miscellaneous and Comments - A form permitting the user to comment on his feelings toward standardization of front-end loader operator stations.

Jan 1, 1973

By A. B. Kinzel

IT is with sincere appreciation and a deep sense of responsibility that I accept the honor of delivering the Howe Memorial Lecture. In our time metallurgical research has delved into phenomena ever more complex in character, so that much of this work must be done by organized teams. We, of the Union Carbide and Carbon Research Laboratories and the Electro Metallurgical Co., are truly a team and the lecture which is to follow is due largely to a few of the outstanding members of that team here cited and well known to you in their own right: John Lamont, whose work deserves special recognition, Walter Crafts, William Forgeng, William Binder, Robert Fowler, and David Swan. Thus it is a fitting pleasure to accept this honor for the research institution which I represent. Henry Marion Howe was truly inspirational and his breadth of view was such that he could use either the method of rigid proof or implied mechanism to further the broad understanding. But perhaps his outstanding characteristic was his sound judgment and his ability to weigh observations. The following is replete with observations and judgment as to their implication. We would like to feel that it is the type of dissertation which Howe, himself, would have enjoyed. Chromium carbides in stainless steel have, in a sense, been a key to the progress of the stainless steel industry. Brearley's first stainless steel, containing approximately 0.30 pct C, was successful because it differed from previous Fe-Cr alloys in that the carbides were present in such amount and form as to provide the stainless and cutting properties. Again, the 12 pct Cr die steels with a carbon content of 1.5 pct or more achieved their properties by virtue of the form and distribution of the chromium carbides. In the straight chromium steels with from 12 to 25 pct Cr, chromium carbide amount and distribution are once more the determinants with respect to physical properties and corrosion resistance. In each of the cases just cited the role of the carbide is in essence similar to that of iron carbide in carbon steel and this story is so well known that we need not dwell on it. In the austenitic stainless steels, the chromium carbide plays only a minor positive role with respect to mechanical properties, and for corrosion resistance the objective is to achieve maximum homogeneity. Thus the tendency has been to ever lower the carbon and correlatively the chromium carbide content of the austenitic steels. Today stainless steels with carbon less than 0.03 pct are beginning to find their true place, and this is being aided in major degree by our Company's introduction of a new type of ferrochromium having negligible carbon. The role of carbon in these austenitic steels has long been recognized, and a truly enormous amount of research has been devoted to the subject. Much has been learned about corrosion resistance, but the attack on the problem from the point of view of carbide precipitation and the nature thereof has left much to be understood. Because of its fundamental importance to the future of stainless steels, further study of the mechanism of carbide formation in austenitic steels is of interest. In the 18-8 Cr-Ni steels, which are typical of the austenitic class of steel, carbon is soluble in the aus-tenite to the extent of approximately 0.15 pct at 1832°F (1000°C). The solubility becomes less with decreasing temperature, so that at 1652°F (900°C) the solubility limit is approximately 0.06 pct, at 1472°F (800°C) it is about 0.03 pct, and at 932°F (500°C) it is approximately 0.01 pct. Obviously this is the ideal situation for the classical precipitation phenomenon. It is true that, in a sense, 18-8 austenite resulting from annealing is in a metastable state, and the phenomenon of ferritic precipitation from this austenite should be considered. However, in a so-called fully austenitic 18-8 steel this is secondary with respect to the carbides and can be neglected. A wide variety of chromium carbides is known. Only one chromium carbide, Cr,,C,,, frequently referred to as Cr,C, has been found in unmodified low carbon 18-8 steel. This, fortunately, simplifies the study of the precipitation phenomena. For this study an A.I.S.I. type 304 stainless steel of commercial quality was selected. Analysis showed C 0.07 pct, Mn 0.49 pct, Si 0.33 pct, Ni 9.34 pct, Cr 18.91 pct, and N 0.033 pct. As expected with this composition the steel was homogeneous after annealing except for a few pools of ferrite, which were disregarded in that their location was not coincident with subsequent carbide precipitation areas. On corrosion testing the steel was normal and typical of good commercial quality. Specimens 1/2xlx4 in. were annealed and subjected to precipitation heat treatment for 100 hr at 100" temperature intervals from 1000° to 1500°F and for ten selected intervals ranging from 5 min to 64 hr at 1300°F. These blocks were then cut into small pieces for metallographic studies. The first problem

Jan 1, 1953

By Victor J. Miller

The Northeast Tripp Slide is one of the larger slope failures that can be attributed to open pit mining. It is a 11.7 million cubic meter (15.3 x l0 6 yd3) wedge failure created by two thick gouge-filled fault zones whose intersection plunges into the Tripp pit at 13°. One unique aspect of this failure is that the slide is now stable despite having moved an estimated 8.5m (28 ft.) towards the pit and having reached a maximum overall velocity of 2lmmlday (0.07 ft/day). The slide started to fail in 1968, accelerated until mining in the slide's toe area was halted in April 1970, then began to decelerate. With the cessation of mining in the toe area, rubblized rock sloughed into the pit bottom which added weight to the failures toe constraints and began the stabilization process. Measurements from a network of survey stations revealed a correlation between the slide's velocity and heavy precipitation. Runoff water accumulating in the tension cracks is believed responsible for this correlation. Moreover, the survey net measurements showed that during the failure's last stage the vertical component of the slide's movement greatly increased. It is proposed that this settling was caused by plastic-like flowage of the clay gouge underlying the failure The fault zone's effective shear strength was estimated by a back- analysis on the wedge's geometry and by performing direct shear tests. The tests, performed on cores taken from samples that were collected on the failure, yielded a residual shear strength of: cohesion 0.61 kg/cm2 (8.7 lb/in2), and a friction angle of 9.6°. This residual shear strength was near1y equal to the back-calculated strength indicating that the controlling fault's effective strength was essentially the same as the residual shear strength of the gouge filling

Jan 1, 1983

Operation of a pilot-scale Floatex density separator was studied using a pyrite concentrate slurry feed. Experimental measurements were further utilized in parameterization of a hindered settling simulation model. The model incorporates particle slip velocity calculation and also a separation efficiency function, adjusted to describe the feed material characteristics. The model was used to compute the overflow and underflow compositions when both fluidization bed pressure and teeter water flow rates were varied. In addition, prediction accuracy and the model validity were evaluated. The simulations gave good results when the teeter interstitial velocities were moderate enough to avoid excessive turbulence in the fluidization. Prediction of the cumulative overflow and underflow size distributions in most experiments yielded essentially less than 10% relative error between the sample data and the estimated value of the 50% passing sizes.

Jan 1, 2011

By Louis Frost

"1N 1951, the Dominion Steel and Coal Corporation, as part of a programme of modernization in their coal mining operations at Sydney Mines and Glace Bay, Nova Scotia, undertook the development of two large cross-measure tunnels, which were to be used for the purpose of raising the coal from the underground workings by conveyor belts direct to wash plants located on the surface.These tunnels were to reach from the surface to central loading points underground located at a depth of 600 and 900 feet, respectively. They were to be driven on a pitching gradient of 11 degrees to the horizontal, and they were to reach a length of 3,600 feet in the Sydney Mines area and 10,000 feet in the Glace Bay area.The longer tunnel, 19 feet in diameter, was driven a distance of 1,500 feet and, for domestic reasons, bas not yet been •completed. The other tunnel, in the Sydney Mines area, 18 feet in diameter, is steel supported and concreted throughout its entire length. This tunnel has been completed and will be in operation at an early date.The development of this particutar tunnel posed an additional problem, as the rock through which it had to pass was considered to be water-bearing. In the sinking of Princess Colliery shafts in 1866, a great deal of water had been encountered, and it had been necessary, while sinking, to line the shaft with cast-iron tubbing through the water-bearing strata. The shaft had to be lined for a depth of 300 feet at the upper end. It was anticipated that we would meet with difficulty in driving the tunnel through this strata."

Jan 1, 1955

By Henry Ferguson

THIS paper is a preliminary statement, intended to present the more important results of the recent studies of the ore deposits of the Alleghany district in advance of the publication of the complete report, now in preparation. The Alleghany district (Fig. 1) has been a producer of gold since the early days of California gold mining. As in other California gold camps, placers yielded the bulk of the earlier production, but during the last 30 years lode deposits have been the only important source of output. The district differs from other California gold-quartz districts in that nearly all the production is from small shoots of high-grade ore. The mining camps of Alleghany and Forest are on the western flank of the Sierra Nevada range about 20 miles west of the crest. The gently sloping lava-covered plateau which controls the upland topography has been trenched by deep canyons of the westward flowing streams, which are cut to depths of as much as 2000 ft. below the crests of the ridges. The district is about 18 miles from Nevada City but the automobile road which connects the two places is 31 miles long. The senior writer first visited the district in 1913 and published a short description of the lode deposits.1 The increasing importance of the district made further study desirable. A detailed topographic map was therefore made by W. B. Upton, Jr., of the U. S. Geological Survey, in the spring of 1924, and the senior writer spent parts of the field seasons of 1924 and 1925 in the district. For the second field season the Geological Survey was fortunate in securing the services of R. W. Gannett, who had already done geologic work in the district for several of the mining companies. Owing to the death of Mr. Gannett in October, 1925, just before the close of his field work, the work of preparation of the entire report fell to Mr. Ferguson. He takes this opportunity of acknowledging how deeply he is indebted to Mr. Gannett for his excellent work in the field, and his realization of how greatly the completion of the report on the district has suffered from lack of Mr. Gannett's assistance in the office. In October, 1.928, after the completion of the first draft of the manuscript, Mr. Ferguson made a short visit to the district.

Jan 1, 1929

By E. J. Berentsen

The Turquoise Ridge gold deposit is located near the Getchell Mine, 72 km (42 miles) northeast of Winnemucca, Nevada. The deposit is owned by First Miss Gold Inc. The host rocks consist of Ordovician and Cambrian age shales and limestones. They have been crosscut by the northeast-trending Turquoise Ridge shear zone and northwest-trending shear zones. Gold mineralization occurs at depths from 427 m (1 400 ft.) to 671 m (2 200 ft.), along the intersection of these structures. Large tonnage, flat-lying deposits occur where folded, favorable beds are intersected by the faults. The carbonate within the beds has been replaced with silica, pyrite and arsenic sulfides.

Jan 1, 1996

During the presentation of the papers, questions were deferred until after the last paper in the afternoon. At that time, the authors and co-chairmen formed a panel for discussion of any pertinent questions. The panel was as follows: Carl Gerity - Ozark Lead Company – Chairman Mike Young - Master Mechanic - Ozark Lead Company Largo Albert - Maintenance Specialist (hoisting) Mines Maintenance Dept. - International Nickel Company, with assistance from Co-author Alan Cameron - International Nickel Company Gary Beerkircher - Product Manager - Rexnord Inc., Nordberg Machinery Group Richard Schaal - Director of Maintenance - Superior Division, Magma Copper Company Rich Hodgson - Assistant Production Superintendent - FMC Corporation The remarks have been edited by the panelists for readability, the basic ideas were not changed. Additional thoughts by the panelists submitted at the time of their revision are indicated by ***.

Jan 1, 1975

By K. A. Lewandowski

Leach heaps are plagued with problems due to fine material migrating downwards, clogging the spaces between the coarser ore particles, resulting in decreased metal recovery rates. The use of a binder in agglomeration would allow the fine material to be immobilized, increasing the availability for the solution to come in contact with all the ore. As much as 1.64*1012 BTU per year would be able to be saved if a 25% increase in copper recovery was experienced, assuming that 70% of heap leaching units converted to using agglomeration technology. This is equivalent to approximately 18% of the energy currently being used in leaching heaps.

Jan 1, 2008

By Joe B. Rosenbaum

New applications and concepts of solvent extraction as a unit operation in extractive metallurgy have been developed in the last decade. Wider use is being made of established solvent extraction techniques in processing spent nuclear fuel elements and in winning columbium-tantalum, hafnium-zirconium, thorium, tungsten, uranium, and vanadium. Commercial application has been extended to recovery of beryllium, boron, bromine, copper, molybdenum, phosphorus, rare-earth metals, and rhenium. Systems in use and proposed typically extract metals or salts as anion, cation, or neutral species from aqueous media by reaction with a recycleable organic reagent.

Jan 1, 1971

By Doss H. White

The value of non fuel mineral production in Alabama in 1991 was $539.9 million, a decrease of $20.7 million, a result of the nationwide recession that was felt throughout the State's economy. The 3.7% decline in mineral sales was due primarily to a slowdown in construction, which had a major impact in the construction minerals mined instate. Production and sales of crushed stone and sand and gravel fell significantly below the value reported in 1990. Despite the recession, Alabama moved from 21st to 20th place nationally in mineral value. Mineral industry developments in 1991 were primarily in the metals manufacturing sector. Reynolds Metals Co., Sheffield, celebrated 50 years of operation in Alabama. During the November celebration, the company dedicated a $121 million expansion that included a new casting facility and sheet rolling operation. Gulf States Steel Corp., Gadsden, completed the first full year of operation of its new multimillion-dollar continuous caster.3 The Foundry of the Shoals Inc., Florence, replaced two coke-fired furnaces with electric furnaces.

Jan 1, 1993

By J T. Carswell, K Sutton

Quality control (QC) of reverse circulation drill sampling commonly involves regular duplicate subsampling to monitor the extent of imprecision or random error in the subsampling system. In gold mining in Australia, the subsampling tool is usually a cone splitter.Other sources of error, including delimitation error and bias (systematic error) in the subsampler are often assumed to be negligible or ignored for being too difficult to QC.This paper describes actual sampling errors and error mitigation based on quality assurance and quality control (QA/QC) of surface and underground reverse circulation drill sampling at a high-nugget gold mine at Sunrise Dam, Western Australia.CITATION:Carswell, J T and Sutton, K, 2014. Sources of sampling error and implications for quality assurance and quality control in surface and underground reverse circulation drilling, in Proceedings Sampling 2014 , pp 65–70 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 29, 2014

By William Kelly

THE objeet of this paper is to describe the electric equipment at the iron-ore mines of Penn Iron Mining Co., Vulcan, Mich. and of Republic Iron Co., Republic, Mich.; to give the result's of tests; and to discuss the methods in use from an operating as well as from an efficiency standpoint. Electricity was introduced at the Penn mines for pumping, hoisting, and compressing air in the spring of 1907, upon the completion of a hydro-electric plant built by that company on the Menominee river about 4 miles from the mines. This plant was described in a paper1 presented before the Lake Superior Mining Institute. A paper2 presented at a meeting of the same Institute describes some of the operating features, and the safety devices of the electric hoist at the Curry shaft were briefly described in the Proceedings of the First Co-operative Safety Congress held under the auspices of the Association of Iron and Steel Electrical Engineers at Milwaukee, Wis., Sept. 30 to Oct. 5, 1912.

Jan 2, 1914