Bakır flotasyon tesisinin kapalı devre test ile simülasyonu
Simulation of copper flotation plant with locked cycle test results
Yazar: ERDEM ÖZDEMİR
Danışman: PROF. DR. ALİM GÜL
İstanbul Teknik Üniversitesi / Fen Bilimleri Enstitüsü / Cevher Hazırlama Mühendisliği Ana Bilim Dalı / Ceza Hukuku Bilim Dalı
Derece: Yüksek Lisans, 2019
Konu: Maden Mühendisliği ve Madencilik = Mining Engineering and Mining ; Metalurji Mühendisliği = Metallurgical Engineering
Dizin: Köpük flotasyonu = Foam flotation

Özet:
Bakır ve çinko metalleri insanlığın ihtiyacı doğrultusunda birçok endüstride geniş bir kullanıma sahiptir. Bu metaller doğada, belirli kimyasal yapıda çeşitli mineraller halinde rezervlenmektedir. Bu mineraller karbonatlı, oksitli veya sülfürlü yapıda olabilir. Yer kabuğunda bakır ve çinko elementleri çoğunlukla sülfürlü mineraller halinde bulunur. Başlıca sülfürlü bakır minerali kalkopirit (CuFeS2 ), ve çinko minerali sfalerit (ZnS) dir. Ancak tez kapsamında çalışılan numune ek olarak kovelin (Cu2S), kalkozin (CuS) ve bornit (Cu5FeS4)içermektedir. Ayrıca üzerinde çalışılan numunede 1,4 g/t altın (Au) ve 36 g/t gümüş (Ag) de bulunmaktadır. Bu tür kompleks yapıda mineral içeren cevherler belirli bir tane boyutuna indirildikten sonra ayrı ayrı zenginleştirilebilmektedir. Bahsedilen tane boyutu genelde 100 mikronun altındadır ve zenginleştirme için uygulanan yöntem flotasyondur. Flotasyon, genel olarak minerallerin, yüzey özellik farklarını kullanarak belirli kimyasallar sayesinde seçimli olarak yüzdürülmesine ya da bastırılmasına dayalı bir ayırma yöntemidir.
Tez kapsamında % 4,03 Cu,  % 1,2 Zn,  %35 Fe ve %40 S içeren bir bakır çinko cevheri üzerinde çalışılmıştır. Deneylerin amacı ağırlıklı olarak bakır ve çinko içeren numunenin verimli bir şekilde selektif flotasyonudur. Verimli bir ayırma gerçekleştirmek için başta farklı bastırıcıların etkisi olmak üzere, reaktif miktarları, reaktif türleri ve tekrar öğütmenin etkisi incelenmiştir. Tez kapsamında çalışılan numune laboratuvar tipi çubuklu değirmende D80: 38 μm boyuta indirilmiştir. Öğütme,  yaklaşık % 60 pülte katı oranında,  17 dakikada yapılmıştır. 
Flotasyon deneyleri Denver D-12 tip standart karıştırmalı laboratuvar ölçekli bir flotasyon makinasında gerçekleştirilmiştir. Deneylerde, 1 kg numune için 2,5 L; 2 kg numune için 4,5 L flotasyon hücresi kullanılmıştır. Flotasyon deneyleri sırasında 2,5 L 1100 d/dak.; 4,5 L için 1350 dev/dak. için karıştırma hızı uygulanmıştır. Flotasyon deneylerinde öncelikli olarak bakır ve çinko mineralleri kaba devrede alınmıştır. Kaba devrede gerçekleştirilen kinetik testlerde, yüksek tenörde bir ön konsantre alınabilirliği ortaya koyulmuştur. Ön ve kaba konsantre rejimlerinin belirlenmesinin ardından çeşitli miktarlarda çinko sülfat ile birlikte sodyum metabisülfit eklenerek selektif bakır çinko flotasyonu yapılmıştır. Selektif bakır çinko devresinde  çok ince saçınımda çinko-bakır bağlı taneleri varlığından dolayı çinko mineralleri bastırılamamıştır. Kaba flotasyon devresinde tüm deneyler pH 12'de yapılmıştır. Farklı türde toplayıcı karışımlarının etkisinin incelendiği deneylerde en yüksek konsantre bakır verim ve tenörü, ditiyofosfin ve ditiyonokarbamat türü toplayıcı karışımlı deneyden elde edilmiştir. Sonraki deneylerde ön ve kaba flotasyon aşamasında bu toplayıcı türü karışımı kullanılmıştır. Daha sonra kaba konsantreden alınan numuneler mikroskopta incelenmiş ve görüntülerde yüksek oranda pirit- kalkopirit bağlı taneleri tespit edilmiştir. Bu bağlı taneleri serbestleştirmek amacıyla kaba konsantreye tekrar öğütme uygulanmıştır. Tekrar öğütmede 20 mikronun altına öğütülen mazlemeye çinko sülfat eklenerek kısmen serbestleşmiş çinko mineralleri bastırılmıştır. Ayrıca pH 12,3'ün üzerinde çıkarılarak kısmen serbestleşmiş piritlerin bastırılması hedeflenmiştir. Daha sonra öğütülen malzemeden iki kademe temizleme ile nihai konsantre elde edilmiştir. Ayrıca birinci temizlemenin artığına, bir kademe süpürme yapılarak sistemden artık olarak atılmıştır. Deneylerde ditiyofosfin, ditiyanokarbamat ve sodyum izopropil ksantat toplayıcıları kullanılmıştır. Optimum sonuçlar ditiyofosfin ve ditiyonokarbamat karışımından elde edilmiştir. Bu toplayıcı karışımıyla yapılan iki temizlemeli kapalı devre deneylerde %16,93 Cu tenöründe bakır konsantresi yaklaşık %80 verimle elde edilmiştir. Kapalı devrelerden elde edilen ham verilerden excel simülasyonu yapılmıştır. Yapılan simülasyon çalışmaları sonucunda akım şeması üzerinde her akışta tonaj ve tenör değerleri hesaplanmıştır. Bu tonaj ve tenör değerlerine göre flotasyon hücre hacimleri ortaya konulmuştur. Tesiste kurulan flotasyon devresi üzerinde yapılan kütle balansı, kapalı devre testlerden elde edilen simülasyon tenör ve tonaj değerlerine yakın çıkmıştır.

Abstract:
Industrially; copper and zinc have a wide area of use. These metals are reserved in compounds. These compounds may contain oxide, carbonate or sulphide-type ore. Copper and zinc metals are often found in sulphide-type on the earth. Sulphide-type copper mineral is generally chalcopyrite (CuFeS2), zinc mineral is sphalerite (ZnS). Aditionally the sample, contains covellite (CuS),  chalcosite (Cu2S) and, bornite (Cu5FeS4). Also, the sample contains traces of gold 1,40 g/t (Au) and silver 36 g/t (Ag). For these type of ores particle liberation happens usually under 100 microns and the applied method is flotation. Flotation is a separation method by floating or depressing minerals according to their surface properties using certain chemicals.
The sample contains % 4,03 Cu, % 1,4 Zn, 35% Fe, and 40% S. The aim of experiments was to efficiently selective flotaion of the copper and zinc minerals. In order to accomplish an efficient separation, reagent dosages, effect of different collector types mixture, effect of depressant and regrind effect were investigated. 20 different representative drilling core sample were crushed to under 2 mm. Crushed material were seperated into 1 kg test pockets. Pockets were stored in the deep freeze in -17˚C for the purpose of prevention any oxidation. During the grinding tests, the sample was ground in the laboratory-type rod mill to the P80:38 microns. Grinding experiments was applied at approximetly 60 % pulp solid contents with the different grinding time. After that, grinding time was calculated as 17' from grinding time versus P80 graph.
Flotation experiments were carried out with Denver D-12 type conventional lab-scale flotation machine. During the flotation experiments, for 1 kg sample 2,5 L cell volume; 2 kg sample 4,5 L cell volume were used. Flotation tests were conducted with the rotor speed of 1100 rpm and 1350 rpm for 2,5 L and 4,5 L. Laboratory type standart hydrogen electrode pH meter were used for the pH measurement. Firstly, copper and zinc minerals were taken at rougher flotation section at pH 12. According to rougher kinetic tests results, high grade of preconcentrate can be obtained. After the determination of preflotation and rougher flotation conditions, effect of depressants were investigated. Effect of depressant experiments showed that there is no selectivity between zinc and copper minerals so that it was focused only to produce copper concentrate. pH was set to 12 during the preflotation and rougher flotation and 100 g/t aerofine 3418A and 50 g/t XD5002 were added. Furthermore, rougher concentrate was prepared for the mineralogical analysis. Microscope images showed that there are still locked and dissaminated pyrite and chalcopyrite minerals at P80:38 microns. For the liberation these kind of locked particles, rougher concentrate were reground to -20 microns. Next, reground material were cleaned with two cleaner stages. Scavenger stage were applied on first cleaner tailings and scavenger tailings were taken as a fine tailings.
For the first cleaner flotation stage zinc sulphate was added as a sphalerite depressant. In the cleaner stages pH was set to around 12,3 for the purpose of depression of liberated pyrite minerals. 
There are more than five locked cycle flowsheet applied on the ore but one has reliable and stable results. The most reliable results were taken from two concentrate and two tailings staged flowsheet. Locked cycle tests were carried out with eight cycle on the sheet. In the first cycle, crushed material was fed into laboratory rod mill for the grinding material P80: 38 micron. Next, ground material were put in to laboratory flotation cell for conditioning. During the condition, pH was set to 12 then 25 g/t aerofine 3418A and 25 g/t MIBC were added. After the five minutes conditioning time, preconcentrate was taken for two minutes. Furthermore, preconcentrate was cleaned with one cleaner stage and filtered as a first cycle preconcentrate. Next, preconcentrate cleaner tailings were mixed with preconcentrate tailings and added 75 g/t Aerofine 3418A and 50 g/t XD5002. After the conditioning, rougher concentrate were taken for ten minutes. Rougher tailings were filtered and named as a first rougher tailings. Moreover, rougher concentrate were reground -20 μm with laboratory type rod mill then fed into first cleaner stage. In the first cleaner stage 5 g/t aerofine 3418A and 5 g/t XD5002 were added. First cleaner stage was carried out in eight minutes. Scavenger flotation were applied on first cleaner tailings. First cleaner scavenger tailings were filtered as first cycle fine tailings. First cleaner concentrate were cleaned with second cleaner section. Second cleaner concentrate were filtered and named as first cycle concentrate. All cycle was held on with this order for the purpose of determination of middlings behaviour. In the initial cycles, accumulation was about 450 g. After the sixth cycle, system  has become stable and accumulation has reaced almost zero point. According to optimum flotation conditions of determined flowsheet, the seventh cycle tests were realized and final copper concentrates were obtained by 16,93 % Cu grade with a recovery of 80%. When system has balanced, raw mass and grade data were processed on the excel simulation. Material balance was calculated using by metallurgical equation. Moreover, flotation circuit cell volume were calculated with using excel simulation. Calculated cells amount were installed to the plant. During the comissioning which was part of the thesis work, some plant trials were applied to the flotation circuit. First of all, flotation streams were investigated and determinated sample points. Plant trial conditions were set to be constant as locked cycle parameters. Sampling was carried out approximately eight hours. Samples were taken from flotation circuit with the time period of 30 minutes. Taken samples were stored in a bucket for the purpose of representativeness. Furthermore, taken sample has reduced, filtered and dried in a oven. Dried sample were prepared for the chemical analysis. Chemical analysis was calculated with ICP method. Flotation circuit mass balance has calculated using by plant trials chemical assays of the plant trial samples. When the comparison of the locked cycle test simulation and mass balance calculation, the main streams such as rougher tailings, preconcentrate and final concentrate were more similar with simulation. On the other hand, differences between simulation and plant mass balance were higher in first cleaner, second cleaner and scavenger sections. The more streams means the more differences between simulation and plant mass balance. 
As a result, plant surveys proved that the simulated mass balance and copper grade results were similar as plant's mass balance and grades. Locked cycle test results give lots of informations such as mass pull amount, grade, conditioner volume, flotation cells volume, tickener area, tailing dam capacity etc. Locked cycle test data deals with not only plant design problems but also current flotation plant performance. Locked cycle flotation experiments are useful method for these type of problematic ores.