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By Edward W. Parker

I. INTRODUCTION. AT the Seventy-sixth meeting of the Institute, held in New York, N. Y., February, 1899, I presented a, paper on this subject entitled, Coal-Cutting Machinery,' which has become somewhat out of date, in view of the decade's development in this feature of the coal-minim industry. It should be explained that the present paper is limited to the use of mining-machines in bituminous mines. Practically all of the anthracite mined is ?shot from the solid," no undercutting being done. Anthracite-mining is therefore not considered in the statistics or other references. The statistical record as compiled by the United States Geological Survey shows that in 1898, the year preceding the presentation of my former paper, there were 32,413,144 short tons of coal undercut by the use of machines. In 1907 the machine-mined coal-product amounted to 138,547,823 short tons. In sympathy with the general decrease in coal-production during 1908, the output of machine-mined coal decreased to 123,183,334 short tons, but the proportion of machine-mined coal to the total output, nevertheless, showed an increase, from 35.1 per cent in 1907 to 37 per cent in 1908. In 1898 only 19.5 per cent. of the total product was undercut by the use of machines. The number of undercutting-machines in use in the bituminous mines of the United States has increased from 2,622 in 1898 to 11,144 in 1907, and to 11,569 in 1908. The total production of bituminous coal in the United States in 1908 was almost exactly double that of 1898, while the machine- * Published by permission of the Director of the U. S. Geological Survey. 1 Trans.. xxix.. 400 to 4,59 (1899).

Sep 1, 1910

By A. Tsekov, Stefan Neufeind, T. Angelov, G. Savov

"A complementary ion exchange/solvent extraction process for the recovery of copper from low-grade ores developed by Iontech Engineering Ltd. is described. It has successfully been implemented on a full-scale plant level at the Buchim Copper Mine in Macedonia, yielding a maximum copper cathode production of 2400 tonnes per year. Technical and economic aspects of three years of operational experience are comprehensively compared with the Asarel Medet copper plant in Bulgaria, which is utilizing conventional solvent extraction in the enrichment/purification step. INTRODUCTION The concentration and separation of copper from primary leachates is one of the key operations in the hydrometallurgical processing of oxide ores (Langner, 2011). Conventionally, this basic operation is conducted by submitting the pregnant leach solution (PLS), with a 1–5 g/L Cu content, from the initial sulfuric acid leach to a solvent-extraction (SX) process in which copper reports to the organic phase (see Figure 1). After phase separation, the metal is stripped from the extractant/diluent mixture employing sulfuric acid to furnish a concentrated metal solution, usually containing 40–50 g Cu/L. The final products of the overall process are copper cathodes, which are obtained from a terminal electrowinning step.Although considered as “technology of choice”, SX to date is still suffering from several drawbacks, such as the risk of poor phase disengagement and the high capital costs of solvent-extraction plants. Especially, the processing of leachates with lower copper tenors (< 1 g/L) is often economically less attractive. As a result, tremendous efforts have been dedicated to the evaluation of alternative technologies to address these challenges."

Jan 1, 2015

By T Graham

The invention of the traditional belt and idler conveyor system is attributed to Thomas Robins in 1892. This basic concept, ubiquitous in most mining operations, has not changed markedly in 125 years and still suffers from five forms of energy losses. Firstly, Wheeler, Roberts and Jones (2004) showed that indentation rolling losses, a consequence of the indentation a roller inflicts on the belt, consume 60 per cent of all energy. Secondly, Wheeler (2006) explained that bulk solid flexure resistance resulting from material interaction and movement absorbs 20 per cent. The remaining energy is lost by idler rotating resistance (5 per cent), belt flexure resistance (5 per cent) and secondary resistances (10 per cent).Traditional conveyor systems are continuous as opposed to batch systems, implying both positive and negative features. On one hand, continuous systems have high availability (available infers ‘ready to run’), but, on the other, utilisation of availability is often low; for example, when a conveyor is running empty or below capacity. This converts into reduced system OEE (overall effective efficiency). Many attempts have been made to mitigate energy losses and low OEE issues, including development of rail-based systems. The key drawbacks have been higher capital cost, low flexibility, poor availability and lack of market acceptance.A recent conceptual development known as the ‘autonomous rail conveyor’ (ARC) promises to address issues of energy loss and poor availability and provide increased flexibility by offering solutions for difficult applications. The ARC uses a series of interconnected skeletal bogies covered with a conventional conveyor belt as a load bed running on relocatable light rail. The bogeys have two axles, each driven by integrated variable speed-geared motor and brake systems, and are connected to previous/next bogeys by specialised couplers. Power is provided by a generator (located in a dedicated bogey) and ‘daisy chain’ plug in cables. Centralised redundant control is provided at a preferred point. Although the system is batch in nature, the number of carriages and trains is flexible, meaning that the system is effectively continuous. Loading is achieved by feeding the ARC from a perpendicularly oriented apron feeder, while material off-loading is accomplished by a twist-dump operation in which the belt is twisted through an angle of 50&#176; below the horizontal, causing material to flow from the belt into a receiving conveyor.The ARC system can run multiple trains over any distance, although benefits are more significant in distances of over two kilometres. Capital and operating costs are predicted to be lower than competitive products, with planned maintenance being simple to perform.CITATION:Graham, T, 2017. The autonomous rail conveyor significantly reduces ore haulage costs, in Proceedings Iron Ore 2017, pp 467–472 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

By D. K. Chibwe, P. Taskinen

"Typical operation of industrial Peirce-Smith converters (PSCs) used in copper smelting results in several common phenomena such as excessive splashing, slopping, tuyere blockage, and low converter campaign life due to high tuyere line refractory erosion. These phenomena give rise to an inefficient operation from both a process and an energy perspective.This investigation presents numerical and physical models of flow patterns, mixing, solid-liquid mass transfer, and slag-matte phase distribution in a slice model of an industrial PSC. For physical simulations, a 0.2-scale cold model was developed. Water, kerosene, air, and sintered benzoic acid compacts were used to simulate matte, slag, injected gas, and solid additions into the PSC. The two- and three-dimensional numerical simulations of the three-phase system were carried out using volume of fluid (VOF) and realizable k- e (RKE) turbulence models to account for the multiphase and turbulence nature of the flow respectively. These models were implemented using the commercial computational fluid dynamics numerical code FLUENT.Both numerical and physical simulations were able to predict, in agreement, the mixing and dispersion characteristics of the system in relation to various blowing conditions employed in the investigation. Measurement of mass transfer characteristics conclusively indicated that fluid flow in the PSC is stratified. Both blowing configuration and volume of slag in the converters were found to have significant effects on mixing propagation, wave formation, and splashing.The splashing and wave motion in these converters are known to cause losses of metal/matte and potential production time due to the requirement for intermittent cleaning of the converter mouth, resulting in reduced process throughput. To prevent these losses and hence to increase process efficiency, we propose a combined blowing configuration using top lance and lateral nozzles. The numerical simulations were conducted on combined, as well as lateral, blowing conditions. In comparison, the recent results of the combined blowing concept were found to be encouraging."

Jan 1, 2014

By Jerry G. Craig

The Bureau of Mines has developed a computerized oil and gas well data retrieval system from studies of the waterflood potential of a producing reservoir. This system stores pertinent information about each well completed in a reservoir and facilitates selection and manipulation of large volumes of data. Use of the system enables reservoir information to be supplied rapidly, accurately, and at less expense.

Jan 1, 1971

By Davis E. Traut

TiC14, AIC13, and VC14 were coreduced simultaneously via a Kroll-type magnesium reduction to form an alloy sponge as part of a research effort by the Bureau of Mines to produce titanium alloy powder. Reduction tem-peratures of 750°, 850°, and 950° C, a range of chloride feed rates from 34 to 152 Ib/(ft2·h), and crucible diameters of 7-3/4 and 10-1/4 in were tested. Microprobe analyses using a 40-um-diam spot showed no effect on alloy microhomogeneity (distribution of Al and V in the Ti, considered critical as a powder property), due to changes in process variables within these ranges. Statistical analysis of these data for alloy sponge indicated greater variation in the chemical composition (AI and V) than that found in commercial Ti-6Al-4V plate or Plasma Rotating Electrode Process (PREP) powder. Therefore, a one-step approach to pow-der, such as grinding the Ti alloy sponge, would not produce a powder product of acceptable chemical homogeneity. An approach that homoge-nized the sponge, such as with a melting step, would produce an accept-able powder product with respect to chemical homogeneity.

Jan 1, 1987

By P. Zhang

Previous studies have shown that separation of clay minerals from non-clay minerals is the key to economic recovery of the phosphate from the waste phosphatic clays (slime) which represent about 20- 300/0 of the phosphate mined. Small hydrocyclones of different sizes (4" and 6") were tested for sizing the slime at +10 microns in order to remove most of the clay minerals. The effect of design variables and operating parameters were evaluated. Multi-stage cycloning and the effect of dispersant were also investigated. Chemical and mineralogical analyses of different size fractions of both the underflow and overflow samples indicated satisfactorily separation efficiency.

Jan 1, 1998

By M. Syamujulu

In 1994, Nkana smelter commissioned a multimillion US Dollar Chilean Teniente converter (CT) in its operations in order to improve its smelting technological base. The main objective was to reduce the overall smelting energy costs and increase the sulphur capture for sulphuric acid production to support the company?s leaching operations at Nchanga tailings leach plant. A decade of CT operations have since ensued following its installation and start-up. This paper provides a technical and operational audit of the performance of the CT in the first 10 years of its operations at Nkana smelter.

Jan 1, 2005

By Jeff Fontana, Daniel Ebin, Eric Westergren, William Currier

The PMott MacDonaldlane Train Tunnel West Extension (PTTWE) Project at HartsfieldJackson Atlanta International Airport includes the extension of the existing automated people mover system and terminal renovations at the baggage claim station. To assist with design, pricing, and assessment of risk on this progressive design-build project, 3D design models were prepared to increase visibility into realities of construction. This paper will focus on the original intent of the models and explore the additional, unforeseen value they provided throughout the design and construction process.

Dec 1, 2022

By K. L. Boyd

"Can the increasing tonnages of ore and waste mined today be accommodated simply by using bigger and better crushers? This paper reviews the latest primary crushing and large cone crusher designs. It highlights how—in addition to using the larger crushers now available—maximizing the utilization percentages in each stage of the rock size reduction circuit will assist in meeting the larger production requirements of existing and new mining operations. Its goal is to help consultants and operators select the optimal equipment and custom-designed items in material handling from the mine to the mill, maximizing operating hours per year.RÉSUMÉ Le recours à des concasseurs plus grands et plus performants suffit-il à gérer l’augmentation du tonnage du minerai et des déchets que l’on extrait de nos jours ? Cet article examine les tous derniers modèles de concassage primaire et de grands concasseurs à cône. Il montre que, pour satisfaire aux exigences de production plus conséquente des exploitations minières existantes et nouvelles, il est important d’optimiser les taux d’utilisation à chacune des étapes du circuit de réduction de la taille des roches, en plus d’utiliser les concasseurs plus grands actuellement disponibles. L’objectif de cet article est d’aider les sociétés d’experts-conseils et les exploitants à sélectionner le meilleur équipement ainsi que des articles fabriqués sur mesure pour la manutention de la mine au concentrateur, en optimisant les heures d’exploitation par année.INTRODUCTIONAre larger crushers the answer to meeting the demands of existing or new mining projects for increased tonnage? Boyd (2002) predicted that larger crushers for primary and fine crushing would be beneficial and contribute to the further development of mineral processing. Larger crushers for primary and fine crushing applications have since become available.The rock size reduction process from the mine to the selected milling process facility has not changed since a person first broke a rock to get at the interesting mineralization within. The rock size reduction process circuit from the mine to the mill can be summarized in four stages:• stage 1: mining (rock breakage and fragmentation);• stage 2: prescreening (scalping);• stage 3: primary crushing; and• stage 4: fine crushing (secondary, tertiary, and quaternary).Plant throughput and ore delivery schedule will form the basis for flowsheet design and equipment selection. Once the project design criteria and the mine plan have been established, the circuit’s utilization percentage (percentage of hours when the circuit is being utilized) can be optimized for each of the four stages. The mining and process departments play a key role in the selection of the final comminution circuit, and both departments should be familiar with and contribute to the design of the circuit."

Jan 1, 2017

Background Chronology 2008 Positions May 8, 2010 Summit Going Forward Say one thing and do another ??Genuine passion to see our customers succeed?, ?A passion for exceeding our customer?s expectations? Self serving actions to gain competitive advantage at the expense of the customer No single supplier or technology can manage the overall mine and its integrated operation - April 2005: Informal meeting at Toronto CIM 2004 -Unanimous interest - July 2005: Formal Calgary Meeting, developed: - Scope and vision statement , - OEM/OTM communication plan, - industry communication plan - Aug?Nov 2005: Aligned with SMART - Nov 2005?May 2006: Direct OEM, OTM communication - May ?Sept 2006: Developed baseline standards - Sept -April 2007: Developed SMART document - Sept 2007: Document distribution and request OEM OTM response - Jan ?May 2008: Acquired OEM OTM feedback and positions May 2008: Results presented to SMART and discussions on next steps

May 1, 2010

By A. Ansari

Standard depressants used in the selective separation of copper and molybdenum sulfides by flotation pose environmental and health hazards. In this investigation, the floatability of chalcopyrite and molybdenite in the presence of six well-characterized lignosulfonates was studied through Hallimond tube flotation and adsorption tests. The results define the physicochemical conditions (lignosulfonate type and dosage, pH, presence of ions) for possible selective separation of molybdenite from chalcopyrite. Dissolved polyvalent cations, especially calcium, appear to serve as "activators" for lignosulfonate adsorption on both minerals. From this point of view, the use of lime as a pH modifier is not always appropriate for optimum selectivity since high dosages of lime to produce high pH tend to indiscriminately depress both sulfides.

Jan 1, 2006

By Karinne Fernanda Menezes de Freitas

O objetivo deste estudo foi avaliar a soldabilidade do aço da classe API 5L X80 de fabricação nacional quanto ao trincamento pelo hidrogênio. Para isto, o aço foi soldado pelo processo arame tubular usando como metal de adição o arame AWS E110-MC com 1,6 mm de diâmetro e como gás de proteção a mistura 75% Argônio+25%CO2. Foram adotados quatro níveis de energia de soldagem, 0,7 kJ/mm, 1,10kJ/mm, 1,60 kJ/mm e 2,00 kJ/mm e o ensaio de soldabilidade utilizado foi o WIC. Para as quatro energia de soldagem foi constatada a presença de trincas que iniciavam na região de grãos grosseiros da ZTA (RGGZTA) e se pronunciavam para a zona fundida (ZF). A ZTA apresentou um comportamento microestrutural com presença de bainita e a ZF com predominância de ferrita acicular. A microdureza na ZF foi superior em relação às demais regiões.

Aug 8, 2018

B. P. Thomson (South Aust. Dept. of Mines): Dr. A. B. Edwards presents in the published discussion on the above paper, his interpretation of the dolerite dyke-lode relationship as a major objection to the syngenetic origin for Broken Hill ore bodies to which Mr. Condon subscribes.Dr. Edwards claims that the dyke-lode relationships indicate clearly that the Broken Hill orebody was formed after the folding and regional metamorphism of the Willyama Series. Field evidence compels m:e to join with H. F. King (1958), in finding this conclusion unacceptable-as follows:(1) Broken Hill type mineralization, as defined by King and Thomson (1953), is clearly confined by the regional sillimanite isograd and does not extend into the lower temperature almandine and biotite zones as does the Thackaringa type mineralization.(2) In the Corruga area the writer has mapped a dolerite dyke apparently cutting a conformable quartz gahnite lode of the Broken Hill type (see Edwards 1958, p. 12).(3) At the Champion mine a similar lode is cut by a fine grained granite dyke.An alternative history of the dolerite dykes to that given by Stillwell &amp; Edwards (1956, p. 26) appears possible. It has been suggested by the writer in an unpublished thesis, (Thomson 1955, p. 30), that the pyroxenite and serpentine plugs and dolerite dykes are rheomorphic end-products of a complex regional metamorphic and metasomatic process involving the formation of amphibolite and pegmatite. The analogy in composition between the dolerite and normal amphibolite is close (see Edwards 1958, Fig. 1, p. 22).

Jan 1, 1959

The Anglo Platinum Mogalakwena North concentrator was one of the first three hard rock mineral processing applications to utilise a HPGR circuit in preference to a SAG milling circuit. This was driven by the extreme competence of the ore that had given rise to throughput issues on the original South Plant. Despite more than adequate capacity and performance of the individual comminution units, the circuit has struggled to obtain full design capacity. A thorough review of the integrated performance of the circuit revealed a wide range of bottlenecks and underutilised parts of the circuit arising from design issues. Interestingly, the major opportunities lie in materials handling issues and mismatch of loads about the circuit. This paper presents the high-level outcomes and the associated opportunities, and is pertinent to all HPGR circuit designs.

Jan 1, 2012

By Brandon Smith, William Bohlen

Riverbank erosion along the Brazos River has modified the structural support conditions for an existing bridge structure in Southeast Texas. The 8-span, 1,200-foot long bridge was completed in 1992 along Highway 99, a key corridor for the Houston area. Foundation support conditions have been altered due to natural, ongoing river morphology for three bents on the affected riverbank. Recent events include three storms within a three-year period that each exceeded the 500-year return probability. This resulted in accelerated erosion along the northern riverbank. Over forty vertical feet of soil has been removed in some locations. Drilled shaft foundations are now functioning as columns. This review considers the behavior of columns and deep foundation elements due to the modified support conditions. The review ultimately presents a rational methodology for the review of the present and anticipated future bridge configuration with respect to current strength, service, and extreme event limits, with corresponding safety factors for the design. Introduction and Background The Fort Bend Grand Parkway Toll Road Authority (Owner) oversees the State Highway 99 Judge Jodie Stavinoha Bridge over the Brazos River. The 8-span, 1,200-foot long bridge was originally designed with four southern approach spans of 115 feet; a three-span continuous river crossing with span lengths of 195 feet-235 feet-195 feet; and a single northern approach span of 115 feet. Figure 1 provides an overview of the northern portion of the bridge. Table 1 provides a summary of the original bridge substructure configuration. When the bridge was completed, the centerline of the Brazos River was located at midpsan of the main span and the northern high bank was originally located at Bent 7. Erosion process has forced the northern high bank to Abutment 9 (also called Bent 9), and the river centerline during normal flow conditions is located south of Bent 8. The overall estimated river excursion at present day is approximately 270 feet to the north from its original location. An erosion wall that was installed in 2017 limits further erosion of the north river bank within the 300-foot width of the roadway corridor right-of-way. This study was developed to better understand the existing conditions, and to forecast the need, nature, and criteria for potential future improvements.

Jan 1, 2019

By R. C. Gegg

The property of MacLeod-Cockshutt Gold Mines, Limited, is in the Little Long Lac area of the Thunder Bay Mining Division, Ontario, about four miles south of the Canadian National railway station of Geraldton. The town of Geraldton is about 170 miles east of Port Arthur and 22 miles west of Long Lac, which is on the main line of the Canadian National railway. The Trans-Canada highway passes through the property and during the winter season the road is maintained between Port Arthur and Long Lac. Electric power is supplied by the Hydro-Electric Commission of Ontario from their Cameron Falls plant and is distributed at 44,000 volts by a local transformer station. Early History The South orebody was discovered by diamond drilling in the spring of 1936. This is a porphyry contact orebody with major pyrite and minor arsenopyrite mineralization. Some gold is present in the &apos;free&apos; state, but for the most part it is locked in the pyrite, with little or none in the arsenopyrite. Preliminary test work was based on comparing straight cyanidation with bulk flotation and cyanidation of the concentrates. The overall results by the two methods were found similar, but for economic reasons the flotation-cyanidation process was adopted. Before the mill came into operation, a new and entirely different type of orebody, consisting of massive sulphide lenses in banded iron formation, was discovered to the north of the existing workings. These bodies are an intricate mixture of pyrite and arsenopyrite, both carying gold.

Jan 1, 1949

By J. Matos e Silva

"Abstract The paper refers five examples of executed diaphragm walls with one to five levels of anchorages. For each example a comparison is made between the results of the design using the traditional criteria and the ones obtained by using the Eurocode 7 and the Portuguese Norm related to the EC7 (NP EN 1997-1 2010), as the Eurocodes consider the possibility of each European country to adopt their own geotechnical parameters as these may vary from one country to another. A comparison is also made with the behaviour of these diaphragm walls during its construction phase; so the values from the design that were adopted in the construction are listed in tables 2 to 6 in the “Construction Options” columns. From the referred examples the author presents his conclusions.1 - INTRODUCTIONIn the so-called “traditional criteria” the soil parameters (dry unit weight, angle of internal friction and cohesion) have values obtained through tests or empirically. It is assumed that the characteristic values of the soil parameters adopted in Eurocode 7 are equal to the soil parameters adopted in the calculations by the “traditional criteria”. In the present paper the design according to the EC7 was made considering only the ultimate limit states and the respective Cases A1 and A2.The partial coefficients used for the actions in the present paper were those on Table 1:2 - CONSIDERED EXAMPLESThe five examples correspond to the situations of figures 1 to 5, and the respective results are summarized in tables 2 to 6.For all the considered examples, that are case histories, a simplified method published by Matos e Silva (March 2000) was adopted. This method attains good performances when compared with other more sophisticated ones as the finite elements and the elasto-plastic methods as described in a paper published by Matos e Silva (2000). All the obtained values were compared between them and with the so-called “Construction Options” that were the values adopted by the designers of the considered examples. For the wall bottom length calculation two possible situations were envisaged:"

Jan 1, 2014

By Walter R. Storms

This paper, which describes the mining methods of the Potash Co. of America in Eddy County, N. Mex., is one of a series being prepared by the Mining Branch, of the Bureau of Mines on mining practices and methods in various mining districts of the United States. 4, T. The mine and plant of Potash Co. of America are situated in sec. 20 S., R.. 30, E.,. N. M. P. M., about 22 miles by paved road northeast of Carlsbad (fig. 1). A spur line of the Atchison, Topeka & Santa Fe Railroad also connects the mine with Carlsbad.

Apr 1, 1948

By Robert L. Goehring, Fred Munene

This case history for three major additions to a tire manufacturing plant in north Georgia discusses the impact of highly variable Karst (sinkhole) geology on both foundation design and construction. For Phase 3, alternatives were considered for a heavily loaded mixing tower with up to 2,000 kip column loads. A non-traditional approach of using auger cast piles was adopted. To satisfy a risk-averse Owner, a pile load testing program including strain gauges was conducted. One test pile was cored full depth to demonstrate effective bearing on steeply inclined rock. For Phase 4, similar methodology was used at another critical mixing tower structure associated with that phase and the installation procedure was refined to reduce grout take. In Phase 5 construction, another mixing tower and the equipment pits were supported on auger cast piles. After pile installation, a sinkhole opened up between two equipment pits. Remedial grouting was performed. Development also required construction of a R&D building over a 96 inch diameter RCP storm sewer. Auger cast piles were used to support the building to avoid overstressing the existing pipe and risk separating joints. This case history outlines lessons learned while constructing auger cast piles in Karst and proving quality was built into the foundations.