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By D. G. Kotlyar

Oxidized sulfide minerals, commonly formed by weathering, increasingly are becoming important resources as primary sulfide deposits are depleted. Oxidation, in general, reduces selectivity and recovery in flotation. The objective of this research was to determine the mechanisms of oxidation and collector adsorption in sulfide systems as a step towards developing improved technologies for beneficiation of oxidized sulfides. Data from microcell flotation tests are combined with electrochemical and spectrophotometric measurements. The adsorption of ethyl xanthate (EX) on oxidized sulfides was characterized using AC impedance techniques and correlated to flotation response. The floatability of chalcocite, chalcopyrite and pyrite was dependent on the extent of oxidation. Oxidation for 1?) d at 65? C inhibited flotation of chalcocite, and chalcopyrite even after the addition of a large amount (1 x 10') M) of collector. Heavily oxidized (27 d) chalcocite remained unfloatable. However, longer oxidation of pyrite and chalcopyrite restored the floatability. Oxidation resulted in an electrically resistive, nonporous surface layer.

Jan 1, 1993

By Kadri Dagdelen

The importance of factoring mining dilution into feasibility stage ore reserve estimates is discussed. An example that relates the issue of mining dilution to the size of smallest volumes that can be mined selectively is given. Relationship between the size of the selective mining units (SMU's) and the variance of their average grade is presented. The concept of variance reduction and its use in factoring mining dilution into geostatistical ore reserve estimation is described. An approach to estimate variance reduction factor by way of conditional simulation is introduced.

Jan 1, 1996

By M. W. Harmer

Laboratory test work, pilot plant testing and plant experience with wet, fine mesh screens is discussed. High frequency, low amplitude, fine mesh, wet screens were installed on the 6th cleaner concentrate stream. These screens produce a 60% oversize recovery split of the molybdenite concentrate that is of leached equivalent quality. The undersize stream is leached with 'hot ferric chloride to remove Pb and Cu contaminates before roasting or shipment to sulfide customers. This plant modification is a good example of using existing technology in an unconventional flow sheet to yield significant cost savings.

Jan 1, 1990

By Lech Bolinski, Joe Shirley

Over two dozen Resin-in-Pulp (RIP) plants have been in operation in the fonner Soviet Union since the beginning of 1970. Currently approximately 50% of the gold produced in Russia, Uzbekistan and Kazakhstan, which equates roughly to 10% of the total world output, is recovered using the RIP technology. An alternative to activated carbon, ion-exchange resin has several advantages for the recovery of gold. Resins are less prone to clay or organic foul­ing and do not require thennal regeneration. They also have poten­tially higher loading capacities, faster adsorption kinetics, and are more resistant to abrasion. The original Russian RIP process was based on engineering expertise developed for uranium ore pro­cessing and has a few major drawbacks. The biggest problem is an 18-30 day resin stripping process, which results in a huge gold lock-up. The problem was recently overcome by the development of a new stripping technology which cuts the time to 12-24 hours. Currently, there is rather limited practical knowledge of the Russian RIP process in the West, but often Western mining com­panies developing gold mines in the fonner Soviet Union are hard pressed by their local joint venture partners to take a serious look into RIP. It is probably only a matter of time before one of these companies will cross Soviet technology with Western engineering and develop a modem RIP process as an attractive option to the current CIP and CIL technologies.

Jan 1, 1996

By E. H. Cho

An acid mine drainage sludge sample containing 3.6% Mn, 0.18% Ni, 0.15% Co and 0.39% Zn was obtained from the central part of West Virginia and leached with sulfuric acid at pH 3 and at room temperature. Air was bubbled through the leach solution to precipitate the iron content from the solution at pH 4.5. The leach solution contained 258ppm Zn, 105ppm Ni, 90ppm Co and 2050ppm Mn. Three distinctive solvent extraction studies were performed to selectively extract these metal ions in a sequential operation with DEHPA, Cyanex 272, and then Lix 84. The AIO ratio was 1 and the extractants were diluted 10 times with kerosene. A study of solvent extraction with DEHPA showed that Zn could be extracted without Ni or Co up to pH 2.75. A pH of 2.75 is considered to be the optimum pH to maximize the selective separation of Zn from Mn. The separation factor of Co over manganese suggests that the two metal ions be extracted together by Cyanex 272. The separation factor of Co over Ni indicates that these two metal ions can be separated most effectively around pH 5.3. Also, Ni can be effectively extracted by Lix 84 above pH 5.5.

Jan 1, 1993

By P. E. Lenz

In Nevada, large deposits of diatomite were formed during the Miocene / Pliocene in deep, alkaline lakes. Associated sediments are predominately fine-grained lacustrine with varying fluvial-clastic layers. Tephra layers are common, and probably served as a source of soluble silica for the extensive diatom colonies. Four companies are currently producing diatomite from Nevada deposits. Nevada ranks second to California in annual domestic production. Nevada diatomite is used in filter aids, fillers, absorbents, catalyst supports, and many other products. Economics and mineral reserves indicate that diatomite production in Nevada will be significant well into the next century.

Jan 1, 1993

By J. R. Dyni

The lacustrine Green River Formation of Eocene age in the Green River Basin, southwest Wyoming, contains the world's large resource of natural sodium carbonate in bedded trona (Na2CO3? NaHCO3? 2H2O). In this study, 126 billion tons are estimated to be present in 22 beds, ranging from 4 to 36 feet in thickness. These 22 beds underlie areas of about 50 to more than 775 square miles at depths ranging from about 650 to more than 3,000 feet below the surface. The total resource of trona ore in the basin for which drilling information is available is estimated to be about 149 billion short tons. The undeveloped deeper trona beds in the southern part of the basin may be exploitable by solution mining. Additional unevaluated sodium carbonate resources are present in disseminated shortite in strata interbedded with the trona and in shallow sodium carbonate brines in the northeast part of the basin.

Jan 1, 1995

By M. S. Klima

This paper will detail the work that has been done in 1990 and 1991 at the Pittsburgh Energy Technology Center (PETC) on the development of the U.S. Department of Energy's (DOE) patented Micro-Mag Process. This process entails the use of a micronized-magnetite (minus-10 microns) and water mixture for cleaning fine coal (minus-600 microns) in a cyclone, along with the production and recovery of the magnetite. Emphasis will be on the results from micronized-magnetite recovery testing using both conventional wet-drum and high-gradient magnetic separators in which several key operating conditions were varied including feed percent solids, magnetite-to-coal ratio, magnetite grade, and feed rate. Also presented will be selected results from testing done on a 76-mm diameter cyclone at a relative density of 1.50 showing the effects of operating pressure, magnetite grade, and medium-to-coal ratio on separation efficiency. In addition, two methods for producing micronized magnetite, along with the corresponding size distributions, will also be discussed.

Jan 1, 1992

By F. F. Peng

For a given flotation system, the rate-controlling step in fine coal flotation can be the attachment of bubble to the fine coal particles. The probability of attachment depends upon the surface characteristics and the collector adsorption on the coal particle surfaces. Thus, to further improve the flotation rate and selectivity is to alter the distribution of hydrophobicity on coal particles efficiently by applying the effective means of collector dispersion. In order to enhance the combustible recovery and selectivity of fine coal flotation, and lower the amount of collector used in the fine coal flotation processes, four modes collector dispersion/addition techniques are tested. They are direct liquid addition in pulp (mechanical agitation), ultrasonic energy dispersed collector into pulp, atomized collector into pulp and gasified collector transported by air and formation of collector-air bubbles in pulp. The effect of four collector dispersion/addition methods on the floatability of -600 µ m (-30 mesh US sieve) coal samples from four ranks of coals were studied. The ultrasonic energy dispersed collector and atomized collector flotation processes enhance flotation yield and most significantly improve the kinetics of fine coal flotation. The gasified flotation process can increase the selectivity of separation and decrease the collector consumption due to starvation flotation is achieved in this process. The flotation efficiency of gasified collector process is as high as that of conventional flotation. It is because that the effect of the decrease in the combustible recovery is offset by the increase of selectivity. In addition to flotation experiments, induction times of bubble attachment to the bed of fine coal particles for four ranks of coals are measured.

Jan 1, 1992

By B. E. Davis

A technique was devised by the Bureau of Nines to recover coarse (plus 1.7 m) North Carolina phosphate now discarded as waste. Concentration utilizes the difference in particle sphericity between the relatively flat calcite and the rounded phosphate pebbles. Sizing, treatment in a continuous hydraulic classification system, and grinding upgraded a 17.0-pct-P205 reject to 24.5 pct P205 with a phosphate recovery of 63 pct.

Jan 1, 1985

By L. Wilfley, G. A. Calboreanu

"Wear resistant 25% Cr hard cast iron is a heterogeneous material containing 1,600 HV hard chromium carbides embedded into martensitic matrix. It is the most used cast alloy in wear operations of mining industry, chemical industry, fracking, etc. During the last decade the hard iron manufacturers increased wear resistance in the following ways: • Reduce / eliminate harmful retained austenite and chromium + carbon depleted areas especially within eutectic carbide cells. In Max5A the improved 25% Cr alloy brittle carbide eutectic cells are removed which increases toughness. • Delay delamination between chromium carbides and matrix particularly during sliding abrasion wear. Delamination commences at the weakest chromium carbide – matrix contact lines exposing chromium carbide to localized bending stresses which gradually chip carbides away. Once the chromium carbide base is removed accelerated wear degradation occurs. • Evenly randomize and avoid discontinuities in the chromium carbide network. When chromium carbides are perpendicularly positioned to the wear surface wear rate is minimized. On the contrary the highest wear rate takes place when 3D oriented chromium carbides are parallel with the wear surface. In this case chromium carbides cover top network discontinuities being just matrix supported which leads to rapid carbide cracking. The new hard iron discussed in this paper fulfils and exceeds the above requirements increasing the casting service life around 250% in various wear operations. Physical and mechanical properties of Max5A are compared with well manufactured Industry Standard 25% Cr hard iron. The paper presents practical accomplishments obtained in the USA, South America and India. INTRODUCTION The chemical composition of Max5A and Industry Standard 25% Cr are identical like specified by ASTM A532 Class III Type A requesting C = 2.0-3.3% and Cr = 23.0-30.0%. Manganese, silicon, nickel, molybdenum and copper contents are limited by maximum accepted quantities.[1] According to published data the eutectic chromium carbides chemical composition of 25% Cr hard iron is Cr = 54% and C = 8% which indicates strong depletion of chromium and carbon along chromium carbide - matrix interfaces. [2] The optimized Max5A microstructure has minimized retained austenite, is devoid of harmful carbide eutectic cells and has the most advantageous chromium carbide and secondary chromium carbide morphology. This favorable microstructure increases wear resistance, mechanical properties and corrosion resistance. The paper compares the Max5A microscopic features and mechanical – physical properties with similar properties of high quality Industry Standard 25% Cr."

Jan 1, 2017

By Miguel Angel Banuet Rodriquez, Hector Conceso Aragon

"Construction of the longest road tunnel in Mexico is part of Acapulco’s alternate roadway to the scenic roadway project. This article provides a brief summary regarding details of the construction since the project began. Work was done by skilled engineers and Mexican workers that includes outstanding execution of the work in an urban and tourist area. This project will have a strong impact on the steady development of the port of Acapulco.Project locationAcapulco’s Alternate Roadway Project is currently under construction in the city of Acapulco, located in the Guerrero state of Mexico. It is an 8-km (5-mile) long roadway that starts at the Icacos neighborhood, heading toward Acapulco’s airport, crossing a drill-and-blast tunnel that is 3,160 m (10,350 ft) long under Veladero Park, a mountainous area. The tunnel overburden reaches a depth of 380 m (1,250 ft) approximately at the central point of the horizontal alignment, and continues with a 4-km (2.5-mile) long elevated road to connect to the existing Viaducto Diamante toll road (Fig. 1).Tunnel construction has already been finished using two portals for drilling and blasting operations; Brisamar Portal on the west side and Cayaco Portal on the east side (Fig. 2).Geometric dataOnly two main excavation cross-sections were considered in the executive design. The first one corresponds to the running tunnel for a three-lane vehicle section, with crosssection areas between 120 and 130 m2 (1,290 and 1,400 sq ft) taking into consideration different geological and geotechnical conditions to be encountered (Fig. 3).The second one corresponds to eight emergency bays, each one 50 m (164 ft) long, laid down within the tunnel outline, located at 400 m (1,312 ft) from each one with excavation cross-sections between 156 to 169 m2 (1,680 to 1,820 sq ft), in accordance to geological and geotechnical conditions to be encountered (Fig. 4).The tunnel’s horizontal alignment is completely straight, having a small curve within the vertical alignment on Cayaco portal, because there were some differences in levels when tunnel excavation started. Tunnel slopes were designed in such a way that rainwater does not run down into the tunnel at any time."

Jan 3, 2018

By Donald C. Violetta

During 1963 more than 200 million tons of coal was consumed by the public utility industry. Ash residual, from the burning of this coal has caused disposal problems. In order to prevent air pollution, dust collection systems have been installed to suppress the fine ash emitted from boilers before it is discharged into the atmosphere. The collected ash, called fly ash, must be removed from the dust collection facilities and disposed in some manner, usually at a substantial cost to the utility. Many years ago fly ash disposal was not a serious problem because the demand for power was relatively low and coarse coal was burned on stoker type traveling grate machines. The low burning rates utilized on these grates did not produce large quantities of air-borne coal ash while the mechanics of this system of burning resulted in a clinkered residue of cinders. Both of these conditions resulted in the production of only small quantities of fly ash. As the demand for electricity increased, a method of burning pulverized coal was developed and was introduced to the public utility industry in 1918 at Milwaukee, Wisconsin.1 This method of firing was considered a significant advance in steam electric generation as it (1) permitted the use of coals that could not be economically fired by other methods and (2) has generally resulted in improved efficiency, greater generating capacity and lower costs.2 The utilization of larger generating units using pulverized coal substantially increased the quantity of fly ash which resulted in a' more severe disposal problem.

Jan 1, 1966

By Manuel Gomez

This, report presents correlative data on ash and sulfur forms for Pittsburgh seam coal from Greene County, Pa. Prediction equations were developed for sulfur form distributions and washability characteristics. The equations developed were evaluated with an independent set of exploration core samples. It was demonstrated-that both sulfur distributions and washability characteristic's can be predicted from exploration core data ahead of mining. This information can improve the efficiency of selective mining techniques for the production of a uniform metallurgical coal product.

Jan 1, 1967

By Michael A. Redfearn

The use of nitrogen in molybdenite flotation circuits is a relatively new concept utilized by only a few mines around the world. This paper briefly discusses circuit chemistry and the reason for its success. It follows the entire testing process which resulted in a sodium hydro-sulphide dosage reduction from 18.5 to 4.4 lbs/ ton (9250 to 2200 g/t) through to the final installation of an inert gas generator which will produce reagent savings in excess of one million dollars annually. A description of the various types of nitrogen generators commercially available is included.

Jan 1, 1983

By N. A. Abdel-Khalek

An experimental study was carried out to compare the results of laboratory column flotation versus conventional flotation for the cleaning of fine coals of different rank and their oxidized samples as well as discrete size fractions. Surface oxidation caused a significant reduction in the flotation yields achieved with both methods. However, the column flotation produced concentrates? of lower ash content. The recovery -size curve achieved with column was similar in shape to that obtained with conventional cell, with superior results in the region of fine fractions. The influence of some column operating parameters was also discussed.

Jan 1, 1990

By Mustafa Akser

Dwindling economic deposits of rutile (TiO2) prompted research on recovering it from secondary sources. One such resource is in Southwestern Turkey feldspar mining rejects that contain 5%-6% TiO2 mostly rutile. Previous studies could only upgrade the TiO2 to 36% TiO2 grade, claiming that the leftover flotation reagents on the surfaces of rutile and host minerals detrimentally affected upgrading of the rutile. As a potential remedy to this concern, a laboratory sample (from an operating feldspar mine in Milas, Turkey) was first caustic scrubbed, then air calcined at 400°C to evaporate the oily reagents. Subsequent physical separation of rutile included wet screening, gravity, magnetic, and electrostatic separation technologies. Such processes indeed improved the grade of the rutile concentrate up to 83 % TiO2 Low recovery of the TiO2 (<20%), was attributed to lack of full liberation of rutile, even at 125x44 micron range; interference with impurity minerals such as sphene and zircon; and simply the process inefficiency in upgrading the fine grains through conventional separation technologies for heavy minerals.

Jan 1, 1999

By R. G. Reddy

Spent potliner is generated in the electrolytic process of reducing alumina to make aluminum metal using molten fluoride electrolytes. Spent potliner (SPL), obtained at the end of the pots life contains a small amount of cyanide and fluoride which present environmental problems. SPL became a listed hazardous waste under RCRA (K088) in July, 1990. One of the possible uses of spent potliner is as a substitute for Calcium Fluoride (CaF2) as a fluidizer in iron-and steel making and in the glass industry. A low viscosity and a low melting point are prime requirements for accomplishing the above. In the present study, viscosity measurements were performed. Effect of various additives, such as sand, limestone, dicalcium silicate dust, and MgO baghouse dust on the viscosity of SPL was studied. The results showed that with the addition of sand and limestone, very high fluidity of SPL melt can be obtained at significantly low temperatures. The results suggest that spent potliner mixed with additives could potentially be used as a substitute fluidizer.

Jan 1, 1999

By Tung Vu, Michael Harrington

"The Regional Connector project is a complex subway light-rail project that runs through the heart of downtown Los Angeles, CA. The design-build contract was awarded in April 2014 to the Regional Connector Constructors (RCC), a joint venture between Skanska USA Civil West California District Inc. and Traylor Brothers Inc., and their designer team consisting of Hatch Mott McDonald (HMM) and subconsultants. The project broke ground on Sept. 30, 2014 and is expected to be completed in winter 2021-2022. The main components of the project consist of 6.4-m (21-ft) diameter twin-bored tunnels, three crosspassages, a 87-m (287-ft) long crossover SEM cavern, three new underground stations, and cut-and-cover tunnels along South Flower, Alameda and 1st Streets. The project map is shown in Fig. 1. The design and construction challenges associated with each of these main components are discussed in the following sections.Project designTunnels. The Regional Connector tunnels are located primarily within the Fernando formation consisting predominantly of extremely weak to very weak, massive, clayey siltstone. About 305 m (1,000 ft) of tunnels on the eastern end are in alluvium and mixed face of Fernando formation and alluvium. The tightest curve of the tunnel alignment is 178 m (583 ft) in radius. The tunnels were designed with a reinforced precast concrete tunnel lining (PCTL) to be used with an earth balance pressurized (EPB) tunnel boring machine. The PCTL ring is 5.5 m (18 ft-10 in.) inside diameter, 27 cm (10.5 ft) thick, 1.5 m (5 ft) long, and consists of five segments and a key. The segments were designed with 6,500 psi concrete, 80 ksi yield strength wire rebar, convex joint surfaces to enhance seismic performance, and a dosage of 1.7 lbs polypropylene microfibers per cubic yard of concrete for fire resistance. The rings are designed with a right ring and left ring pattern and a taper of 3.8 cm (1.5 ft) to allow for alignment curve negotiation. Figure 2 shows typical sections of the PCTL. The structural lining design was modeled using FLAC 3D. Four consecutive rings were modeled with 32 dowel connection. Loading considered in the lining design includes temporary ground load during excavation, long-term ground loads, seismic loads and train loads. The seismic loads on the PCTL were simulated with the racking deformation applied at vertical boundaries of the model. Due to variation of geologic condition along the tunnel alignment and different surcharge requirements, three different types of reinforcement (typical, heavy and extra heavy) were designed. The typical type PCTL was required in Fernando formation, the heavy type was required in alluvium or mixed-face condition, and the extra heavy type was required in alluvium and underneath the Japanese Village Plaza where a surcharge of 1,000 psf was required."

Sep 1, 2019

By William C. Cassun

The important role of pit slope design in the overall economics of open-pit mining ventures has long been recognized and well documented.1 2 Any economic analysis of a pit slope design must reflect the trade-off between the benefits and the increased risk of slope failure inherent in steeper slope angles. The conventional practice in pit slope design is based on the factor of safety analysis.&apos; The approach uses average values to arrive at a deterministic solution to slope stability. Past analyses based on this approach generally have had several drawbacks: 1. The risk from slope failure could not be quantified. 2. Only relatively simple pit wall geometries were considered. 3. The analyses were performed solely for the benefits of steeper slopes, seldom discussing the potential costs of slope failures. 4. Analyses were seldom considered in a year-by-year mine planning context. Recognizing that average values do not reflect the uncertainty in geomechanical parameters, a probabilistic philosophy to slope design has evolved. 4,5 However, this approach had not been fully developed to include complex slope geometries and mine planning consideration. In order to develop an integrated, pit slope design procedure, a research project was initiated at the University of Arizona during 1973 in conjunction with the Canada Center for Mineral and Energy Teleology. This research resulted in the development of a probabilistic design procedure. The remainder of this paper gives an overview of the probabilistic design procedure and a discussion of the requirements for the benefit-cost analysis of a pit slope design. Finally, the results of a case study undertaken at an operating mine are given.

Jan 1, 1977