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[Separates of all the Technical Publications published in 1930 are available at Institute headquarters. All the papers are on file in public, university and technical libraries, and when so indicated in the list, may be found in the Transactions.] Metal Mining Technical Publi- Title Author Meeting Tran~. cation Volume No. and Page* 264-A.29 Quicksilver Industry in 1929. Im- L. H. Duschak San Francisco T30, 283, 319 provements in the Metallurgy of C. G. Maier and Oct., 1929 Quicksilver and Symposium on The others Present Status of the Quicksilver Industry (With Discussion) (37) 271-A.30 Observation on Ground Movement R. E. Palmer New York T30. 168 and Subsidences at Rio. Tinto Feb.. 1930 Mines, Spain (20) 274-A.31 Development and Installation of the C.L.Berrien New York T30, 139, 166 Hawkesworth Detachable Bit (29) Feb., 1930 276-A.32 Protective Measures . Against Gas 0. A. Glaeser New York T30, 129. 167 Hazards at United Verde Mine (11) Feb., 1930 287-A.33 Use and Cost of Compressed Air (12) R. S. Lewis New York Feb., 1930 314-A.34 Miami Copper Company Method of F. W. Maclcnnan New York T30, 39, YO Mining Low-grade Orebody (44) Feb., 1930 319-A.35 How Human Beings Respond to W. J. McConnell New York Changing Atmospheric Conditions Feb., 1930 (9) 324-A.36 Vertical and Incline Shaft Sinking at Arthur B. Foote San Francisco T30, 87 North Star Mine (21) Oct., 1929 32bA.37 Governmental Control of the Pro- W. E. Colby New York duction and Sale of Mineral Re- Feb.. 1930 sources (With Discussion) (16) 333-A.38 Some Recent Developments in Open- E. E. Hunner New York T30, 106, 117 pit Mining on the Mesabi Range Feb.. 1930 (25) 334-A.39 Mining Methods and Costs at Presi- V. D. Howbert El Paso dio Mine of American Metal Co. of R. Bosustow Oct., 1930 Texas (15) 339-A.40 Operation of Pressure Fans in Series W. 8. Weeks New York (8) V. 6. Grishkevich Feb., 1931 364-A.41 Top Slicing with Filling of Slices as H. Willey El Paso Used at the Charcas Unit of Com- Oct., 1930 pania Minera Asarco, S. A. (17) 373-A.42 Drill Sampling and Interpretation of H. T. Matson New York Sampling Results in the Copper G. A. Wallis Feb.. 1931 Fields of Northern Rhodesia (17)

Jan 1, 1930

By S. Gregory

Many gold treatment plants have sulphides and coarse particle gangue components that contain gold in their tailings (cyanide leach residues and flotation tailings). Coarse and very fine liberated gold may also be present in some tailings streams. Two issues usually govern the economics of recovering additional gold from gold occluded in sulphide and gangue particles. The liberation characteristics of the gold in the sulphide and coarse gangue (gold liberation versus grind size) is the first and most important of the two. If fine grinding can liberate a reasonable amount of the gold, then the next important issue is how effectively these two constituents can be concentrated. Flotation will achieve high sulphide recovery but is usually costly. Selectivity is also an issue because of entrainment. Continuous centrifugal gravity concentration offers a viable solution to the recovery of both sulphides and coarse gangue however selectivity is also likely to be a problem. The possibility of improving concentrate grade by using a Falcon B separator was investigated. This paper describes a program of test work carried out on a gold leach tailings using a laboratory Falcon '13' separator. The program involved three different bowl angles: 10, 12 and 14 degrees and two bowl speeds (giving 250 and 350G accelerations). A number of tests were also performed on samples that had free gold present to ascertain the size-by-size recovery of gold in these samples.

Jan 1, 2005

By R. A. Elliott

This paper describes the development of an infra-red ore dryer at the Mines Branch. The dryer utilizes an infra-red heat source mounted over a stainless-steel belt to dry ores and concentrates of varying particle size and moisture content. It has been found to ?be successful in meeting Mines Branch requirements of minimum dusting loss and ease of cleaning to decrease ,contamination between various batches of dried products. It is described with the hope that it might .prove useful in other laboratories where there are similar problems. It might have an application in the drying of concentrates containing high values where dust losses would result in a considerable loss of revenue; also for small batches of material or an intermittent drying problem.

Jan 1, 1958

By Vishal Shukla, Ashok Kumar, C. Raghu Kumar, D. P. Chakraborty

"Traditional iron ore beneficiation plants using washing, screening, and jigging processes require large volumes of process water. Generated slime during processing, its disposal and impact on water usage, is an important concern for any mining company. The challenge today, during extracting iron ore, is how to dispose of slime such that it is contained and stable, while maximizing water reuse and minimizing surface footprint. The slime disposal methods are sensitive to the environment. Two major technologies presently being practiced for disposing the slime across the world are pressure filtration and paste thickening. In this perspective an attempt has been made to do comparative study between paste thickening and pressure filtration for selection of suitable slime disposable technology for Noamundi processing plant, Tata Steel. From the studies, it is evident that by adopting pressure filtration technique it was possible to produce filtered slime, whereas filtration, rate reduces significantly with the presence of finer fraction. Further , the greatest advantage of paste thickening is volume reduction of slime. By considering the technology limitation and comparative lab test work result, it is opined that combination of filtration and paste thickening technology is favorable in terms of capital and operating cost and slime >45 µ should be filtered and < 45 µ slime should go for paste thickening. The study and the results are presented in details.INTRODUCTION The beneficiation processes always end up with an enriched portion of desired mineral - “Concentrate” and an enriched portion of undesired minerals - “Tailing”. The proportion and nature of tailings generated in every beneficiation process differ from ore to ore and reflects the characteristics of the feed processed. The traditional method of tailing disposal in many mineral processing industries has been to dispose them as thickener underflow at around 35% solids. Though, it had certain advantages of transportation with ease but contribute to a greater volume occupied in the tailing dam. This sustained practice and failure to utilize the space available in the tailing dam judiciously endangered the mining industry. The tailing management has become the toughest task to the mining industry and most of the mines are running out of space for disposing the fresh tailings as many of these are at the verge of getting filled."

Jan 1, 2018

By J. Rostami, J. Brune, S. Slouka

Respirable dust poses long-term health issues to personnel working in mining and civil projects where mechanized mining and tunneling machines are used for hard rock excavation. Machines, such as longwall shearers, roadheaders, or continuous miners, use consumable picks to cut rock. However, rock dust is generated in the work environment each time the pick tip contacts the rock surface. Additionally, the generated dust can be reintroduced into the air further down in production, such as at transfer points. The purpose of this study is to characterize respirable dust particles generated at the pick tip at two stages of pick wear: new and moderately worn. Understanding dust characteristics in terms of concentration, size distribution, and particle shape throughout the life-cycle of the pick will provide a basis to mine and tunnel operators for changing out picks to mitigate dust generation at the source. This paper discusses the dust characteristics generated from preliminary full-scale cutting tests in a laboratory of an igneous rock block cut with conical picks at new and moderately worn levels of wear. The moderately worn pick generated a size distribution that favored smaller particle sizes compared to the new pick. The moderately worn pick also generated more overall dust, as the concentration level in mg/m3 was over double the amount of the new pick. Lastly, both picks generated particles that had similar aspect ratios and roundness measures.

May 18, 2022

By F. W. Jessen

The production of salt through brining operations and the recovery of potash salts by means of solution with water constitute, in large measure, processes which have been used for many years. And although the salt industry has utilized dissolution techniques for years and years it has been only recently that a better understanding of the mechanism of solution has emerged. Greater adaptability of solution mining processes for recovery of valuable salts from ore bodies found at greater and greater depth has resulted in increased adoption of such methods. Because of the high costs of mining in beds of limited thickness, modifications probable when extracting mixed halites, and restrictions (mostly self-imposed) regarding the extent of cavern development to minimize or eliminate subsidence more information is needed on the solution mechanism. It is the purpose of this paper to present the fundamental aspects of the problem of total solution of salts and to point out the developments of certain solution mining techniques.

Jan 1, 1971

By João F. C. L. Costa, Vanessa C. Koppe, Clayton V. Deutsch, Marcel A. A. Bassani, Pedro H. A. Campos

Blending inevitably occurs to some degree during mining and mineral processing operations. Traditional mine scheduling assumes that all variables average linearly, which allows the calculation of the blended material properties as a linear weighted average. However, metallurgical properties such as metal recovery and Bond Work Index (BWI) are known to average nonlinearly. The nonlinear properties of the blended material can be calculated using nonlinear blending models. A simulated annealing algorithm is developed and applied to a short‐term mining face optimization problem to maximize the total metal recovered considering nonlinear blending. Compared to a standard schedule plan, the optimized one shows relevant improvements in synergistic and antagonistic blending behaviors.

Jun 25, 2023

By Balasubramani M., Ramana P. V., Bairagi K.

This paper is a case study of an earthen Cofferdam, constructed for a deep excavation at Intake location in river Aaundha, Maharashtra, India. A 12 m (39 ft) deep excavation was made to construct the Intake well. The natural soil profile at this location is clay followed by rock. An earthen Cofferdam of maximum height 7 m (23 ft) was made around the periphery of the excavation. Cofferdam embankment and excavation slopes were analyzed using the finite element software PLAXIS-2D. Clayey sand was proposed as filling material during design, but clayey soil was used for the embankment formation during execution. The embankment was built up to 5m height in July 2017 and completely submerged due to monsoon. In February 2018 embankment height raised to 7m (23 ft) by filling the soil over the existing submerged embankment. After a few days of completion of the Cofferdam construction, the inside slope of the embankment had started sliding due to continuous dewatering. The design and execution of deep excavation with earthen Cofferdam for waterfront structures are challenging. The selection of filling material for embankment formation and dewatering techniques plays a vital role in the safety and stability of the embankments. Although all the safety measures were taken during the design, execution challenges are unpredictable.

Nov 1, 2022

By Ltd. (CYMCO) Technology Center Yunnan Metallurgical Group Co., S. Jian, T. Song, J. D. Miller

"As sulfide zinc ore deposits become exhausted, nonsulfide zinc ore deposits become the other resource for zinc production. However, flotation for zinc recovery from nonsulfide deposits is much more difficult than the flotation of sphalerite from sulfide deposits and may become an option for zinc production only if a suitable processing technology is developed. In this regard, research efforts are being devoted to the flotation of nonsulfide zinc oxide minerals. Some common methods which have been used for the flotation of zinc oxide minerals include sulfidization-flotation, activation, and direct flotation with anionic and cationic collectors. The traditional flotation techniques have low efficiency and difficulty in separation and recovery of the zinc minerals from gangue minerals, for example, smithsonite (ZnCO3) from other carbonate gangue minerals and hemimorphite (Zn4Si2O7 (OH)2·H2O) from silicate gangue minerals. In this study potassium lauryl phosphate was evaluated as a collector for the flotation of nonsulfide zinc minerals. Results from contact angle and microflotation experiments at pH 6–9 show that hemimorphite (Zn4Si2O7 (OH)2·H2O) can be floated from quartz and that smithsonite (ZnCO3) can be floated from calcite. However, when dodecylamine, which is a common collector reported in literature for the flotation of nonsulfide zinc minerals, was used as collector, no significant difference between nonsulfide zinc minerals and gangue minerals was observed at any pH value. Initial evaluation suggests that potassium lauryl phosphate is effective at low concentration and may be a promising collector for zinc oxide mineral flotation.INTRODUCTIONBefore the twentieth century, zinc was mainly produced from nonsulfide ores (Large, 2001; Gilg, Boni, & Cook, 2008). With the development of differential flotation processes in the late 19th century major production switched from nonsulfide zinc sources to sulfide zinc ores (Hitzman, Reynolds, Sangster, Allen, & Carman, 2003; Gilg et al., 2008). However, as it is becoming more difficult to find new sulfide ores, recovery from nonsulfide deposits is being considered. Nonsulfide flotation for zinc recovery is much more difficult than the flotation recovery of sphalerite from sulfide deposits and may become an option for zinc production if a suitable technology is developed. Zinc and lead oxide ores are difficult to beneficiate ores due to their similar floatability with gangue, high solubility in water, and the influence of slimes (Fa, Miller, Tao, & Li, 2005). In order to improve the performance of oxidized minerals, lots of effort is being devoted to this field of study (Ejtemaei & Mahdi, 2014)."

Jan 1, 2016

By J. H. Potgieter

This paper describes an investigation to separate and selectively recover silver, gold and platinum from jewellery and laboratory wastes, and produce technical grade material suitable for training purposes. Nitric acid was employed to dissolve and separate silver and other impurities from gold and platinum, which were then dissolved in aqua regia and precipitated selectively. The current work quantified the effects of temperature, stirring rate, leaching agent concentration and solid: liquid ratio on the recovery of the various noble metals and in summary yielded the optimum treatment conditions in each case. It was found that increased leaching reagent concentration (in the case of Ag), temperatures and stirring rates enhanced the recovery kinetics and optimum yield of the three noble metals. An increased liquid to solid ratio also enhanced the recovery of the silver. Recovery yields 97.1%, 99.9%and 99.4% for silver, gold and platinum, respectively, were obtained. The recovered silver, gold and platinum had purities of, respectively, 98.3%, 99.1% and 99.0%. A cost saving of approximately R10 000 per kilogram of scrap refined could be achieved by the in-house treatment of the waste, not to mention the time-saving benefit.

Jan 1, 2004

By Shane McDonald

Australia produces approximately 84 million tonnes of coal per annum from 30 longwalls in New South Wales and Queensland, operating at an average depth of around 300m. Industry trends and expectations place increasing emphasis on the reliability of these operations; for example, the current average face width is 245m and this is expected to increase to approximately 270m by 2010. In recent years, Beltana No.1 Mine has consistently been Australia?s highest longwall producer, with ROM output of 7.2Mt in YEJ2007. Strata management has become a key feature of the operation of Australia?s longwall mines, as a tool to manage variances in the geotechnical environment. This involves an understanding of the impacts of the geotechnical environment on ground behaviour and the implications for mine design, stability and associated support, interaction and horizon selection. The anticipated ground control impacts of these factors form the assumptions that drive the format of plans specifying both triggers and actions, with quantified, monitoring-focussed experience used to refine the plan in a continuous improvement (ie ?plan-do-check-act?) process. This paper summarises the Beltana No.1 Mine strata management methodology and experience to-date.

Jan 1, 2008

By Christian O’Keefe, Alex Van der Spek

"Evaluation of inequalities of the partial sums defining cumulative grade and recovery prove that cumulative grade has an upper bound not dependent on the bank number and cumulative recovery possesses an upper bound proportional to the bank number. This explains one observed characteristic of cumulative grade versus cumulative recovery curves. The existence of such maxima combined with the observed shape of cumulative grade-recovery curves leads us to postulate that these curves can be well described by rectangular hyperbolae. Rectangular hyperbolae are characterized by asymptotes at right angles to each other. It is proven that this property can be used to derive an equation for the cumulative mass flow of the concentrate in terms of the asymptotic values of cumulative grade and cumulative recovery. Fitting hyperbolae to experimental cumulative grade cumulative recovery plant data then results in estimates of the maximum grade, the maximum recovery and a constant which may be taken as a performance coefficient of a row. Normalization of the cumulative grade and the cumulative recovery by their respective maxima gives rise to a graphical representation that can be used to evaluate the operating performance of a row of floatation bank cells. The normalization procedure removes the bias that is introduced by inevitable variation in mass flow of the feed and variation in feed mineralogy or feed particle size distribution. Reduction of the normalized data collapses all cumulative grade-recovery data onto one and the same curve. This removes the differences caused by the variation in row performance. The reduced cumulative grade-recovery data scatters around a common hyperbola. This can be used to gauge the validity of the assumptions made. Normalized cumulative mass pull is determined from the fitted values of the maximum recovery and maximum grade. This may serve as an alternative performance indicator of the floatation row."

Jan 1, 2014

By Joel Ochoa, Miguel A. Montiel

El artículo da cuenta del estudio de un prototipo de placa de desgaste para chutes de transferencia con un nuevo sistema de sujeción. El sistema fue validado mediante el software Ansys. La innovación no solo está constituida por un nuevo diseño sino también por el tipo de proceso productivo para la fabricación del producto mejorado. Se incluyeron ensayos con torquímetro, que presentaron resultados favorables.

Sep 30, 2019

By I. A. Menshikh, S. A. Kotov, N. L. Shinkarkin

All the cycle for conversion of natural and reprocessed uranium of ROSATOM’s “TVEL” Fuel Company is concentrated at the Siberian Chemical Plant. Conversion capacities are located on two sites, i.e. the Radiochemical Plant, which refines feedstock, and the Conversion Plant, which produces uranium oxides and uranium hexafluoride. Liquid radioactive waste arising in the process, after pretreatment, is transferred to a dedicated national operator for underground disposal. Environmental strategies pursued by the Fuel Company involve the introduction of process solutions intended to eliminate liquid radioactive waste and convert it into a solid phase. In the 5–7 years coming, work will be performed to refurbish and upgrade conversion facilities of the Siberian Chemical Plant that will principally change its radioactive waste management scheme. This paper describes the current waste management scheme and sets out planned schemes for management of natural and reprocessed uranium waste, which will be implemented in the refurbishment of conversion facilities of the Siberian Chemical Plant. The new technologies for the management of liquid radioactive waste from uranium conversion described in the present paper have an undeniable advantage over the current technology mainly due to significantly less environmental harm.

Jan 1, 2020

By G. L. Oldright

The data presented in this paper were collected at the smelter of the Consolidated Mining & Smelting Co. of Canada, Ltd., Trail (Tadanac), British Columbia. Many thanks are due to S. G. Blaylock, vice president and general manager, James Buchanan, smelter superintendent, and George Murray, assistant smelter superintendent, for their broad general policy in permitting these data to be taken and published. life are also grateful for the help given by the staff members stated in Paper XI, Report of Investigations Acknowledgment is given several excerpts from experiments by F. E. Lee, head of the research staff. INTRODUCTION The past reputation of zinc as an ingredient in lead blast furnace slags has been bad. Zinc is thought of as a cause of accretions in the shaft of the blast furnace, of hot tops, of the formation of "zinc mush" in the crucible and forehearth, of lead-rich slags, of viscous slags, and. of low capacity in smelting.

Jan 1, 1934

By C. R. Larsen, R. G. Bradburn, A. L. Mular, B. C. Flintoff

"Following the acquisition of grinding- circuit data at Brenda Mines Limited for simulation purposes, various methods to mass balance the circuit were evaluated. Unfortunately, different answers were obtained for different balancing methods. A data-adjustment procedure was therefore developed. The procedure assumes that all ""best-fit"" criterion. A search is conducted with the minimum set of water masses and narrow size ranges of solids around the circuit (called search variables), which permit the computation of any of the measured data points. Search variables are used to calculate measured values for an objective function of the form: [ ] where W, is a weighting factor•, M1 is a measured value and M. is a corresponding . value calculated from search vanables. The best values of M, are those which minimize R. The above method is compared with others that have been utilized for mass balancing circuits."

Jan 1, 1976

By J P. Barkas

World demand for iron ore is far from peaking, and the substantial price decline that has captivated commentators since early 2014 is not the result of widespread misreading by producers of the prospects for global market growth. Recent price movements provide little useful insight into the industry’s longer-term fortunes, apart from reminding mining companies that their businesses need to be robust against the volatility and inescapable cyclicality of mineral commodity markets. It is timely to take a strategic view of the history and longer-term prospects for the iron ore industry and of Australia’s place in it.CITATION:Barkas, J P, 2015. Perspectives on the long-term outlook for iron ore, in Proceedings Iron Ore 2015, pp 3–6 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

By Batric Pesic

A hydrometallurgical process for recovery of Zn and Mn from Bunker Hill Water Treatment Plant sludge was developed. The process is unique because it treats an unusual combination of metals, Zn and Mn. Another non-standard feature of the process is that leaching takes place with ~S04 in the presence of a reductant (SO2). Zn and Mn are recovered simultaneously by electrowinning, Zn on a cathode and Mn on an anode, as MnO2 The major purpose of the process is to detoxify the sludge by removing toxic heavy metals. The recovered metal values could be marketed to offset the overall costs of water and sludge processing. This work is presented in two separate parts, 1. Leaching and Solution Purification, and II. Electrowinning. This paper covers only the first part of the hydrometallurgical process: Leaching and Solution Purification. Initially, the paper briefly discusses the history of the pollution problem at the Bunker Hill Superfund Site, the origin of heavy metals in the mine waters, and the development and use of the water treatment plant. The leaching of sludge was examined as a function of the following parameters: constant vs. initial pH, initial vs. delayed addition of SO2, temperature, and pulp density. It was found that leaching with H2SO4 at pH 2.0 for 15 minutes (at 60°C), followed by S02 addition for another 15 minutes, could completely dissolve Zn and Mn. Solution purifIcation was done in three stages. In the first stage iron was removed by adjustment of pH (to pH 5.0) and oxidation offerrous to ferric iron (with H2O2). In the second stage, Cu and Cd were removed by Zn dust cementation. Finally, Co and Ni were precipitated with the organic reagents 1-2 nitroso-naphthol and nitroso, respectively. Electrowinning of Zn and Mn is discussed separately in Part II of this paper.

Jan 1, 1996

By Kevan Ford, Getulio Junior, Rob Henderson, Wayne Philips, Chris Martin, Rob Thorpe

"Plant 2 at the Kinross Paracatu gold operation in Minas Gerais State, Brazil, was commissioned in late 2008. Designed to be the largest tonnage, lowest grade gold milling operation in the world, Plant 2’s throughput has steadily increased towards its design capacity of 5087 tons per hour.The plant employs a gold-specific flotation scheme to recover the majority of gold, which is present as discrete free gold particles. Additional gold is recovered by jigs in the grinding circuit, and the jig concentrates are reground, upgraded and subjected to intensive cyanide leaching. The flotation cleaner concentrate is reground and leached using a conventional CIL circuit.One year after start-up, with the plant stabilizing operationally, plant circuit surveys were completed by a joint team from Kinross Paracatu and Blue Coast Metallurgy Ltd. The surveys audited the flotation and concentrate treatment circuits to investigate the metallurgical performance.This paper describes the Plant 2 operation and how the surveys were designed and executed. Some of the quality control systems that were employed to provide meaningful data from such a low grade gold ore are described and findings and actions taken to improve metallurgy as a consequence of the surveys are discussed.INTRODUCTIONThe Paracatu mine is located two kilometres north of the city of Paracatu, in the north western portion of the state of Minas Gerais, Brazil, 230 km southeast of the national capital Brasília (shown in Figure 1). The present day Paracatu mining operation, fully owned by Kinross Gold, through its Brazilian subsidiary Rio Paracatu Mineração (RPM) is designed to be the largest tonnage, lowest grade free milling gold operation in the world.The deposit currently hosts some 1.2 billion tonnes of ore at a grade of ~0.41 g/t gold, hence contains a total gold reserve of over 20 million ounces. At a designed plant throughput rate of 61 Mt/year the mine life goes beyond 2030."

Jan 1, 2011

By Marcelo B. Mourao

Interactions of carbonaceous materials with liquid Fe-C melts have been investigated experimentally by determining the volatile yield due to rapid devolatilization (Flash Pyrolysis) and rates of dissolution at temperatures ranging from 1623 to 1873 K. The volatile yield increased with increasing the temperature. The comparison of results of different types of carbonaceous materials was possible by expressing the volatile yield in terms of Enhancement Factor, fE. The rates of dissolution of spectroscopic graphite and an industrial coke seemed to obey the correlation for natural convection under turbulent conditions. The experimental data for the graphite suggested that the rate of dissolution was controlled by the mass-transfer in the liquid boundary layer adjacent to the solid sample. The value for the empirical parameter correlating the dissolution coefficient and the operating variables was found to be 0.19 which was close to that reported in the literature. The comparison of the results obtained for the coke samples with those for the graphite revealed that the impurities and the porosity of the samples can effect the dissolution rates.

Jan 1, 1992