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By Gregory Molinda

NIOSH has recently completed a study of the interaction between polyurethane (PUR) and coal mine roof in order to determine the mechanism of reinforcement, in both highly fractured rock and unfractured rock. Four case studies of PUR rein­forcement are presented. At a West Virginia site, a borehole camera revealed the location of roof voids and guided the PUR injection. By injecting polyure­thane into a zone from 0.6 to 1.8 m (2?6 ft.) high in the roof, a roof beam was created and ongoing intersection falls were halted. In another highly fractured roof in a western Pennsylvania mine, a total of 5.8 cm (2.3 in.) of rubbleized rock was found in a zone up to 3 m (10 ft.) into the roof. Untargeted PUR injection filled approximately 1/2 of the fractures. These two case studies showed that it is not necessary to fill up 100% of the void space to create stability. In the other two field sites, polyurethane was injected into weak, but unfractured roof. Post-injec­tion video monitoring showed that weak bedding planes were hydraulically wedged open and polyure­thane injected along bedding. The reinforcement value of this injection method is limited because of the wafer-thin layer of PUR introduced along bedding, and the lack of a PUR ??webbing?? which would serve as a structural framework to provide strength. It was determined that video inspection prior to PUR injection can aid in identifying the fracture zones to target, and minimize ??blind?? pumping and loss of PUR.

Jan 1, 2008

By W. A. Boyer

If the safety factor of a mine hoisting rope is checked for the lowest depth, is the rope then safe for all levels? The answer here is no. A new set of volues is proposed. HOISTS for metal mines are seldom designed for one particular depth. They are intended for an ultimate load and depth with a given speed but are first used to hoist from a shallower depth, gradually working to the ultimate as development progresses. Sometimes the mine is not worked on the upper levels because of the kind and grade of ore, the mining of which depends on market conditions. In this case the hoist may be used in the same working shift to lift ore, for example, from the 1200 and the 4400-ft level. This means that the safety factor of the rope should be considered not only when it hoists from the greatest depth, but also when it hoists from the upper levels. According to present safety factor standards, a hoisting rope can have ample safety margin when hoisting from the lower or lowest level and still be shy of sufficient safety margin when hoisting from the upper levels. This will be shown by curves presented in this paper. The safety factor of a hoisting rope as defined by the U. S. Bureau of Mines' is the ratio of the static load to the tensile strength of the rope. It is calculated by dividing the breaking strength of the rope* In Fig. 1 curves A and B are plotted against the vertical suspended distance, which is the abscissa, and the factor of safety, the ordinates. Curve C is a smooth curve drawn through the midpoints of the steps of curve A and would more closely represent the minimum factor of safety for new rope for any particular depth. Curve D is a smooth curve drawn through the midpoints of the steps of Curve B and would more closely represent the minimum factor of safety when ropes should be discarded for any particular depth. If a new rope were applied with a minimum factor of safety, then the vertical distance or ordinate under curve D for any particular depth, divided by the vertical distance or ordinate under curve C, would represent the minimum percent of remaining area intact that could be allowed before the rope would be discarded. Also, that section of the ordinate between curve C and curve D is proportional to the section of the area of the new rope that could be allowed for wear and broken wires. Considering that point at 2500 ft on the chart, the ratio do to co would be the minimum percent of rope area remaining at the point of discard. The ratio of that portion of the ordinate cd to co would represent maximum percent of area of a new rope available for wear and broken wires.

Jan 1, 1955

By Ed Coulter

Nevada has a population of 1.2 million people of which only 21% or 312,rIXl are native born Nevadans. he population between 1980-1990 grew a record of 389%. and ls the fastest growing state in the union. While only 2.5% of the total population are direct employees of the mining industry, excluding Reno and Las Vegas, one out of every three people living in Nevada work for a mining related industry.

Jan 1, 1991

The last one decade in India had witnessed phenomenal growth in the coal, lignite, iron ore, limestone and many other minerals. To meet the production of minerals, the energy requirement is also very high both in terms of petroleum products and electrical energy. Because of spiralling prices of crude oil and also of the electrical energy, attempts have been made by various mining companies to select energy efficient mining machineries (in underground and surface) mines, together with energy efficient mining and mineral processing operations. Mineral sector of India is working seriously in tune with the Planning Commission stipulations to conserve energy.Apart from energy use (both electrical and fossil fuels) in mining, it is also significantly contributing to the cost of mineral processing depending on the quality of ore/rock being processed. Over the years a number of studies have been initiated and implemented to reduce cost of energy bill. An attempt has been made to present the various aspects of energy conservation, Impact of Energy Conservation Act 2001, Relevance of energy audit, areas of energy consumption and conservation efforts, etc in this paper. The paper also deals with some of the case studies to show how energy conservation programs has been implemented in the Indian mineral sector.

Dec 6, 2010

Richard Moldenke, New Pork, N. P. (communication to the Secretary*):—Iu following the discussion of the specifications for cast-iron and finished castings, I mas strongly impressed with two points which might be ventilated to advantage more fully from the view of the practical foundry-man. First, the reasons why only silicon and sulphur were specified and not phosphorus, manganese and carbon; and second, why but little was said concerning the presence of iron oxide in the material.

Jan 1, 1905

By Monserrat Rebolledo, John L. Uhrie, Candice Simms

Jetti Resources has developed a catalyst which overcomes the passivation of primary sulfides, enhancing the heap leach recovery of copper from chalcopyrite. It is important to emphasize that heap leaching of copper sulfides, whether with or without the Jetti catalyst, is a biooxidative process. We have demonstrated the efficacy of Jetti’s technology in bioreactors, with a chalcopyrite-dominant porphyry ore. In the presence of Jetti’s technology, optimal bioleaching conditions significantly increase copper extraction by 2.6 times that of the baseline.

Feb 1, 2025

By J. Brent Hiskey

Thiourea is a reagent capable of complexing gold and silver. With the proper oxidant, thiourea can effectively dissolve metallic gold and silver. A thiourea leaching system possesses potential advantages over conventional cyanidation including acidic solutions instead of alkaline, cationic complexes in contrast to metal-cyanide anoionic complexes, and faster leaching rates with acidic solutions of thiourea. Generally, the rate of metal dissolution in thiourea solution is fastest when ferric ion is employed as an oxidant. Acidic thiourea solutions for certain applications could offer a nontoxic and noncorrosive option for processing gold and silver. This paper provides information regarding the fundamental chemistry of the acidic thiourea system, and reviews the dissolution of both silver and gold in acidic solutions of thiourea. The recovery of gold and silver from acidic thiourea solution is discussed.

Jan 1, 1981

By Khosrow Ebrahimi

We document the occurrence, mineralogy and firing properties of magnesite from hydrothermal deposits in the upper Cretaceous ophiolites of Eastern Iran. These deposits occur as thin veins of nearly pure cryptocrystalline magnesite with only trace amounts of quartz, calcite, and dolomite. Ratios of CaO/SiO2 in the bulk ore are variable. The composition of dead-burned magnesite from these deposits, calcined at different temperatures and durations, is predictable from phase equilibria. Textures of the calcined magnesite were studied by SEM; results show that the principal impurites produced during calcination are forsterite, monticellite, merwinite, dicalcium and tricalcium silicates, and that these are located principally at triple junctions of the periclase crystals. The size of periclase crystals increased with temperature and duration of calcination and the amount of periclase-periclase grain contacts was reduced by increasing the concentration of impurities in the starting products. EPMA show: 1) that CaO and FeO contents in periclase increase steadily with an increase in the CaO/SiO2 ratio and the FeO content of the bulk ore, respectively, and 2) that limited solid solution exists between impurites. This study demonstrates that high quality dead-burned magnesia refractories can be produced from the Iranian magnesite with low CaO, SiO2 and FeO contents.

May 1, 2004

By Karl Hoover, Cassandra Spence, Brent Bachewich, Jaakko Levanaho

"Since year 2000 the gold recovery into copper concentrate at Hudson Bay Mining and Smelting has reduced considerably during the summer months. This initiated a testwork program in the fall 2003 to find reasons and solutions for such change. During the testwork mineralogical factors, water quality effects and different reagents were evaluated. The reason for the lower recovery appeared to be the sensitivity of some collectors to pulp temperature. As a result of the work different collectors are being used in the mill in the summer than in the winter. A modification to the water system was also made to provide cooler water to the mill in the summer months.INTRODUCTIONHudson Bay Mining and Smelting (HBMS) is located in Northern Manitoba, just on the border of Manitoba and Saskatchewan. The complex consist of four different mines, Callinan, Trout Lake, Konuto Lake and 777 mine including copper smelter and zinc hydrometallurgical plants producing metallic copper and zinc. Gold is a very important byproduct providing about 30% of the mine revenue. In early 1990s when the company was facing declining ore reserves and was starting to gradually close down the operations, the 777’th exploration drill hole discovered a new orebody, which initiated a $0.5 billion expansion program. The main projects were a new 777 hoist with underground development and a new zinc cell house. About $30 million was used to increase the mill throughput to the present 2.27 million tonnes or 2.5 million short tons per year. The expansion programs were finished by the end of the year 2003.Historically the gold recovery at HBMS has varied between 60 and 70%. The recovery has been somewhat lower during the summer months. Gold recovery improvement projects were conducted throughout the 1990s, but were not focused on the seasonal gold recovery variance. However, since 2000, when the water system modifications to meet the increased water demand were completed, the gold recovery during the summer has been much lower than in the past. In August 2003 gold recovery fell below 30%, which initiated a metallurgical research program to discover the reasons behind the low gold recovery in the summer."

Jan 1, 2005

The mineralogy and textures of a portion oC ore from the 25 level northern orebody, North mine, Broken Hill are described. This ore, in places, contains up to 70 per cent cubanite (CuFe2S3 ) which has crystallized directly with other sulphides and associated gangue minerals.Several breakdown reactions involving the cubanite are also considered.

Jan 1, 1969

By John J. Dran

[This paper summarizes the more popular mineral valuation techniques being used by the dining industry, and attempts to demonstrate how they are Interrelated and where some are in error when compared to modern methods of financial analysis. The academic community and the larger and more sophisticated raining companies, have largely rejected the older valuation methods such as the, Hoskold and Morkill concepts and replaced them with "discounted cash flow" (DCF) techniques. The relevancy and usefulness of the older methods have been questioned primarily on the basis of their underlying assumptions that mining companies do not have numerous alternative investments and must use sinking funds deposited at market or ?safe" rates of return to recover the purchase price of mineral reserves. The ability to reinvest In numerous (often non-mining} alternative business ventures, the changing purchasing power of the dollar, corporate income and other taxes, depreciation and depletion allowances, and modern financial management practice, all mitigate against the continued use of the older valuation methods, especially for the larger mining companies. Snail mining companies and Individuals still use these methods largely out of habit and familiarity. This paper attempts to point out the weaknesses and strengths of the valuation methods In current use in the hope that these techniques may be better understood and [rcrn] appropriately applied.]

Jan 1, 1973

By L. G. Austin

A simple finite difference algorithm has been used to compute a simulation model for unsteady-state ball milling. The model considers the mill to act as if it were three fully-mixed reactors in series (one-large/two- small), it incorporates the-Rogers-Austin mass transfer law relating mill hold-up to flow rate through the mill, and it allows for the reduction in breakage rates caused by over-filling of the mill at high flow rates. The computation is rapid and requires only small storage space. Examples are given of use of the model.

Jan 1, 1984

By Yoginder P. Chugh

The control of soils has been found to be one of the highest priority factors influencing the return of reclaimed prime farmland to production. The compaction of soil is primarily a result of the movement of high ground pressure wheel scrapers and dozers over soils during the soils replacement and regarding process and occurs to a maximum depth of about 1.2 m (4 ft) The deleterious effects of soil compaction can be controlled by deep soil loosening to a depth of 1.2 m (4 ft). Most farm tillage equipment can only loosen soils to a maximum depth of 0.33 m (1.0 ft). Deeper soil loosening equipment, in the 0.33-1.2 m (1.0-4.0 ft) range is currently avail- able in the U.S. and abroad which may have potential for application on reclaimed prime farmland. This study has appraised such available equipment and the results indicate that a few manufacturers in the U.S., West Germany, and England market static and vibratory soil loosening equipment which have potential for deeper soil loosening of reclaimed prime farmland.

Jan 1, 1985

It is possible to closely simulate the product sizing distribution from an industrial rod mill by the correct choice of parameters in laboratory batch milling. For laboratory batch rod milling, changes in per cent critical speed and per cent rod volume cause the product sizing distributions to move in parallel at a constant slope. For ball milling, larger sized media in batch milling grind more coarse material giving a closer simulation to the industrial rod mill product. Changing the feed size distribution by removal of coarse particles or addition of fine particles changes the product sizing distribution, but would not normally be practiced in laboratory testing. These data should be applicable to many instances where laboratory simulation of an industrial product size distribution is required.

Jan 1, 1998

By Wilbur A. Haley

One of the purposes of the investigation was to determine and report on how coal augers are used to mine bituminous coal that is near outcrops and may not be mined otherwise by any other present-day method. Such increased overall recovery results in the conservation of coal resources. Other reasons for the investigation were to determine and report the production rates of such operations and the types of equipment used. Auger mining of coal as a means of coal recovery supplementary to strip mining has been introduced and developed since World War II. This type of coal mining has grown rapidly, and bituminous-coal production by auger mining has been estimated at 6,000,000 tons in 1955. Mines were selected for study to give reasonably representative information on auger mining in different types and thicknesses of coal beds in different geographic areas and on different types and sizes of equipment operated under various conditions. This report is based on field studies of 8 auger-miming operations in Pennsylvania, West Virginia, and Ohio and in 5 different coal beds. Operations were observed in active strip mines, in strip mines reopened for auger mining, and along benched coal-bed outcrops. At the time of this study, no augers were reported to be operating in underground mines. The diameters of holes drilled ranged from 26 to 52 inches, and the lengths of holes ranged up to 208 feet. At some operations auger holes often penetrated abandoned underground mine workings.

Jan 1, 1957

By Kent E. McElhattan

The selection of efficiency machinery for an underground mining system is a highly skilled art. Modern techniques of mining, based on sound engineering principles, increase the probability of success. These techniques alone, however, will not guarantee maximum productivity. Human judgement, based on experience with mineral seams and geological conditions, remains the single most important input of the decision-making techniques in selecting high production equipment. The myth of a "Scientific Selection Factor" is absurd. The belief that such a factor exists can waste the time and resources of mining management. Until now, mineral extraction from the original desposit has defied all attempts to systematize the knowledge. Computerized programs are excellent for solving underground transportation and ventilation problems. The variables can be defined by mathe¬matical techniques in this type of problem. By the same, token, "Trial and Error" is not an acceptable method in today's technology-minded world. The candid truth of the matter is that practically nothing is known of the phenomenon of extracting a mineral from the solid. Likewise, local conditions make it nearly impossible to predict accurately the reaction of surrounding strata after the mineral has been removed. The unpredictability of the environment resulting from extracting minerals underground will remain a challenge to mining engineers for many years. In fact, we are barely beyond the trial and error stage at this point in time.

Jan 1, 1969

By L. A. Pursell

The case history, which is the subject of this paper, involves the expansion of a lime plant in Texas. The major new production unit is to be a 545.45 tonne per day coal fired rotary kiln. Auxiliary facilities include additional quarry haulage equipment, modifications to the crushing plant, a hydrator, storage facilities, and a loading system. The existing plant consists of a high calcium limestone quarry, two coal fired rotary and 10 gas fired shaft kilns, nth crushing, screening, stockpiling, hydrating, storing and loading facilities. The proposed modification would increase quarry volume from 545,000 tonnes per year to 805,000 tonnes per year. The plan also includes the addition of dolomitic lime to the product line which requires the purchase and handling of 186,000 tonnes of screened, washed dolomite yearly. A decision to undertake such an expansion was previously a simple and straight forward process. Such a decision now faces a major and complicating obstacle which can delay the project, raise the expected cost, or prevent it entirely.

Jan 1, 1980

By J. A. Herbst

Until the past decade, the energy-size reduction relationship was the dominant form of mathematical model used in the description of tumbling mill grinding processes (1,2). Typically, in these models some single measure of product fineness (e.g. the 80% passing size, the size modules, the specific surface area) is chosen as the dependent variable and the energy input per unit mass of material being ground acts as the independent variable. In some instances these models have been useful for the correlation of experimental data, but, invariably, they have been found to be inadequate for meaningful process simulation. During the last decade considerable advances have been made in the development of phenomenological grinding models derived from population balance considerations (3-7). In these models a set of dependent variables is employed, corresponding to the set of mass fractions which constitute the product size distribution, and energy input is abandoned in favor of "time in the mill" as the independent variable. The phenomenological models have been shown to provide accurate simulations of the entire size distribution of ground products produced in batch (8-10), locked cycle(11)and steady and unsteady state, open and closed circuit continuous grinding(12,13).

Jan 1, 1973

By P. B. Ritter

The feasibility of redesigning and adapting the all- secondary-explosive, low-voltage, electric safe detonator for use as a delay detonator has been demonstrated. Experimental results have shown that operational characteristics can be adjusted to allow direct competition of safe delay detonators with commercial delay blasting caps. Five different experimental delay period models have been developed and tested. Test results are given and results are evaluated. Subsequent simplified and improved designs for practical and economical mass production are described. Design concepts to meet Bureau of Mines voltage withstand, incendivity, and spark sensitivity requirements are included. Recommendations are made for experimental procedures to ready the safe delay detonator for mass production and extensive use in mining and other industries.

Jan 1, 1977

By R. K. Rath

Adsorption, electrokinetic and flotation studies were carried out. The maximum density of adsorption of guar gum was obtained at pH 7.5 on sphalerite and at pH 11.5 on galena. Flotation recovery of sphalerite was reduced only marginally by the presence of guar gum in the pH range investigated, whereas recovery of galena was reduced to about 20% at pH 12. Differential flotation tests on binary mixtures indicated that 10 ppm of guar gum can depress galena and separate it from sphalerite at around pH 12.1. Co-precipitation tests confirmed the formation of a metal ion-guar gum complex in the bulk solution, suggesting that this is the precursor to adsorption. The adsorption isotherms for both minerals followed Langmuirian behaviour, but adsorption of guar gum on galena was significantly higher than on sphalerite and dissolution tests indicated that the concentration of Pb ions released from galena was an order of magnitude greater than that of Zn ions dissolved from sphalerite

Jun 21, 1905