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By S. T. Hall

The authors have previously described a method of enhancing the magnetic susceptibility of pyrite by a mild alkaline oxidative pressure leach (Hall et al, 1985). This finding was successfully applied to enhance the magnetic desulfurization of an Eastern Canadian coal (Hall and Finch, 1984). This paper reports on enhancement of the magnetic susceptibilities of iron-containing sulfides (arsenopyrite, chalcopyrite and marcasite) upon similar treatment. It also explores the application of this pretreatment in the magnetic separation of fine sulfides from oxides and other sulfide minerals. In the latter case galvanic interference effects have a serious impact on the potential use of this method.

Jan 1, 1985

By Lukas Erkurt, Wolfgang Holzer, Stefan Auderer, Richard Gradnik

SEM/ NATM Tunnelling contracts in Austria and Germany are usually tendered applying unit price contracts. The purpose of this paper is to show how contractual problems are resolved in these countries. The focus will be put on describing the theoretical approach, several case studies will show how contractual problems were dealt with on site.

Jun 23, 2024

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

I found the article most interesting and timely. I consider it imperative that mine case studies like this one are undertaken and published so that the practicing mining engineer and geologist can obtain a better idea of what is the preferred estimation and grade control technique to use, given the site-specific geology, grade distribution, etc. Mining companies, with operating gold mines, need to be encouraged to either initiate such a study in-house or support a research project by financing a graduate student at a local university. The latter method is preferable since it not only provides research support but also an excellent training opportunity for young mining engineers and geologists. Ideally, what is needed is exploration drill-hole, blasthole, and mine production data that can then be used to assess the efficiency of different estimation and grade control methods by comparing the results obtained using the drill/blasthole data with the actual production data. It is hoped that more mining companies will realize that the rewards from this kind of research greatly outweigh the sometimes perceived problem with releasing data to a university. I noticed that the authors used the blasthole data to represent the actual values, the validity of which relies heavily on the two key assumptions with regards to blasthole assays and in-pit ore-waste selection procedures. Could the authors comment in more detail on these assumptions and what criteria they used to check the validity of the assumptions? Since the article was published, have the authors had the opportunity to compare their estimates with the actual production data? It would be interesting to compare other estimation methods with OK and IK. For example, how well did the polygonal method used at the mine compare with the results given in the paper? I would like to mention that B.L. Kwa completed a similar study (Kwa and Mousset-Jones, 1986) using exploration drill-hole, blasthole, and production data from the Alligator Ridge mine in Nevada. The study compared a variety of estimation methods, including IK, and the final results are shown in Fig. 1. The IK method exhibits the smallest difference between estimated and production results and is closely followed by the indicator moving window (Ind MW) method. Reply by Y.C. Kim We do concur with Professor Mousset-Jones with respect to the merit of such case studies performed at a local university. In answer to his request for comment on the validity of the two key assumptions made in our paper, we can only say that these are reasonable assumptions that are very difficult to validate. The assumption of blasthole assays being accurate can be partially validated by a careful analysis of the nugget component of blasthole variograms. The assumption that the in-pit ore-waste selection procedure in use was an effective one can only be validated through comparison between the mine estimates of ounces of gold versus the actual recovered ounces of gold at the mill. Neither one of these two assumptions has been validated to date due to several reasons. For example, the mine went through a change in management since the study. Similarly, no compelling reasons have been presented to perform detailed variogram analysis of the blasthole data. We do, however, believe that the obtained results (i.e., indicator kriging performs better than OK in a highly skewed deposit) will remain the same in relative terms, regardless of how valid these two assumptions may be. The most recent com¬parative study by D. S. Lindsey of the Montana College of Mineral Science and Technology also shows a similar ranking between ordinary kriging versus probability kriging, which is an extended version of indicator kriging.

Jan 1, 1987

By C. E. Jordan

The Bureau of Mines investigated a simultaneous elutriation and flotation technique for separating polyethylene terephthalate (PET) from polyvinylchloride (PVC). The Tuscaloosa Research Center, through a cooperative agreement with a recycling company, devised an elutriation-flotation method for beneficiating the mixed heavy plastics portion of municipal solid wastes which is composed of polystyrene (PS), PET, and PVC. A sodium carbonate brine was used to remove the lighter PS from the heavier PET and PVC. Over 98 pct of the PS was recovered in a high purity PS concentrate containing 99.7 pct PS, The remaining PET and PVC mixture was washed to remove the brine salts that would be harmful to the flotation process. The clean PET and PVC mixture was conditioned in a solution of 45 g/t gelatine and floated in a short (l-m) column flotation cell in an upward flow of water. In a single flotation step, over 75 pct of the PET was recovered in a high purity product containing 99.8 pct PET. The bubbles attached to the PET lowered the effective density of the PET-bubble agglomerate and allowed successful separation of these agglomerates from the PVC plastic pieces. This method separated two-thirds of the mixed plastics fraction into high purity PS and PET products suitable for recycling and left only one-third of the plastics in the mixed plastics product.

Jan 1, 1992

By J. G. Groppo

Eastern U.S. oil shale contains approximately 10 to 12 gallons of oil per ton and 7 to 10 percent pyritic sulfur. Physical beneficiation of these low grade shales using froth flotation provides a high grade material which will make extraction of oil more economically attractive. The Alabama oil shale used in the present study contained 16.46 percent organic carbon and 14.5 gallons of oil per ton as determined by Fischer assay. Liberation size of the kerogen was estimated at 10 /-Lm. 'Ken-Flote' column flotation technology provided a high grade concentrate containing as much as 60 percent organic carbon, however, carbon recovery was about 60 percent. Carbon recovery was improved to 75-85 percent while maintaining concentrate grades in excess of 40 percent carbon by providing longer retention time in the column which facilitated effective bubble-particle collision and attachment.

Jan 1, 1993

By Jameel Menashi

A process has been developed for recovering the refractory metal values contained in mixed secondary source materials such as contaminated grinding wastes, oxides and sludges. The wastes are calcined with sodium carbonate in air to convert the refractory metals to water soluble Na2CrO4, Na3V04, Na2Mo04 and Na2W04. The base metals, Fe, Ni, Co and Cu, are converted to water insoluble oxides. An aqueous leach of the calcine effectively separates the refractory metals from the insoluble base metals. Co, Ni and Cu can be recovered from the oxidic residue by conventional processing methods. The leach solution is purified and then treated to recover the refractory metals as Cr203, NH4VO3, (NH4)6Mo7024?4H2O and H2WO4. The purity of the products is sufficient for their use in many metallurgical applications. The Na+ values in the leach solution are recovered as Na2CO3(or NaHC03) and are recycled to the calcination step. It is shown that the process can accommodate a wide range of feed compositions and generates a minimum of aqueous effluents.

Jan 1, 1984

By C. Bazin

The fundamental error associated to the sampling of heterogeneous ore particles is unavoidable and can only be minimized by an adequate selection of the sample mass and/or size reduction of the sample before division into sub samples for testing or assaying. The preparation of ore batches to carry out the mineral processing evaluation of an ore is a sampling process. The fundamental error associated with the division of the ore into batches is responsible for the differences in the composition of the batches. This paper attempts by various means to assess the reproducibility of the gold content of test batches prepared from a gold ore sample. The results of the investigation shows that for fine gold grain sizes the variability in the composition of the batches is less than that predicted by Gy?s formula and that the gold grain size is a strategic information in the assessment of the gold content variability of test batches.

Feb 23, 2014

By W. K. Dressel

Over 2 million tons of steel furnace dusts are produced each year in the United States. Recycling of these dusts rather than disposal would conserve mineral resources, reduce land requirements for disposal, and decrease pollution. In addition to the iron content, the dusts contain appreciable amounts of zinc and lead. Removal of lead and zinc with simultaneous reduction of Iron oxide has been achieved by inert atmosphere roasting of carbon-containing pellets prepared from steel furnace wastes. High- and low-carbon wastes were blended to obtain mixtures with sufficient carbon to serve as a reductant during subsequent roasting. The mixtures were pelletized with the addition of bentonite and water, and the dried pellets were roasted in a stream of nitrogen at temperatures ranging from 900" to 1,150° C. Zinc and lead contents were reduced to the 0.1 percent level or less by roasting mixtures containing as little as 7 percent carbon for 1 hour at 1,100° C. ton metallization increases with available carbon content; 50 percent metallization is achieved with 7 percent carbon, 85 percent metallization with 12.5 percent carbon, and over 95 percent metallization when the pelletized mixtures contained in excess of 20 percent carbon.

Jan 1, 1973

By Roger C. Wilmoth

The most common method of treating mine drainage to prevent water pollution consists of neutralization and aeration. This process removes the acidity, iron, aluminum, and some other heavy metals, but does not render a water suitable for domestic or industrial use because of the high dissolved solids, hardness, and sulfate in the treated water. The Environmental Protection Agency (EPA), has been sponsoring and conducting research since 1966 on the use of reverse osmosis (R.O.) for the treatment of mine drainage. This process has been found to be highly effective in the removal of nearly all the dissolved solids. In addition, since R.O. is a concentrating process, potentials for recovery of valuable heavy metals are possible. This paper is a review of the state-of-the-art on the treatment of mine drainage by reverse osmosis.

Jan 1, 1972

By S. Tuylu, D. Adiguzel, A. Bascetin, O. Ozdemir, U. Cinar

"In recent years, use of mineral processing tailings as a paste backfill has been significantly increased due to the environmental problems occurred during the disposal of the tailings in mine area. The paste material used to fill the stopes excavated is generally consisted of tailings (solid ratio: 70–85%) and cement (3-7%w/w). Meanwhile, the strength of the paste material is completely related to ratios of these materials. Cemented Paste Backfill (CPB) system is used as support depending on the roof pressure and also as an underground disposal method. Therefore, temporary strength value of the paste material gains a great importance. For example, the paste backfill material must compensate the minimum 4 MPa strength value in order to procure the permanent fortification, and the mixture cost must be minimal as well. For this reason, in this study, the Pb-Zn tailings were used to prepare for the paste backfill material using cement in order to perform the uniaxial compressive strength (UCS) tests. Additionally, a statistical experimental design and linear programming techniques was used to find the optimum mixture in terms of strength and cost. The pulp solid ratios (PSR) of the tailings and cement (by the solid) were varied as 75-85% and 3-8%, respectively. The results of this study; the optimum mixture is ~79.5%PSR and 6.8% cement ratio (by the solid) for 4 MPa UCS (28 day) and 16-17 cm slump value determined by this method with statistically reviewed experimental data. INTRODUCTION Fine-grained tailing waste materials resulting from mining activities are an important environmental problem. In terms of environmental protection, the use of tailings with paste backfill is a very attractive approach and a desirable solution to tailings management. Paste backfill, which is used to fill underground openings, is made by primarly mine tailings mixed with hydraulic binders and water. Its main benefits are not only disposal of mine tailings in underground, but also providing a certain strength to ensure a safe working environment. Therefore, it is necessary to use hydraulic binders due to durability of the system (Benzaazoua et al., 2004). Paste backfill, formed from the mix of dewatered plant tailings (75-85 wt. % solids), water and binder (3-8 wt. %) (Brackebusch, 1994; Landriault et al., 2005).Water is used to determine the chemical reactions between tailing and binder materials for strength as well as increasing pumpability. Portland cement and ash can be used as binder materials (Douglas and Malhotra, 1987). The factors which affect strength and stability of paste backfill separate into two groups called interior and exterior factors. Interior and exterior factors determine the strength, stability and deformation features of backfill in short and long terms and also affect the transportation of backfill to underground (Benzaazoua et al., 2004; Erçikdi et al., 2006; Yilmaz et al., 2010; Nasir and Fall, 2010). Material properties of tailings as well as site operations to fill the underground openings are the most important steps for the design purposes. Grain size distribution, chemical, and physical properties of tailings, and mechanical properties of paste fill material are used to understand the suitability of tailings as paste fill material (Ozturk et al., 2014). In addition, Fig. 1. is also proposed to apply a typical design for the operational stages of paste fill applications on the site."

Jan 1, 2016

By S. M. Khandrika

Most of the coal presently used by the utility industry is cleaned at preparation plants employing wet processes. While water is the mainstay of coal washing, it is one of the least desirable components in the final product causing a reduction of the heating value as well as handling difficulties. Currently available dewatering techniques are inefficient for dewatering very fine coal (minus 325 mesh). The US coal industry has expressed serious concern about utilizing advanced cleaning processes such as column flotation because an efficient and economical dewatering process is not available.

Jan 1, 1992

By H. W. Newcomen, D. R. Brox

Assessments of the stability performance of open pit highwalls have traditionally been based on consideration of surface movement rates. Although generally accepted in industry, this approach presents challenges for establishing appropriate threshold values to allow for safe mining operations and does not address the issue of total strain. Surface movement monitoring data from a limited number of open pit mining operations in North and South America has been evaluated in terms of highwall strain, rock mass quality and inferred mechanism of instability. Threshold strain levels for key stages of highwall stability proposed from previous work have been reviewed and are generally endorsed based on evaluation of additional data. An empirical highwall strain criteria for the assessment of highwall stability performance is postulated based on the observed trends in highwall strain data from the case histories presented.

Jan 1, 2004

By William Gleason

"It has been more than 10 years since Montana Resources’ copper/molybdenum mine has had a lost-time injury at its site in Butte, MT. With more than 350 employees and an average production of about 15.5 Mt/a (17 million stpy), which includes 29.5 kt (65 million lbs) of copper and 4 kt (9 million lbs) of molybdenum, the mine’s journey to a decade of exemplary safety is an impressive feat that has required hard work, commitment and, most of all, leadership from every level of the operation.It has not always been an easy journey, nor has it been one without scrapes and bumps along the way, but it is one that has produced stellar results. Mike McGivern, vice president of human resources, and Pete Steilman, safety manager at the mine, spoke with Mining Engineering about how the operation transformed from a site that was accustomed to collecting citations for safety violations to one that is now collecting awards for its safety record.Simply put, Montana Resources has made safety a value and the people its top priority.The operations faced many predictable challenges in 2003, when the mine was being brought back online after three years of sitting in care-and-maintenance status. It had been shuttered in 2000 because of low commodity prices and high operating costs, and, in the three years that the mine was idled, the equipment became dilapidated, and some of it was no longer in compliance. Less than half of the workforce returned to the site for the restart, and the team was challenged with getting the gears turning again to produce copper and molybdenum. But the most daunting challenge of all of was keeping the workforce safe during the transition.In the first few years of operation the mine racked up citations from the U.S. Mine Safety and Health Administration (MSHA) at an alarming pace, so much so that in May 2006, Montana Resources received a letter from the MSHA District Office informing the company that the operation was being placed on the Potential Pattern of Violations list due to its excessive citation history and high rate of significant and substantial citations. It was clear that something significant had to be done to change the course of action at the mine."

Oct 1, 2019

By M. Z. Rashad

The purpose of this paper was to compare the covariance and multiquadric functions in same mining and geological applications. The basic principals of the two techniques are covered briefly. Similarities and dissimilarities of the covariance and multiquadric functions are discussed geometrically, mathematically and practically. A 14 by 16 kilometer square area in Baharia Oasis, Western Desert of Egypt, was selected for a detailed comparative test. The gravity anomalies of forty-eight test stations were computed by using the data of sixty-three control stations. The results indicated that the covariance function is quite better than the multiquadric one. On the other hand, the multiquadric function more simple in computation than the other one, which needs to precalculations such as semivariogram and covariogram. Moreover, multiquadric equations are based on geometrical and physical considerations rather than covariance functions.

Jan 1, 1990

By Steve Kral

The concept of sustainable development is not new to the mining industry. Many people involved with mining have argued that the industry throughout the world has always supported the communities where they operated. They say that the mining industry was employing sustainable development practices before the phrase became fashionable. The mines employed residents and supported the local economy through tax payments and the purchase of goods and services from local businesses. In many cases, mining companies have built medical facilities and contributed money, equipment and services to local schools. In modern times, mining companies have donated reclaimed land to the local communities for use as parks and recreational facilities.

Jan 1, 2004

By Lancelot E. Kun

The high-rate floc blanket clarification process developed by the Anglo American Corporation of South Africa is capable of continuously clarifying gold and uranium pregnant solutions containing 200-2500 mg/L suspended solids to below 25 mg/L solids at an average rising velocity of 8 m/h at the low process cost of R0,001-0,002 per ton of liquid treated. Overall process cost savings are mainly due to increased polishing filtration cycle times.

Jan 1, 1984

By James R. Dunn

The mineral industry" has made and will continue to make important contributions to the growth of our economy. However, the industry is presently under continuous and increasingly severe attacks by organizations and individuals interested in the preservation of our environment. Yet neither the public interest value of preservation nor the public interest value of our mineral industry is known. In the absence of such values, judgments about best land use where minerals are involved are nearly always made on emotional grounds. An example of the sort of problem that has been created is in the New York Metropolitan Area. Because of institutionally-created shortages stemming largely from zoning decisions and building over deposits, the effective cost of fine aggregate delivered at dockside at Manhattan rose from $2. 65/cubic yard in July, 1969, to $4.10/cubic yard in May, 1970, this in spite of approximately ;l0,000 years of reserves on Long Island. The cost of fine aggregate for New York City alone has "Increased over $S million annually within one year, with the cost increase for the Tri-State Metropolitan Area (coastal Connecticut, New Jersey, and New York)" exceeding $10 million annually. The ab0ve situation has created a cost increase which is clearly the result of non-conservation of a mineral resource. That is, in readily measurable values, it Is costing the public about $10 million per year because good conservation practices were ignored-a public-interest loss. Admittedly, this is not the whole story because the public is simultaneously obtaining other values from the land surface that would otherwise have had to be dedicated, temporarily to be sure, to mineral recovery-a public-interest gain. However, the point is that we do not know where the balance lies-the net public interest value; it may lie one way for deposits zoned out of existence by a growing suburb, another for areas designed as permanent greenbelts, and still another for areas under new highways

Jan 1, 1971

By W. H. A. Timm

Lake Dufault Mines, Limited, is located ten miles north of Noranda, Quebec. Following discovery of a copper-zinc orebody late in 1961, metallurgical investigations of diamond drill core were conducted by Lakefield Research Limited during 1962 and 1963. These were followed by pilot-plant tests in May and June of 1964, utilizing underground development ore. Milling operations commenced on August 16, 1964, and reached rated capacity of 1,300 tons per day by October 1, 1964 (Figure 1). Since that time, tonnage was gradually increased to a maximum of 1,400 tons per day and since January 1, 1968, gradually decreased to the present rate of 1,100 tons per day due to an increase in the hardness of the ore. The ore milled from start-up until January, 1968, was mainly massive sulphides, consisting of a complex mixture of iron, copper, zinc and minor amounts of lead sulphides with important associated values in gold, silver and cadmium (aver- age specific gravity 3.8). From January, 1968, until the present, the ore being processed is described geologically as a boxwork of sulphide stringers of chalcopyrite, pyrite and pyrrhotite and some local dissemination in siliceous rhyolite (average specific gravity 2.9) supplemented by the massive copper-zinc ore. The limited size of the orebody and the mining method employed (open stoping by blasthole and slash and bench methods) would have made grade control from the mine difficult and expensive. It was decided to try to cope with grade variations in the mill. The resulting rapid and wide swings in mill heads, hardness and specific gravity have presented many and varied challenges for the mill personnel.

Jan 1, 1970

By T. R. Ellis

"In August 2013, the Appraisal Practices Board of the U.S. Appraisal Foundation issued its “APB Valuation Advisory #4 - Identifying Comparable Properties,” which contains analysis and examples that may suggest a broader meaning of the term “comparable properties” than is current convention in real property value appraisal in the United States and internationally. The advisory is directed at the appraisal of real property of all kinds, including interests in mineral properties, as indicated by the inclusion of two mineral property appraisal papers in the appendix that lists suggestions for further reading.This paper firstly reviews as background the historical meaning of a comparable property, with its perceived constraints, against real-world examples of transacted mineral properties that are adjusted to subject properties to provide market value indications. The advisory’s findings are then examined and evaluated as to whether they may influence and benefit the work of mineral property appraisers, particularly in situations where their sales comparison analyses may be challenged. The conclusion is that the combination of constraints that the advisory’s findings specify for use of the term “comparable” will for mineral properties be met infrequently or rarely. As a result, mineral property appraisers are advised to avoid using the terms comparable or “comp” when referring to transactions of mineral properties that they analyze to derive data for appraising the value of their subject mineral properties. IntroductionA mineral property appraiser rarely finds a sale of a mineral property that meets the literal meaning of the term “comparable sale,” by being directly comparable to the subject property of the appraisal in all factors of significance in the determination of the value of the property interest being appraised. Real estate appraisers have long used a more liberal interpretation of the meaning of comparable sale and its commonly used abbreviation, “comp.” Real estate appraisers compare sales to their subject property on the basis of price per hectare or acre of land and price per square meter or square foot of building floor area. Adjustments totaling to within the range of +30 to -30 percent from the sale property to the subject, before factoring in the property area, have generally been viewed as appropriate for the sale to be considered a comparable sale."

May 1, 2016