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By J. M. Demarthe

MINEMET RECHERCHE (IMETAL Group, formerly Le Nickel-Penarroya-Mokta) has studied on a laboratory scale and patented a new process to recover metal values -copper- from sulphide ores and concentrates. It combines a cupric leaching, solvent extraction and copper sulphate electrowinning according to the main following steps - A direct and selective leaching of copper sulphide ores or concentrates at atmospheric pressure with cupric chloride to form cuprous and ferrous chlorides and elemental sulphur. Pyrite remains in the sulphur residue. - The leach solution is separated into two parts -In the first one, air is injected to precipitate iron as goethite. The oxidation step can be easily realized at atmospheric pressure by maintaining an excess of cuprous and-ferrous chlorides in the liquor. Copper is extracted from the second part as cupric ion by an LIX reagent. Direct injection or air in the mixing chamber maintains good extraction conditions, so that only one mixer settler is required. - Iron, in the form of ferrous ion, is not coextracted. - Copper is stripped from the organic phase, transferred into a sulfuric phase and sent to sulphate electrowinning cells, which are now conventional in copper metallurgy. - Only electricity and air are required. After a feasibility study, the next step will be to develop it on a pilot and commercial scale.

Jan 1, 1976

By CHAS. A. R. LAMELY

In the extreme northeast corner of the State of Washington, on the Canadian border, lies the Metaline mining district. This district is old in history, but young in production. The Metaline district is a zinc-lead area of the replacement type in dolomite and limestone. The ore bodies of the Josephine horizon are in many ways similar to the ore bodies of the famous Tri-State zinc fields. The beds are faulted and folded and have varyng low dips in varying directions, and underlie large areas of the district. Production started in 1927 on a very limited basis. The property is now mining and milling 700 tons per day. The mine is opened by adit tunnels and a vertical shaft. As the ore horizons gained depth, it was necessary to sink inclines to follow the ore horizon (see Fig 1). From 1927 to date, approximately 600,000 ft of diamond drill was put down This work indicated that sufficient tonnage existed to justify a redesigning of the whole operation, surface and under- ground. After four years of general study, the following program was planned :

Jan 1, 1949

By Liu Liching, Yan Hongjiu

"In China, the electromagnetic induction initiating system f or blasting engineering has been greatly developed in recent years. With this system, the total number of detonators that may be blasted in each round has been increased from 420 to 940 as a result of the invention of a powerful exploder, and the length of leg wires of a detonator has reached 12 metres. These developments make the new system suitable for blasting engineering of medium scale. Another attractive feature of this system is a set of testing instruments which have been invented that enable a final check of the system to be made before firing as in normal electric blasting. This paper reports on these developments.IntroductionThe initiating system for blasting explosives is one of the key factors to the success of blasting engineering, along with the progress of explosives and blasting technology. Today, we have initiating systems which may be divided into two groups: electric and non-electric. The earliest initiating system was non-electric. Many mines, as well as many projects where rock excavation was involved, were blasting non-electrically before the new nonelectric initiating system came into use in the 1970s. Electric and non-electric initiating systems are the two main initiating systems which are now being widely used, each having its advantages and disadvantages. In electric blasting, the blasting supplies and the blasting circuit can be checked before it is used or fired and, therefore, is regarded to be more reliable. In many large-scale blasting engineering where reliability of the initiation is of great concern, a checkable initiating system is preferable. The electric initiating system is not considered to be as safe as the non-electric initiating system due to its inability to choose proper ignition energy. Therefore, a non-electrical initiating system would be preferable where a hazard of stray currents or static electric energy exists. At this stage of development, one would say it would be advantageous to develop a new initiating system."

Jan 1, 1988

By James J. Scott

"In the two years since their introduction, Friction Rock Stabilizershave replaced40 per cent of the previously used fixtures in the metal and non-metal mines in the United States. The Friction Rock Stabilizer, or ""Split Set "" , is a mechanical rock stabilizing system which uses friction to prevent rock movement . The Split Set is a tapered, rolled, steel tube that is slotted along its length. As it is driven into a drilled hole smaller than its diameter, it compresses, pre-loading the surrounding rock. This paper discusses the uses and acceptance of the friction rock stabilizer in mines in the United States. IntroductionFriction Rock Stabilizers, commonly called Split Set™· Stabilizers, have been under development by the author and Ingersoll-Rand Company since early 1973. Five years of intensive laboratory and field testing led to their introduction to the metal mines in the United States in the fall of 1976. Actual sales started in February of 1977and, to date, over five million units have been marketed to metal mines. The present rate of usage is near 3 V2 million units per year, which is approximately 50070 of all roof fixtures used in the United States metal mining industry,Richard Hoppe, senior editor of Engineering & Mining Journal. , in an article in the February 1979 issue entitled ""Winning the Battle Against Bad Ground"", said: ""In hardrock mining operations in the western U.S., particularly underground uranium mines where exceedingly difficult ground is common, no roof support system has been accepted more quickly than the Split Set friction bolt invented by Dr. James J . Scott and developed by Ingersoll-Rand Co."" This paper describes how Friction Rock Stabilizers function and presents technical data to show the uniqueness of such systems."

Jan 1, 1980

By R D. Skaggs

A good blast is dependent on a well engineered design. However, a good design's performance depends on the control of the drilling and loading of the blast holes. The key stone to good blasting is the drilling of straight holes of the correct orientation in the required positions.

Jan 1, 1991

By Alireza Javadi Nooshabadi, Kota Hanumantha Rao

Formation of hydrogen peroxide (H2O2), an oxidizing agent stronger than oxygen, by sulphide minerals during grinding was examined. It was found that pyrite (FeS2), chalcopyrite (CuFeS2), sphalerite ((Zn, Fe) S), and galena (PbS) generated H2O2 in pulp liquid during wet grinding and also when the freshly ground solids are placed in water immediately after dry grinding. Pyrite produced more H2O2 than other minerals and the order of H2O2 production by the minerals was found to be pyrite > chalcopyrite > sphalerite > galena. The pH of the water influenced the extent of hydrogen peroxide formation with greater amounts of H2O2 produced at highly acidic pH. Furthermore, the effect of mixed sulphide minerals, i.e., pyrite-chalcopyrite, pyrite-galena, chalcopyrite-galena and sphalerite-pyrite, sphalerite-chalcopyrite and sphalerite-galena on the formation of H2O2 showed increasing H2O2 formation with increasing pyrite fraction. There is clear correlation of the amount of H2O2 production with the rest potential of the sulphide minerals; the greater the rest potential of a mineral the greater the formation of H2O2. This study highlights the necessity of revisiting the electrochemical and/or galvanic interactions between sulphide minerals, and interaction mechanisms between pyrite and other sulphide minerals in terms of their flotation behavior in the context of inevitable H2O2 existence in the pulp liquid.

Jan 1, 2014

By Ravi Sahu, Kris Jeremiah

This paper aims to study the benefits of using Post Processed Kinematics(PPK) surveying methodology combined with Drone photogrammetry to achieve highly precise positioning (X, Y, Z) of drill holes based on the drill pattern and to highlight innovative workflow of transmitting the drill plan to smart drills for accurate drilling.

Feb 1, 2020

By E V. Man, O N. Savassi, J-P Franzidis

In flotation, the property that is exploited to effect the separation of a mineral from an ore is referred to as its floatability. Assessment and quantification of ore floatability is of fundamental importance in modelling and simulation of flotation circuits. This paper describes the design, construction and operation of a new laboratory flotation rig that has been developed at the Julius Kruttschnitt Mineral Research Centre (JKMRC) to determine the floatability of different ores on a small-sale. The rig is designed for continuous operation. The major features of the rig are highlighted in this paper. An unique characteristic of the rig is a 9.1 L conditioning vessel. This provides a mixing regime close to plug-flow which avoids complications such as different collector coverages for similarly sized particles (which would considerably increase the complexity of ore floatability determination). The main advantages of this rig when compared to conventional batch flotation cells include better control of the chemical environment and the ability to determine froth recovery, entrainment and water recovery with increased accuracy. Preliminary tests undertaken with a high-grade lead/zinc ore from BHP Cannington are presented demonstrating the good level of reproducibility obtainable in operating the equipment.

Jan 1, 2000

By G. Barbery, P. Dauphin

"In order to obtain reproducible and reproducible results in batch flotation, researchers have tried various approaches, from normalized testing, to modifying existing cells, and even designing new machines. The paper reviews the basic concepts that must be fulfilled in a laboratory flotation machine, and present the FLOTOMATEC design, which results from five years of testing and development. In addition to providing accurate measurements of air flowrate, power consumption, agitation speed, the cell eliminates most operator induced errors. Results of experiments comparing FLOTOMATEC to a conventional cell, as well as reproducibility tests are given. Introduction The use of batch flotation machines in mineral processing research and development is very common. These machines have been used for many years without almost any change in their design, instrumentation or automation. Various authors since Raffinot (1960) have criticized the available designs, especially with respect to the large errors that are introduced by the manual mode of operation. Barbery et al. (1986) have presented a review of the problems involved in laboratory testing in flotation research.Batch laboratory machines should have the following characteristics:"

Jan 1, 1989

By H. R. Green, D. M. Lamb, A. D. Taylor

The design of slurry launders has usually been based on strictly empirical concepts. An examination of the most common procedures reveals that they do not account for many of the variables that are recognized as significant for slurry transport. These may include flow rate, volume concentration of solids, solids specific gravity, solids size distribution, particle shape, launder geometry, and roughness of the wetted surface. It was decided to develop a design procedure, which would accomplish two things:

Jan 8, 1978

By H. R. Green

The design of slurry launders has usually been based on strictly empirical concepts. An examination of the most common procedures reveals that they do not account for many of the variables that are recognized as Significant for slurry transport. These may include flow rate, volume concentration of solids, solids specific gravity, solids size distribution particle shape, launder geometry, and roughness of the wetted surface. It was decided to develop a design procedure, which would accomplish two things: First: Take into consideration most of the known significant variables and systematize the procedure to assure consistent results. Second: Provide a rational basis for examining and utilizing operating data to refine and improve the system. To accomplish this, it was necessary to develop a basic design concept. This concept has been developed through a complex development history, and yet still appears workable and technically sound. That concept may be outlined as follows: First: The solids transport velocity is the fundamental basis for slurry launder design. Second: The stream configuration, hence the launder size, is integrated with the solids transport velocity so that the actual stream velocity exceeds the solids transport velocity. Third: The launder slope is that which will achieve the required actual stream velocity. It was originally planned to utilize a series of charts and nomographs for the design procedure. A nomograph, based on data given in Taggart's Handbook of Mineral Dressing (1), was developed to determine the solids transport velocity. The velocity thus determined was used with Manning charts to determine the launder slope .and configuration as though the fluid were water. It was planned to utilize a "slope adjustment factor" to increase the slope to compensate for the apparent viscosity of the slurry. All attempts to develop a suitable correction factor failed. This failure was principally due to the complex relationship of viscosity to the required slope. It was then decided to revise the entire approach. Specifically, the present procedure is based on the Darcy-Weisbach flow equation, which is rationally preferable to the Manning equation. The Camp (2) equation is used for the solids transport velocity. The necessity of reference to the Moody curves for the Darcy friction factor is avoided by using the Colebrook and White (3) equation. The resulting mathematics require iterative solutions for several of the unknowns.

Jan 1, 1978

By Bernard J. Scheiner, R. J. Wesely, Cy E. Jordan, A. D. Zunkel, Gary M. Wilemon

The U.S. Bureau of Mines has investigated leaching of off-grade western United States phosphate deposits using sulfuric acid (H2SO4) in the presence of methanol. A combination of ore grinding followed by sulfuric acid in methanol leaching extracted over 80 pct of the phosphate. This leaching technique significantly limited the dissolution of iron, aluminum, and magnesium impurities and produced a phosphoric acid that was well within the desired limits for these impurities. The dissolution rate of phosphate rock in the sulfuric acid methanol system was significantly slower than the conventional aqueous sulfuric acid dissolution. Initially dissolution of the phosphate particle was controlled by the amount of surface area available to the sulfuric acid reactant. After about an hour of leaching, phosphate dissolution was slowed almost to a stop, limited by the rate of ion diffusion through the calcium sulfate barrier layer surrounding the unreacted phosphate particle. Consequently, the ore was ground below 38-pm to increase the particle surface area available and improve the overall dissolution kinetics and extraction. In addition, the sulfuric acid was slowly metered into the reactor to prevent any excess heat from the mixing of acid with the methanol from simulating the formation of methyl sulfates. The formation of methyl sulfate ultimately lowered the overall phosphate extraction by exhausting the supply of available sulfate ions for the dissolution reaction. Subsequent evaporation of the methanol from the product acid decomposed the methyl sulfates leaving undesirable sulfate ions in the product acid. Slow acid addition to the reactor provided high phosphate extraction and limited the amount of sulfate in the final product acid.

Jan 1, 1993

Today, the employment of the modern hammer drill, using hollow steel, with air or water to expel the cuttings from the drill hole, is the rule rather than the exception in practically all rock-drilling operations. There is considerable carelessness in the matter of keeping the hole in the drill steel properly opened. Plugged steels wholly or partially, prevent proper operation of the drill, whereby drilling speed is retarded, water tubes are bent or broken, and time is lost in tink-ering, or, as often happens, the drilling gang is idle while someone makes a trip to the shop after another drill steel or water tube. To relieve this condition, the Ingersoll-Rand Company has developed, as an auxiliary to the Leyner sharpener, a new device-the Leyner shank and bit punch.

Jan 4, 1918

By C Simpson, J M. F Clout, G Plummer

Detrital iron deposits (‘DID’) in the Hamersley Province of Western Australia have been successfully mined and processed as small quantities of saleable standalone or blended lump and fine ore products over many decades. Despite their past, present, and potential future exploitation, they have received far less geological attention than the economically more significant higher grade bedded iron deposits (BID) or channel iron deposits (CID). DID are composed of Pliocene to Quaternary unconsolidated colluvial/alluvial hematite gravels with minor sand, silt, and ferruginous clay, as well as minor iron rich hematite clast conglomerates. DID mineralisation at Solomon has been subdivided into six distinct stratigraphic units, DID-5 (lowermost) to DID-0 (uppermost) based on differences in clast and matrix textures and mineralogy, degree of consolidation/lithification, particle size distribution and replacement textures. The most significant accumulations of DID-1 to DID-5 are in tributary valleys that have been cut through the Joffre Member of the Brockman Iron Formation on either side of the main Kings-Queens valley which contains the CID. DID-5 and DID-4 are ferruginous conglomerate, referred to as canga, with four subtypes recognised. DID-4 is composed of BID hematite clasts cemented by hematite or goethite. DID-5 is characterised by partial to complete replacement of BID hematite clasts and matrix by the yellow ochreous and brown to black vitreous forms of goethite. DID-3 to DID-0 are unconsolidated hematite detritals composed of largely (>80 weight per cent) fine to coarse gravel, minor sand and little silt or clay. The gravel clasts are largely composed of hematite and hydrohematite that replace BID, micro-nanoplaty hematite-martite-ochreous goethite (m-nplH-M-oG) BID, reworked canga, hematite pisoids, siliceous hematite BID (micro-nanoplaty hematite-martite-quartz, m-nplH-M-quartz), chert (~10–20 per cent) and minor (<5 per cent) Banded Iron Formation (BIF). Hematite abundance decreases from DID-3 to DID-0 whilst siliceous hematite BID is highest in abundance in DID-0 and lowest in DID-3. A key finding of the study is that >90 per cent of mineralised detrital clasts are derived from hypogene altered BID micro-nanoplaty hematite-martite with subsequent later supergene ochreous goethite mineralisation hosted by the Joffre Member. The DID genesis model entails interpreted Palaeoproterozoic siliceous hematite hypogene microplaty BID mineralisation of the Joffre Member undergoing Cenozoic supergene leaching of quartz and replacement by ochreous goethite to form low phosphorous BID micro-nanoplaty hematite-martite-ochreous goethite clasts. BID martitenanoplaty hematite-martite-ochreous goethite clasts and detritus eroded from hill slopes were deposited in the valleys as an initial single DID-5-DID-4 unit over a weathered pre-CID basal conglomerate. DID-5 and DID-4 underwent goethite cementation of the matrix followed by replacement by hydrohematite, hematite, and maghemite. DID-3 formed during subsequent periods of erosion of largely DID-4 hematite clasts and pisoids. Groundwater flow-through hematite canga resulted in partial to complete hydration replacement of DID-4 hematite clasts and matrix by ochreous and vitreous goethite to form DID-5. Deposition of DID-2 to DID-0 was via reworking and deposition of DID-5 to DID-4 canga and DID-3 as clasts and pisoids/ooids, as well as siliceous hematite and BIF sourced from the Joffre Member in the hill slopes.

Nov 8, 2021

The Kalgoorlie Golden Mile is proposed as an Archaean volcanogenic-type deposit, modified and redistributed along later-developed structures. Rich pyritctelluride and Green Leader gold mineralization is associated with the end phase of basic volcanism and, in a general sense, is confined to a broad interface zone between the products of marine basic volcanism and later acid volcanism. There is virtualIy no ore associated with the main body of acid volcanics.Green Leader is the local name of a type of ore that occurs at some igneous rock contacts in association with graphitic sediments and layered acidic tuffs. It is characteristicalIy green, of various shades, and commonly yields very high gold assays, the reasons for which are not apparent in hand specimen.Carbonates, carbonaceous shales, tuffites, rich gold ores, and acid and intermediate dykes, alI indicating a common heritage, are features common to the main ore hosts, viz. the Paringa Basalt (Calc Schist) and theGolden Mile Dolerite (&quot;Quartz Dolerite Greenstone&quot;).Several probably related fracture systems have reconstituted most, but not all, of the original deposits into a large number of steep tabular lodes.Includes two sets of discussions and comments from later Proceedings volumes.

Jan 1, 1974

By R Berry, N Bonnici, K Nguyen, S G. Walters, M Kamenetsky, C L. Evans

Mineral maps produced via optical and scanning electron microscope û mineral liberation analyser (SEM-MLA) automated mineral identification systems are being used to examine the relationships between mineral phases in spatially well-constrained samples of intact and crushed rock. The use of sophisticated image analysis software allows data pertaining to individual or groups of mineral phases to be extracted from the mineral maps. Analyses of the extracted data can be used to identify populations of phases. Data extracted from the images can be compared with mineral processing data, eg hardness, grindability, degree of liberation, from the same sampling intervals, enabling ore characterisation to be linked to processing behaviours in a spatially constrained, quantified manner that can potentially assist with mine planning and optimisation.

Jan 1, 2008

By E. R. Borcherdt

RECENT years have seen great progress in drilling practices. The partial change from heavy drills and stopers to lightweight airleg machines has had great impact on stoping and drifting methods. The change was sparked, of course, when the Germans applied tungsten carbide to drill bits during World War II. This invention made it possible to drill nearly cylindrical blastholes, with a very mini- mum of drill steel changes, so that small fast-cutting bits* could be used without sacrifice of powder-loading density at the bottom of the hole. This is one of the few instances of increased efficiency through reduction in size. Usually the opposite is true. Application of these lightweight drills and small bits has lessened the vibratory stresses set up in the drill steel rod, and rod breakage has been greatly reduced. During the past several years phenomenal reduction in drill steel breakage has resulted from the use of carburized rods. These rods are now supplied by several manufacturers, with carburization of both the exterior and interior surfaces of the hollow drill steel. The rods are heat-treated over their entire length after the completion of bit connections, collars, or lugs (in the case of 7/8 Hex and round steel). This eliminates breakage from the usual metallurgical notch existing in conventional rods-there are instances of more than 2000 ft of drilling in hard ground before breakage. Steel rod failure results from fatigue checks, at right angles to the rod axis, progressing to a point where parting or breaking of the rod occurs. There are many of these checks in a used rod. Only the most advanced check causes the break, but with a small additional amount of drilling another check advances to the breakage point, and most of the value of reconditioning is lost. Since the development of carburized rods, the practice of discarding drill rods after the first break may be economically justified.

Jan 4, 1957

Current models for the origin of banded iron-formation (BIF) derived in situ iron ore deposits assume that the typical chert and iron oxide banded BIF is the primary rock from which enriched iron ore deposits arise by the selective leaching of quartz and the residual accumulation of the iron oxides. Weathering extends to the base of all the hematite deposits of the Hamersley Province, which lie within the saprolitic weathering zone, as shown by the preservation of the primary textures, and are overlain by a hydrated goethite zone representing the solum. Since groundwater typically contains dissolved silica, derived from the hydrolysis of silicate minerals, greatly in excess of the saturation solubility of quartz, it is incapable of removing quartz from below the watertable, and quartz is preserved in all saprolites derived from quartz bearing rocks. Quartz may be dissolved in the solum due to the influx of meteoric water having little or no dissolved silica, but soil-forming processes that dissolve quartz likewise alter the oxides by hydration and transposition with consequent destruction of primary textures, and there is no suggestion of soil-forming processes having produced the deposits, which typically extend to considerable depths below the water-table. A characteristic feature of the typical hematite orebodies is the unaltered nature of the iron oxides in the ore compared to those in the adjacent cherty BIF. Although the solubility of ferric iron at >10 ppb is much less than quartz, any process capable of dissolving the huge volume of quartz must also affect the iron oxides, but evidence of this is completely lacking in most ore deposits. Hydrothermal alteration at Koolyanobbing in the Southern Cross Province affects both cherty BIF and ore, and clearly shows equal mobilisation and recrystallisation of hematite within quartz/hematite veins and in the surrounding cherty BIF, which shows no evidence for the selective leaching of quartz. Similarly, any metamorphic process capable of removing all quartz from BIF would leave characteristic evidence in the remaining oxides. However, a hydrothermal process may account for at least portions of certain ore deposits, in which the chert bands have been replaced by carbonates at comparatively low temperatures, to produce a quartz-free magnetite/ carbonate rock that is readily upgraded to high-grade ore by supergene dissolution of the carbonates. This process is interpreted to have formed high grade ore at Mt Gibson, Koolyanobbing æAÆ Deposit and Mt Tom Price North Deposit. In addition to carbonate replaced cherty BIF at Mt Gibson in the Murchison Province, there are extensive sequences of magnetite BIF without chert bands, and this chert-free facies of BIF is interpreted as the precursor to the majority of in situ hematite ore deposits, through oxidation and weathering of magnetite and iron silicate and carbonate minerals. The world-class martite/microplaty-hematite deposits of the Hamersley Province are a special case of BIF-derived iron ore deposits that were initiated during the Proterozoic era. Evidence from overlying Wyloo Group conglomeratic horizons containing clasts of Hamersley Group BIF indicates that microplaty-hematite formed within the clasts before deposition, and that Proterozoic subaerial oxidation and weathering of uplifted Hamersley Group formations produced the microplaty-hematite in both chert-free and cherty BIF.

Jan 1, 2002

By R. W. Culbertson

During the 1920&apos;s, approximately 11,000 small gasifiers were in operation in the United States. They converted about 15 million tons of coal per year into low Btu gas. In some cases, hot raw gas was used for applications such as firing open hearth furnaces or brick kilns. In other cases the gases produced in the gasifiers were cooled, cleaned and desulfurized for various utility and industrial purposes. The steel industry was by far the largest user. The most important fact was that the conversion of coal to a usable fuel gas was demonstrated to be commercial. During the late 1940&apos;s and early 1950&apos;s, this form of energy was largely replaced by natural gas as the pipeline net work was expanded. Producer gas was no longer economically competitive. Over the past twenty-five (25) years, government controls on the interstate price of natural gas have maintained this situation. At present only three commercial small coal gasifier plants are still in operation in the United States. These facilities are operated by the Glen-Gery Corporation which is located in eastern Pennsylvania; the National Lime and Stone Company which is located at Carey Ohio; and Holston Defense Corporation which is located at Kingsport, Tennessee. There are, however, numerous operating gasifiers in other parts of the world. Recent events indicate that this situation is changing. The quantity of domestic natural gas available to the interstate market has been declining at the same time that demand for all forms of energy has been increasing. The current potential market for gas with a heating value ranging from 100 to 500 Btu per cubic feet has been estimated at sixteen (16) quads (16 x 1015 Btu) per year. This is expected to grow to twenty (20) quads by the year 2000.

Jan 1, 1978

By Ernest M. Spokes

Starting in 1980, a committee of the National Research Council studied the question of why some underground coal mines are safer than others and what could be done to improve the ones with poorer performance. The results indicated that safety is improved by management commitment to safety, by cooperation between labor and management on safety matters, and by quality training of employees and managers. Younger mines have higher accident frequencies, but are less likely to be killed or permanently disabled by a lost-time accident. Methods for improving safety performance are proposed.

Jan 1, 1983